DRAFT Environmental Impact Assessment Kupol Gold Project …documents.worldbank.org/curated/en/182011468296685218/... · 2016-07-19 · DRAFT Environmental Impact Assessment Kupol

E1200, Vol.1

DRAFT Environmental Impact Assessment

Kupol Gold Project Fareast Russia

June 2005

Prepared by:

BEMA GOLD CORPORATION

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Table of Contents 1.0 Executive Summary 1

1.1 Geology..................................................................................................................... 1 1.1.1 Exploration........................................................................................................ 3 1.1.2 2005 Exploration Program................................................................................ 3 1.1.3 Resource Statement........................................................................................... 3

1.2 Mining....................................................................................................................... 4 1.2.1 Open Pit ............................................................................................................ 4 1.2.2 Underground ..................................................................................................... 5 1.2.3 Waste rock ........................................................................................................ 6

1.3 Process Description................................................................................................... 7 1.4 Tailings Facilities...................................................................................................... 9 1.5 Ancillary Facilities.................................................................................................. 10

1.5.1 Access Roads .................................................................................................. 10 1.5.2 Airport facilities .............................................................................................. 11 1.5.3 Mill and Services Building ............................................................................. 11 1.5.4 Permanent Camp............................................................................................. 11 1.5.5 Power Generation............................................................................................ 12 1.5.6 Fuel Storage and Distribution ......................................................................... 12 1.5.7 Wastewater treatment plant............................................................................. 12

1.6 Environmental......................................................................................................... 13 1.6.1 Baseline Information....................................................................................... 13

1.7 Impact Summary..................................................................................................... 14 1.8 Environmental and Social Action Plan (ESAP)...................................................... 17 1.9 Health and Safety Plan............................................................................................ 17 1.10 Spill Prevention, Control, and Countermeasures Plan (SPCC) .............................. 18 1.11 Waste Management Plan (WMP) ........................................................................... 19 1.12 Reclamation and Closure ........................................................................................ 19 1.13 Monitoring .............................................................................................................. 20 1.14 Environmental and Social Management (EH&S)................................................... 21 1.15 Public Consultation and Disclosure ........................................................................ 21 1.16 Charitable Foundation............................................................................................. 22 1.17 Implementation Schedule........................................................................................ 22

2.0 Policy, Legal, and Administrative Framework 24

2.1 Purpose and Need ................................................................................................... 24 2.2 Project Investment .................................................................................................. 24 2.3 Legal Matters .......................................................................................................... 25

2.3.1 License for Exploration and Production ......................................................... 25

Kupol Environmental Impact Assessment i

2.3.2 Environmental Covenants............................................................................... 27 2.3.3 Occupational Health and Safety...................................................................... 28

2.4 Taxation .................................................................................................................. 28 2.5 Russian Permitting Requirements........................................................................... 29

2.5.1 Russian Permitting Agencies .......................................................................... 29 2.6 Russian Permitting Process..................................................................................... 32

2.6.1 Preliminary Assessment.................................................................................. 32 2.6.2 Feasibility Studies........................................................................................... 32

2.7 IFC/EBRD/WBG Environmental Assessment Requirements ................................ 35 2.7.1 OP 4.01 Environmental Impact Assessment................................................... 35 2.7.2 OP 4.04 Natural Habitats ................................................................................ 35 2.7.3 OP 4.11 Management of Cultural Properties.................................................. 35 2.7.4 Assessment and Management of Cumulative Impacts (DRAFT)................... 36 2.7.5 Hazardous Management Guidelines ............................................................... 36 2.7.6 Public Consultation and Disclosure ................................................................ 36 2.7.7 OP 4.10 Indigenous Peoples (forthcoming).................................................... 36 2.7.8 OP 4.37 Safety of Dams.................................................................................. 36 2.7.9 Non-applicable policies .................................................................................. 36

2.8 Public Consultation and Disclosure ........................................................................ 37 3.0 Project Summary 38

3.1 Mining..................................................................................................................... 38 3.1.1 Open Pit Mining.............................................................................................. 38 3.1.2 Underground Mining ...................................................................................... 40

3.2 Milling..................................................................................................................... 40 3.2.1 Ore Stockpiling ............................................................................................... 41 3.2.2 Ore Receiving/Coarse Ore Storage................................................................. 41 3.2.3 Primary Crushing ............................................................................................ 41 3.2.4 Crushed ore storage......................................................................................... 42 3.2.5 Grinding and Classification ............................................................................ 42 3.2.6 Gravity Concentration..................................................................................... 43 3.2.7 Grinding Thickener and Leach Circuit ........................................................... 43 3.2.8 Counter Current Decantation (CCD) Circuit .................................................. 44 3.2.9 Merrill Crowe and Refining............................................................................ 45 3.2.10 Cyanide Destruct............................................................................................. 46

3.3 Reagent Handling and Mixing ................................................................................ 47 3.3.1 Sodium cyanide (NaCN)................................................................................. 48 3.3.2 Lime (CaO) ..................................................................................................... 48 3.3.3 Flocculant (AF-305)........................................................................................ 49 3.3.4 Lead nitrate Pb(NO3)2 ..................................................................................... 49 3.3.5 Zinc dust.......................................................................................................... 49 3.3.6 Antiscalant ...................................................................................................... 49 3.3.7 Sulfuric Acid................................................................................................... 49 3.3.8 Sodium Hydroxide .......................................................................................... 50 3.3.9 Copper Sulfate (CuSO4*5H2O)....................................................................... 50

Kupol Environmental Impact Assessment ii

3.3.10 Calcium hypochlorite...................................................................................... 50 3.3.11 Diatomite......................................................................................................... 50 3.3.12 Borax, Fluorspar, Soda Ash and sodium nitrate ............................................. 50

3.4 Kupol Mill Water Balance ...................................................................................... 50 3.5 Tailings Facility ...................................................................................................... 52 3.6 Ancillary Facilities.................................................................................................. 54

3.6.1 Explosives Storage.......................................................................................... 55 3.6.2 Mancamp......................................................................................................... 56 3.6.3 Chemical Storage ............................................................................................ 59 3.6.4 Fuel and Lubricant Storage............................................................................. 60 3.6.5 Power and Heat Supply................................................................................... 61 3.6.6 Facility Water Supply ..................................................................................... 62 3.6.7 Tailings Line ................................................................................................... 62 3.6.8 Wastewater Treatment Plant ........................................................................... 62 3.6.9 Solid Waste Disposal Area ............................................................................. 63 3.6.10 Pevek Service Center ...................................................................................... 64

3.7 Transportation ......................................................................................................... 64 4.0 Baseline Conditions 66

4.1 Topography............................................................................................................. 66 4.2 Geology................................................................................................................... 67

4.2.1 Property Geology............................................................................................ 69 4.3 Permafrost ............................................................................................................... 71 4.4 Climate.................................................................................................................... 71 4.5 Surface Water Hydrology ....................................................................................... 73 4.6 Surface Water Quality............................................................................................. 75 4.7 Vegetation ............................................................................................................... 76 4.8 Soils......................................................................................................................... 79 4.9 Seismicity................................................................................................................ 79 4.10 Fauna....................................................................................................................... 81 4.11 Hydrobiology .......................................................................................................... 82 4.12 Ichthyology ............................................................................................................. 83 4.13 Archeology.............................................................................................................. 83 4.14 Protected Areas ....................................................................................................... 83 4.15 Land Use ................................................................................................................. 84 4.16 Socioeconomics ...................................................................................................... 84

4.16.1 Demographics ................................................................................................. 85 4.16.2 Health and Welfare ......................................................................................... 89

4.17 Economic Baseline Conditions ............................................................................... 90 4.17.1 Russia.............................................................................................................. 90 4.17.2 Russian Far East.............................................................................................. 91 4.17.3 Chukotka Autonomous Okrug ........................................................................ 91

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5.0 Alternatives 97

5.1 No Action Alternative............................................................................................. 97 5.1.1 Socioeconomic Impacts .................................................................................. 97 5.1.2 Environmental Impacts ................................................................................... 98 5.1.3 Justification for rejecting the No Action Alternative...................................... 98

5.2 Mining..................................................................................................................... 98 5.2.1 Justification of rejecting large open pit alternative......................................... 99

5.3 Tailings Basin Alternatives................................................................................... 100 5.3.1 Environmental Impacts ................................................................................. 100

5.4 Site Power ............................................................................................................. 101 5.4.1 Socioeconomic Impacts ................................................................................ 101 5.4.2 Environmental Impacts ................................................................................. 101 5.4.3 Justification for choosing on-site power ....................................................... 102

6.0 Impacts 103

6.1 Air Quality Impacts............................................................................................... 103 6.1.1 Stationary Sources ........................................................................................ 104 6.1.2 Fugitive Sources............................................................................................ 104

6.2 Topography and Land Disturbance....................................................................... 105 6.2.1 Existing impacts............................................................................................ 106 6.2.2 Construction impacts .................................................................................... 106 6.2.3 Operational surface disturbance.................................................................... 106

6.3 Soil Impacts .......................................................................................................... 107 6.3.1 Acid Rock Drainage...................................................................................... 107 6.3.2 Potential ARD Impacts ................................................................................. 110

6.4 Permafrost ............................................................................................................. 110 6.5 Hydrogeological Impacts ...................................................................................... 113 6.6 Surface Water Quality Impacts............................................................................. 113 6.7 Vegetation Impacts ............................................................................................... 115 6.8 Wildlife Impacts.................................................................................................... 115 6.9 Aquatic Biology Impacts ...................................................................................... 116 6.10 Aesthetics/Visual Resources................................................................................. 116 6.11 Noise Impacts........................................................................................................ 116 6.12 Socioeconomic Impacts ........................................................................................ 117

6.12.1 Population and Demographics ...................................................................... 118 6.12.2 Housing ......................................................................................................... 118 6.12.3 Economic Impact .......................................................................................... 118 6.12.4 Indigenous Population .................................................................................. 119

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7.0 Bibliography 120

List of Tables Table 1. Exploration Programs ........................................................................................... 3 Table 2. Indicated Resource, All Veins, All Zones ............................................................ 3 Table 3. Inferred Resource, Vein, All Zones ...................................................................... 4 Table 4. Open Pit Production Schedule .............................................................................. 5 Table 5. Underground Production Schedule ....................................................................... 6 Table 6. Estimated Mill Reagent Consumption .................................................................. 9 Table 7. Impact Matrix for the Kupol Project................................................................... 15 Table 8. Cost and implementation schedule for Kupol EMP ........................................... 23 Table 9. Cost and implementation schedule for Kupol SMP............................................ 23 Table 10. Operating Costs (USD). ...................................................................................... 24 Table 11. Regulatory Waste Classification......................................................................... 34 Table 12. Production Sequence for Probable Underground Reserve.................................. 40 Table 13. Estimated Mill Reagent Consumption ................................................................ 47 Table 14. Description of the reagents and packaging stored at reagent storage ................. 59 Table 15. Regulatory Waste Classification......................................................................... 63 Table 16. Annual Solid Waste Generation.......................................................................... 64 Table 17. Exploration Programs ......................................................................................... 66 Table 18. Monthly Runoff Coefficients Kupol................................................................... 74 Table 19. Summary of estimated background water quality in the Kaiemraveem River and

its contributing catchments near the project site............................................................. 75 Table 20. Russian Inflation and GDP (real % growth) ....................................................... 90 Table 21. ChAO Industrial Output (in million Rubles) ...................................................... 92 Table 22. Industrial Output – Food Industry....................................................................... 93 Table 23. Agricultural Output of ChAO. ............................................................................ 93 Table 24. Import-Export Structure of the Chukotka Autonomous District ........................ 93 Table 25. Chukotka Autonomous District Quality of life of the population, 2000-2003 ... 94 Table 26. Emissions from diesel generators ..................................................................... 104 Table 27. Land disturbed during exploration.................................................................... 106 Table 28. New disturbance occurring during construction ............................................... 106 Table 29. Proportions of acid-generating, uncertain, and acid-neutralizing waste rock in the

ARD domains from waste rock samples of the ARD database .................................... 108 Table 30. Model predicted tonnes of waste rock mined from the open pit and underground

and not used for back fill. ............................................................................................. 109 Table 31. Noise Levels from various sources of Noise pollution..................................... 117

List of Figures

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Figure 1. Location Map of the Kupol Deposit ..................................................................... 2 Figure 2. Kupol Mill Process Flow Sheet. ........................................................................... 8 Figure 3. Bema Gold Kupol Deposit Acquisition Flow Chart ........................................... 26 Figure 4. Government of the Russian Federation............................................................... 31 Figure 5. Overall Site Layout ............................................................................................. 39 Figure 6. Kupol Mill Water Balance. ................................................................................. 51 Figure 7. Mancamp Layout ................................................................................................ 57 Figure 8. Kupol Regional Geology .................................................................................... 68 Figure 9. Regional Geology ............................................................................................... 70 Figure 10. Vegetation map ............................................................................................... 78 Figure 11. Soil map .......................................................................................................... 80 Figure 12. Location of the project in relationship to the larger geographical boundaries ...

......................................................................................................................... 86 Figure 13. Typical Permafrost Zone............................................................................... 112

List of Appendices Appendix A Baseline Data Appendix B Geochemical Characterization of Waste Rock and Ore Materials Appendix C AMEC Tailings Report Appendix D AMEC Wasterock Management Report Appendix E AMEC Site Water Balance Appendix F AMEC Site Water Management Plan Appendix G AMEC Surface Water Quality Modeling Report

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1.0 Executive Summary On December 18, 2002, Bema Gold announced that it had completed the terms of a definitive agreement with the Government of Chukotka, an autonomous Okrug (region) in northeast Russia, to acquire up to a 75% interest in the Kupol gold and silver project (the “Kupol Deposit”). The Kupol Deposit is located in the Northwest part of the Anadyr foothills on the boundary between the Anadyr and Bilibino Regions in the Chukotka Autonomous Okrug. The geographical coordinates for the site are 66°47’northing and 169°33’ easting. The location of the deposit is shown in Figure 1. Bema Gold completed a Preliminary Assessment of the property in 2004. The results of this assessment indicated that the project should advance to the feasibility level. The feasibility and corresponding Environmental and Social Impact Assessment (ESIA) were completed in June 2005. The results of this feasibility study indicate that the development and production of precious metal ores from the Kupol Deposit is feasible. The initial production life of the deposit is approximately seven years based on Probable Mineral Reserves, with mill production beginning in June 2008. The exploitation of the inferred mineral resources may extend the life of the project.

1.1 Geology Regional and local geology, exploration completed by others and Bema, data collection and management of the database are described in Section 4 of this report. The Kupol deposit is located in a 3000 km-long Cretaceous volcanogenic belt and is centered within a ten-kilometer wide caldera. The succession of volcanic rocks in the area is 1300 meters thick and is comprised of a lower sequence of felsic tuffs and ignimbrites, a middle sequence of andesitic to basaltic-andesitic flows and fragmentals capped by felsic tuffs and flows. Mineralization is associated with a north-south trending, steeply east dipping (75-90 degrees) fault splay of the Kaiemraveem regional fault of similar orientation. The andesitic volcanic units and mineralization have been intruded by massive to weakly banded rhyolite dykes, rhyolite and dacite flow-dome complexes, and basalt dyke. Gold and silver mineralization is hosted by colloform to crustiform-banded quartz-adularia veins and polyphase breccias. Seven vein textures were logged including massive vein, banded colloform and crustiform veins, vein breccia, quartz breccia, stockwork, stingers, wall rock breccia and yellow siliceous breccia. Polyphase brecciation within the vein system is believed to be principally hydrothermal. The predominant gold and silver minerals are electrum, native gold, silver-rich tetrahedrite (freibergite), acanthite, and a several sulphosalt minerals. Gold occurs within or is rimmed by sulphosalts. Kupol is categorized as a low sulphidation fissure vein type epithermal deposit or a quartz-adularia-sericite type epithermal deposit.

Kupol Environmental Impact Assessment 1

Figure 1. Location Map of the Kupol Deposit


1.1.1 Exploration The exploration history at Kupol starts with the discovery of quartz veins in 1966 by a Soviet-sponsored regional mapping program. In 1995, the main Kupol system was found. From 1995 to 2001, the state-funded Anyusk Geological Expedition completed about 3,000 meters of drilling and about 8,000 linear meters of trenching. Bema acquired the property in December 2002, and completed extensive drilling, trenching and testing of the deposit in 2003 and 2004. The sampling programs are summarized in Table 1.

Table 1. Exploration Programs Year No. of Diamond

Drill Holes Drill Hole Meters (m)

No. of trenches

Trench Meters (m)

No. of Channel Samples

Channel Sample Meters

2003 166 22,200 15 2500 -- -- 2004 309 53,800 2 226 87 700

1.1.2 2005 Exploration Program Infill drilling and exploration drilling along strike, at depth and on the flanks of the deposit will be completed in the 2005 drill season. A total of 45,000 meters of core drilling using one CKB-4 Russian drill, three Longyear 44 drills and two Longyear 38 drills is planned. Drilling will be focused on better defining the geological model and increase confidence in grade estimation, increase drill density as required for Russian reserve reporting, and additional condemnation and metallurgical drilling. Stripping, mapping and sampling will be conducted in the South and North zones. 1.1.3 Resource Statement Indicated and Inferred Resources are summarized in Tables 2 and 3, respectively, by gold grade cutoff. Mineable reserves are included in these tabulations. Complete sets of grade- tonnage tables and graphs by resource classification and zone are included in Section 5.8 of this report.

Table 2. Indicated Resource, All Veins, All Zones Au (g/t) Cut off

Tonnes X(1000) Au(g/t) Au (Ounces X 1000)

Ag(g/t) Ag (Ounces X 1000)

0 7,661 17.52 4,317 222.08 54,702 2 7,351 18.22 4,305 230.52 54,481 4 6,865 19.29 4,258 244.06 53,866 6 6,403 20.33 4,184 257.02 52,911 8 5,673 22.02 4,017 277.86 50,677

10 4,942 23.95 3,806 300.45 47,741 12 4,224 26.16 3,553 326.25 44,311 14 3,616 28.38 3,299 351.50 40,859 16 3,069 30.76 3,035 377.96 37,288


Table 3. Inferred Resource, Vein, All Zones Au (g/t) Cut off Tonnes X(1000) Au(g/t) Au

(Ounces X 1000) Ag(g/t) Ag (Ounces X 1000)

0 7,033 8.52 1,928 120.46 27,240 2 6,049 9.75 1,895 136.47 26,540 4 5,105 10.97 1,801 152.91 25,096 6 4,090 12.45 1,637 171.39 22,539 8 3,093 14.18 1,410 191.18 19,010

10 2,293 15.99 1,178 208.35 15,357 12 1,659 17.90 955 223.05 11,895 14 1,141 20.21 742 238.91 8,768 16 840 22.10 597 250.82 6,773

1.2 Mining The mine planning was completed by Bema with assistance from Wardrop Mining and Minerals. The probable reserve was completed by Bema Gold staff. The mining report is provided in Section 6 of the Feasibility Study. The open pit will deliver 1.42 million tonnes over a mine life of 4 years at an average grade of 20.4 g/t gold and 193 g/t silver. The underground will deliver 5.7 million tonnes over a mine life of approximately 7 years at a grade of 16.0 g/t gold and 219 g/t silver. The production schedule has the open pit and underground mines operating at the same time. 1.2.1 Open Pit The ultimate pit depth in the Big Bend area will reach approximately 100 meters and the pit depth for the majority of the Central and North zone areas will be approximately 45 to 50 meters. From current thermistor data it is known the entire pit is within permafrost and therefore, does not include any groundwater influences in the pit slope design. Final bench geometry of 24-meter bench height (70º bench face angle) with a minimum bench width of 10 meters can be successfully and safely mined. The average overall slope angle of the pit varies depending on geotechnical parameters but averages 50 degrees. The selectivity for the open pit is constrained by the minimum mining width of 2 meters. The Kupol open pit will be mined as a standard truck/loader operation, with the crusher located at the processing plant, and the waste dump located approximately 2 km to the south at the tailing impoundment. The strip ratio 12:1 is consistent over the life of the pit. Loaders (4.3 m³) will be the main loading units in waste and will be used for loading a high percentage of the ore. The loader units required for the open pit mining effort are presently on site at Kupol and are being used in the construction effort. Excavators (4.3 m³) will be used in the ore grade control efforts in the pit. The excavators will clean the waste from the top of the shot vein under the guidance of grade control personnel. Next the excavators will pull the ore from the face of the bench either loading an available truck or placing to the side


for later loading into a truck. The excavators will clean the ore until the footwall waste is reached. The loaders and excavators are matched with 35-tonne mining trucks that will be used for both waste and ore haulage. The trucks used for the construction effort will be used for ore and waste haulage from the pit. Based on cycle times for average haul distances to the crusher and waste dump, and on 340 scheduled days per year, this will require a fleet of 9 trucks. Drilling requirements in waste will be handled by one large diesel powered rig equipped with a drill capable of single-pass drilling of 6 m benches with an additional .9 m for sub-grade, with a hole size of 140 mm. In addition, two ECM diesel-powered drills are available for the open pit operations. These smaller drills are presently being used at site for the construction effort. The production schedule for the open pit assumes mining commences in 2006 to provide construction materials for the tailing impoundment and other structures. Ore from the open pit will be mined for four years, 2007 through 2010. Table 4, Open Pit Production Schedule, shows the amount of material mined and the average grades per year from the open pit. The production schedule assumes that the pit operators work 340 days per year on two 11-hour shifts.

Table 4. Open Pit Production Schedule Year 2007 2008 2009 2010 Total

High Grade (tonnes) Au Grade (g/tonne) Ag Grade (g/tonne)

178,628 21.2 182

229,895 26.8 230

268,454 36.9 352

273,222 22.7 229

950,199 27.4 255

Low Grade (tonnes) Au Grade (g/tonne) Ag Grade (g/tonne)

117,030 6.3 57

119,304 6.3 64

56,194 6.8 69

182,968 6.5 79

475,496 6.5 69

1.2.2 Underground The Kupol mineralization is typical of a high-grade vein hosted deposit: single and multiple veins; globally continuous mineralized vein with localized discontinuous segments and variable thickness. Veins on surface are extremely well exposed. Over 75% of the veins at the surface are washed clean and mapped extensively. Two nearly 70-meter sections of the vein are drilled on a very tight spacing (10 meters along strike and 5 meters along dip). A detailed drilling program in 2004 and detailed surface mapping and sampling quantified local vein geometry. From this information ore loss and dilution criteria are developed for the underground mine. The initial stope layout is done using an 8 g/t undiluted grade and a 20 gram/meter (grade thickness) criteria. During the 2004 drilling program several thermistors were installed. Permafrost exists to at least a depth of 250 meters, well below the depth of the feasibility mine plan. Based on the geometry of the mineralization and the results of the geotechnical studies, longhole stoping is the mining method:

• Sills are driven on 15-meter spacing approximately 4 meters high. • Longhole stopes (panels) are drilled between the sills (approximately 11 meters).


• Stopes are filled with waste rock when required. • No sill pillars are left between mining fronts. A concrete sill pillar is constructed on

the first (lowest) sill cut of a mining front if there is an expectation ore will be mined up to this sill from below.

Two major declines driven 5 meters wide by 5.5 meters high exploit the underground reserves, one in the south end of the mine and one in the north. Development of the south decline begins in 2006 and the north decline system begins in 2008. Table 5, Underground Production Schedule, shows the amount of material mined and the average grades per year.

Table 5. Underground Production Schedule Mining location

S=South, N=North Production

(tonnes) Au Grade (g/tonnes)

Ag Grade (g/tonne)

2007 (S) 71,879 23.3 290 2008 (S) 234,821 24.6 312 2009 (S & N) 648,173 19.3 190 2010 (S & N) 902,172 16.7 235 2011 (S & N) 1,043,364 16.7 223 2012 (S & N) 1,004,288 15.7 219 2013 (S & N) 1,043,003 12.6 213 2014 (S & N) 713,504 13.2 193

The waste cross-cut access off the main decline will be 5 m by 5 m to the vein therefore allowing the use of trucks to deliver backfill into the stope. Development in ore will be 4 m to 5 m high and the full width of the ore. All development will use two boom jumbos. The backfill cycle is an integral part of the production cycle and on an annual basis approximately 1500 tonnes per day of backfill is required to be placed to maintain the underground production schedule. It must be noted the last sill cut and associated panels at the top of a mining front will likely not need to be filled. Backfill will be a combination of run-of-mine waste either directly from underground development (50% of backfill requirements), with open pit waste (preferentially taking acid-generating or potentially acid-generating material) or waste from a borrow source located on surface (unlikely). The waste for backfill that is obtained from the surface sources may need to be sized to below 0.3 m. Waste from the surface sources will be trucked with the open pit mine trucks to a stockpile area near the portal, then reloaded onto the underground mine trucks (40 tonnes) and back-hauled to the stopes requiring backfill. Backfill requirements first exceed waste from underground development in 2010. In 2010 the open pit operation is coming to an end and surface trucks are therefore available. Ore and waste haulage will be accomplished using 40-tonne articulated trucks. Development of declines and access ramps will be completed using 8 yd3 LHD’s. 1.2.3 Waste rock The total amount of waste produced is approximately 18.6 million tonnes. Mine waste rock will be generated primarily by the open pit (16.9 million tones or 91%). Underground mine


development waste totals 1.7 million tones. Up to 3.3 million tones of the total waste may be used to fill underground stoping areas. The tailing impoundment will be constructed of non-acid and potentially acid-generating waste. Potentially acid generating waste rock may be used only in the core of the dam where thermal modeling indicates very fast freezing of this material. In addition, the liner on the face of the dam greatly reduces the possibility of any water reaching the potentially acid-generating material used in the core. Acid-generating mine waste will be placed in the tailing impoundment basin (upstream of the impervious liner). The mine waste will eventually become covered by tailings or water and freeze very soon after the end of mine life. An extensive wasterock characterization program was developed during 2003 and 2004. Approximately 276 samples were selected for acid generation potential and other geochemical testing. The results of the geochemical characterization indicated that approximately 42% of the material may be acid-generating, approximately 47% of the material may be non-acid generating, and the remaining portion (~11%) is potentially acid-generating and must be classified in the field. The non-classified material has been assumed to be non-acid generating for this study. The assay will be done by relatively fast index sulfur testing, such as the Leco S test and tests for neutralization potential. Following testing, blasted waste materials will be flagged to identify their ARD classification, and haul trucks directed to the appropriate dump location in the tailings impoundment. Non-acid generating waste will be used, for example, for any general site fill requirements during construction and operation.

1.3 Process Description The mill is a conventional gold/silver cyanidation plant that will incorporate a CCD thickener washing circuit and Merrill-Crowe zinc precipitation because of the high silver ore grade. There will be a cyanide recovery circuit using AVR (acidification-volatilization-recovery) and cyanide destruction using calcium hypochlorite. The mill is designed to have a maximum throughput of 3,191 tonnes per day (at 100% availability) at a grind size of 80% passing 53 microns (averages 3,000 tonnes per calendar day at 94% availability). This equates to an annual throughput of 1,095,000 tonnes per year. An overall flow sheet for the process is shown in Figure 2, Kupol Mill Process Flow Sheet.


Figure 2. Kupol Mill Process Flow Sheet.


The estimated reagent consumption for the mill is provided in Table 6, Estimated Mill Reagent Consumption.

Table 6. Estimated Mill Reagent Consumption Reagent Consumption (kg/tonne milled) ConsumptionA (tonne/year)

Sodium cyanide 1.64B 2239 Lime 4.18 5706 Flocculant (AF-305) 0.23 314 Lead nitrate 0.33 450 Zinc Dust 0.25 341 Antiscalant 0.00001 0.014 Sulfuric acid 2.2 3003 Sodium Hydroxide 1.69 2307 Copper sulfate 0.95C 1297 Diatomite 0.29 400 Calcium Hypochlorite 1.25 1699 kg/day tonne/yearD

Borax 210 96 Fluorspar 35 16 Sodium Nitrate 259 118 Soda Ash 56 25.5 A Based on an average daily throughput of 3000 tonnes/day over a 15-month period B Assumes an AVR cyanide recovery circuit C Assumes a cyanide recovery circuit reduces consumption of reagents D Based on a daily production of 797 kg (dry) of doré. The maximum fresh water make-up for the mill processing needs is approximately 137.5 m3/hour. This is based on a maximum dry feed rate of 3191 tonnes per day. Assumptions in the water balance include:

• Ore feed contains 3% moisture; • Solids specific gravity is 2.65; • Water retained by the tailings is 30%; • CCD wash ratio is 3:1; and, • The percent solids to the dam is 50%.

Process water for the mill water balance can either be sourced from the tailings dam or from process water wells located downstream of the tailings facility.

1.4 Tailings Facilities Feasibility design of the tailings disposal facility was completed by AMEC Earth & Environmental (AMEC). The feasibility design is based on a total tailings volume of 12,000,000 tonnes. At this time the indicated mineable reserves total approximately 7 million tonnes. Feasibility-level engineering/design was performed on two tailing impoundment alternatives. Both alternatives are for conventional deposition of tailing material. Investigations into dry


stacking showed the tailing material was too difficult to filter and dry stacked tailing disposal was not a viable alternative. The preferred impoundment location alternative is contingent upon the results of condemnation and geotechnical drilling that is being performed in May and June of 2005. The conventional tailings impoundment will be initially constructed with a starter dam to provide storage for the first 3 years of tailing production, storage of acid-generating rock and a portion of the potentially acid-generating rock, and the necessary process water. The starter dam will be constructed of non-acid generating and potentially acid-generating material from pre-production open pit and underground mine development. It is expected the open pit and underground mines will provide sufficient construction materials. The dam will be raised in the downstream direction during the first three years of operation with material from the open pit mining activity. Storage capacity will be increased by extending the geomembrane up the slope and also extending the upstream foundation seepage cutoff up the abutments. The seepage barrier will consist of a bituminous geomembrane liner placed on the upstream face and anchored to bedrock at the toe of the dam. Bituminous membranes are considered to have much greater permanence than plastic liners. Bituminous membranes have high durability at low temperatures, can tolerate a wide range of pH values, and can be installed at temperatures down to -25°C. Tailings, waste rock (acid-generating and potentially acid-generating), and water management will be carried out using procedures that allow effective operation in cold climates. The primary strategy will be to maintain a water cover over the tailings, manage the impoundment to provide reclaim water during the winter, and deposit the acid-generating and potentially acid-generating material in the impoundment basin such that the waste rock will ultimately be covered by tailings. To avoid water surplus in the pond, diversions will be installed to divert most of the catchment area upstream of the tailings facility.

1.5 Ancillary Facilities Due to its remote location, the Kupol project must include all ancillary support facilities to include access roads, airport facilities, permanent camp, power generation, fuel storage and distribution, wastewater treatment plant, and other necessary facilities. 1.5.1 Access Roads The main access road in the winter of 2003 – 2004 was a 765 km route from Pevek through Bilibino to Kupol. During the summer of 2004 a route using an all-season road, Pevek to Dvoynoye, then a winter route due south to the Kupol site was investigated. Currently, a 430 kilometer winter road is used. It starts in Pevek and travels southwest around the bay and then almost directly south to Kupol. It is envisioned that there will be at least 4 road camps along the way and one stopping point that will have shower and eating facilities.


Bema Gold established management facilities, a staging area, and a truck shop in Pevek in the second half of 2004, and was extremely successful in shipping approximately 650 containers and over 10,000 tonnes of fuel to the site. Currently the route for the majority of the supplies required at Kupol will be:

• Summer and fall o Ship by sea to Pevek (port open from mid-July to mid-September) o Truck nonperishable and non-hazardous freight to staging yard in Dvoynoye

(~530 km)

• Winter (January through April) o Truck freight from staging yard in Dvoynoye to Kupol (~125 km) o Truck fuel from Pevek to Kupol by traveling due south of Pevek along a

winter road (~430 km) o Truck any remaining freight in Pevek to Kupol along the winter road (~430

km) 1.5.2 Airport facilities The airstrip has been designed and construction began in mid-2005. It is being constructed approximately 10 kilometers north of the mine site along a plateau in Starichnaya valley. The airstrip will initially be 1500 meters long (including approaches) by 150 meters wide and is expected to be expanded to accommodate larger planes (AN26, AN12). The airstrip will also have a 1000 m3 fuel tank for aviation fuel and a 50 m3 fuel tank for surface vehicles. There will also be a small building for shift change inspection and airstrip support. 1.5.3 Mill and Services Building The Mill and Services building combines six distinct areas:

1. Mill Area; 2. Power House; 3. Mine rescue; 4. Service Complex; 5. Tank Building; and, 6. Truck Shop.

These distinct areas are contained in directly adjoining pre-engineered buildings. The Mill and Services Building is currently being fabricated in North America. 1.5.4 Permanent Camp The 11,250 m² camp has been designed as a “Permanent Camp.” The size was established at 606 persons nominal and can be comfortably expanded to 656 persons. All areas of the camp will be heated using waste heat from the mill central Power House, supplemented when necessary by the mill boiler system. The living quarters include VIP units, single occupancy


rooms, double occupancy rooms and senior staff quarters. Due to the extreme weather conditions at Kupol great care has been taken to provide adequate recreational facilities for use after work, such as games rooms, a gymnasium, and an exercise room. The kitchen and dining area has a 3.5-meter ceiling. The camp has a centralized fire alarm system, with the control panel located in the permanently staffed camp security office. The Permanent Camp is currently being manufactured in North America. 1.5.5 Power Generation General The Kupol power house will operate as an “Island Installation,” producing electricity but not connected to an external power grid. The installed generating capacity is approximately 25 MW with an anticipated demand of 15.5 MW. All generator units are equipped with jacket water and exhaust heat recovery systems producing a heated water/glycol medium at approximately 91°C. The waste heat system will recover the equivalent of approximately 15MW. The hot medium will be used to heat the mill building complex and the camp facility. Generating Units A total of eight generating units will be installed. Four medium-speed units rated 4,940 kW at 720 RPM will be the base load engines. It is anticipated three units will be operating at all times, the fourth unit being a spare. Four medium-speed Cat 3516 generating units rated 1,450 kW continuous at 1200 RPM will provide power equivalent to a second spare base load engine. One or two of the Cat 3516 generator sets will be used for peak shaving allowing a close match between site operational power requirements and working generating capacity. 1.5.6 Fuel Storage and Distribution The tank farm will hold a combined 30,000 m³ of diesel fuel, 800 m³ of aviation fuel, and 300 m³ of gasoline. All the fuel for the site will be trucked from Pevek over the winter road. The fuel handling facility, located adjacent to the tank farm, will allow the unloading of two fuel trucks simultaneously. Fuel distribution to the power house day tank, camp day tank, and fueling station will be through welded, arctic-grade steel pipe. The tank farm is surrounded by a 1.1-meter high compacted earth berm with 0.8-meter high internal berms. Berms and enclosed area are lined with a 2-mm thick, high-density polyethylene liner. Surface water and possible contaminants will be collected in lined sumps and pumped to a water treatment facility for cleaning. 1.5.7 Wastewater treatment plant


The wastewater treatment plant has a treatment capacity of 160 m³/day. This is based on an estimate of 265 m³/day per person for a 600-person camp. The system can also accommodate a temporary increase for 300 people, which could occur during summer exploration.

1.6 Environmental 1.6.1 Baseline Information The Kupol Property is located within the watershed of the Anyui and Anadyr upland regions, in the east foothills of the Anyui mountain ridge. This area is characterized by prevailing low, rounded hills with occasional flat, midland areas. The watersheds are flat-shaped or convex-plane, with rounded hilltops elevated from 100 to 200 meters above the riverbeds. The tops are divided by wide (200 to 300 m) but shallow (20 to 30 m) saddles. The absolute elevations of the hills surrounding the property do not exceed 700-1050 m (1034.2 m for Malankhai Mountain and 815.0 m for Kupol Mountain). Permafrost is distributed throughout the Kupol Property area. Depending on geomorphology, thickness of permafrost layer goes down from surface to 200 to 320 meters and reaches its maximum deep under riverbeds (and may reach 450 meters based on SRK criteria). Thickness of seasonal melting varies from 0.02-1.5 meters in river valley terrains, to 12.4 meters on watersheds. Annual temperature range line never goes 20 - 30 meters under surface. Rock temperature within this range can vary from -14.0˚С at the bottom to -5.8˚С at the 2-meter depth. Temperature gradient within permafrost rock is 0.023˚C/m. In accordance with the closest weather stations, the average annual air temperature at the Kupol site, with only minor variances, is near -13°C. The total amount of precipitation does not exceed 277 mm. The absolute minimum average monthly temperatures occur in January and February (-58°C). During the warmest months (June-August), the average air temperatures are 8.3; 11.3; and 10°C; respectively. Snow cover appears in the mountainous regions in the middle of September and achieves a maximum depth in March. The average depth of snow cover is 38-45 cm. The duration of a stable snow cover is approximately 237 days. Wind patterns for the region around the Kupol site are defined primarily by the trade winds that are characterized by atmospheric circulation. The average annual wind speed is 2.1-2.6 m/s with a maximum wind speed of 20 m/s. The maximum wind speed recorded at the closest government weather stations is 24 m/s. The maximum wind speed recorded by the weather station installed at site (2003) is 30 m/s. The territory around the Kupol site belongs to the watersheds of the Srednyi Kaiemraveem and Malyi Anyui Rivers. The Srednyi Kaiemraveem drains into the Mechkereva River. The Mechkereva River is a right tributary of the Anadyr River. The Anadyr River is one of the largest waterways on the Chukotka peninsula and the waters flow from the West to the East thru the middle part of Chukotka and drains into the Bering Sea in the Pacific Ocean. The Malyi Anyui River is a right tributary of the Kolyma River. The Kolyma is one of the largest waterways of the far northeastern part of Russia and flows from the south to the north to the Eastern Siberian Sea in the Arctic Ocean. The Srednyi Kaiemraveem River runs north to


south and is situated just east of the deposit. It has an average bed width of 5-10 meters and an average depth of 0.3 meters. The river is primarily fed from surface water runoff (90%) that includes rain, snowmelt, and seasonal thawing of the active permafrost layer. Based on analogous rivers, the maximum amount of flow occurs in June and July. During the spring and summer, the river will experience 97% of its annual flow (3% occurs in the fall months). Based on geobotanical classification, the Kupol deposit is located in the Anyusk geobotanical district of the sparsely forested area of the western part of the Anyusk-Chukotka foothills. The forest composition of the Kupol deposit is represented by 73 species that are typical for the Omolon and Anyusk geobotanical regions. The area is not populated with any rare or protected species. There is no commercially valuable vegetation located within the site boundaries. The existing vegetation can be used as reindeer pastures. The value as a reindeer pasture is no higher than the multitude of other areas surrounding the site with similar plant densities and speciation. The area around the Kupol deposit is populated by wildlife that is typical for the cold-weather, topography, and mountainous terrain that surrounds the deposit. Animals that are typically found in western Chukotka include 20 species of mammals, 1 amphibian, and 34 species of nesting birds. The territory can be divided into 3 distinct regions for the mammals: tundra, taiga, and mountains. There are a total of 34 bird species, belonging to 8 orders, within the Srednyi Kaiemraveem and Mechkereva River basins. Of these species, the Passeriformes and the Charadriiformes are the dominant species encountered and have the largest populations. Among rare and endangered species that could be located within the area include:

1. mammals protected under the Russian Red Book for Far Eastern Russia: a. Yakutian snow sheep; and, b. Wild northern reindeer.

2. birds protected under the Russian Red Book for Far Eastern Russia: a. Peregrine Falcon; and, b. Gyrfalcon.

The site was surveyed during the 2003 field season by an archeologist. This included walking the site and surrounding areas to look for any identifying artifacts or areas that may have archeological significance. There were four areas that were identified for further investigation. None of these areas were located within the construction footprint. There are no nature preserves or protected environmentally sensitive areas in the vicinity of the Kupol site.

1.7 Impact Summary Impacts describe the potential effect that a risk source may have on one or more environmental receptors. Receptors can include affected humans (mine personnel, local


communities) as well as natural ecosystems. Potential environmental impacts have been assessed by characterizing the natural receiving environment, mine employees and local communities affected by the mine and combining this with theoretical knowledge of typical effects of exposure to the identified risk sources. The potential environmental impacts from the project are listed in Table 7, Impact Matrix for the Kupol Project.

Table 7. Impact Matrix for the Kupol Project Component Impact

Description Mitigation Measures Risk

Air quality

Short term, episodic, and localized impacts

Baghouses for dust, regular maintenance, catalytic converters where feasible, watering of fugitive dust sources, minimizing land disturbance

HIGH for stationary sources, LOW for mobile sources, and HIGH for fugitive sources

Topography and land disturbance Changes in topography, removal of vegetation, reduction in surface water quality, and changes in hydrobiological characteristics

Minimizing land disturbance, interim reclamation, and stockpiling of topsoil

INTERMEDIATE for topography and land disturbance

Soil Compaction, soil structure loss, potential chemical changes due to chemical composition

Minimize land disturbance, interim reclamation, extensive testing of potential ARD and implementation of waste rock management plan.

Potential impacts from ARD, if not properly managed are HIGH


Component Impact Description

Mitigation Measures Risk

Surface water quality Changes in river productivity, increased sedimentation, no long term impacts

Proper erosion control measures, minimization of land disturbances, treatment of domestic wastewater, minimize potential of filtration from tailings impoundment

Potential impacts from erosion and run-off are LOW. Potential impacts, if tailings facility leaks, are HIGH. None of the impacts are considered long term.

Vegetation Reduction in ground cover, loss of habitat and feeding area for wildlife

Minimize land disturbance, interim reclamation

Potential impacts from vegetation loss are LOW. Impacts are considered to be short term.

Wildlife impacts Loss of wildlife through poaching and disturbance of natural habitat

Minimize land disturbance and implement strong anti-poaching policies at the mine.

Potential impacts for wildlife loss are LOW. Impacts are considered to be SHORT term.

Aesthetics and visual resources Loss of aesthetics and visual resource through land disturbance and creation of new land formations

Minimize land disturbance and conduct interim reclamation

Potential impacts are LOW. Impacts for land formation are considered LONG term.

Socioeconomics Increase revenue for the region through taxation and job creation. Overall positive impacts for the region.

Maximize opportunities for IP and women. Maximize number of employees hired from the region. Maximize potential for local purchases and potential development of small to medium enterprises to support the operations.

Potential positive socioeconomic impacts are considered HIGH and LONG term.


Component Impact Description

Mitigation Measures Risk

Archeology Loss of cultural monuments during land disturbance

Conducted survey of area. Create an cultural monument response plan

Potential impacts are LOW. Impacts are considered to be LONG term.

1.8 Environmental and Social Action Plan (ESAP) The ESAP assesses mitigation, monitoring, and other measures that will be taken during construction and operations to eliminate adverse environmental and social impacts, offset them, or reduce them to acceptable levels. The mitigation measures are divided into those that will be implemented during the construction and operation phases of the development. Mitigation measures recommended during construction will emphasize protection from future degradation due to existing conditions. Mitigation measures designed for the operation phase will emphasize reduction in land disturbance, aggressive reclamation, erosion and sedimentation control, spill prevention and control, and waste management. All mitigation measures are discussed within the context of whether or not the option is technologically feasible, socio-economically acceptable, and financially viable. Mitigation measures will include:

• Minimizing new land disturbances; • Minimizing air pollution and fugitive dust; • World Class Acid-Base Accounting Program • Implementation of a sound surface water run-off diversion program; and, • Implementation of proper environmental management at the mine site including a

comprehensive monitoring program.

1.9 Health and Safety Plan A full Health and Safety Plan (HSP) has been created for this project. The framework of the HSP was developed based on the World Bank’s Occupational Health and Safety Guidelines, Russian health and safety requirements, and the U.S. Mine Safety and Health Administration guidelines. The HSP includes information on:

• CMGC policy statement; • Worker responsibilities; • Emergency procedures; • Accident investigation procedures; • Evacuation procedures; • General health and safety guidelines;


• Personal protection; • Cyanide handling and transportation • Laboratory safety; • Employee information and training; and, • Transportation safety.

During construction, CMGC has full-time health and safety personnel at the site to ensure compliance with the guidelines set forth in the HSP. Additionally, CMGC has contracted a Western H&S specialist to work with the Russian specialists on a regular basis. The health and safety department reports directly to the construction manager and indirectly to CMGC General Director. During operations, CMGC will provide a full time health and safety person for the mine site. This person will be responsible for implementing the Health and Safety Plan and compliance with Russian Health and Safety Requirements. The health and safety department will report directly to the Mine Manger and indirectly to the General Director and corporate health and safety personnel. The full HSP is provides as Attachment A of the ESAP (Volume 2).

1.10 Spill Prevention, Control, and Countermeasures Plan (SPCC) The Spill Prevention, Control, and Countermeasures Plan (SPCC) identifies the procedures required to ensure that the potential for, and risk resulting from, accidental releases of potentially hazardous materials is minimized to the greatest degree possible. The materials used at the mine and mill are common to other mining facilities of this type located elsewhere and include diesel fuel, lubricants, petroleum solvents, miscellaneous laboratory chemicals, and a variety of process reagents. A list of the major chemical reagents to be used at the Kupol mine and processing plant, including the estimated rates of consumption, is provided in the SPCC. The SPCC includes information on:

• Materials, properties, quantities, and containers; • Transport of materials; • Spill responsibility; • Readiness; • Accidental release notifications, and protocols; • Spill prevention; • Spill control and countermeasures; and, • Employee training.

The full SPCC is provided as Attachment B of the ESAP (Volume 2).


1.11 Waste Management Plan (WMP) The Waste Management Plan was created within the framework of mining in Far Eastern Russia. Some standard practices available in the West are not available in Russian Far East. This management plan details the various aspects of waste disposal including management supervision, waste classification, and arrangements for regular waste removal. The Waste Management Plan applies to construction, operation, and closure of the project. It includes management of wastes generated during construction, operations, and closure. The following types of wastes are covered in the plan:

• Domestic liquid effluent; • Domestic solid wastes; and, • Industrial solid wastes.

Disposal methods proposed include collection and sorting of materials and then placement of the materials within a secured and appropriate landfill designed to meet World Bank Group guidelines. Additionally, general management guidelines are provided in this plan. The full Waste Management Plan is provided as Attachment C of the ESAP (Volume 2).

1.12 Reclamation and Closure The purpose of reclamation is to return the disturbed areas to stabilized conditions following mining and ore processing activities as well as ensure long-term protection of land and water resources in the area. The long-term re-vegetation goals for the areas reclaimed at the project site will be to achieve perennial plant cover similar to those measured in an adjacent vegetative community or reference area. Acceptable levels for reclamation success are defined as:

• re-establishment of a vegetative cover having a density of at least 75% of the surrounding undisturbed areas; and,

• an average species diversity of approximately 30 to 40% of the surrounding undisturbed area at the end of 4 years following re-vegetation.

The current reclamation plan covers a total of XXX ha that will be subject to the following reclamation measures:

• removal of equipment and structures; • portal closure; • regrading and recontouring of all areas; • revegetation in accordance with the natural vegetation structure; and, • long-term water management measures to prevent acid mine drainage and soil

erosion.


Note that the pit walls will not be re-sloped and that existing roads will be left in place to minimize long-term impacts to permafrost. Newly reclaimed areas will be managed consistent with the project reclamation goals. The sites will be evaluated periodically during the first several years after re-vegetation to determine the effectiveness of the reclamation plan. The success of the reclamation plan will be monitored during this time to assure erosion has been prevented and that the species reestablishment has been occurring. Maintenance will be conducted on the site as necessary to assure reestablishment of vegetative species. Any interim reclamation completed will have the goal of stabilizing a given area as soon as possible after mining activities are completed in that area. CMGC will have the policy of reclaiming any portion of the project that can no longer be identified as being needed for the project. Total costs for reclamation, including recontouring, rehabilitation of reindeer pastures, biological stabilization (erosion protection and seeding of perennial grasses) and physical stabilization (anti-erosion measures) are approximately US $ 9,000,000. This amount will be corrected depending on the results of the reclamation plan that will be submitted prior to commencement of operations. The full Preliminary Reclamation and Disclosure Plan is provided as Attachment D of the ESAP (Volume 2).

1.13 Monitoring The primary purpose of the ESAP is to ensure that the project complies with Russian operating permits and environmental regulations as well as World Bank Group policies and guidelines and to evaluate the effectiveness of the environmental mitigation measures as described herein. The results of the monitoring program will be reviewed by project management on a periodic basis. If adverse environmental changes occur because of the project, appropriate remedial measures will be implemented to reduce or eliminate project-related effects. Specific details of any mitigation measures associated with unforeseen project-related effects will be developed based on the results of the monitoring. Environmental monitoring is proposed for the following environmental components and mine facilities:

• Site meteorology/air quality; • Groundwater quality; • Surface water quality; • Surface water hydrology; • Aquatic resources; • Tailings impoundment;


• Acid rock drainage; • Waste management; • Socioeconomic impacts; • Noise levels; • Health and safety; and, • Reclamation success.

1.14 Environmental and Social Management (EH&S) A full team of EH&S specialists are already in place for managing environmental, health, safety, and social issues associated with the Kupol project. This includes a corporate level EH&S specialist with extensive experience in developing natural resource projects throughout the world. The corporate level EH&S Permitting Manager has the following responsibilities in association with the Kupol project:

• setting the EH&S policies for the project; • training local persons; • ensuring subcontractors complete work to international standards; • hiring of the local environmental staff in Russia; and, • assisting the corporate office with any other environmental, social, health or safety

problem at any of the other Bema Gold operations. Two dedicated environmental persons have been hired by CMGC to work on rotation at the mine site. These people will report to the Site Construction Manager and the General Director in Magadan. The responsibilities of these people are:

• monitor all environmental parameters; • ensure compliance with all regulatory guidelines and permits; • develop annual reports as required under Russian and IFC guidelines; • develop annual budgets for the environmental department at the mine; and, • develop long-term programs to ensure protection of the environment.

Additionally, two Health and Safety specialists work at the mine site on rotation. These people will report to the Site Construction Manager and the General Director in Magadan.

1.15 Public Consultation and Disclosure As a key component to determining the feasibility of the project and in accordance with Russian legislative requirements and international norms, Bema Gold Corporation is undertaking an Environmental Impact Assessment (EIA) for the project. Public consultation is an integral part of the EIA process and provides an opportunity for all Interested and Affected Parties (I&APs) to identify opportunities and concerns associated with the development of the project. The primary purpose of public consultation is to provide the



public, technical specialists, and regulatory authorities a chance to work with the developer to produce a better project based on informed decisions. As part of the public consultation process, Bema Gold Corporation will conduct consultation in three phases:

1. Preliminary Assessment (1st quarter 2005); 2. Development of the impact assessment (2nd quarter 2005); and, 3. Concluding phase (2nd quarter 2005).

To date, the Preliminary Assessment has been completed. A full description of Bema’s Public Consultation and Disclosure Plan in English and Russian can be found at www.kupolgold.com. Additionally, the full Public Consultation and Disclosure Plan is provided at Attachment E of the ESAP (Volume 2).

1.16 Charitable Foundation CMGC is establishing a Charitable Foundation that will support developments in health care, training and education, indigenous peoples’ traditional activities, and sustainable small to medium (SME) enterprises within the ChAO. This fund will become operational after the 1st full year of operations (2009) and will be funded by CMGC in the amount of US $3,000,000 over the life of the project. This includes US $1,000,000 invested in the foundation in the 1st quarter 2009 and US $250,000 per year beginning in 2010 and lasting 8 years. Additionally, CMGC will solicit other organizations to join the foundation and contribute to the development of the region. The full DRAFT Charter for the Charitable Foundation is provided in Attachment F of the ESAP (Volume 2).

1.17 Implementation Schedule ¼ page text here

http://www.kupolgold.com/

http://www.kupolgold.com/

l Environmental Impact Assessment 23

Table 8. Cost and implementation schedule for Kupol EMP

Timing of Implementation

Estimated Costs (US $1000’s)

Item No.

Budget Description Detailed Studies

Date Provision Recurrent C-1 EMP C-2 EMP C-3 EMP C-4 EMP C-5 ERP C-6 RCP C-7 WMP C-8 SPCC C-9 EAP C-10 EMP C-11 EMP

Table 9. Cost and implementation schedule for Kupol SMP Timing of

Implementation Estimated Costs

(US $1000’s) Item No.

Budget Description Detailed Studies

Date Provision Recurrent S-1 SMP S-2 SMP S-3 SMP

Kupo

2.0 Policy, Legal, and Administrative Framework The Kupol Gold Deposit is located in the Anadyrskii Region of the Chukotka Autonomous Okrug. It is remotely located (the closest village is more than 90 kilometers away) and requires full, on-site infrastructure (power, mancamp, airstrip, and other utilities). In addition to the Anadyrskii Region, several facilities will be located within the Bilibinskii Region (airstrip and explosives storage), and the Chaunskii Region (shipping center and winter road).

2.1 Purpose and Need The Chukotka Autonomous Okrug has been developed based on resource extraction. Base metals have been at the heart of the Chukotka economy since 1941. The gold sector has traditionally made up the largest portion of the industrial output for the region. However, since the Russian economic crisis of 1998, the region remains underdeveloped and has serious problems with aging infrastructure, a severe negative population migration away from the region, and developmental issues with other sectors that could provide long-term stability to the region. The current administration has invested a large amount of money into the region to temporarily mitigate many of the problems of the region. It is not clear what will happen after the 2005 governor’s elections to the region. If the current administration does not continue within the region, it is unclear how the region will continue to develop. There are no other major industries within the region that can provide for the region. The largest two gold mining projects (Kupol and Maiskoye) are currently under development. Without these projects, the socioeconomic situation of the region will likely decline over the short-term.

2.2 Project Investment The total estimated preproduction capital cost estimate is US $407 million, which includes working capital for supplies, taxes, and owners costs. The operating cost estimate was prepared by Bema Gold and engineering contractors supporting the feasibility study and is based on actual or estimated supply costs, actual and estimated logistic costs, engineered productivity / production rates, and equipment operating and maintenance costs from other operating mines and equipment vendors.

Table 10. Operating Costs ($US).

Cost/Tonne Milled Activity or Cost Center $7.52 Surface Mining

$21.69 Underground Mining $24.54 Processing $3.23 Site Services $3.13 General and Administrative $1.27 Reclamation


Cost/Tonne Milled Activity or Cost Center $35.50 Taxes

$135.68 Total Operating Cost

2.3 Legal Matters The rights to explore and develop the Kupol deposit are currently owned by the Closed Stock Company Chukotka Mining and Geological Company (CMGC). These rights were granted pursuant to the Exploration and Production License (АНД 11305 БЕ, and former License АНД 00746) that was issued by the Ministry of Natural Resources and the Administration of the Chukotka Autonomous Okrug on October 4, 2002. CMGC was initially set up as a wholly owned subsidiary of the Chukotka Unitary Enterprise (a wholly owned government enterprise). Based on the agreement reached between Bema Gold Corporation and Chukotka Unitary Enterprises (Framework Agreement dated December 18, 2002 and amended August 7, 2003), Bema Gold Corporation (through their fully owned subsidiary, Kupol Ventures Limited) can acquire up to 75% ownership of CMGC. The remaining 25% will remain with Chukotka Unitary Enterprises. Kupol Ventures Limited (KVL) can acquire up to 75% of CMGC in the following stages (see Figure 3):

1. Initial 20% - paying US $8 million in cash (paid December 2002) and expending a minimum of US $5 million in exploration by December 2003 (completed);

2. 10% - paying US $12.5 million in cash by December 2003 (completed); 3. 10% - paying US $10 million in cash (paid December 2004) and expending an

additional US $5 million by December 2004 (completed); 4. 35% - completing a bankable feasibility study and paying US $5.00 per ounce for

75% of the gold identified in the proven and probable reserves categories in the feasibility study within 90 days of completing the study.

2.3.1 License for Exploration and Production The Exploration and Production License (АНД 11305 БЕ) is owned by CMGC. It was registered October 4, 2002 and expires on March 16, 2024. The validity term of the license may be extended by the government of Chukotka, if the license holders provide a substantiated application for an extension of the license terms 6 months before the expiration date. Under the license agreement, CMGC must make regular payments for the right of subsoil use as provided by the existing regulatory and legal instruments to include: (1) During exploration:

• 1.0% of the costs of the geological and exploration works; and, • 2.0% of the value of metal mined during the geological exploration works


It is believed that this license condition is no longer valid and has been superseded (2003) by the mineral extraction tax of 6% for gold sales and 6.5% for silver sales. For the purpose of the feasibility study, the 6% and 6.5 % sales tax for gold and silver, respectively, have been assumed.

Figure 3. Bema Gold Kupol Deposit Acquisition Flow Chart

(2) During operations:

4) Will own an additional 35% of CMGC/Kupol Deposit by completing a bankable feasibility study and by paying US $5.00 per ounce for 75% of the gold identified in the proven and probable reserve categories in the feasibility study (within 90 days of the completion of the feasibility study).

Owns 0% of

December 18, 2002: Definitive agreement reached between Bema GolDecember 18, 2002: Definitive agreement reached between Bema Gold and d and Government of Chukotka: Bema can acquire up to 75% interest in CGovernment of Chukotka: Bema can acquire up to 75% interest in CMGC/Kupol DepositMGC/Kupol Deposit

The Exploration and Production License was The Exploration and Production License was issued by the Ministry on Natural Resources issued by the Ministry on Natural Resources

and the Administration of the Chukotka and the Administration of the Chukotka AutonomousAutonomous Okrug Okrug to Chukotka Mining and to Chukotka Mining and

Geological Company (CMGC) on October 4, 2002.Geological Company (CMGC) on October 4, 2002.

1) Owns an initial 20% of CMGC/Kupol Deposit by paying US $8 million cash (paid December 2002) and expending a minimum of US $5 million (expended) on exploration on the Kupol property by December 2003.

2) Owns a further 10% of CMGC/Kupol Deposit by paying US $12.5 million in cash by December 31, 2003 (paid December 2003).

3) Owns an additional 10% of CMGC/Kupol Deposit by paying US $10 million in cash (paid December 2004) and expending an additional US $5 million (expended) on exploration by December 2004.

Owns 100% of

White Ice Ventures Ltd.White Ice Ventures Ltd.

CMGC & Kupol DepositCMGC & Kupol Deposit

Bema Gold CorporationBema Gold Corporation Chukotka GovernmentChukotka Government

Chukotka Unitary EnterpriseChukotka Unitary EnterpriseOwns 80% of

CMGCCMGCOwns 20% of

Owns 70% ofChukotka Unitary EnterpriseChukotka Unitary EnterpriseCMGCCMGC

Owns 30% of

Owns 60% ofChukotka Unitary EnterpriseChukotka Unitary EnterpriseCMGCCMGC

Owns 40% of

Owns 75% of Owns 25% ofChukotka Unitary EnterpriseChukotka Unitary Enterprise

CMGCCMGCOwns 100% of

Chukotka Unitary EnterpriseChukotka Unitary Enterprise

Owns 100% of

Kupol Ventures LimitedKupol Ventures Limited

Owns 100% of

• an amount established during State Geological Commission of Experts (GKZ Expertiza) examination; and,

• when metal losses exceed acceptable norms, double the normal payment rate.


Additionally, Amendment 1 to the license agreement (dated August 7, 2003) requires CMGC to:

• complete exploration works and submit a report with results and calculations of gold and silver C1 & C2 reserves to the GKZ Expertiza no later than December 31, 2005;

• start development of the deposit, after holding the GKZ reserves Expertiza no later than 2006; and,

• have an annual throughput of no less than 40,000 tonnes per year with a total gold recovery of 85% or more, and a total silver recovery of 70% or more.

2.3.2 Environmental Covenants Pursuant to the conditions set forth in the license, CMGC must:

• present pre-project and project documents concerning environmental protection measures, calculations of environmental quality requirements, documents authorizing the use of water, information on ecological validation of the allotment, calculations of fees for disposal of wastes, and release of pollutants to the Department for State Ecological Inspection;

• present pre-project and project documentation of the production of minerals, performance of works in water basins and protected areas to the Department and Chukotka District Centre of State Sanitary and Epidemiological Inspectorate for regulatory inspection and approval;

• present to the Department documents required to obtain the license for water use; • perform construction and reconstruction of production and other facilities on the site

of the deposit, subject always to the plans approved by the Department; • promptly make payments for any acts of environmental pollution under the regulatory

and legal acts of the Russian Federation, compensate any damage to hunting, agrarian, fishing and water industries;

• promptly report to the Department any escape of pollutants; • envisage ways and means of mitigating environmental pollution in the event of a

release; • implement measures of technical and biological reclamation of lands provided to the

lessee for temporary use. Indicate the amount of funds necessary for effective reclamation and restoration of the land to a state suitable for further effective economic use;

• construct facilities for mitigating industrial, household, and surface water runoff discharges as well as air pollution mitigation measures;

• present to the Department, together with the feasibility study for the deposit, an Environmental Impact Assessment;

• perform monitoring of water sources, including laboratory monitoring as approved by the Department; and,

• comply with the requirements set forth by the government agencies of the Chukotka Autonomous Okrug concerning solutions for the social and economic problems of the territory to include relations with the reindeer husbandry industry.


It should also be noted that reclamation and closure of the facility is only deemed complete upon the signing of the Act of Liquidation or Abandonment by the Ministry of Natural Resources and Rosgorteknadzor. 2.3.3 Occupational Health and Safety Pursuant to the conditions set forth in the license, CMGC must:

• maintain all workplaces in accordance with the requirements, norms and standards of occupational health and safety set forth in the Russian Federation; and,

• perform the development of the deposit in accordance with the Uniform Rules of Protection of Subsoil in Hardrock Mining (Moscow, 2003), and standards (norms and rules) for occupational safety and environmental protection.

2.4 Taxation Taxes payable by gold producers in Russia to federal, regional, and local budgets include: export duty, profits tax, value added tax, tax for extracting minerals, and other miscellaneous taxes.

• Export duty - Current rates of export duty are 0% of the sales value of gold (reduced on February 14, 2002 from 5%) and 0% of the sales value of silver (reduced on August 15, 2002 from 6.5%).

• Profits tax - On January 1, 2002, the profit tax was reduced from 35% to 24%, of

which 5%, 17%, and 2% is payable to federal, regional, and local governments, respectively. Russian legislation provides for certain exemptions from profits tax, which include charter capital contributions, use of funds for capital investments, and transfer of funds free of charge from foreign investors to finance capital investments of production designation, conditional upon their use within one year of receipt.

• Value added tax (VAT) - Current rate of VAT is 18%, which was wholly transferable

to the federal budget in the year 2002. VAT is also charged by custom authorities for the import of goods into the territory of Russia. The rates of import VAT are the same as the rates of VAT applicable to the sale of goods in Russia. The most important exemptions from import VAT include the import by foreign investors of technological equipment as charter capital contributions to a Russian company. Import VAT is offset once the goods are accounted for on the books of the recipient.

• Tax for extracting minerals – The current rate of tax for extracting minerals is 6% of

the sales value for gold and 6.5% of the sales value for silver.

• Miscellaneous taxes - There are also some other regional and local taxes and royalties that include property tax (2.2%).


2.5 Russian Permitting Requirements The government of the Russian Federation is organized into federal, territorial, and local (municipal) units. Territorial units include Autonomous Republics, Krais, Oblasts, and Okrugs. Local (municipal) units are called Raions. Each unit has some decision making capabilities but the ultimate approval for the Kupol project must be received at the Federal level. This is because the project involves foreign investment of more than US $500,000 and involves development of a natural resource. 2.5.1 Russian Permitting Agencies The Kupol deposit is located in the Northwest part of the Anadyr foothills on the boundary between the Anadyr and Bilibino Regions in the Chukotka Autonomous Okrug. The geographical coordinates for the site are 66°47’northing and 169°33’ easting. Because the Kupol deposit is located along the boundary between the Anadyr and Bilibino Raions it will require extensive coordination between the two regions. Additionally, the project will deal with the Chaunsk Region on issues related to the winter road and freight in Pevek. The Russian federal government has been restructured several times in the last 24 months. The latest reorganization attempts to establish clear responsibilities and remove duplication of functions. The primary agencies that will be involved in this project include: Federal Level: Federal Agency for Oversight of Environment, Technology, and Nuclear in Russia (Rosgorteknadzor): The main approving agency at the federal level for all phases of project development. They must be involved in licensing, technical assignments, approval of feasibility studies, environmental assessments, detailed engineering, and operations. This agency is located in Moscow and has a regional office in Magadan and Anadyr. This is now the main permitting body and incorporates decisions by Rosprirodnadzor into their conclusions. Government Committee for the Environment (Rosprirodnadzor): A main approving agency at the federal level for all aspects of environment. They are responsible for all aspects of water management and use issues. They are also responsible for reviewing the Feasibility and Environmental Impact Assessment. This agency is located in Moscow and has regional offices in Bilibino and Anadyr. Territorial Level (Okrug): Fareastern Regional Mining and Industrial Inspectorate (Rosgorteknadzor): This is the regional office of Rosgorteknadzor and is located in Magadan. This office covers mining and industrial activities located in Kamchatka, Chukotka, Magadan, and parts of Yakutia. This agency is responsible for review and enforcement of all phases of the project.



Chukotka Division for Technology and Environmental Supervision (Rosgorteknadzor): This is the Chukotka regional office of Rosgorteknadzor and is located in Anadyr. They are responsible for industrial activities located in Chukotka. This agency will primarily be responsible for enforcement for the project. Territorial Committee for the Environment (Rosprirodnadzor): This is the regional office of Environmental Committee and is located in Anadyr. This office covers all environmental aspects of activities located in Chukotka. This agency is responsible for review and enforcement of all environmental aspects of the project. Municipal Level (Raion): Raion Administration (Raionnaya Administratsiya): Because the project is located in different raions, the Bilibinskii, Chaunskii and Anadyrskii Raions must be involved. The administration is responsible for coordination of all permitting activities that occur in the region. The administration (and all appropriate sub-agencies) is responsible for approving land allotments, Declaration of Intent, Act of Site Selection, and licenses in the region. Raion Environmental Inspectorates (Raionniye Inspektsii): Agencies based in Bilibinskii, Anadyrskii, and Chaunskii Regions. These agencies are responsible for enforcement of Russian rules during exploration, construction, operations, and closure. They are the lead agencies in notification of violation and citations. Local inspectors are also involved in approval of the Declaration of Intent, Act of Site Selection, and can be asked to participate in review of the feasibility documents. Other agencies that may be involved during the project (either through regulatory review) or site inspections that may impact environmental, health, and safety approvals include:

• Department of Fisheries; • Forestry Service; • Watershed Management Division; • Association of National Minorities of the North; • Land Use Committee; • Sanitary and Epidemiological Committee (SES); • Department of Archeology; and, • Territorial Commission for Emergency Situations; and, • Fire Safety Inspectors.

A full layout of the primary agencies that must approve the Feasibility Study and EIA is provided in the following figure.


Figure 4. Government of the Russian Federation Government of the Russian

Federation

Ministry of Healthand Social

Welfare, RussianFederation

Ministry of NaturalResources,

RussianFederation

Ministry ofAgirculture,

RussianFederation

Federal Bureau forConsumer

Oversight andPersonal Welfare

(Moscow)

Federal Bureau forNatural Resource

Protection(Moscow)

Federal Agency forWater Resources

(Moscow)

Federal Agency forFisheries(Moscow)

Federal Agency forAgriculture(Moscow)

Federal Agency forOversight ofEnvironment,

Technology, andNuclear (Moscow)

Territorial office ofHuman WelfareManagement for

the MagadanOblast (Magadan)

Territorial office ofHuman WelfareManagement forChAO (Anadyr')

Lena WatershedManagement

Bureau (Yakutsk)

Amur WatershedManagement

Bureau(Khabarovsk)

Federal Division ofFisheries and

Wildlife (Magadan)

FederalFareastern

Fisheries Bureau(Petropavlosk-Kamchatsky)

Center forHygiene and

Epidemiology,Magadan

Oblast(Magadan)

GovernmChukot

ent ofka

Center forHygiene andEpidemiologyin Chukotka

(Anadyr')

EnvironmentalProtection

Agency in theMagadan

Oblast(Magadan)

EnvironmentalProtection

Agency in theChukotka

Okrug(Anadyr')

LenaWatershed

ManagementBureau inMagadan

Oblast(Magadan)

AmurWatershed

ManagementBureau inChukotka

Okrug(Anadyr')

MagadanWildlifeDivision

(Magadan)

ChukotkaWildlifeDivision(Anadyr')

FareasternDivision forTechnology

andEnvironmental

Supervision(Magadan)

ChukotkaDivision forTechnology

andEnvironmental

Supervision(Anadyr')

DepartmentCulture, Yout

SportTourism

PoliciesChukot(Anady

forh,

s,, and inkar')

ChukotkaFisheries

Inspectorate(Anadyr')

RegionalHead Doctorof Hygiene

andEpidemiology

(Bilibino,Ugol'ny Kopi,and Pevek)

RegionalGovernmentInspector forthe EPA inMagadan

(Omsukchan)

RegionalGovernmentInspector forthe EPA inChukotka(Bilibino,Pevek)

RegionalFisheriesInspector

(Omsukchan)

RegionalOffice of the

MagadanWildlifeDivision

(Omsukchan)

RegionalOffice of the

ChukotkaWildlifeDivision(Bilibino,Pevek,

Ugol'ny Kopi)

RegionalOffice of theFareastern

Div. for Tech.and Env.

Supervision(Magadan/

Omsukchan)

WesternChukotkaDivision ofTech. and

Env.Supervision

(Pevek,Bilibino)

EasternChukotkaDivision ofTech. and

Env.Supervision

(Anadyr')

RegionalFisheriesInspector(Bilibino,Pevek,

Ugol'ny Kopi)

RegionalHead Doctorof Hygiene

andEpidemiology(Omsukchan)

Kupo

2.6 Russian Permitting Process The permitting process in Russia is executed using steps that are similar to those in the West (except more detailed) that include preliminary assessment, environmental assessment, detailed engineering, and operational permitting. Each of those steps is further detailed in the following sections. 2.6.1 Preliminary Assessment During development of a project, the preliminary assessment must include a Basis for Act of Site Selection and a Declaration of Intent. These documents are provided to the local regulatory agency (in Bilibino, Anadyr, and Pevek) to provide the regulators with an idea of the direction that the project is going and helps avoid costly delays in project design. The information should include enough detail to allow a reviewer to determine the environmental risks associated with the proposed design and should evaluate the various alternatives associated with the design. The Declaration of Intent is used for the local and/or federal regulators to give preliminary approval of the project design and is typically based on analogous sites, literature, and available design data. The Act of Site Selection is provided to the local and/or federal regulatory authorities. This is a drawing with all of the facilities sited at proposed locations and should include potential alternatives for siting. These documents were submitted and approved in January 2005. Based on these documents, the company will prepare a Scope of Work (Technicheskaya Zadaniya - TZ) for an Environmental Assessment. The TZ should incorporate all the requirements of development of an EA and include instructions to address any potential issues identified in the Act of Site Selection and Declaration of Intent. This will be completed in June 2005. The most recent EA guidelines (Order #372, Section 4) require that the approved Declaration of Intent and Act of Site Selection be disclosed, along with the TZ, during a public consultation to local interested and affected parties. The public consultation should generate a protocol that should be incorporated into the TZ to address any concerns or requests that arose during the public consultation. The 1st round of public meetings occurred in January 2005. It is anticipated that the next round of public meetings will occur in the summer of 2005. 2.6.2 Feasibility Studies In Russia, the feasibility process (and accompanying environmental impact assessment) is divided into two separate categories: Investment and Construction. In the event of a foreign investor or joint-venture project, both documents are required and must be submitted to the federal government for review and differ only in their level of detail. The following sections outline the environmental requirements for each level.


Investment-level feasibility (TEO-I) An Investment-level Feasibility (TEO-I) is required to be completed in accordance with CП 11-101-95, “Order of Development, Approval, Confirmation, and Composition for the Basis of Investment in Construction of a Facility, Building, or Equipment” (Moscow, 1995). This feasibility study requires a preliminary design highlighting the basic engineering and safety designs that are proposed for construction and operation of the project. A separate volume is required for the TEO-I called an OVOS. The OVOS is a preliminary environmental impact assessment that is more detailed than the initial Declaration of Intent and Act of Site Selection. It can be based on government background material, analogous sites, and field studies. The document must undergo review at the local level and is then provided for review (known as Expertiza) at the federal level (Moscow). Local regulatory agencies that review the OVOS are the same groups that approved the Act of Site Selection and Declaration of Intent. Additionally, public consultation must be completed at the local level and a meeting protocol attached to the document prior to submittal at the federal level. The conclusions from the local review are provided with the OVOS to the Federal Ministry of the Environment and the Federal Mine Supervision Department. Positive Conclusions from the Federal-level Expertiza indicate that the basic design is acceptable and the project can continue to develop. Because of the “fast-track” nature of this project, Bema Gold is planning on combining this level with the Construction-level feasibility (TEO-C). Construction-level feasibility (TEO-C) A Construction-level Feasibility (TEO-C) must be completed in accordance with СНиП 11-01-95, “Instructions on Development, Approval, Confirmation, and Composition of Project Documentation for Construction of a Facility, Building, or Equipment” (Moscow, 1995). This document is much more detailed than a TEO-I. The information is based on equipment and design parameters that will be installed during construction. The TEO-C requires a section (volume) on Environmental Protection. The Environmental Protection Section will be very detailed and contain information regarding the effective use of resources and reclamation/closure. The document is reviewed at the local (including public consultation) and federal level. A positive conclusion allows the company to begin construction. Bema Gold plans to develop the TEO-C in Russian by July 2005. This document will contain the same engineering design parameters and environmental solutions that are presented in the Western feasibility study. Detailed engineering/operational permits During detailed engineering, all drawings are finalized and equipment parameters are provided to the regulatory agencies. As part of detailed engineering, the company is required to develop the following documents:

• Maximum Allowable Emission Limits;


• Water Use and Discharge Limits; • Solid Waste Disposal Limits; and, • Reclamation and Closure Feasibility.

Maximum Allowable Emissions Limits The air pollution discharge permit limit is developed in accordance with the Federal legislation “Ambient Air Protection” (N96-F3, 5/4/99). It requires limits on a tonnes per year basis and grams per second. Air pollution discharges are reported on an annual basis (Form 2-TP, Air). The reported amounts are based on parametric monitoring and stack sampling. Pollution payments are paid on a quarterly basis based on discharge amounts calculated in the permit application. Water Use and Discharge Permit Surface water use permits must be received in accordance with the Russian Water Code (No. 167-F3, 11/16/95). The Russian Water Code requires that a user identify sources, discharges, uses, recycling, and requires limits on tonnes per year usage. Water discharge limits are developed in accordance with surface water protection laws (1991). Water pollution discharges are reported on an annual basis (Form 2-TP, Water). Pollution payments are paid on a quarterly basis based on discharge measurements and pollution loading. Waste Management and Disposal The Solid Waste Management Plan must be developed in accordance with federal law “On industrial consumption and generation of wastes” (No. 89-FZ, 6/24/89) and governmental decree “Guidelines for development and approval for waste generation and disposal” (No. 461, 6/16/00). These guidelines require that a user identify wastes, classify them in accordance with their toxicity and volumes, and develop a disposal strategy. Solid waste placement is reported on an annual basis (Form 2-TP, Wastes). Pollution payments are paid on a quarterly basis based on actual waste generation and disposal. The regulatory guidelines require that a user identify wastes, classify them in accordance with their toxicity and volumes, and develop a disposal strategy. The waste classification was completed in accordance with the requirements set forth in Order No. 511 (promulgated on 6/15/01). This decree requires the categorization of all wastes in accordance with the requirements set forth in Table 11, Regulatory Waste Classification.

Table 11. Regulatory Waste Classification Waste

Classification Potential Environmental

Impact Comments

Class I Very High Environmental system can be irreversibly impacted. No mitigation measures available.

Class II High Environmental system can be severely impacted. Impacts can last longer than 30 years.

Class III Average Environmental system can be impacted. Impacts can last longer than 10 years.


Waste Classification

Potential Environmental Impact

Comments

Class IV Low Environmental system can be impacted. Impacts can last longer than 3 years.

Class V Very Low Very low possibility of environmental impacts. Requires special testing to be classified as a Class V waste.

2.7 IFC/EBRD/WBG Environmental Assessment Requirements In the event that an international lending institute chooses to participate in the project, there will be several environmental and social policies that will be applicable to the Kupol project impact assessment. For the basis of this preliminary assessment, the IFC/World Bank Group (WBG) policies have been included. 2.7.1 OP 4.01 Environmental Impact Assessment IFC’s policy on environmental assessment (EA) states that all projects proposed for IFC financing require an EA (including a social assessment) to ensure that they are environmentally and socially sound and sustainable. The breadth, depth and type of environmental assessment vary according to the type of the project. The policy requires all IFC projects to be categorized. Under IFC categorization, the Kupol project is a Category A Project that will require a full Environmental Impact Assessment (EIA). This impact assessment must include an Environmental and Social Action Plan (ESAP). These documents must be completed 60 days prior to loan approval and be disclosed in the World Bank Info Shop. The EBRD policy (April 2003) is similar to the IFC policy in scope and content. Typically, one document can be created that satisfies requirements for both EBRD and IFC. 2.7.2 OP 4.04 Natural Habitats This policy affirms IFC’s commitment to promote and support natural habitat conservation and improved land use, and the protection, maintenance, and rehabilitation of natural habitats and their functions in its project financing. IFC does not support projects that involve significant conversion or degradation of critical natural habitats. 2.7.3 OP 4.11 Management of Cultural Properties This World Bank policy is designed to assist in preservation of cultural properties. The management of cultural property in a country is the responsibility of the host government. However, before proceeding with any project that may entail destruction of a cultural property bank staff must (1) determine what is known about the cultural aspects of the proposed project site and, if necessary, (2) conduct a brief reconnaissance survey of the proposed project area.


2.7.4 Assessment and Management of Cumulative Impacts (DRAFT) The successful management of cumulative impacts is essential to the sustainability of IFC projects. In conjunction with the EA process, it is necessary to consider cumulative impacts of existing, proposed, and anticipated projects. 2.7.5 Hazardous Management Guidelines These guidelines apply to facilities and activities involving the transportation, production, handling, storage, and disposal of hazardous materials. The key requirements of these guidelines are screening, management, preventative plans during operations, response plans in the event of an accident, and community involvement. 2.7.6 Public Consultation and Disclosure In accordance with IFC requirements, the Kupol project needs to develop a Public Consultation and Disclosure Plan (PCDP) based on technically sound and culturally appropriate measures that can be implemented during project life to ensure timely and effective communications with affected stakeholders. The PCDP should outline local and WBG requirements for public consultation, identify key stakeholders, outline consultation schedules, describe implementation, and describe documentation and reporting. 2.7.7 OP 4.10 Indigenous Peoples (forthcoming) Pending finalization of this OP, IFC projects must comply with the World Bank’s OD 4.10, Indigenous Peoples. This directive describes bank policies and processing procedures that affect indigenous peoples. The directive provides policy guidance to (a) ensure that indigenous peoples benefit from development projects, and (b) avoid or mitigate potentially adverse affects on indigenous peoples caused by bank-assisted projects. 2.7.8 OP 4.37 Safety of Dams This policy sets forth IFC’s requirements for projects where dams are to be constructed. For large dams (over 15 meters high), a review by a panel of independent experts is required along with preparation of detailed plans and periodic safety inspections. 2.7.9 Non-applicable policies Additional IFC policies that do not specifically apply to the Kupol project include:

• OP 4.12 Involuntary Resettlement; • IFC/MIGA Policy Statement on Forced Labor and Harmful Child Labor; • OP 7.50 Projects on International Waterways; • OP 4.09 Pest Management; and, • OP 4.36 Forestry.


2.8 Public Consultation and Disclosure The development of the Kupol property represents one of the most significant developments within the Chukotka Okrug. As a key component to determining the feasibility of the project and in accordance with Russian legislative requirements and international norms, Bema Gold Corporation is undertaking an Environmental Impact Assessment (EIA) for the project. Public consultation is an integral part of the EIA process and provides an opportunity for all Interested and Affected Parties (I&APs) to identify opportunities and concerns associated with the development of the project. The primary purpose of public consultation is to provide the public, technical specialists, and regulatory authorities a chance to work with the developer to produce a better project based on informed decisions. As part of the public consultation process, Bema Gold Corporation will conduct consultation in three phases:

1. Preliminary Assessment (1st quarter 2005); 2. Development of the impact assessment (2nd quarter 2005); and, 3. Concluding phase (2nd quarter 2005).

To date, the Preliminary Assessment has been completed. The entire Public Consultation and Disclosure Plan is provided in ESAP, Appendix X.


3.0 Project Summary The Kupol deposit is located in the northwest part of the Anadyr foothills on the boundary between the Anadyrskii and Bilibinskii Regions in the Chukotka Autonomous Okrug. The geographical coordinates for the site are northing 66°47’ and 169°33’ easting. The area is accessible only by helicopter during the summer months and over a winter road from February until early May. It is located 200 km southeast of the town of Bilibino. Phase 1 of the proposed operation is designed to be a combination open pit and underground operation commencing in 2008 and finishing in 2014. The total proposed throughput in the mill is projected to be 1,095,000 tonnes per year (a daily throughput of 3000 tonnes/day). The principle mining method will be longhole stoping. The total number of employees at the site will peak at around 850 and level off to around 600. Due to the remote location of the project, a man camp and complete infrastructure will be constructed to support the mine. The man camp will be built to house 606 people and is located 600 meters west of the mill site. A sewage treatment plant (capacity of 160 m3/day) will be constructed to treat sanitary effluent. Electrical power will come from a combination of diesel generators (24 MW) that will be constructed at the main plant site. Fuel will be stored in ten 3000-tonne capacity fuel tanks located adjacent to the mill building.

3.1 Mining The mine is designed to be an open pit and underground operations with a current mine life of 7 years. The open pit will deliver 1.43 million tonnes over a mine life of 5 years (including pre-stripping in 2006) at an average grade of 20.4 g/t gold and 193.2 g/t silver. The underground will deliver 5.7 million tonnes over a mine life of approximately 7 years at a grade of 16.0 g/t gold and 219 g/t silver. 3.1.1 Open Pit Mining The Final Pit will be approximately 1500 meters along strike with a width of 190 meters and will ultimately reach a depth of 104 meters. From current thermistor data, it is known the entire pit is within permafrost and therefore, does not include any groundwater influences in the pit slope design. It was assumed that a final bench geometry of 24-meter bench height (70º bench face angle) with a minimum bench width of 10 meters can be successfully and safely mined. The average overall slope angle of the pit will varies depending on geotechnical parameters but averages 50 degrees. The Kupol open pit will be mined as a standard truck/loader operation, with the crusher located at the processing plant, and the waste dump located approximately 2 km to the south at the tailing impoundment. The strip ratio 12:1 is consistent over the life of the pit. The open pit will generate approximately 16.9 million tonnes of waste rock. Open pit mining is designed to operate two shifts per day (10 hour shifts), 340 days per year.


Figure 5. General Site Layout


3.1.2 Underground Mining Mining takes place from the bottom of the open pit downward and outward to the maximum extent of economic Indicated resources. Twin declines from the two portal locations, South and North, provide access to the orebodies approximately 150 meters on the footwall side of the vein zone. Spiral ramps branch off of the ramps at strategic locations to provide vertical development of each orebody. The stoping method is drifting on sill and longhole stoping. The method addresses issues of vein geometry, ground conditions, dilution, permafrost mining, productivity, and general safety. The vertical sublevel interval chosen is 15 m. Primary mechanized underground equipment will consist of electric-hydraulic two-boom jumbos, 40 tonne capacity truck, 2 sizes of LHDs (a small 3 - 5 m3 unit for narrow stopes, and a 6.5 m3 unit), and mechanized longhole drills. The production sequence for the probable underground reserve is provided in Table 12.

Table 12. Production Sequence for Probable Underground Reserve Mining location

S=South, N=North Production

(tonnes) Au Grade (g/tonnes)

Ag Grade (g/tonne)

2007 (S) 71,879 23.3 290 2008 (S) 234,821 24.6 312 2009 (S & N) 648,173 19.3 190 2010 (S & N) 902,172 16.7 235 2011 (S & N) 1,043,364 16.7 223 2012 (S & N) 1,004,288 15.7 219 2013 (S & N) 1,043,003 12.6 213 2014 (S & N) 713,504 13.2 193 The underground mine will generate approximately 1.7 million tonnes of waste. Of these 1.7 million tonnes, 50% will be used as backfill. Additionally, all non-acid generating (NAG) material will be used to build pads. Any acid-generating (AG) material that cannot be placed back underground, will report to the waste rock facility located upstream of the tailings impoundment.

3.2 Milling The design of the ore processing plant is based on the results of the metallurgical investigations and on throughput rate, head grade information and plant availability. A nominal throughput rate of 3,000 tonnes per calendar day was established; the mill utilization was set at 94 percent resulting in a design operating rate of 3,191 tonnes per operating day. Based on ore characteristics, the following processing scheme will be used:

• Ore stockpiling; • Ore receiving/coarse ore storage;


• Primary crushing; • Crushed ore storage; • Grinding and cyclone classification; • Gravity Concentration; • Thickening and Leaching; • CCD Circuit; • Merrill-Crowe and Refining; • Mercury Removal and Precipitate Drying; and, • Cyanide Destruct.

3.2.1 Ore Stockpiling During construction of the tailings facility and pre-stripping in the open pit, the ore stockpile will grow to approximately 300,000 tonnes. This material will be fed to the mill over the life of the project. During operations, it is envisaged that a coarse ore stockpile facility holding of 50-70,000 tonnes of Run-of-mine ore stockpiles near the crusher will be used for blending purposes or to supply the mill in “white out” or extreme weather conditions when the pit is shut down. Run-of-mine ore will be fed by a 980 Cat front-end loader into the 150-tonne live capacity coarse ore bin. 3.2.2 Ore Receiving/Coarse Ore Storage Run-of-mine ore from the open pit may be dumped directly by Komatsu HM350 or equivalent sized truck into a 150 live-tonne capacity coarse ore bin. Additionally, a front-end loader operating from stockpile(s) may also dump into the uncovered bin. It is envisaged that a total of 50,000-70,000 tonnes of coarse ore stockpiles near the crusher will be used for blending purposes or to satisfy the mill in extreme weather conditions when the pit is shut down. A grizzly with a 500-700 mm (20-28”) slot opening is seated over the bin to prevent oversize rock from entering the crusher. The grizzly oversize will be reduced by a track mounted rock hammer or removed to an area for future breakage. 3.2.3 Primary Crushing Ore is removed from the coarse ore bin with a NICO 1.52 m x 6.71 m long (60”x 22’) heavy duty variable speed apron feeder which feeds a 150 kW (200 hp), 1220 x 1524 mm (48”x 60”) double toggle Birdsboro Buchanan jaw crusher with a capacity of 500 tonnes per hour. The crushed product falls onto No. 1 conveyor, a 1220 mm ( 48”) wide belt ; an Eriez Model SE-7525 SC-2 self cleaning belt magnet is installed near the head pulley of this conveyor to remove tramp steel. No. 1 conveyor discharges on No. 2 conveyor. This 1067 mm (42”) wide and 115 m long belt transports the minus 150 mm crushed ore to the crushed ore storage bin. A dual idler belt scale is installed on the conveyor to measure the crushing rate


and tonnage. As it leaves the crusher building, No. 2 conveyor is enclosed in a 3350 mm (11’) diameter steel tubular gallery to the top of the ore storage bin. The jaw crusher and associated equipment are housed in an insulated and heated building. However, inside ambient temperatures in cold weather are expected to be 0 to -10°C due to air infiltration at the jaw crusher feeder entry point into the building. A 4.72 m/s (10,000 cfm) bag type dust collector is provided to collect fugitive dust at the crusher and transfer points. Additionally, the building houses a mechanical control room and a small control room is provided for the operators with the necessary instrumentation and monitoring devices to control the plant and is serviced by 10-tonne capacity overhead crane. 3.2.4 Crushed ore storage Crushed ore is conveyed to the top of an un-insulated steel bin with a live storage capacity of 3,500 tonnes. Two 1830 x 6900 mm (72” x 22’) apron feeders equipped with variable frequency drives discharge the ore from the bin to No. 4 conveyor, this 1067 mm (42”) wide belt transports the ore to the SAG Mill. The conveyor is equipped with a three idler belt scale to measure the SAG Mill tonnage and is enclosed in a 3350 mm (11’) diameter steel tubular gallery before it enters the mill. If the bin feeders have plugging problems due to the presence of clay material or freeze/thaw conditions, No. 3 conveyor, a moveable 1067 mm (42”) wide belt, will intercept the feed to the bin and discharge the ore outside the bin where it can be picked up with a loader and dumped in an external 15-tonne capacity hopper. A 1220 x 4500 mm (48” x 15’) long apron feeder with a variable feed drive (VFD) under this hopper will feed ore directly onto No. 4 conveyor. An 8,000 cfm bag house, located in the mill, will collect dust from the apron feeder discharge chutes. 3.2.5 Grinding and Classification The ore reclaimed from the crushed ore bin is fed via No. 4 conveyor to a SAG/Ball Mill grinding circuit sized to produce a P80 product size of 53µm at the design mill feed rate of 3,191 tonnes per day (133 tph). The grinding equipment includes: A 7010 mm diameter x 3000 mm EGL (23’x 9.8’) Koppers (Metso) SAG mill, which is driven by a low speed 120 rpm 3,432 kW General Electric synchronous motor equipped with a GE variable speed drive. The motor is connected to the pinion with an Eaton Air Flex Model 51VC1600 clutch. The SAG mill will typically operate at 65-80% critical speed. The SAG Mill discharge grate openings will have 20 mm slots. Extra grates with larger slots will be available for testing to optimize the grate opening. A rubber-lined trommel with replaceable screen sections with 9.5 mm slotted openings will be installed on the mill discharge to extract a recycle product which will be recycled to the SAG mill via 2, 24” wide belt conveyors.


The mill will be charged with 125 mm (5 inch) grinding balls to a design volumetric loading of 8 percent. A single idler belt scale will be installed on one of the recycle conveyors to measure pebble production. The SAG mill trommel undersize slurry is pumped to a single deck Tyler, twin shaft 1830 x 4880 mm (6’x 16’) screen with a 1.5 mm slotted panel openings. The screen oversize returns to the SAG mill feed chute while the undersize flows by gravity to the cyclone feed pump box. A 5029 mm diameter x 9200 mm long (16.5’x 27’) Koppers ball mill which is driven by a General Electric low speed 120 rpm 3,432 kW synchronous motor performs secondary grinding. The motor is connected to pinion with an Eaton Air Flex Model 60VC1600 clutch. The mill will be charged with 50 mm (2-inch) grinding balls to a design volumetric loading of 38 percent. The discharge from the ball mill and the SAG screen undersize are fed with VFD equipped pumps to a bank of Krebs 250 mm (10”) cyclones. The cyclone manifold contains a total of 11 cyclones, with 9 operating and 2 spares. Cyclone underflow returns to the ball mill and the cyclone overflow is pumped to a 2-hour surge tank. Lime is added to the SAG mill feed along with the process water. 3.2.6 Gravity Concentration The gravity circuit consists of a centrifugal gravity concentrator and a Deister shaking table. About 75 tph of solids are pumped by a dedicated pump from the cyclone feed pump box to a 15 hp Knelson KC-XD30 centrifugal gravity concentrator complete with automatic control package to regulate and control the unit operating/flush cycle for concentrate production. A gamma gauge and magnetic flowmeter are installed on the Knelson feed pump discharge line to measure and control the mass flow rate to the circuit. The gravity concentrator tailing flows back to the cyclone feed pump box and the concentrate discharges into a 2-tonne capacity storage hopper located in the refinery area after passing over a 380 x 305 mm Eriez 500 Gauss magnetic drum separator for removal of tramp iron. The hopper discharge is processed daily by a half-size Deister shaking table. The table concentrate, expected to contain about 250 kg gold per tonne, is refined separately. The table tails are pumped back to the Knelson concentrator feed. 3.2.7 Grinding Thickener and Leach Circuit The grinding thickener, the leach tanks, the CCD thickeners and the solution tanks are located in the tank building adjacent to the grinding bay. The grinding circuit cyclone overflow product at 35% solids flows by gravity to an agitated surge tank with 2 hours of storage capacity. Slurry from the surge tank is pumped by one of two horizontal slurry pumps to a Heath & Sherwood Model 1530 automatic linear primary sampler which discharges into a secondary Model 810 Vezin type sampler to produce a shift sample of leach circuit feed. A gamma gauge and magnetic flowmeter are installed on the surge tank pump discharge line to measure and control the mass flow rate to the grind thickener.


The sampled products flow by gravity to the 14.0 meter diameter EIMCO Deep Cone grinding thickener, equipped with a 18.7 kW (25 hp) hydraulic rake drive. A thickener bed level sensor is installed to detect the slurry/solution interface. Flocculant is added at a controlled rate to obtain the desired thickener overflow clarity and solids settling rate. The grinding thickener overflow is routed to the grinding circuit as make-up water. The grinding thickener underflow is fed to the pre-aeration tank at 55 % solids by one of two variable speed horizontal slurry pumps; the density and flow rate are measured and controlled by a gamma gauge and a magnetic flowmeter. About 1.5 times the nominal discharge volume of the thickener underflow is recycled back to thickener discharge tank to reduce the viscosity of the slurry by shear-thinning. The pre-aeration tank, 13.2 meter in diameter and 19.8 meters high, is equipped with a Lightnin dual impeller agitator with a 112 kW drive for slurry suspension. The tank is covered and vented. The discharge from the pre-aeration tank gravity flows to the first of a series of five agitated leach tanks. The five leach tanks are 18 meters in diameter and 17.5-19.1 meters high and are equipped with Lightnin dual impeller agitators with 187 kW drives for slurry mixing and suspension, and provide a total of 120 hours leach retention time. All leach tanks are covered and vented to atmosphere. The leach circuit is designed to allow the slurry to flow by gravity from one tank to the next by an elevation difference of the tank discharge piping between tanks of 0.41 meters. The piping layout also allows individual tanks to be by-passed. The discharge from the last leach tank is pumped by one of two Goulds/ITT horizontal slurry pumps to the first CCD thickener. The slurry mass flow rate to the CCD circuit will be measured with a combination gamma gauge and magnetic flowmeter. Air is supplied to the leach circuit by one of two Aerzen Model VLM95 low pressure screw compressors and introduced into the tanks through hollow agitator shafts. The design air addition rate is 170 cubic meter per hour per 1000 m3 tank volume. Cyanide solution and lime slurry are added at controlled rates to the first and third leach tanks. The leach pH is controlled by lime addition, while cyanide is added at preset levels based on silver grade and maintained at target free cyanide concentration levels by online cyanide analyzers. 3.2.8 Counter Current Decantation (CCD) Circuit The pregnant solution from the leach circuit is separated from the leach residue in a five stage CCD circuit consisting of five EIMCO 14-meter diameter Deep Cone Thickeners, equipped with 18.7 kW (25 hp) hydraulic rake drives. Bed level sensors are installed to detect the slurry/solution interface in the thickeners. The CCD thickener underflow from each stage is pumped to the mix tank of next stage thickener at 50 percent solids by one of two, variable speed horizontal slurry pumps; the density and flow rate are measured and controlled respectively by a gamma gauge and a magnetic flowmeter. About 1.5 times the


nominal discharge volume of the thickener underflow is recycled back to thickener discharge tank to reduce the viscosity of the slurry. Solution from the AVR/hypochlorite cyanide destruction circuit or process water are added as wash water in the solution/slurry mix tank above the last CCD thickener at a rate of three times the solution volume in the thickener underflow, to achieve a design wash ratio of three to one, and to mix with the underflow of the preceding thickener. The thickener overflow solution is advanced in the circuit in a counter current mode to achieve a high washing efficiency. Controlled amounts of flocculant solution are added to the thickener feed to flocculate and settle solids in each thickener. The overflow from the first CCD thickener is the pregnant solution, containing 99.5 percent of the gold and silver extracted in leach circuit, flows by gravity to the pregnant solution tank. The 14.0 meter diameter x 19.0 meter high pregnant solution tank provides 7.3 hours storage at the design flow rate, of 399 cubic meters per hour. The barren solution tank is of the same size and provides equal retention time. The pregnant solution is pumped to one of two automatic US Filter Auto-Jet solution clarifiers with a design capacity of 432 cubic meters per hour each. The filters are pre-coated with diatomaceous earth supplied to the filters from the pre-coat mixing system as a slurry at 2% solids by one of two Goulds ITT Model 3196 horizontal solution pumps. During filtration diatomaceous earth is added to the clarifiers by high pressure Milton Roy positive displacement pumps as 5 percent slurry at a design rate of 0.07 kg per cubic meter in order to prevent filter media blinding and to maintain a high rate of filtration. The clarifiers filters are fully automated and can be switched on line; when a filter is full the feed is shut off and switched over to the standby filter and the filter cake is then automatically removed by internal high pressure water jets. The cake slurry is pumped to No.4 CCD thickener. The clarified pregnant solution feeds directly to the Merrill-Crowe de-aeration tower. 3.2.9 Merrill Crowe and Refining Oxygen is removed from the pregnant solution in the Crowe de-aeration tower, which is 2.45 meters in diameter and 6.125 meters high, by the vacuum supplied by one of two Somarkis liquid ring vacuum pumps with a capacity of 440 ACFM. The de-aerated solution discharges into a sump from which it is pumped by one of two vertical high pressure pumps to the precipitate presses located in the refinery room. Zinc dust is added in dry form to the sump at a controlled rate by a volumetric dry feeder. The zinc dust is added with a volumetric feeder to precipitate the gold and silver from the pregnant solution. The barren solution exiting from the Shriver precipitate presses is pumped to the storage tank to the cyanide recovery circuit. The refinery process equipment which includes precipitate presses, precipitate dryer, a flux mixing system, two induction furnaces, slag processing equipment and a concrete bullion storage vault with steel door are located in the enclosed refinery room serviced by a 5-tonne overhead crane. A small shaking table to upgrade the


gravity concentrate is also located in the refinery. The room is isolated from the rest of the mill building and has a specially designed security system. The refinery is also equipped with dedicated ventilation equipment to ensure regulated air quality standards are met. During normal operation one of three Shriver Eimco Model M1200FB-83DB semi- automatic horizontal plate and frame filter presses are on-line. The design capacity of each filter press is 386 cubic meter per hour. The press filter cloth is pre-coated with diatomaceous earth by one of the two clarifier pre-coat pumps described earlier. After a press is filled with precipitate, as indicated by an increase in feed pressure, it is taken off line and opened to remove the filter cake while the standby filter is brought on line. The cake from the filters drops into precipitate trays situated below the filters. The time required to fill a press is about 100 hours at design mill feed rate and head grades. The estimated amount of precipitate and diatomaceous earth in a press is about four tonnes, the contained gold and silver content about three tonnes; the estimated cake production is about 1.0 tonnes per day. The filter cake is dried in a drying oven. The dried filter cake is removed from the oven and measured amounts of flux are then added to each tray and the mixture transferred to a tray tipping device connected to a screw conveyor which feeds the flux mixer. The mixture is then moved by screw conveyor to one of two furnace feed hoppers. Refining takes place in one of two 500 kW Ajax Model M-24 MSF 5000 induction furnaces with 12 ft3, 340 L silicon carbide crucibles. The furnaces are equipped with Ajax Model CPX cooling systems, hydraulic tilting mechanisms and exhaust systems with bag house scrubbers. Additional ventilation fans are provided in the refinery room for air exchanges. Approximately five pours per day are required to satisfy design production of 798 kg of Dore. In addition, gravity concentrate produced by the shaking table is also smelted once a week. The hot slag pots containing about 40 liters of slag each are cooled in a designated ventilation area of the refinery. The cooled slag is removed from the pots, broken and hoisted to the jaw crusher platform and manually fed to a Kue-Ken Model SN-2599005L slag crusher. The crushed slag is collected in a bin and then fed to the SAG Mill on an intermittent basis. 3.2.10 Cyanide Destruct The relatively high concentration of cyanide in the leach circuit required for optimum silver recovery has made cyanide recovery an attractive process option as it will significantly reduce the addition of new cyanide to the leach circuit and lower cyanide destruction costs. The cyanide recovery circuit is located, for hygiene and safety reasons, in an enclosed room adjacent to the leach and CCD circuit. Weak acid dissociable (WAD) cyanide will be recovered from the barren solution by reducing the solution to pH 4.5-5.0 with the addition of sulfuric acid to convert WAD cyanide to HCN. The reaction takes place in a covered


acidification tank, 5.0 m in diameter and 5.5 m high, providing a retention time of about 10 minutes. The tank is equipped with a 7.5 kW Lightnin agitator with 316 SS wetted parts. The plant consists of two parallel HCN stripping and absorption circuits. The acidified solution will be pumped at a rate of 200 cubic meters per hour to each circuit which consists of a packed bed stripping/de-sorption tower, 4.3 meters in diameter and 22.0 meters high, to remove HCN from the solution, and a packed bed absorption tower, 4.3 meters in diameter and 13.0 meters high, to recover the HCN into a caustic solution. The stripping tower solution feed pumps, equipped with double mechanical seals and stainless steel wetted parts, are driven by 50 hp VFD motors. Air will be blown into the bottom of the stripping towers at a rate of 135,000 cubic meters per hour per tower by 300 hp low pressure air fans to remove HCN gas and transfer it to the bottom of the absorption towers. The HCN gas is absorbed from the gases in a sodium hydroxide solution sprayed into the top of the absorption towers. The sodium hydroxide solution is re-circulated through the towers by 50 hp solution pumps, equipped with double mechanical seals and stainless steel wetted parts, rated at 230 cubic meters per hour until the required cyanide solution strength is attained and transferred to the mill cyanide solution tank. Stand-by recirculation pumps are also installed. Two 25 hp “Bleed” fans with a design capacity of 6500 cubic meter per hour remove surplus entrained air from the stripped gas flow and maintain the correct pressure in the system. The stripped solution exiting the stripping towers will be pumped to a hypochlorite cyanide/thiocyanate destruction tank to remove the remaining cyanide and thiocyanate prior to its use as wash water in the CCD circuit. The chlorination tank, 6 m in diameter and 6 m high, providing 25 minutes of retention time is equipped with a 18.7 kW agitator. Calcium hypochlorite solution is added at a controlled rate by a variable speed pump to the reactor tank. A sulfuric acid storage tank, 5 m in diameter and 5.0 m high, will be installed near the plant; sodium hydroxide solution will be supplied from the mill reagent mixing area. If the AVR system is down, the #5 CCD underflow is pumped to the hypochlorite cyanide destruction tank for treatment of the slurry stream. From there, the slurry overflows to the neutralization tank and then is pumped to the tailings head tank. Extensive HCN detection instrumentation will be installed in the AVR building and the building will be evacuated at a pre-determined HCN level.

3.3 Reagent Handling and Mixing The estimated reagent consumption for the mill is provided in Table 13, Estimated Mill Reagent Consumption.

Table 13. Estimated Mill Reagent Consumption Reagent Consumption (kg/tonne milled) ConsumptionA (tonne/year)

Sodium cyanide 1.64B 2239 Lime 4.18 5706 Flocculant (AF-305) 0.23 314 Lead nitrate 0.33 450


Reagent Consumption (kg/tonne milled) ConsumptionA (tonne/year) Zinc Dust 0.25 341 Antiscalant 0.00001 0.014 Sulfuric acid 2.2 3003 Sodium Hydroxide 1.69 2307 Copper sulfate 0.95C 1297 Diatomite 0.29 400 Calcium Hypochlorite 1.25 1699 kg/day tonne/yearD

Borax 210 96 Fluorspar 35 16 Sodium Nitrate 259 118 Soda Ash 56 25.5 ABased on an average daily throughput of 3000 tonnes/day over a 15 month period BAssumes an AVR cyanide recovery circuit CAssumes a cyanide recovery circuit reduces consumption of reagents DBased on a daily production of 797 kg (dry) of doré. 3.3.1 Sodium cyanide (NaCN) Daily consumption of sodium cyanide will be approximately 4.9 tonnes per day. Sodium Cyanide, in the form of briquettes, will be received in 1000 kg containers. The bags will be removed from the container by overhead crane and placed on the grating of the receiving hopper equipped with pyramid-type knife to open the bottom of the bag. The receiving hopper is located above the 10-cubic meter solution mix tank equipped with solution recirculation pumps. The solution mix tank is filled with the volume of barren solution required to produce a 25 percent strength sodium cyanide solution; this solution is re-circulated through the receiving hopper until the cyanide is dissolved; the solution is then transferred to a 25 cubic meter cyanide solution storage tank.

The cyanide solution is distributed to the leach and Merrill-Crowe circuits by variable speed positive displacement reagent pumps. 3.3.2 Lime (CaO) Daily consumption of lime will be approximately 12.5 tonnes. Crushed pebble lime (CaO) will be received in one- or two-tonne bulk bags as a minus 20 mm product; the contents of the bags will be unloaded into a 15-cubic meter lime storage bin holding about 20 tonnes of lime.

Lime will be removed from the storage bin at a controlled rate of 1,500 kg per hour and fed to a 760 mm diameter by 300 mm wide California pebble roll crusher prior to being processed in a dual cell Denver attrition lime slaker equipped with 15 hp attrition agitators. The discharge from the slaker is fed to a screw classifier in closed circuit with a 5 hp ball mill. The classifier overflow density is controlled to 15 percent solids; the grit is removed by the trommel screen on the ball discharge.


3.3.3 Flocculant (AF-305) A polyacrylamide flocculant is used in the thickener for improved settling of the solids and water clarification. Daily consumption will be approximately 0.7 tonnes. Flocculant, in granular form, will be supplied in 1000 kg super sacks. The polymer will be batch mixed in a fully automated system, supplied by the flocculant vendor, to produce a 0.5 % strength solution. The polymer will be unloaded from the bags into a receiving hopper from where it is transferred at a controlled rate with a variable speed screw conveyor and an air blower to the “Jet Wet” mixer installed above the 16 cubic meter mixing tank equipped with an 800 mm diameter dual prop 10 hp agitator. The solution, after mixing and aging, is pumped with a progressing cavity pump to a 130 cubic meter storage tank from where it is distributed to the process with variable speed progressive cavity metering pumps. 3.3.4 Lead nitrate Pb(NO3)2 Lead nitrate is consumed at a rate of approximately 1 tonne/day. Lead nitrate, in crystal form, will be received in 25 kg bags. Forty individual bags will be dumped into a 10-cubic meter agitated mix tank equipped with a 7.5 hp mixing pump to produce a 10 percent solution. Process solution will be used for mixing the lead nitrate. The solution will be transferred to 10-cubic meter storage tank from which the solution is distributed at controlled rates to the leach and Merrill-Crowe circuits by variable speed positive displacement reagent pumps. The storage tank holds sufficient solution to supply the process for more than one day 3.3.5 Zinc dust Zinc dust is used in the Merrill Crowe process at a total rate of 0.25 kg/tonne. Daily consumptions will be approximately 1.1 tonnes and is provided in one-tonne supersacks. 3.3.6 Antiscalant Antiscalant is added to the chemical mixtures to prevent scale deposition in the tanks, pumps and the pipes. It is provided in liquid form in metal 200-liter barrels and is added directly from the barrels to the mixtures. It is used at a rate of 0.00001 kilogram per tonne of ore. 3.3.7 Sulfuric Acid Sulfuric acid is used in the AVR cyanide recovery circuit. The estimated consumption is 2.2 kg/tonne of ore. This is a daily consumption of approximately 6.6 tonnes. It will be provided in 340 kilogram drums.


3.3.8 Sodium Hydroxide Sodium hydroxide is used in the AVR cyanide recovery circuit. The estimated consumption is 1.69 kg/tonne of ore. The daily consumption rate is approximately 5.1 tonnes day. It will be provided in 1 tonne supersacks. 3.3.9 Copper Sulfate (CuSO4*5H2O) Copper sulfate is used as a catalyzer in the cyanide destruct process and for settling of iron-cyanide complexes. It is provided in one tonne bags. It is used as needed but the approximate dosage is 0.95 kilograms per tonne or 2.9 tonnes per 24-hour period. 3.3.10 Calcium hypochlorite Calcium hypochlorite is used in the cyanide destruction circuit. It is provided in 60 kilogram carboys and is used at a rate of 1.25 kg/tonne (or 3 tonnes/day). 3.3.11 Diatomite Diatomaceous earth is used as a filter pre-coat in the Merrill-Crowe process. It is provided in 23 or 45.4-kilogram bags. It is used at a rate of 0.29 kilogram per tonne (0.9 tonnes per 24-hour period) 3.3.12 Borax, Fluorspar, Soda Ash and sodium nitrate Borax, fluorspar, soda ash, and sodium nitrate are components of the flux and are used in the process of refining the gold-silver precipitates and gravity concentrates. Consumptions per day are 210, 35, 56, and 259 kilograms, respectively. They are provided in plastic or paper bags of 23 and 50 kilograms.

3.4 Kupol Mill Water Balance Assuming an average 3000 tonnes per day (125 tonnes/hour) throughput for the mill, the mill needs 121.3 m3/hr for processing. It is anticipated that the majority of this process water (92%) will be recycled from the tailings impoundment. The rest will be made up from the potable water wells (9.4 m3/hr). Assumptions in the water balance include:

• Ore feed contains 3% moisture; • Solids specific gravity is 2.6; • Water retained by the tailings is 30%; • CCD wash ratio is 3:1; and, • The percent solids to the dam is 50%.

The Mill Process Water Balance is provided in Figure 6, Kupol Mill Water Balance.


Figure 6. Kupol Mill Water Balance.


3.5 Tailings Facility The tailings will be typical, finely-ground gold mine tailings, with 100% finer than 150 microns. Gold will be recovered using a cyanide leach. Following gold recovery, the tailings will pass through a cyanide destruction circuit to reduce the total cyanide concentration to a typical level of about 5 mg/L and less than 10 mg/l thiocyanate. Tailings will be transported to the pond in a tailings pipeline as a slurry with a slurry density of 50% by weight. The Kupol Project is scheduled to have a nominal plant throughput of 3000 tonnes per day (tpd) of tailings during Years 1 through 6 (open pit phase 1 through 3 and the first 3 years of underground operations) and 2000 tpd for the remaining mine life (underground mining phase). Bema provided an upper bound for the ore reserve of 12 Mt, for a mine life of about 13.5 years. For design purposes, the in-situ dry density of deposited tailings is estimated at 1.15 tonnes/m3 based on the experience at Bema’s Julietta Mine. This value of in-situ density is somewhat lower than typical gold tailings values, but is selected as being representative of tailings that will be deposited entirely sub-aqueously and hence will have relatively low density and high void ratio. The design tailings volume based on this density is 10,500,000 m3. Additionally, the open pit and underground mining phases are scheduled to produce approximately 13,000,000 million tonnes of acid-generating (AG) and potentially acid-generating (PAG) waste. At an assumed density of 2 tonnes/m3 the waste rock will require 6,500,000 m3 of storage in the tailings impoundment. The total amount of solids to be stored in the tailings impoundment is expected to be about 17,000,000 m3. The following design criteria have been adopted for design of the tailings dam:

• The dam should have sufficient freeboard at all times to store the projected tailings volume and AG and PAG waste rock, plus 4 m depth of water (at least 2 m of free water and up to 2 m of ice in winter);

• A minimum of 1.5 m freeboard should be provided above the resulting tailings and

water level, comprising 1.0 m for storage of a Probable Maximum Flood (PMF) Event, plus 0.5 m for waves above that level;

• The tailings starter dam will be constructed of non-acid generating (NAG) and zoned

placement of PAG rock waste rock and be sized to provide storage for the first three years of tailings production, acid-generating and remaining potentially acid-generating waste rock plus the design water storage and freeboard requirements;

• The tailings and impounded waste rock are sulfide-bearing and are therefore

potentially acid-generating. At closure, given that it is neither feasible (given water balance constraints) nor advisable (given dam safety and long term liability considerations) to maintain a permanent water cover over the tailings, pond water removed and the tailings (which will have covered the impounded waste rock) will be frozen in a saturated state. The impoundment will then be capped with sufficient


cover material to prevent the tailings from thawing, vegetated, and graded to encourage runoff and minimize infiltration; and,

• The tailings slurry water will contain total cyanide, at initial concentrations in the

order of 5 mg/L. Total cyanide within tailings pore spaces may persist at these concentrations. The impoundment should be constructed so that there will be minimal seepage through the dam and foundations. It is assumed that there will be no seepage via deep groundwater flow, as the tailings basin will be rendered essentially impervious by permafrost, and will remain frozen. The dam face will be fully lined with a bituminous geomembrane to minimize seepage loss through the dam and foundations. The membrane will be tied to bedrock at the upstream toe to form a seepage cutoff.

The dam design selected for the Kupol tailings facility is a rockfill embankment with an impervious geomembrane on the upstream slope. The selection of this dam section is dictated by the climate and availability of materials. A rockfill embankment can be constructed in cold weather conditions, so the dam construction season will not be limited by the short summer season. The dam will be constructed of NAG and PAG rock from mine waste or quarried. The dam will be initially constructed with a starter dam to provide storage for the first 3 years of tailings production, AG and a portion of the PAG waste rock from the open pit and process water. It is planned that the starter dam shell will be constructed with NAG and PAG material from pre-production open pit development, or if this cannot be scheduled, then quarried NAG rock would be used. The dam will be raised in the downstream direction. It is anticipated that the downstream dam shell will be raised near to its projected ultimate crest elevation during the first three years of mine operation, when the open pit will be in production and waste rock will be available. The storage capacity would be increased above the starter dam level in later years by extending the geomembrane up the slope and also extending the upstream foundation seepage cutoff up the dam abutments. The seepage barrier will consist of a bituminous geomembrane liner placed on the upstream face and anchored to bedrock at the upstream toe of the dam. The geomembrane will be extended up the dam face as the dam is raised. Bituminous membranes are mainly composed of needle-punched, non-woven, polyester geotextile (300 gm/m2) impregnated and faced with bituminous mastic (mixture of bitumen and filler) creating a liner product with a typical thickness of about 4.8 mm. Bituminous membranes are considered to have much greater permanence than plastic liners. The material has high durability under extremely low temperatures. Bituminous geomembrane can also tolerate pH values in the range of 2 to 11 and can be installed in temperatures as low as -25º C. Tailings, AG and PAG (the portions of which not used in the starter dam) waste rock and water management will be carried out using procedures that will allow effective operation in the cold climate at Kupol. The primary strategy will be to maintain a water cover over the tailings surface during winter, the impounded waste rock will mostly be placed above the active pond for first several years of production and ultimately be covered by tailings.


Tailings will be discharged underwater, beneath the ice cover, to avoid freezing on exposed beaches, which could lead to large losses of water. The water cover will also allow for process water reclaim during winter. During winter, there will be no input of runoff, and water will be depleted through losses to storage in tailings voids. The minimum depth of water will be maintained throughout the winter by either:

a) providing enough excess water in storage at the beginning of winter, through storage of runoff during the summer months, or

b) adding fresh makeup water from the process water wells during winter. To avoid water surplus in the pond, diversions will be installed to divert most of the catchment area upstream of the tailings facility. The total ditch length will be 4200 m. Regular monitoring of the tailings pond water depth and volume will be required to ensure that the operating criterion is being met. The water balance will be updated on a monthly basis throughout the year. Decisions on water additions will be made based on the ongoing water balance monitoring. During the spring freshet, side discharges on diversion ditches may be opened to allow discharge into the pond on a controlled basis, in order to achieve target water depths at the beginning of winter, or conversely water balance observations may indicate that maximum diversion should be achieved to prevent water surpluses in the pond. The entire tailings report (as prepared by AMEC) is provided as Appendix C.

3.6 Ancillary Facilities Because the Kupol mine will not be located near any existing town or facilities, the project will be completely independent of existing utilities. The following facilities will be constructed/utilized in support of the mine, mill, and tailings facility:

• explosive storage; • man camp; • chemical storage; • fuel and lubricant storage; • power and heat supply; • administrative complex; • facility water supply; and, • waste water treatment plant.

Each of the facilities listed above is described in greater detail in the following sections.


3.6.1 Explosives Storage Currently, the explosives storage is located 850 meters northwest of the proposed permanent mancamp along the right bank of Kaiemraveem Creek. The explosives storage area has been permitted in accordance with the CMGC Explosives Storage Project (Проект Базисного поверхностного расходного склада взрывчатых материалов ЗАО «Чукотская Горно-геологическая Компания», 9.25.04). However, the explosives storage area will be moved during the summer 2005 field season due to the proximity to the construction and operational area. The new location is approximately 2.2 kilometers from the existing construction camp in the upper reaches of Starichnaya Creek. It will be accessed via the airport road (already construction) and be on the Eastern side of the road (access will be located approximately 300 meters from the road). There are no facilities within the blast safe zone (calculated to be 1700 meters) at the new location. The facility will be made up of 20-foot containers and can store up to:

• 180 tonnes of explosives material containing ammonium nitrate or 62 tonnes of explosive material containing trinitrotoluene on three (Pads 1-3) pads (fourteen 20-foot containers each);

• 400,000 sticks of blasting sticks on a separate pad (Pad 4) in (seventeen 20-foot containers);

• 30,000 blasting caps on Pad 4 in one 20-foot container) • 400,000 meters of detonating cord on Pad 4 in two 20-foot containers; and, • 115 tonnes of ammonium nitrate on a separate pad (Pad 5) in five 20-foot containers.

This material is stored on the same pad as the ANFO mixing plant (one 20-foot container). The ANFO plant is a Dyno Nobel Amix 25 capable of producing 25 kilograms per minute of ANFO.

The area for storing explosive materials (Pads 1-4) and ANFO mixing (Pad 5) are separated by fencing. The explosives mixing area is located in a separate container (near Pad 5) and also contains an explosives laboratory. The entire complex is surrounded by a fence that is 2 meters high and topped with barb wire. The site is under constant security (in accordance with Russian requirements) and is accessed via a road that deadends at the explosive storage facility and passes the mill construction area. There is a guard post near the entrance of the facility made out of a 20 foot container. Heat is provided to the containers by electric radiators. The facility is lit by two projector lights atop towers within the boundaries of the explosives storage area. Power is provided to the site via a diesel generator located outside the boundaries of the facility. All of the pads are made up of gravel material with a thickness of approximately 300 mm. The pads are large enough to hold all of the containers.


Pevek Explosive Storage Because the port at Pevek does not have storage capacity for explosive materials, CMGC will be required to offload all explosive materials from the port facility within a 24-hour period. Therefore, CMGC is investigating a storage facility outside of Pevek. There is currently an abandoned facility available approximately 23 km from town (at the head of the winter road) that has already been permitted for storage of explosive material. This facility has been abandoned for quite some time and offers excellent storage potential. It would be fenced and guarded similar to the on-site facility, is located several kilometers away from the closest waterway and is not near any other facilities. A decision will be made regarding this facility (or a similar facility) prior to receiving shipments of any explosive material. 3.6.2 Mancamp The mancamp is an 11,520 m2 facility designed to last 15 years. It is based on a maximum occupancy of 606 persons. The design consists of a main hallway running east-west with 17 wings on either side of the main corridor that contain sleeping areas, wash facilities, and mechanical rooms. Additionally, the main corridor contains recreational areas, kitchen facilities, a first aid station, and laundry facilities. Additionally, there will be a tent camp located on the western side of the facility to accommodate overflow during the construction period. A layout for the mancamp is provided in Figure 7.

The wings are broken down into separate living units as follows:

• Five wings with single occupancy; • Nine wings with single/double occupancy; • Two wings with personal rooms; and, • One wing for VIP rooms.

Taking into account the extreme weather conditions at Kupol and the long Arctic winter, great care has been taken to provide adequate recreational facilities for use after work. The recreational areas are mainly located in the approx. 245 m x 9 m wide main corridor, known as “The Relaxation Village.”


Figure 7. Mancamp Layout

In this village, there are 6 billiard game rooms, 6 table tennis rooms, 10 table soccer tables, 2 table hockey games, 7 television rooms, 1 library/reading area, and 1 computer room complete with internet access. Each television area is set up to accommodate 24 persons at any time and includes both smoking and non-smoking rooms. Additionally, the common area will have:

• A gymnasium/exercise room; • A laundry room; • A first aid station; • A kitchen and dining area; and, • A camp security office.

The gymnasium and exercise room are located on the east end of the facility and an area of 743 m2 and 258 m2, respectively. The main gymnasium is suitable for basketball and volleyball. The exercise room contains a variety of equipment that include cardiovascular (treadmills and rowing machines) and weights. The gymnasium can be entered/exited from the arctic corridor or two separate entrances. The laundry is accessible from the main corridor, set into the main kitchen/dining block, and includes 7 industrial dryers and 5 industrial washers. Laundry will be done by designated laundry staff. The staff will also have access to stations for ironing, sewing, and two tables for folding clothes. The First Aid room has two treatment rooms complete with single beds and washrooms/shower and an exam room complete with a single bed. There is also a separate


room with three single beds. Also included are a secure prescription storage area, a reception area and a private office. The main dining area will have an 11-ft high ceiling and will be divided into three distinct eating areas that can accommodate 344 people at a time. It includes areas for self-serve, short-order and dish drop-off. Windows on three sides of the dining area will allow an outside view. The kitchen area has coolers for meat, fish, dairy, and vegetables, as well as a storage freezer. There are designated bakery, meat preparation, and vegetable preparation areas. The camp has its own bakery complete with a cooler and freezer. The camp is heated by the heat recovery system on the power plant. During extreme cold events or as needed, the camp will be supplemented by the boiler assembly adjacent to the Camp. This assembly consists of two main 1,354 kW boilers, totaling 2,708 kW, and a series of modular boilers totaling 3,767 kW. Because the boiler plant is only intended to supplement heat as needed, the boilers were chosen for their ability to meet varying loads. The modular boilers have steps of 90 kW, meaning that they can increase or decrease their heat output by 90 kW at a time. All boilers are computer controlled, with the controller triggered by a temperature sensor on the supply water. Fire Protection Each room in every wing will be equipped with a smoke alarm and heat detector. In the hallways, each wing will have 4 wall mounted smoke detectors at evenly spaced intervals. Each wing will have 7 separate smoke alarms with a relay wired to the forced air fans’ power supply and 8 high temperature heat detectors. These will be located inside of each of the eight mechanical rooms for every wing. Every single occupancy wing will be equipped with 2 or 3 (depending on the length of the corridor) manual pull fire alarms (one at each end of the hall and one in the middle) and 2 fire alarm bells (one at each end of the hallway). Emergency lighting will be provided in each single occupancy dorms by 5-36 watt, battery powered lights that will be evenly spaced throughout the hallway. Exits will be clearly marked by exit signs and an evacuation plan will be provided for each room. All exits will be equipped with portable fire extinguishers (25 total) and there will be an additional 55- 4.5 kg ABC type fire extinguishers. These extinguishers will be wall-mounted and evenly spaced throughout the complex. The corridor, mudroom entryway, exercise room kitchen and gymnasium are all equipped with emergency lighting (36 watt battery pack), wall mounted smoke detectors, heat detectors, high temperature heat detectors in the mechanical rooms, manual fire alarm pull stations, and fire alarm bells. Additionally, the kitchen is equipped with 2 Ansul R-102 Fire


Suppression Systems with ABC type extinguishers. These systems are automatic with heat and smoke sensors located above the ranges and deep fryers. The control panel for the centralized alarm system is located in the permanently staffed camp security office. The fire alarm system is divided into zones. Each wing comprises one zone, along with the kitchen, the gymnasium, and the entry / exercise room; the main corridor is divided into five zones. When one of the heat or smoke detectors dispersed throughout the building is triggered, the control panel in the office will indicate in what zone the alarm occurred, set off the alarm bells, and turn off the appropriate HVAC fans. There will be a fire hose cabinet located adjacent to the entrance to each of the wings with a 5080 single valve and 30.5 meters of hose (17 total). An additional 8 fire hose cabinets with 30.5 meters of hose will be dispersed throughout the rest of the facilities in strategic locations (25 total). 3.6.3 Chemical Storage The following reagents will be used during milling at Kupol and destruction of process reagents (Table 13, Estimated Mill Reagent Consumption). The packaging will be in accordance with Russian standards and are transported to site in 40-foot containers. The reagents are stored on site in accordance with the following plan (Table 14).

Table 14. Description of the reagents and packaging stored at reagent storage Reagent Container Total

(tonne) Type of Package Amount NoA. of

Containers Sodium cyanide, solid granular 1-147 3947.5 1000 kg packets 3948 147 Anionic flocculent, liquid 148-164 438.8 23 kg drum 19033 17 Lime, solid 165-284 3216.3 45 kg bag 71473 120 Sodium metabisulfite, solid 285-489 5512.5 23 kg bag 239674 205 Copper sulphate, solid 490-504 395 45 kg bag 17174 15 Lead Nitrate, solid 505-516 380 25 kg bag 15200 12 Antiscalant, liquid 517-519 72.5 220 kg drum 330 3 Zinc Powder, solid 520-525 146.3 23 kg pail 6359 6 Diatomite, solid 526-528 73.4 23 kg bag 3206 3 Borax, solid 529-537 238.8 23 kg bag 10381 9 Potassium Nitrate, solid 538-540 68.8 50 kg bag 1375 3

Total 17,674.9 540 ANumber of containers assumes that each 40 foot container holds 27 tonnes of reagents. The reagents are shipped to site in manufacturer’s packaging, either plastic sealed bags or metal drums. The packages are stored in 40-foot sea containers, each reagent in a separate contained except for compatible cargos of small amounts. Upon arrival on site, the containers are placed at the storage area. The containers have storage for 15 months reagents stock. The on-site reagents storage area (with the exception of sodium cyanide) is an open container pad located on the mill pad, 50 meters from the main mill building and approximately 100 meters from the camp (See Figure 5, General Site Layout). The foundation of the mill laydown area is bed rock; the container pad, located on the eastern part of the mill site, is


filled with deluvial rock. The pad is accessed via an internal roadway from the camp to the mill. There is also a fenced in area in the storage area for highly poisonous substances (sodium cyanide) located approximately 4 km from the mill in the area previously used for the exploration camp. The fenced area is designed strictly for the storage of cyanides and other highly active poisonous substances (HAPS). The area is designed to keep up to 1500, 40-foot containers. All reagents are stored at ambient temperature inside the container in such a way to prevent tipping. For the fenced area, a 2-meter high barbwire fence surrounds the containers and is 2 meters away from the nearest containers. Additionally, the containers of the first level are placed side-by-side, doors to the inside. The containers of the second level are placed doors to the inside. The containers are handled by the 40-tonne KATO 450 crane. From the cyanide storage area, reagents will be transported to the mill in full, closed containers. The mill will have the capacity to store a full 40-foot container inside the mill building. In all cases, the manufacturer’s package is left intact until the reagents are moved to the dedicated reagents mixing areas in the mill. Empty 40-foot containers are stored within the fenced area until they are used. The fenced reagents storage will be covered with 0.4 meters of well draining material Surface water is diverted around this facility. The fenced area is provided with lighting, armed security rounds and 24-hour operating security camera that transmits the view to the central control desk in the mill. Pevek Storage Because the port at Pevek does not have storage capacity for hazardous materials, CMGC will be required to offload all hazardous materials within a 24-hour period. Therefore, CMGC is investigating a storage facility outside of Pevek. There is currently an abandoned facility available approximately 23 kilometers from town (at the head of the winter road). This facility has been abandoned for quite some time and offers excellent storage potential. It would be fenced and guarded similar to the on-site facility, is located several kilometers away from the closest waterway and is not near any other facilities. A decision will be made regarding this facility (or a similar facility) prior to receiving shipments of any hazardous material. 3.6.4 Fuel and Lubricant Storage It is estimated 30,000 m3 of diesel fuel will be required to operate the powerhouse, camp, mill, and mobile equipment at Kupol for 15 months. Grade No. 2, winter and arctic grade diesel fuel, will be stored in a lined and contained fuel farm about 200 meters from the mill building and fed to the powerhouse by gravity. The volume of containment will be 110% of the largest tank within the containment area. The fuel farm is currently designed to accept up


to ten 3000 m3 fuel storage tanks. All of the fuel tanks will be fabricated in Russia and will be suitable for the arctic conditions. All tanks will be certified to meet Russian safety and environmental standards. Each tank will sit on a concrete ring foundation that will be filled with compacted gravel and flooded with liquid bitumen. Each tank will have a staircase to access the tank top and have a fire control foam generator system and suitable lightning arrestors. The tank farm piping system will be designed to allow fuel transfer between tanks. This will allow the blending of arctic and winter fuel and emptying of tanks for clean out. All fuel for the site will be trucked from Pevek over the winter road. Fuel will be unloaded on a concrete slab with a center sump to control spillage. Two fuel trucks (the maximum fuel load per truck 16 m3) may unload at the same time. There will be an additional fuel transfer pump that will be used as a standby. It will take the fuel pump approximately 10 minutes to off-load a fuel truck. The pump will connect to the truck using a suction hose with kamlock type ends. Two fuel pumps will be installed at the fuel farm, one dispenser for mining equipment, the other for refueling other mobile equipment. The dispensers will be placed on an elevated concrete slab with a center sump to control spillage. They will be protected against mechanical damage by 150 mm diameter concrete filled steel pipes. A 26-hour capacity powerhouse day tank will be placed just outside the powerhouse. This tank will be in a lined and bermed (height of external berm is 1.1 meters + internal berms sized at 0.8 meters) area that will be large enough to hold 110% of the entire volume of the tank. A camp day tank, sized to hold the camp fuel supplies for 26 hrs, will be placed near the camp. This tank will also be in a lined and bermed area that will be large enough to hold 110% of the entire volume of the tank. Both tanks will be manually filled once per day. Aviation fuel will be trucked from the fuel farm to the airstrip. Therefore, there will be two additional 400 m3 tanks that store aviation fuel. There will also be one 300 m3 fuel tank that will store gasoline. Additionally, there will be a 50 m3 fuel tank that can be used by the surface equipment that is servicing the airport facility. Both tanks will be located in a lined and bermed area that will hold 110% of the volume of the aviation fuel tank. 3.6.5 Power and Heat Supply The Kupol power house will operate as an “Island Installation,” producing electricity but not connected to an external power grid. Installed generating capacity is approximately 25 MW with an anticipated demand of 15.5 MW. The installed capacity is based on a have two spare generators on hand. Generating voltage will be 4160V, 3-phase, 60 Hz; plant voltage will be 480V, 3 phase, 60 Hz. All generator units are equipped with jacket water and exhaust waste heat recovery systems producing a heated water / glycol medium at approx. 91°C. The waste heat system will recover the equivalent of approximately 15 MW. The hot medium will be


used the heat the mill building complex and the camp facility. A common combustion air intake system sized to provide sufficient air to all engines is provided. It incorporates pre and final air filters. The external air intake system allows radiant heat from the generator sets to be used for building heat. The generating sets will operate on No. 2 fuel oil, either winter or arctic grade. A level sensing system will be installed in the day tank with low and high level alarms. Incoming fuel will pass through a fuel oil heater to raise the temperature above the clout point to prevent the forming of wax crystals during cold temperatures. A coalescing type filter will be used to maintain water and sediment levels below 0.1 %, as required by the engine manufacturer. The powerhouse air emissions will be discharged through a 32-meter exhaust stack. Two double-walled tanks, ½ m3 each, will be installed inside the powerhouse for engine return fuel. An automatic fuel level system with a display mimic diagram will pump hot return fuel back to the day tank using a positive suction gear pump. The day tank will also act as a heat sink for the hot return fuel. Note that any building located more than 500 meters from the mill complex will have to be heated using either small boilers or electrical heaters. This includes: the mine portal building, explosives storage/ANFO facilities, security checkpoint, the well pumphouses, and the airport facilities. 3.6.6 Facility Water Supply Fresh water supply is from well(s) located on Kaiemraveem Creek, about 6 km downstream of the mill. Water will be pumped 6 km in an insulated 7” steel and HDPE pipeline to a tank at the mill site that will also serve as a firewater tank. If required fresh water can be pumped into the tailings impoundment area. 3.6.7 Tailings Line The proposed tailings dam is about 3 km from the mill at a lower elevation. An insulated (3” insulation) 305mm (12”) line transports tailings to the dam. A spare line is also installed. About 200m of the line may need heat tracing near the mill, but after reaching a high point, tailings flow by gravity to the dam. Choke(s) may be required near the end of the line. Process water is reclaimed from the dam using a submerged pumping system to pump through a 10” heat traced and insulated (3”) steel and HDPE pipeline. A spare line will also be installed. 3.6.8 Wastewater Treatment Plant The sewage treatment plant has a treatment capacity for up to 42,000 US gal/day (160 m3/day). The system can also accommodate a temporary increase of 300 people, for summer exploration. The wastewater treatment plant will be a series of treatments designed to treat both sewage and kitchen wastes. It consists of a grease trap, an aerobic section, a biological


contacting unit, a settling tank, and disinfection by UV light. The unit will be located near the mancamp and discharge into the tailings facility. The wastewater treatment facility is located approximately 70 meters northeast of the camp and will be built inside of containers. The Waste Water Treatment Plant is arranged in two rows of 11 containers with one container on the end to act as a windbreak. Each container has dimensions 2.45 m x 2.45 m x 6.1 m (8’ x 8’ x 20’). The structure measures 32.8 m x 15.7 m x 6.984 m high. The containers will be topped by a 15.25 m (50’) wide Quonset roof structure. 3.6.9 Solid Waste Disposal Area The Solid Waste Management Plan must be developed in accordance with federal law “On industrial consumption and generation of wastes” (No. 89-FZ, 6/24/89) and governmental decree “Guidelines for development and approval for waste generation and disposal” (No. 461, 6/16/00). These guidelines require that a user identify wastes, classify them in accordance with their toxicity and volumes, and develop a disposal strategy. Solid waste placement is reported on an annual basis (Form 2-TP, Wastes). The regulatory guidelines require that a user identify wastes, classify them in accordance with their toxicity and volumes, and develop a disposal strategy. The waste classification was completed in accordance with the requirements set forth in Order No. 511 (promulgated on 6/15/01). This decree requires the categorization of all wastes in accordance with the requirements set forth in Table 15, Regulatory Waste Classification.

Table 15. Regulatory Waste Classification Waste

Classification Potential Environmental

Impact Comments

Class I Very High Environmental system can be irreversibly impacted. No mitigation measures available.

Class II High Environmental system can be severely impacted. Impacts can last longer than 30 years.

Class III Average Environmental system can be impacted. Impacts can last longer than 10 years.

Class IV Low Environmental system can be impacted. Impacts can last longer than 3 years.

Class V Very Low Very low possibility of environmental impacts. Requires special testing to be classified as a Class V waste.

The Kupol solid waste landfill will be facility located along the airport road in the upper reaches of Starichnaya Creek. It will be designed to accommodate wastes for the life of the project. It will include a sorting area; incinerator, standard landfill area, lined hazardous waste facility, and an area to landfarm soils containing petroleum products. The area will also have a section designated for recycling of scrap metal, used batteries, and products containing mercury. Table 16 provides estimates for the amount of material to be generated at Kupol, their hazard classification, and their disposal method.


Table 16. Annual Solid Waste Generation Material Est. Amount

(tonnes) Hazard

Classification Disposal Method

Waste Rock 18,000,000 5 Waste dump/portal construction or construction material

Tailings 12,045,000 4 Tailings impoundment Reagent Packing 31 5 Neutralize in mill, burn, landfill Crucibles 10 N/A Put in crushing circuit Used oil/petroleum products

125 2 Burned in used oil heaters

Mercury lamps 1 1 Container for off-site recycle Batteries 2 2 Container for off-site recycle Automobile tires 65 5 Place in waste dump/landfill Scrap Metal 169 5 Stockpiled in boneyard for recycle Soil contaminated with petroleum products

40 3 Landfarmed in designated area of landfill

Wood Products 40 5 Landfill/incinerate WWTP Sludge 4 5 Chlorinate landfill/tailings impoundment Household wastes 114 5 Landfill/incinerate 3.6.10 Pevek Service Center A service center has been leased in Pevek that is capable of housing up to 50 persons and has a small shop for repairs to the trucks. It also has a very small container yard. It is located on the edge of Pevek approximately 4 kilometers from the port. Potable water is trucked to the service center daily (approximately 10 m3 daily). Sewage is routed to a septic tank located on the property. Electricity and heat are provided by the city.

3.7 Transportation The main access road in the winter of 2003 – 2004 was a 765 km route from Pevek through Bilibino to Kupol. During the summer of 2004 a route using an all season road, Pevek to Dvoynoye, then a winter route due south to the Kupol site was investigated. Bema Gold established management facilities, a staging area, and a truck shop in Pevek in the second half of 2004 and was extremely successful in shipping approximately 650 containers and over 10,000 tonnes of fuel to the site. At the present time the route for the majority of the supplies required at Kupol will be:

• Summer and fall o Ship by sea to Pevek (port open from mid-July to mid-September) o Truck nonperishable and non-hazardous freight to staging yard in Dvoynoye

(~530 km)

• Winter (January through April) o Truck freight from staging yard in Dvoynoye to Kupol (~125 km)


o Truck fuel from Pevek to Kupol by traveling due south of Pevek along a winter road (~430 km)

o Truck any remaining freight in Pevek to Kupol along the winter road (~430 km)


4.0 Baseline Conditions The exploration history at Kupol starts with the discovery of quartz veins in 1966 by a Soviet sponsored regional mapping program. In 1995, the main Kupol system was found. From 1995 to 2001, the state-funded Anyusk Geological Expedition completed about 3,000 meters of drilling and about 8,000 linear meters of trenching. Total disturbed area within the exploration area (to include site roads and preliminary mill preparation area was 45 ha. Bema acquired the property in December 2002 and completed extensive drilling, trenching and testing of the deposit in 2003 and 2004. The sampling programs are summarized in Table 17.

Table 17. Exploration Programs Year No. of Diamond

Drill Holes Drill Hole Meters (m)

No. of trenches

Trench Meters (m)

No. of Channel Samples

Channel Sample Meters

2003 166 22,200 15 2500 -- -- 2004 309 53,800 2 226 87 700

The site currently has approximately 110 ha of disturbed area.

4.1 Topography The Kupol Property is located within the watershed of the Anyui and Anadyr upland regions, in the east foothills of the Anyui mountain ridge. This area is characterized by prevailing low, rounded hills with occasional flat, midland areas. The geomorphological features within its boundaries include erosion-tectonic, erosion-glacial, accumulative water-glacial, accumulative fluvial and relict types and sub-types of the topography, showing different distribution and hierarchy. The erosion-tectonic and erosion-glacial types of the terrain feature spatial interrelation and sometimes are juxtaposed onto each other. These displacements occurred in different ways, and are fixed by distribution areas of the Miocene-Pliocene flatland (peneplain), volcanic plateau, quaternary plains and upland terraces. The watersheds are flat-shaped or convex-plane, with rounded hilltops elevated from 100 to 200 meters above the riverbeds. The tops are divided by wide (200 to 300 m) but shallow (20 to 30 m) saddles. The watershed slopes are not likely to exceed 15-20 degrees; their foothills and riverbeds are overlain with young talus-solifluction and solifluction sediments. The absolute elevations of the hills surrounding the property do not exceed 700-1050 m (1034.2 m for Malankhai Mountain and 815.0 m for Kupol Mountain). The valley of the Kaiemraveem River was formed and developed along the fault zone of the same-named fault. Other erosion-tectonic and erosion-glacier landscape patterns around the Kupol Property area also include the relicts of old peneplane surfaces represented by flat or slightly sloping south and southeast plains of 2 - 3 km length and of 0.3 - 1.5 km width can be identified that occur at 650 - 750 m elevations.



The accumulated water-glacier and fluvial landscape patterns are manifest in the low horizons of macro-relief at the Kupol Property area. Absolute elevations of the low horizons in intermontane troughs and in the river valley beds of the Starichnaya River and the Middle Kaiemraveem River range from 423.0 m to 673.5 m. The relative elevations of the watersheds above the riverbeds range from 600 to 1,000 m. The average elevation of the upper boundary of the low tier of the terrain is 50 - 250 m. The erosion depth in the river valleys relative to the local watersheds is 213 m. Slopes of valley terrains look very flat (10-25 degrees) in head and low waters, but they are very steep at trough-like areas of the deep erosion (over 25 degrees). The patterns of the accumulated fluvial landscape are represented by the floodplain and above floodplain terraces of different age. Modern floodplain terrains are well developed along the Middle Kaiemraveem River valley and downstream of some of its tributaries. Their height varies from 0.3 m to 3-4 m. Accumulative and cap-type terraces can be identified among the floodplain terraces of 5-7 meters. The alluvial sediments in such terrains vary from 0.6 to 3.0 m in width. The terrace areas are 50-500 m wide. The edges of such terraces are typically sharp with steep benches and flat surfaces flatly sloping toward the riverbed and comparably steep hillsides.

4.2 Geology The Kupol deposit is located in the 3000-km long Cretaceous Okhotsk-Chukotka volcanogenic belt (Figure 8). This belt is interpreted to be an Andean volcanic arc type tectonic setting, with the Mesozoic Anyui sedimentary fold belt in a back-arc setting to the northwest of the Kupol region. Russian 1:200,000 scale mapping indicates that the Kupol deposit area is centered within a ten-kilometer wide caldera, along the western margins of the 100-kilometer wide Mechkerevskaya volcano-tectonic ‘depression’, an Upper Cretaceous bimodal nested volcanic complex. The volcanic succession in the area is 1300 meters thick and is comprised of a lower sequence of felsic tuffs and ignimbrites, a middle sequence of andesitic to basaltic-andesitic flows and fragmentals capped by felsic tuffs and flows. These sequences are cut and discordantly overlain by basalts of reported Paleogenic age. The volcanic rocks unconformably overlie and intrude folded Jurassic sediments. Mineralization is associated with a north-south trending splay (this is the Kupol structure) off a regional fault (Kaiemraveem fault) of similar orientation. The Kaiemraveem structure terminates 25 kilometers to the north at the Malyi Anyui River fault. The Malyi Anyui is interpreted as a major east-west trending strike slip structure. The magnitude of displacement along the Kupol structure is unknown but the direction is inferred to be normal-right lateral due to fault geometry. Russian interpretation suggests that the Kaiemraveem fault intersects a volcanic subsidence ring structure (Kovalevsky caldera) within the Kupol deposit area. The Kaiemraveem fault and Kupol structure are the locus for felsic dome and dyke intrusions.


Figure 8. Kupol Regional Geology


4.2.1 Property Geology The Kupol Property is situated in the Cretaceous Okhotsk-Chukotka volcanogenic belt. The property is underlain by a bimodal sequence of shallow dipping andesite and basaltic-andesite flows and pyroclastic units, rhyolite dykes and flow dome complexes. Mineralization is hosted in a north-south trending, steeply east dipping dilatant splay off a large regional fault structure of similar orientation. The Kupol deposit consists of one or more polyphase quartz-adularia quartz veins of an epithermal low sulfidation character that are sporadically cut by rhyolite dykes. Gold and silver mineralization is associated with sulphosalt-rich bands and pods within brecciated colloform banded veins Alteration associated with the Kupol structure shows up as a broad (<= 400-metre wide) zone of magnetite destruction. The bulk of the alteration appears to be associated with the structural hanging wall of the zone. Alteration types include: argillic, clay-acid sulfate, and propylitic. Based on geological setting, vein textures, mineralogy and alteration assemblages, the Kupol deposit can be classified as a low sulfidation fissure vein type epithermal deposit (Hedenquist, Arribas and Gonzalez-Urien Classification, 2000), or a quartz-adularia-sericite type epithermal deposit(Sillitoe Classification, 1993). The Kupol deposit has similarities to many large, low sulfidation epithermal deposits including Hishikari (Japan), Comstock Lode (Nevada, USA), Martha Hill Mine (Waihi District, New Zealand), Kubaka (Russia); El Penon (Chile), and Ken Snyder (Nevada, USA).


Figure 9. Property Geology

Kupo

4.3 Permafrost Permafrost is distributed throughout the Kupol Property area. Depending on geomorphology, the thickness of the permafrost layer goes down from surface to 200 to 320 meters, and reaches its maximum deep under riverbeds (based on literature reviews). Thickness of seasonal melting varies from 0.02-1.5 meters in river valley terrains to 12.4 meters on watersheds. Annual temperature range line never goes 20 – 30 meters under surface. Rock temperature within this range can vary from -14.0˚С at the bottom to -5.8˚С at the 2-meter depth. Temperature gradient within permafrost rock is 0.023˚C/m. Seasonal thaw in the deposit area begins in early June and lasts through September. In the bald peak areas, its depth largely depends on slope exposure and is 0.8-2.0 m; in talus patches, the thaw depth is limited by thickness of talus. In arctic tundra patches, the thaw depth depends on mechanical structure of sediments and varies between 0.8 and 1.8 m. The seasonal thaw is 0.3-0.4 m in typical hummocky and hillocky (mound) tundra areas and 2.5 m in local well-drained areas. Within temporary water flows occurring in intensive snowmelt period and at creek sources, the thickness of melting zone is 0.4 - 2.0 meters. In patches of inter-ridge depressions, the thickness of seasonal melting zone does not exceed 0.5 meters. River valley facies comprised of alluvial water seepy sedimentary material extends down to 1.0 meter. Below this limit, all sediments are cemented with permafrost rocks. Cave and fracture types of underground ice formations occur, predominately, on hillsides inside eluvia, diluvia, and proluvial-deluvial massives, as well as within fractures of hard rock due to the diluvial processes of runoff re-distribution (accumulation of runoffs in different types of ground hollows with further frosting). Volume ice content for host rock at the property area can reach 30 - 70%.

4.4 Climate The climate at the Kupol mine site is defined by the site’s geographical location at the northeastern extremity of Eurasia, because of the influence of two oceans and the vast continental mass of Yakutia. Atmospheric conditions include complex circulation patterns that vary considerably over both the warm and cold times of the year. The climate of the region around the Kupol site belongs to the continental climatic region of the subarctic climate belt with extremely severe weather consisting of long and cold winters (8-8.5 months), overcast weather, and short summer periods (2.5 months). In order to characterize the climatic conditions around the Kupol territory, data was used from the closest multi-year weather stations that are located in the Malyi Anyui River basin:

• Weather station “Ileryny” is situated in a lake depression that has a diameter of 10-30 kilometers. It is built adjacent to Ileryny Lake in the upper reaches of the Malyi Anyui watershed. The absolute elevation of the station is 425 meters above sea level (masl). The station is located 2-4 kilometers from a mountain ridge that belongs to the


northeastern portion of the Anyui foothills that have an absolute elevation of 500-900 masl.

• Weather station “Ostrovnoye” is located along an old terrace of the Malyi Anyui

River approximately 60-90 meters from the village. The absolute elevation of the weather station is 98 masl. The valley has a width of up to 10 kilometers in the vicinity of the weather station. The closest mountain structures with an elevation of 300-800 masl and steep slopes of up to 20-45° are 5-6 kilometers to the north, 8-9 kilometers to the east, and 1.5-3 kilometers to the south and east of the weather station.

• Weather Station “Markovo” is located 200 kilometers south of the property.

Markovo is the only climate station with recorded data that coincides with the two summer seasons of data recorded at the mine in 2003 and 2004. Markovo is better suited to establish a direct correlation with the site data. Other reasons for employing Markovo (discussed in detail in the following paragraphs) include the north-south movement of wind currents, long-term monthly patterns and conservatism.

Additionally, a Western “datalogging” weather station was installed in July 2003. This data is also included in the climate baseline section.

In accordance with the closest weather stations, the average annual air temperature at the Kupol site, with only minor variances, is near -13°C. The total amount of precipitation does not exceed 380 mm (Appendix A). The total number of days with an average daily temperature above zero does not exceed 50 days. Positive average daily temperatures are first noted in the first 10 days of June. The transition from positive average daily temperatures to negative average daily temperatures occurs in the middle 10 days of September. The absolute minimum average monthly temperatures occur in January and February (-58°C). During the warmest months (June-August), the average air temperatures are 8.3; 11.3; and 10°C; respectively (see Appendix A). The relative humidity values at the Kupol site are not large and, on an annual average, reach 71% and are an indicator of the high continental climate of the region. The approximate amount of evaporation from surface water sources is 280 mm during warm periods. This also bears witness to the continental climate of the region. Snow cover appears in the mountainous regions in the middle of September and achieves a maximum depth in March. The average depth of snow cover is 38-45 cm. The duration of a stable snow cover is approximately 237 days. As a result of the wind blowing, the valleys are filled with snowdrifts and the tops of the mountains and steeps slows are blown bare. The average snow density reaches 160 kg/m3 with a water content of 107 mm. During the winter months, approximately 116 mm of precipitation falls. This is approximately 46% of the annual amounts. Wind patterns for the region around the Kupol site are defined primarily by the trade winds that are characterized by atmospheric circulation. The average annual wind speed is 2.1-2.6


m/s with a maximum wind speed of 20 m/s. The maximum wind speed recorded at the closest weather stations is 24 m/s. The maximum wind speed recorded by the weather station installed at site (2003) is 30 m/s. Seasonally, the weather at the site follows the following patterns: Winter (September-May) – The average air temperature in January and February reaches -33.1°C. The average absolute minimum temperature observed in January-February is -58°C with an average monthly temperature of -32.9 and -33.3°C, respectively. Strong winds, frequent whiteouts, the most number of foggy days, and days without any sun are observed during the winter period. This period also has approximately 96-133 mm of precipitation. The average snow cover is 0.5 mm. This snow cover is unevenly distributed and is very compacted. Spring (May-June) – The Spring is cold with average temperatures near 0°C with frequent whiteouts, snowstorms, and frosts. Summer (July-August) – The average monthly temperature of the warmest period (July-August) is near 10.1°C, with the average absolute maximum temperature (in July) reaching 30-33°C. Anytime during summer, the temperature can drop and snow can fall. The amount of precipitation to fall during the summertime is between 94 and 131 mm. During summer, daylight hours are approximately 23 hours per day. Fall (August-September) – The transition to winter is extremely short. Falls are cold, cloudy, and humid. The development of snow cover occurs during the second 10 days of September.

4.5 Surface Water Hydrology The territory around the Kupol site belongs to the watersheds of the Srednyi Kaiemraveem and Malyi Anyui Rivers. The Srednyi Kaiemraveem drains into the Mechkereva River. The Mechkereva River is a right tributary of the Anadyr River. The Anadyr River is one of the largest waterways on the Chukotka peninsula, and the waters flow from the west to the east through the middle part of Chukotka and drains into the Bering Sea in the Pacific Ocean. The Malyi Anyui River is a right tributary of the Kolyma River. The Kolyma is one of the largest waterways of the far northeastern part of Russia and flows form the south to the north to the Eastern Siberian Sea in the Arctic Ocean. The site of the ore body and the proposed operations will be located within the northern part of water catchment basin of the Srednyi Kaiemraveem. There is not much known about the catchment basins in this region. There are 17 riverways that are Class 1-2, and a there is a density of waterways of 0.75 km/km2. The area of the watershed of the Kaiemraveem up to the monitoring point located 500 meters above the mouth of the last, largest water way, the Tret’iy Creek, captured in the surface source that will catch the planned construction is 79.1 km2. This area includes the catchment basins of the three largest waterways around the area that is planned for the facility and includes: Pervyi, Vtoryi, and Tret’yi Creeks.


The Srednyi Kaiemraveem River runs north to south and is situated just east of the deposit. It has an average bed width of 5-10 meters and an average depth of 0.3 meters. The river is primarily fed from surface water runoff (90%) that includes rain, snowmelt, and seasonal thawing of the active permafrost layer. Based on analogous rivers, the maximum amount of flow occurs in June and July. During the spring and summer, the river will experience 97% of its annual flow (3% occurs in the fall months). The three small creeks that feed the Kaiemraveem run only during summer months and are a product of precipitation runoff and snowmelt. Peryyi Creek is located north of the deposit. It has a watershed catchment of 4.0 km and a straight creek length of 2.7 kilometers (with an overall length of 5.2 kilometers). Vtoryi Creek is located within the catchment area of the proposed waste dump area. It has a watershed catchment area of 1.7 kilometers and a straight length of 1.3 kilometers (with an overall length of 2.0 kilometers). Tret’yi Creek is located within the catchment area of the proposed tailings facility. It has a watershed catchment area of 5.0 kilometers and a straight length of 3.9 kilometers (with an overall length of 9.6 kilometers). The hydrologic characteristics of the watersheds in the vicinity of Kupol mine are similar to northern Canadian watersheds. Extreme flood events are caused by a large rainfall event combined with the spring snowmelt. Flow data has been collected at two recording sites along an unnamed creek, which is an upper catchment tributary of Kaiemraveem Creek, west of the mine camp. The unnamed creek alignment meanders through the site just west of the proposed tailings facility (in the basin for the alternative tailings location). The first station located immediately upstream of the proposed dam location, while the second is located towards the upper limit of the ultimate impoundment limit. The maximum-recorded stream discharge at the first station is 0.54 m³/s, while the maximum-recorded flow at the second station is 0.24 m³/s. Runoff coefficients are calculated using the recorded flow and rainfall data collected at the mine site. Initially, negligible groundwater effects and only storm events are considered in computing the runoff coefficients. However, the spring freshet in May increases the runoff coefficient substantially for the month with drainage occurring over frozen ground. Additionally, recorded hydrograph data suggest that saturated ground surface condition and subsequent groundwater releases increase the runoff from rainfall events in August and September. The resulting coefficient estimates therefore include these assumptions and are summarized in Table 18 below.

Table 18. Monthly Runoff Coefficients Kupol Month Coefficient

June 0.80 July 0.60 Aug 0.60 Sept 0.90

These coefficients will be reviewed and revised as more site-specific stream flow and rainfall data are collected.


4.6 Surface Water Quality Background concentrations in the Kaiemraveem River: Water quality samples were collected upstream of the project site at stations K101 and K102 on the Kaiemraveem River during 2004. Sampling dates at each station were as follows: K101 – June 21, July 17, September 13, and October 1; K102 – July 17, September 13, and October 1 (no sample on June 21 was taken at site K102). For the purposes of modeling, the average background concentration in the Kaiemraveem River was taken as the arithmetic mean of these samples for each parameter (Table 19). Background concentrations in contributing catchments of the Kaiemraveem River and from natural surface runoff: Water quality samples were taken near the project site in contributing catchments of the Kaiemraveem River at stations K202, K302, K401, K402, and K601 during 2004. As with samples K104 and K105 on the Kaiemraveem River, samples from sites K202, K302, K401, and K402 appear impacted by current activities at the project site (elevated calcium, magnesium, and sulfate concentrations, and depressed pH values), and are not representative of unimpacted background runoff contributing to the Kaiemraveem River. Samples from K601 were collected on July 17, September 13, and October 1 (no sample on June 21 was taken at site K601).

Table 19. Summary of estimated background water quality in the Kaiemraveem River and its contributing catchments near the project site.

Kaiemraveem

River Background Natural

Surface Runoff Russian Class 1 Fisheries

Standarda Bulk Parameters pH 7.29 7.39 6.5-8.5 Turbidity (NTU) 4.9b 4.8b 0.25 Total Dissolved Solids (mg/L) 74 66 n/ac Petroleum Products (mg/L) <0.05 <0.05 0.05 Detergents (mg/L) <0.01 <0.01 0.5 Phenols (mg/L) <0.0005 <0.0005 0.001 Major Ions Sodium (mg/L) 3.8 3.9 120 Potassium (mg/L) 0.65 0.23 50 Calcium (mg/L) 4.0 7.8 180 Magnesium (mg/L) 1.4 3.0 40 Bicarbonate (mg CaCO3/L) 21 27 n/ac Carbonate (mg CaCO3/L) <1 <1 n/ac Chloride (mg/L) <1 <1 300 Sulfate (mg/L) 10 19 100 Nutrients BOD (mg/L) 0.52 0.50 n/ac Ammonia (mg/L) 0.51b 0.23 0.5 Nitrate (mg/L) 0.18 0.23 40 Nitrite (mg/L) 0.040 0.037 0.08 Phosphate (mg/L) <0.05 <0.05 0.2 Metals Aluminum (mg/L) n/ad n/ad 0.04 Antimony (mg/L)c <0.01e <0.01e 0.005f Arsenic (mg/L) <0.05 <0.05 0.05


Kaiemraveem

River Background Natural

Surface Runoff Russian Class 1 Fisheries

Standarda Cadmium (mg/L) <0.01e <0.01e 0.005 Chromium (mg/L) <0.02 <0.02 0.07 Cobalt (mg/L) <0.01 <0.01 0.01 Copper (mg/L) 0.0029b 0.0012b 0.001 Iron (mg/L) 0.59b 0.052 0.1 Lead (mg/L) <0.002 <0.002 0.006 Manganese (mg/L) 0.019b <0.01 0.01 Mercury (mg/L) <0.0015e <0.0015e 0.00001 Molybdenum (mg/L) n/ad n/ad 0.001 Nickel (mg/L) <0.01 <0.01 0.01 Selenium (mg/L) <0.02e <0.02e 0.002 Silver (mg/L) n/ad n/ad 0.01 Strontium (mg/L) <0.05 <0.05 0.4 Thallium (mg/L) n/ad n/ad 0.0001 Zinc (mg/L) 0.0080 0.011b 0.01

* Applicable ambient receiving water quality limits are shown for comparison. a Russian Class 1 fisheries limits are shown for all parameters except where noted (VNIRO. 1999). b Background concentrations in the Kaiemraveem River and/or its contributing catchments exceed Russian Class 1 fisheries standard guidelines. Assessment of potential project impacts on ambient water quality must be considered in this context. c No applicable guideline is available under the given regulatory class. d Concentrations in the Kaiemraveem River and/or its contributing catchments are not available for this parameter. e Method detection limits (MDLs) for background water quality samples in the Kaiemraveem River and/or its contributing catchments exceed Russian Class 1 fisheries standard guidelines. Assessment of potential project impacts on ambient water quality must be considered in this context. Values at one-half the MDL were used for modeling for these parameters. f There is no Russian Class 1 fisheries limit for antimony (VNIRO, 1999). The drinking water/recreational use limit is shown (VNIRO, 1999).

4.7 Vegetation Based on geobotanical classification, the Kupol deposit is located in the Anyusk geobotanical district of the sparsely forested area of the western part of the Anyusk-Chukotka foothills. The forest composition of the Kupol deposit is represented by 73 species that are typical for the Omolon and Anyusk geobotanical regions. The area is not populated with any rare or protected species. Structurally, the area is similar to all other areas that are found within the Anyusk-Chukotsk foothill region and can be divided into 3 elevation belts:

1. Mountain-arctic deserts and arctic tundra. These occur at the top of watersheds, ridges, and crests and along slopes with no vegetation/ fragmented vegetation. The elevation is 600-1000 meters.

2. Tundra with typical subarctic mottling and tussocky features that are featured within the rolling, hilly areas between 500-600 meters; and,


3. Tundra that is found within the river valleys and draw areas that consists of marshes and fields.

The vegetative association of the mountain-arctic deserts and arctic tundra is made up of sparsely populated groves of arctic and subarctic dwarf shrubs and lichen mixed with arctic sedge and grasses. The ground cover can be estimated at being between 10 and 30% of the total surface area for this belt. The vegetative association for the subarctic tundra has a much more consistent grouping then the mountaintops. However, the vegetative ground cover in these areas does not exceed 70%. The vegetative associations found within the river valleys are, typically, more complex than those found within the subarctic tundra areas. Although, the density of the ground covering can be lower due to the relatively short-term development of the alluvial substrate. There is no commercially valuable vegetation located within the site boundaries. The existing vegetation can be used as reindeer pastures. The value as a reindeer pasture is no higher than the multitude of other areas surrounding the site with similar plant densities and speciation. The area is not part within the limits of any reindeer winter foraging areas (the closest location is more than 20 kilometers from the site) but does come close to the summer route used by the reindeer herding group station in the village of Lamutskoye. Their route includes Starichnaya River, Kaiemraveem Creek and Kopytochnaya Creek. Their route brings them no closer than 10 kilometers from the site. A vegetation map is provided in Figure X.


Figure 10. Vegetation map


4.8 Soils The soil cover within the contours of the land allotment is characterized as complex with non-homogenous and contrasting structures. The soil cover can be characterized by a combination of mosaics, combinations, mottled, and cryogenic micro-complexes that are all located within the region of mountain tundra and sub-arctic tundra. The following six types of soil associations can be found at the Kupol site and are characteristic for the sub-arctic region of the Northern Fareast:

1. weakly developed soils that are peaty and turf covered soils formed under the dynamic conditions of the 1st stages of ground-cover of the neoelluvial formations;

2. sod organic accumulative soils that are formed primarily under grassy vegetative cover on bedrock of alluvial origin;

3. gleyzem soils form inside the boundaries of the accumulating formations of the meso- and micro-relief within the mountain arctic and sub-arctic tundra, within saddle areas and on the sandy-loamy deluvial areas that are underlain with permafrost;

4. cryozem soils are formed in combination with gleyey soils in heteronomous conditions in the layers of detrial sandy deluvial and is underlain with permafrost;

5. cryoturbation podburs are formed within the mountain arctic tundra on the slopes with rocky soils and weathered bedrock with sections of permafrost and cryogenesis during soil development process; and,

6. peaty soils are formed in sedge-peat moss communities under marshy conditions in the accumulative relief forms.

Among the established soils types within the Kupol boundaries, the most widely distributed soil types are cryoturbated cryozems and gleyzems (36%), peaty soils (10%) and cryoturbated podburs (6%). A soil map is presented in Figure X.

4.9 Seismicity From seismic zoning map OSR-97-B, the seismic activity level of this region is classified as 6 points for 5% probability of exceedence in 50 years. This seismic level applies to the towns of Anadyr, Bilibino and Ilirney, which are all located in the same region as Kupol. According to SNiP II-7-81*, for this level of regional seismic activity, seismicity does not need to be taken into consideration in design of these structures.


Figure 11. Soil map


4.10 Fauna The area around the Kupol deposit is populated by wildlife that is typical for the cold-weather, topography, and mountainous terrain that surrounds the deposit. Animals that are typically found in western Chukotka include 20 species of mammals, 1 amphibian, and 34 species of nesting birds. The territory can be divided into 3 distinct regions for the mammals: tundra, taiga, and mountains. Near the Kupol site there are 6 orders of mammals (Carnivora, Rodentia, Insectivora, Leporidae, Bovidae and Artiodactyla). Of these species, the Carnivora and Rodentia play the most the key role (both in numbers and commercial value). The Carnivora order is composed of 4 species belonging to 3 families and includes: the grizzly bear (Ursulus arctos), the ermine (Mustela erminea), the wolf (Canis lupus), and the red fox (Vulpes vulpes). The Rodentia order is represented by 5 species belonging to two families and includes: the arctic ground squirrel (Citellus parryi), lemming vole (Alticola macrotis), the red-backed vole (Clethrionomys rutilus), Siberian lemming (Lemmus chrysogaster), and the tundra vole (Microtus oeconomus). The Insectivora order consists of 1 species of Soricidae, the masked shrew (Sorex caecutiens). The Leporidae order consists of two families of species containing two species: the arctic hare (Lepus timidus) and the whistling hare (Ochotona hiperborea). The Artiodactyla order consists of 3 species from two families: the wild northern reindeer (Rangifer tarandus), the moose (Alces alces), and the snow sheep (Ovis nivicoloa). In addition to the species listed above, the following animals can be found in the habitats present at the Kupol site: the tundra shrew (Sorex tundrensis), gray-sided voles (Clethrionomys rufocanus), the collared lemming (Dicrostonyx torquatus), the wolverine (Gulo gulo), the least weasel (Mustela nivalis), the arctic fox (Alopex lagopus), and the siberian salamander (Salamander keyserlingii). There are a total of 34 bird species, belonging to 8 orders, within the Srednyi Kaiemraveem and Myechyerova River basins. Of these species, the Passeriformes and the Charadriiformes are the dominant species encountered and have the largest populations. The Passeriformes consist of the Northern House Martin (Delichon urbica), the yellow wagtail (Motacilla flava), the white wagtail (Motacilla alba), the Red-throated pipet (Anthus Cervinus), the Bluethroat (Luscinia svecica), the Northern wheatear (Oenanthe oenanthe), the Dark-throated thrush (Turdus ruficollis), the Naumann’s thrush (Turdus naumanni), the Arctic warbler (Phylloscopus borealis), the Pallas’s Reed Bunting (Emberiza pallasi), the Snow bunting (Plectrophenax nivalis), the Hoary redpoll (Acanthis hornemanni), and the Northern raven (Corvus corax). The Charadriiformes consist of the rarely seen: the Golden plover (Pluvialis fulva), the Greater ringed plover (Charadrius dubius), the Wood sandpiper (Tringa glareola), the Gray tailed tattler (Tringa brevipes), the Temminck’s stink (Calidris temminckii), the Black-tailed godwit (Limosa limosa) and commonly seen: the Common Sandpiper (Actitis hypoleucos),


the Long-tailed Jaeger (Stercorarius longicaudus), the Pomarine Jaeger (Stercorarius pomarinus), and the Herring gull (Larus argentatus). The Anseriformes that could possibly habitat the area include: the Harlequin duck (Histrionicus histrionicus), and more rarely, the Green winged teal (Anas Crecca), the Greater scaup (Aythya marila), the White winged scoter (Melanitta deglandi), and the Black scoter (Melanitta americana). The birds of prey (Falconiiformes) could use the territory as a hunting ground although their numbers are scarce in the area. The Rough-legged buzzard (Buteo lagapus) uses similar territory for nesting. The Peregrine Falcon (Falco peregrinus) and Gyrfalcon (Falco rusticolus) feed along the middle reaches of Kaiemraveem Creek approximately 5 to 10 kilometers from the site. The Gaviiformes, Galliformes, and the Strigifomres are represented by one species each: the Red-throated loon (Gavia stellata) nests far from the territory in the middle reaches of Kaiemraveem Creek, the Willow ptarmigan (Lagopus lagopus) is seen on territory similar to the site, and the Sandhill crane (Grus Canadensis) and the Short-eared owl (Asio flammeus) could potentially (rarely) been seen. Among rare and endangered species that can be located within the area are:

1. mammals protected under the Russian Red Book for Far Eastern Russia: a. Yakutian snow sheep; and, b. Wild northern reindeer.

2. birds protected under the Russian Red Book for Far Eastern Russia: a. Peregrine Falcon; and, b. Gyrfalcon.

4.11 Hydrobiology The hydrobiological make-up of the benthic communities of the waterways near the Kupol site is based on analogous sites within the Anadyr basin. In general, there are 13 systematic groups of organisms present within the water systems: Turbellaria (planaria), Nematoda (nematodes), Oligochaeta (oligochates), Mollusca (mollusks), Hydracarina (water mites), Amphipoda (gammarids), Ephemeroptera (may flies), Plecoptera (stone flies), Heteroptera (water bugs), Coleoptera (water beetles), Megaloptera (sialid flies), Trichoptera (caddis flies), and Diptera (dipteran). In first and second order streams, there are usually 9-10 groups of organisms (planaria, nematodes, oligochates, mollusks, gammarids, mayflies, stone flies, caddis flies, dipteras, and rarely, water bugs). In mountain waterways and in the make branches of the river there are the same species mentioned above (except for water bug, beetles, and sialid flies - these are characteristics of secondary streams).


4.12 Ichthyology An ichthyological investigation of the rivers around the Kupol deposit was conducted to determine the fisheries status prior to construction and reserves of the stream. The rivers and streams of the area are of submountainous type with fast flow, cold and clear water. The main waterways that can be impacted by the site include the Srednyi Kaiemraveem. This is a Category 1 Fisheries river. The remaining waterways (1st, 2nd, and 3rd creeks) are season creeks and are not flowing for the majority of the year (and do not contain fish). During the investigation there were only two types of fish caught:

• The Siberian Grayling (Thymallus arcticus); and, • Slimy sculpin (Cottus cognatus).

It is believable that such a sparse population of fish species exists in the region due to the short vegetation season, the low water temperatures and the limited amount of food available for the fish, and the fact that the lengthy duration of total freeze-up of the river during the winter months.

4.13 Archeology The site was surveyed during the 2003 field season by an archeologist. This included walking the site and surrounding areas to look for any identifying artifacts or areas that may have archeological significance. There were 4 areas identified for further investigation that includes:

1. Srednyi Kaiemraveem 1: 4.5 kilometers south/southwest of the exploration site for the Kupol deposit;

2. Srednyi Kaiemraveem 2: 3 kilometers to the north of the exploration site for the Kupol deposit;

3. Srednyi Kaiemraveem 3: the left bank of the river, 2.5 kilometers southeast of the exploration site for the Kupol deposit; and,

4. Ittiil’iveem 1: 5.5 kilometers northeast of the exploration site for the Kupol deposit. None of these sites are located in areas where construction is scheduled to occur. All of these sites will be re-visited during the 2004 summer field program.

4.14 Protected Areas There are no nature preserves or protected environmentally sensitive areas in the vicinity of the Kupol site. The closest nature preserve to the mine site is more than 500 kilometers northeast from the property. The closest sanctuaries are located more than 280 kilometers away from the site. The closest sanctuary is located approximately 280 kilometers to the


south of the property in the confluence of the Main and Anadyr Rivers. This location is a nesting sanctuary for several species of protected birds.

4.15 Land Use The land surrounding the Kupol site currently is within the land used by the Lamutskoye agricultural community for reindeer herding and supporting traditional indigenous activities for hunting and gathering. The land is owned and administered by the municipality of Anadyr, region of the Chukotka Autonomous Okrug. Indigenous groups that use the surrounding land include: Chukchi, Lamut, and Chuvanyets. All of these indigenous persons belong to the Anadyr Regional Association for Indigenous Persons of Chukotka.

4.16 Socioeconomics The Russian Federation is country that occupies almost twice the land mass of the United States. It extends approximately 9,000 kilometers (from Europe to Northern Asia) and covers an area of 17 million km2. It is divided into 89 administrative territorial divisions including: 21 republics, 6 territories, 49 oblasts, 1 autonomous oblast, and 10 autonomous okrugs. Additionally, Moscow and St. Petersburg have independent status at the Oblast level. It is also sometimes split into Northern Asia (East of the Urals) and Europe (West of the Urals). A smaller geographical boundary within the Russian Federation is the Russian Far East. The Russian Far East (RFE) is made up of 10 administrative territories and occupies a land mass of 6.63 million sq. km (the entire eastern 1/3 of Russia). These territories are chosen due to their economic ties with the Pacific Rim. The 10 administrative territories of the RFE are: the Amur Oblast, the Chukotka Autonomous Okrug, the Jewish Autonomous Oblast, the Kamchatka Oblast, the Koryak Autonomous Oblast, the Khabarovsk Krai, the Magadan Oblast, the Primorskii Krai, the Republic of Sakha, and the Sakhalin Oblast. The Chukotka Autonomous Okrug is the farthest northeastern part of Russia (approximately 7300 kilometers from Moscow). The Chukotka Region was formed in 1940 as an independent national unit but later was included in regions of Kamchatka, Khabarovsk, and Magadan. In 1980, Chukotka obtained a status of autonomous region within the Magadan Oblast. In 1992, the region gained full, independent status as an Autonomous Okrug. The Chukotka Autonomous Okrug occupies an area of 737,700 sq. kilometers and is the 6th largest area in Russia. It is bordered by the Magadan and Koryak Regions and the Sakha Republic (Yakutiya) and is divided into eight administrative districts: Chukotkskii, Providenskii, Schmidtovskii, Iul’tinskii, Bilibinskii, Chaunskii, Anadyrskii, and Beringovskii. Additionally, the administrative center of the Chukotka Autonomous Okrug is the town of Anadyr. It is not subsumed within any district and is situated on the eastern shore of Chukotka.


The three regions that will have activity related to the Kupol Project include the Chaunskii Region (transportation and logistics), the Anadyrskii Region (deposit and most facilities), and the Bilibinskii Region (Airstrip and explosives storage). The Chaunskii Region is situated in the northern region of the Chukotka Autonomous Okrug. It is bordered by the Bilibinskii Region to the west, the Schmidtovskii Region to the east, the Anadyrskii Region to the south, and the Arctic Ocean to the north. The total area occupied by the Chaunskii Region is 58,000 km2. The regional center is the town of Pevek. The Anadyrskii Region is situated in the eastern region of the Chukotka Autonomous Okrug and is bordered by the Bilibinskii Region, Chaunskii Region, the Schmidtovskii Region, the Kamchatka Oblast (to the south) and the Bering Sea. The total area occupied by the Anadyrskii Region is more than 250,000 km2. The regional center is the town of Ugol’ny Kopi. The closest population center to the site within the Anadyrskii Region is the native village of Lamutskoye. It is 140 km south of the Kupol site. The Bilibinskii Region is situated along the western edge of the Chukotka Autonomous Okrug and is bordered by the Magadan Oblast (southwest), Yakutia (northwest), the Chaunskii Region (northeast), the Anadyrskii Region (east), and the Kamchatka Oblast (south). The total area occupied by the Bilibinskii Region is 175,000 km2. The regional center is the town of Bilibino. The closest population center to the site within the Bilibinskii Region is the native village of Illirny. A map showing the location of the project in relationship to the larger geographical boundaries is provided as Figure 12. 4.16.1 Demographics The overall population trend in Russia is decreasing in population due to migration out of the country and a death rate that is outpacing the birth rate. The current population estimate of Russia is 144 million persons (8.44 persons/sq. km). Most of the population in Russia (78%) is located in the European part of Russia. The largest population centers in Russia include: Moscow (~9 million), St. Petersburg (~4.5 million), Novosibirsk (~1.5 million), Yekaterinburg (~1.4 million), and Samara (~1.2 million). The current population is a decrease of 0.33% over reported 2003 numbers and represents a continuation of the population decline in Russia that began in 1991. It is predicted [2] that the population in Russia will continue to decline at least through 2050 to an estimated 104,000 million people. The total population of the Russian Far East (based on 2001 numbers) is approximately 7.1 million persons. The ten largest cities hold almost 40% of the population within the region. For the last 10 years, the population of the RFE has steadily declined due to declining birth rates throughout Russia and the migration back towards European Russia. The population density within the RFE is significantly less than European Russia (1.1 persons/sq. km).


Figure 12. Location of the project in relationship to the larger geographical boundaries


The population of the Chukotka Autonomous Okrug is currently estimated at 51,410 persons. This is a reduction of almost 70% compared to the 1989 population, 25% reduction from 2000 population figures, and 4.5% reduction from 2002 population figures. Of the 51,410 inhabitants of Chukotka, more than 65% live in the urban centers that include: Anadyr (7,324 persons), Bilibino (5,946 persons), Pevek (4,780 persons), Ugol’ny Kopi (3,863 persons), Egvekinot (2,413 persons), Provideniya (2,723 persons), and Beringovskii (1,998). The administrative center of the Chukotka Autonomous Okrug is the town of Anadyr. It is independent of the 8 regions and is the base for all the Okrug government agencies. The population of Anadyr is 10,873 persons. This is a very slight decrease over 2002 numbers (~1.5%). The population of the Chaunskii Region is 6,962 persons. This is a reduction of more than 33% from the 2001 population and a 9% reduction from the 2002 population. Of the 6,962 persons living in the Chaunskii Region, almost 3/4 of the population lives in the port town of Pevek. The remaining population is spread out in 6 population centers: Baranikha (15 persons), Komsomol’skiy (508 persons), Apapel’gina (5 persons), Rytkuchi (482 persons), Yanranai (226 persons), and Aion (330 persons). The population density for the region is 0.11 persons per sq. km. This number drops to 0.03 persons per sq. km when excluding the port town of Pevek. The population of the Bilibinskii Region is 8,553 persons. This is a reduction of more than 25% from the 2001 population and a 3% reduction from the 2002 population. Of the 8,553 persons living in the Bilibinskii Region, approximately 70% of the population lives in the town of Bilibino. The remaining population is spread out over 9 population centers: Aliskerovo (5 persons); Vstrechnyi (7 persons); Anyusk (535 persons), Illirny (301 persons), Keperveyeem (461 persons), Omolon (936 persons), Ostrovnoye (355 persons), Vesennyi (4 persons); and Dal’nyi (3 persons). The population of the Anadyrskii Region (excluding the administration center of the town of Anadyr) is 7,324 persons. This is a reduction of more than 8% from the 2002 population. Of the 7,324 persons living in the region, almost 50% live in the town of Ugol’ny Kopi. The remaining population is spread out over eight population centers that are mostly centered along the coast and the winter road between Anadyr and Lamutskoye: Shakterskii (93 persons), Markovo (865 persons), Vaegi (457 persons), Kanchalan (635 persons), Krasneno (118 persons), Lamutskoye (203 persons), Snezhnoye (305 persons), Ust’ Belaya (869 persons), and Chuvanskoye (217). Indigenous Population There are currently 40 ethnic groups within the Russian Federation designated as “small numbered native peoples”. They have occupied scattered areas throughout Russia for at least 20,000 years. This includes areas around the steppes of Southern Siberia, Lake Baikal, and Yakutia (livestock herders and crop-growers); north-east Siberia (nomadic reindeer hunters and fishers); and far north-east Russia in Chukotka (settled fishing communities). Between the 10th and 13th centuries, human migration from the south into Central Asia forced some


original inhabitants north and eastward. Currently, the population of indigenous peoples in Russia is estimated at more than 200,000. Within the RFE, paleo-Siberians are considered the 1st indigenous peoples to inhabit the area. This includes the Chukchi, Yukagir, Koryak, Kerek, Eskimos, Kamchadal, Nivkhi, and the Ainu. These tribes were later (3rd century, A.D.), joined by Evenks, Udege, Uiltra, Evens, Nanai, and Yakuts. According to the 1989 census, there were approximately 89,000 indigenous people located within the RFE. The largest populations were in Sakha, Khabarovsk, and Chukotka. Within the Chukotka Autonomous Okrug, indigenous people make up almost 1/3 of the population. The largest indigenous group is the Chukchi (~13,000 persons), Eskimos (~1,500 persons), Chuvantsi (~1,000), Evens (~800), and Yukagirs (~150). The remaining groups (Evenk, Lamut, Komi, Itelmen, Kamchadal, Nenets, and Orochi) all number less than 100 persons for the region and are generally integrated into other indigenous communities. The Chukchi (also known as the Luoravetlans) live mostly in indigenous-dominated villages are derived of two groups: the nomadic reindeer breeders and the sedentary (approximately 70%), coastal sea mammal hunters (approximately 30%). The two groups have been traditionally linked by trade (exchange of sea mammal and reindeer products). Intensive reindeer breeding is the economically most important occupation. Their herds often exceed 1,000 animals, graze the open tundra in the summer and migrate between protected areas in the winter. In addition to reindeer herding and sea mammal hunting, fishing and fur trapping are important parts of their traditional lifestyle. Fishing is developed in the Anadyr, Kolyma and Chaun River mouths. The Eskimos (also known as the Yupik) live mostly along the Chukotkan coast and Vrangel Island. They are the native population of the Bering Sea coasts and previously lived farther inland in eastern Chukotka. The traditional subsistence lifestyle is very similar to the Chukchi and includes marine hunters and other hunting/fishing as secondary occupations. Sea mammal hunting techniques are considered to be the most advanced in the world. The Chuvantsi are ethnically derived from the Yukagir clans, which reside in Western Chukotka, along the Anyui, Palyavaam, and the upper parts of the Amguema Rivers in the 17th century. Additionally, some groups have emigrated to the upper Anadyr River and live in the small indigenous settlements of Markovo, Chuvanskoye, Lamutskoye, and Tavaivaam, Slautnoe, and Aianka. The Chuvantsi are traditionally nomadic reindeer breeders, hunters, trappers, and dog breeders. Additionally, the people of Markovo work in fish processing and green house gardening. Anadyr Region The population of indigenous persons within the Anadyr Region was 3033 based on 1/1/03 figures. Of this total, more than 60% are Chukchi, 25% are Chuvants, 5% are Lamuts, and 4% are Evens. The indigenous population is scattered through 10 population centers within the Anadyrskii Region with the largest indigenous population centers in Ust’ Belaya (~685


indigenous persons), Kanchalan (~540 indigenous persons), Markovo (~476 indigenous persons), Vaegi (~379 indigenous persons), Snezhnoye (~317 indigenous persons), Chuvanskoye (~222 persons), Lamutskoye (~212 persons), Ugol’ny Kopi (~93 persons), Krasneno (~92 persons), and Shakhterskii (~17 persons). The closest indigenous settlement to the Kupol property located within the Anadyrskii Region is the native village of Lamutskoye. It is located 140 kilometers to the south of the site. Of the 212 persons living in Lamutskoye, there are Chukchi (~111 persons, Lamuts (~63 persons), Chuvants (~28 persons), Evens (5 persons), and Yakuts (5 persons). Bilibino Region The indigenous population in the Bilibino Region number approximately 2242 persons based on data provided on 1/1/2004. Of this amount, more than 41% are Evens, 40% are Chukchi, and the remaining 19% are Yukagirs, Nenets, Koryaks, Evenks, and Eskimos. The indigenous population is scattered throughout six different population centers with more than 1/3 (~774 persons) living in the indigenous settlement of Omolon. The rest are located in Bilibino (~337 persons), Anyusk (~334 persons), Ostrovnoye (~331 persons), Illirny (~233 persons), and Keperveyeem (~219 persons). The closest indigenous settlement to the Kupol property within the Bilibino Region is the native village of Illirny. It is located 80 kilometers northwest from the site. The total population of the village is 301 persons with 233 indigenous persons. This population includes: Chukchis (217 persons), Evens (10 persons), Koryaks (5 persons), and 1 Yukagir. Chaunsk Region The indigenous population of the Chaunskii Region is 846 persons with an overwhelming number of Chukchis (~841 persons) based on data provided on 1/1/2004. The remaining population consists of Eskimos (2 persons), Mansi (2 persons), and Even (1 person). The indigenous population is located in three separate native villages: Rytkuchi, Aion, and Yanranay. The closest indigenous settlement to the Kupol property within the Chaunskii Region is the native village of Pytkuchi. It is located 270 kilometers to the north of the site. Additionally, along the Kupol-Dvoynoye winter road there is a former storage area of the farming combine “Pytkuchi” (currently reorganized as the farming combine “Chaunskoye”) and is a transfer point for reindeer and fish products. 4.16.2 Health and Welfare The current life expectancy rate for the total population of Russia is 66.39 years (up from a 65.8 life expectancy in 2002. This includes a 59.91 life expectancy for Russian males and a 73.27 life expectancy for Russian females.


In Russia in 2003, the infant mortality rate is 16 (this is the probability of dying between birth and exactly 1 year of age expressed per 1000 births). The under 5 mortality rate (the probability of dying between birth and five years of age as expressed per 1000 births) for Russia was 21 in 2003. This is significantly better than in 1960 (64) and ranks Russia 115th out of 284 countries surveyed by UNICEF [Statistical Tables]. Russia has an excellent program for immunization. As a result, more than 95% of all children have been vaccinated for TB, DPT3, Polio3, measles, and hepB3. The prevalence of HIV in Russia in 2003 was on par with the world numbers. The adult prevalence rate (15-49 years old) was estimated to be 1.1%. Mortality rates are high for able-bodied males in rural areas. This is because the population is only served by poor health care centers and lacking basic sanitary facilities, thus causing the farming communities to transform into care pockets for the elderly, the indigent and the sick. There is concern that some indigenous nationalities such as the Evenks and the Nenets have incurred catastrophic declines in life expectancy, high rates of sickness that lead to death and there are predictions that some of the indigenous populations may become extinct. The primary natural cause of death is diseases of the circulatory system, cancer and respiratory diseases respectively. Among people of working age, 41% of deaths are related to unnatural causes. The number of alcohol related deaths also has climbed, in some areas alcoholism has reached epidemic proportions.

4.17 Economic Baseline Conditions 4.17.1 Russia Since the 1998 economic crisis when the exchange rate of the US dollar increased from 6 to 24 rubles in less than 6 weeks, Russia has had five straight years of growth (based on 2003 data). During this time, the Russian Gross Domestic Product real growth rate (GDP) has averaged more than 6.6% and inflation has averaged more than 29.5% (see Table 20). Currently (2003), the GDP per capita is estimated to be US $8900, which is significantly less than that of the developed countries in the West.

Table 20. Russian Inflation and GDP (real % growth) Year Inflation (%) GDP (% of Change) 1999 85.7 6.3 2000 20.8 10.0 2001 21.6 5.1 2002 16.0 4.7 2003 13.6 7.3

The GDP by sector is made up of agriculture (5.2%), industry (35.1%), and services (59.8%). More than 25% of the population is considered to be below the poverty line. Of the 71.68 million people that make up the work force, at least 8.5% are categorized as unemployed. This number may be even higher as unemployed Russians often do not register for unemployment benefits (which are used to calculate unemployment numbers). Additionally, a large number of persons are underemployed.


Russia is one of the richest countries in terms of natural resources. It has considerable deposits of oil and gas, coal, precious metals, iron, copper, zinc and large tracts of timber. The industrial production growth rate is 7%. The primary export products include: petroleum and petroleum products, natural gas, wood/wood products, metals, chemicals, and a wide variety of military hardware. Strong oil export earnings have allowed Russia to increase its foreign reserves from only $12 billion to some $120 billon at the end of 2004. Oil, natural gas, metals and timber make up more than 80% of exports and the primary imports are machinery and equipment, consumer goods, medicines, meat, sugar, and semi-finished metal products. An estimated 472,845 families live “in extreme poverty” in Russia, according to Russian statistics. The Labor Ministry estimates that they are 866,000 children in these families. This translates to about 70% of families with children live below the poverty line in Russia. 4.17.2 Russian Far East The economy of the RFE is hindered by terrain, poor weather, and lack of federal attention. Infrastructure is poor due to permafrost conditions and most population centers are isolated during the warm months of the year. There are two major railroads that run parallel to each other (in an east-west direction) that support industry. The northern regions have no railroads and construction is expensive due to climatic conditions and permafrost. Most regions are served by a regional airport and towns along the coast have a port. Major international airports in the RFE include Vladivostok, Khabarovsk, Sakhalin, Magadan, and Anadyr. The largest ports in the Russian Far East are Vostochny (Primorskii Krai), Nakhodka (Primorskii Krai), Vladivostok (Primorskii Krai), Vanino (Khabarovsk), Sovetskaya-Gavan (Khabarovsk), Kholmsk (Sakhalin), and Petropavlosk (Kamchatka). The RFE economy is based on the export of four primary natural resources: precious metals and gems, marine products, timber, and fuel. These exports account for 80% of all export earnings. The primary trade partners (60% of all trade) include: China, Japan, South Korea, and the United States. Imports are primarily manufactured goods and are primarily from China and Japan. Although the RFE economy is under-developed and produces about 5% of the total industrial output it produces about 75% of Russia’s fish and marine products, more than 95% of Russia’s tin and diamonds, 50% of Russia’s gold, 10% of all timber products, 25% of Russia’s platinum, and a large percentage of Russia’s oil and gas. 4.17.3 Chukotka Autonomous Okrug Due to the remote location, climate, and lack of labor force, the Chukotka Autonomous Okrug (ChAO) is based primarily on resource development. From 1992 – 2001, the ChAO (along with most of Russia) did not generate enough revenue to fund the okrug budget


commitments. This period was characterized by the decline of the manufacturing sector, large scale unemployment, and non-payment of wages. Base metals have been at the heart of the Chukotka economy since 1941. The gold sector has traditionally made up the largest portion of the industrial output for the region. This includes 463 gold deposits (457 placer and 6 lode). From 1960 – 1990 about 700 tonnes of gold were recovered (mainly from placer mining) with annual outputs reaching 30 tonnes per year. This number has dipped to approximately 6.4 tonnes of gold in 2002. This is due to the area’s remoteness and lack of investment in the region. Additionally there are tin, tungsten, copper, and coal deposits. Total resources of coal are estimated to be 57.5 million tonnes. There are also deposits of molybdenum, boron, bismuth, titanium, lithium, beryllium, iron, arsenic, nickel, cobalt, and semi-precious stones in the ChAO. Sibneft (a large petroleum company managed by the governor of Chukotka) is currently exploring the ChAO for oil. To date, Sibneft has invested more than US $70 million in the region. It is estimated that the reserves found by Sibneft in the Verkhne-Telekaiskoye field could be as large as 6.6 million tonnes. In addition to gold mining, the region has the following industries: power, fuel, publishing, food and food production. The ChAO total industrial output (in million Rubles) is provided in Table 21.

Table 21. ChAO Industrial Output (in million Rubles) Sector 1999 2000 2001 2002 2003

Total Industry 1900.6 2419.7 2928.9 3674.6 4465.2 Electric Power 512.5 579.4 753.0 1120.4 1412.4 Fuel Industry 136.6 163.9 180.2 238.7 291.1 Mining of Precious Metals 1162.1 1488.5 1646.2 1714.3 1998.2 Chemical Industry 0.1 0.0 0.0 0.0 - Machine Manufacturing 3.1 3.5 4.3 0.0 - Timber Industry 1.4 1.6 1.4 0.0 - Building Materials 0.6 0.7 0.6 0.0 - Light Industry 0.7 3.5 1.8 0.0 - Food Industry 77.8 174.8 337.4 597.7 758.8 Publishing Industry 2.8 2.8 3.2 3.1 3.0 Other Industries 2.9 1.0 0.8 0.4 1.7

The power sector of the Okrug serves the internal needs of Chukotka and the structure is determined by geography. The power generation system includes the large electric power plants and the nuclear power station located in Bilibino. Remote villages operate on diesel and coal-fired power plants that are not connected to the power grid. The food industry represents 18% of the industrial output of the region. Bread, meat, milk, confectionary, fish products, and beer are made locally. Indicators showing the output of the food industry are provided in Table 22.


Table 22. Industrial Output – Food Industry Item Measure 1998 1999 2000 2001 2002 2003

Food Product Output: Full Milk Product Tonnes 361 290 515 553 513 469 Bread Products Tonnes *1000 2.0 2.2 2.7 2.5 3.3 3.1 Meat and Meat Products Tonnes 652 630 508 490 337 312 Confectionary goods Tonnes 68 72 53 34 54 73 Beer Deciliters *1000 12 9 14 14 11 10.5 Fish Products Tonnes 709 776 2271 4500 22895 43725 Agriculture in the district is centered primarily on the traditional agriculture of the native peoples – reindeer herding, fishing, and hunting. Since 2001, there have been concerted efforts to raise the number of reindeers in the region. In 2003, the reindeer herds were in excess of 110,000 head (more than a 120% increase). Agricultural output is shown in Table 23.

Table 23. Agricultural Output of ChAO. Units 1998 1999 2000 2001 2002 2003 Agricultural

Products MMrubles 64.7 112.0 85.6 107.7 136 69.1 Milk Tonnes 233 107 149 71 107 133 Eggs MMeggs 0.4 0.4 0.3 0.7 2.7 3.8 Cattle/Poultry Tonnes 1510 1476 887 882 766 626 Vegetables Tonnes 100 100 200 209 130 100 Potatoes Mtonnes 0.1 - 0.1 0.1 0.1 -

The primary mode of transportation within the region is air travel. There are 10 airports within the Okrug, including two international airports (Anadyr and Provedeniya). The airports of Chukotka are linked with Moscow, Khabarovsk, Bratsk, Omsk, and Magadan. The air travel industry within Chukotka employs 769 persons. Over the last several years the Chukotka Autonomous District established a system of external trade, relying on the geographical position and the social-economic development of the area (see Table 24).

Table 24. Import-Export Structure of the Chukotka Autonomous District Structure 1998 1999 2000 2001 2002 2003

Export USD * 1000 35.8 31.2 307.0 51.1 132.8 138.6 Import USD * 1000 686.2 3313.4 1783.2 14525.0 14366.2 35345.8

The basic part of the trade balance of the territory was made up of the purchase of goods from the Asian Pacific Region (USA, Canada, Japan, and Republic of Korea). From 2001 imports were added from Germany, Turkey, Finland, Netherlands, Switzerland, and Taiwan. The basic items of import are produce, petroleum-chemical and petroleum products, manufactured equipment and machine-building equipment; transport equipment, computer equipment and construction material.


The Chukotka Autonomous District has customs facilities and a customs post in the Airport at Anadyr and a Marine Customs Post in Provideniya. The number of enterprises involved in external trade activities varies from seven to nine organizations and these enterprises represent differ types of ownership structure (private and government owned). The increase in trade since 2001 is brought about by the increase in construction activities involving foreign contractors from Turkey, USA, and Canada, and the import of construction materials and construction equipment. Housing construction is increasing in the Native villages, the municipal regional centers and in the City of Anadyr. There is reconstruction and construction of utilities, power plants, and buildings of public importance. The Government of the Chukotka Autonomous Okrug is interested in encouraging the external trade sector and wants to direct the region to develop export-oriented industries. The basic economic sectors of the District at present is the mining industry and, in the agricultural sector, the reindeer herding and marine mammal and hunting industries, as well as fishing and fish processing. Quality of Life of the Population The quality of life of the population has been raised due to the efforts of the Chukotka government to strengthen the economic base of the region. The quality of life from 2000-2003 is characterized by a growth in the monthly wages and the increase in income for the population. Increase in income and purchasing power of the population has increased the overall spending of the population in goods and services. In 2003, the average monthly wages of a worker increased 3.2 times compared to 2000. In evaluating the material well-being of the population, the poorest section of the population is identified and people are given individually directed assistance utilizing social normative standards, the size of the minimum allowable wage and the average income of a single person. Table 25. Chukotka Autonomous District Quality of life of the population, 2000-2003

2000 2001 2002 2003 Average wages of a single worker rubles 5 687 8 216 13 502 17 982 * Average income of the population rubles 3 995 6 581 9 193 7 845 * Average expenditures of the population rubles 1 971 4 419 6 856 4 970 * Minimum survivable wage, average per person, 4th quarter figure rubles 3 801 4 167 5 548 5 744

Cost of minimum basket of food goods, year end rubles 2 302 2 205 2 453 3 155 * considered preliminary and will be updated with the 2004 report. Anadyrskii Region The working population of Anadyrskii is approximately 8,800 persons. Of the workforce, approximately 19% work in large or medium sized transportation companies, 14% work in agriculture, 8% work in the fuel industry, 18% work in education or culture, 18% work as


public servants, 8% work in government, and 7% work in the health industry. The total number of registered unemployed is 170 persons. Within the Anadyrskii Region, the highest wages are paid for transportation and the fuel industry (an average of ~10,200 rubles based on 2002 data) and the lowest wages are paid for agricultural jobs (an average of ~3100 rubles based on 2002 data). During this time the minimum wage necessary to cover basic living costs was 7,016 rubles and the cost of the minimum basket of food goods was calculated to be 2452 rubles. The main transportation routes for the region are air and sea. The administrative center (Ugol’ny Kopi) has an international airport. The town of Anadyr has a sea port. Additionally, during the summer period, some of the rivers can be navigated with barges. The primary industries in the Anadyrskii Region are coal mining, fishing, and production of food products. In 2004 the amount of coal mined was 272,000 tonnes. There are also 4 reindeer herding brigades within the region: Markovskii, Kanchalanskii, Vaezhskii, and Imeni 1st Revkom of Chukotka. Bilibinskii Region The Bilibinskii Region has lost more than 50% of the population since 1996 due to migration out of the region. Of the total population, approximately 1450 persons are working and 167 persons have registered as unemployed. The average wage in the Bilibinskii Region is 11,481 rubles (based on 2002 data) with the highest salaries belong to those that work at the Bilibino nuclear power plant. Transportation in the region is primarily via airplanes. There is only one all-season road and that is between the town of Bilibino and Keperveyeem (approximately 35 kilometers). The rest of the villages within the region are only accessible on the ground by winter road. Agriculture within the region has suffered a severe downturn with milk production no longer occurring in the region, eggs no longer produced within the region, and meat production reduced to 1/3 the 2002 levels. Chaunskii Region Within the population of the Chaunskii Region, approximately 75% (5200 persons) live in the regional center of Pevek. The average monthly salary is approximately 11,200 rubles (based on 2002 data) and is 4.5 times higher than the 1997 levels. The average salary in agriculture is ~5550 rubles. The region is connected with the rest of the Okrug and Moscow by the airport located approximately 14 kilometers from the regional center of Pevek. The seaport at Pevek is the largest port in Chukotka and is typically open from mid-July through September.


The largest industries within the region are two large placer artels. Additionally there are two small mines (Dvoynoye and Sopka Rudnaya) along with other small placer mines. There is also a large trucking company located in Pevek. Agriculture remains underdeveloped in the region. Within the region there is only one farming brigade “Chaunskoye.” Their main activities are reindeer herding, fishing, and fur trading. There are currently more than 15,000 head of reindeer within the region. The total number of persons working within the farming industry is 185 persons.


5.0 Alternatives This section lists and compares feasible alternatives to the proposed project design, including the “without project” situation. The “without project” alternative is also known as the No Action alternative. Because of the remoteness of the facility and the fact that the ore body is in a fixed location, most alternatives were immediately rejected as non-economical. The components that were evaluated for alternatives were:

• Mining methods; • Cyanide Destruction; • Tailings disposal; • Power generation; and, • Potable water.

5.1 No Action Alternative The evaluation of a “without project” or “No Action” alternative assumes a cessation of all project activity. The No Action alternative assumes that the mine would not be approved and that mining and development would be halted immediately. Additionally, there would be no more physical disturbance to the area. 5.1.1 Socioeconomic Impacts The impacts from the No Action alternative are almost completely negative to regional and Okrug-level socio-economics. While the overall economic situation has somewhat improved under the latest administration, Medium term socio-economic trends (10 to 15 years) within the Anadyr, Bilibino, and Chaunsk regions have continued a downward spiral that includes decreases in the quality of: health care, school systems, recreation, living accommodations, and job opportunities. The Chukotka Okrug has been developed using mining and natural resource development as an economic base. Mining has been severely curtailed within the region over the last 20 years and most support industries have closed. Natural resource development has been a cornerstone of the current administration’s plan for developing the region. The Kupol project will be the largest natural resource development project in the region. Without the long-term prospects of the Kupol mine, the region will most likely continue to see a long-term downward trend in the economic situation of the region. The Kupol mine will employ more than 1,000 persons at full strength. Wages will be among the highest in the region. Additionally, a large multiplier effect is anticipated for collateral employment opportunities generated by Bema’s practice of using as much local support as possible. These employees will come primarily from the Chukotka Okrug and Magadan


Oblast. The overall socioeconomic impacts will be positive for the region. This includes: 25% of the profits from the project, US $5.00 per ounce of gold in the mineable reserves, taxes to the region in excess of 13 million dollars, a charitable fund that will provide more than US $3 million to the region, and taxes on wages earned within the Okrug. 5.1.2 Environmental Impacts The Kupol Mine has been explored for more than 10 years. Currently there are more than 250 ha of disturbed land. Disturbance of this area has impacted:

• Air quality; • Terrain; • Soils; • Permafrost; • Surface water quality; • Vegetation; • Fauna; and, • Aesthetics and visual resources.

The current design maximizes use of the existing cleared areas for construction and operations. The No Action Alternative will not allow for any restoration or reclamation of the existing impacts. Bema intends to reclaim all of the disturbed areas as quickly as possible during the course of construction and operation of the facility. The No Action Alternative would maximize existing impacts while preventing any remediation measures from being implemented. Therefore, the No Action Alternative is rejected based on the current state of the facility and the company’s intention to remediate existing impacts at the site. 5.1.3 Justification for rejecting the No Action Alternative The No Action Alternative was rejected on both an environmental and social basis. Environmentally, the No Action Alternative would continue to degrade the environment through erosion and sediment transport due to the existing land disturbance. The No Action Alternative would not allow for the reclamation of any of the existing land disturbance that was created during exploration. Additionally, the No Action Alternative was rejected on a socioeconomic basis. Without the project there will be no additional 1000 jobs, and no spin-off jobs for the region.

5.2 Mining The original mining design called for a large open pit and underground development. Initial production would have been from an open pit operation, overlapping with and replaced by underground production. Open pit operation was scheduled for 5 years with the underground mining beginning in the third year of operation. The stripping ratio on the upper benches


would have reached 37:1, with an overall stripping ratio of 20:1. Mining would have been performed with three faces using 4 loading units and 12 Cat 777D trucks. The peak mining rate was to be 22 million tonnes per year, with an average of 60,000 tonnes per day. Total waste rock generated in open pit operation would have been approximately 100 million tonnes. Of this material, approximately 40% was estimated to be acid-generating, 40% was estimated to be non-acid generating, and 20% was potentially acid-generating. The waste dump was to be located immediately west of the pit. The waste management strategy was to encapsulate the acid-generating material within the non-acid generating material and minimize infiltration, while promoting freezing. The entire facility would have been covered with a 2.5-meter cover of non-acid generating material to prevent thawing of the materials during the warmer months. Cold temperature leach column testing shows that acid generation is slowed down, but not necessarily reduced to negligible levels at temperatures approaching freezing. It is generally argued that both the acid-generating and acid-neutralizing reactions are retarded by low temperature and therefore the onset of potential ARD is delayed, but not eliminated. While this method was necessary due to the large volumes of waste, there still existed the potential for seepage to occur from the wasterock pile. The new mining design will generate approximately 13,000,000 tonnes of AG and PAG waste. At an assumed density of 2 tonnes/m3 the waste rock will require 6,500,000 m3 of storage. Under the new design, all material that has been determined to be acid-generating will be deposited upstream of the tailings impoundment, within the tailings basin, and will be eventually covered with tailings. The tailings starter dam will be constructed of non-acid generating and zoned placement of PAG rock waste rock and be sized to provide storage for the first three years of tailings production, AG and remaining PAG waste rock plus the design water storage and freeboard requirements. 5.2.1 Justification of rejecting large open pit alternative The new design represents best practices for placement of acid-generating materials. It minimizes waste generation and land disturbance. Reduction of the disturbance in this area has the potential to positively impact:

• Air quality; • Terrain; • Soils; • Permafrost; • Surface water quality; • Vegetation; • Fauna; and, • Aesthetics and visual resources.

Therefore, the mining method has been chosen, which minimizes the open pit while maximizes economic return on investment.


5.3 Tailings Basin Alternatives Two alternative methodologies were considered for tailings disposal for the Kupol Project:

1. A filtered, dry stack tailings storage system; and, 2. A conventional slurry tailings system with tailings impounded by a tailings dam.

Design and economic analyses have been completed by Bema that indicates the dry stack system to be overly costly, in terms of both capital and operating cost. As a result, the dry stack system is no longer being considered for the project. A slurry tailings system has been selected as the most appropriate for the Kupol Project. In addition to tailings storage, the impoundment will also provide permanent submerged storage of mine waste rock that is classified as acid-generating. Site selection studies were undertaken in 2003 and 2004 to identify potential tailings impoundment sites. This report presents feasibility design for two alternative sites that have been identified. Alternative #1 is located southwest of the plant site in the valley of a right tributary of the Middle Kaiemraveem River. Alternative #2 is located immediately south of the plant site on the southern part of the Granichnaya Mountain, in the valley of the Tretaya Creek. The key criteria for site selection were as follows:

• In a valley with a relatively level valley bottom gradient, and a confining point for dam construction, to achieve an acceptable ratio of tailings storage volume to dam fill volume;

• In a valley with modest drainage area, to avoid the necessity of handling large diversion flows; and,

• Located relatively near the plant site to minimize the length of tailings pipelines, and located at a lower elevation to minimize or avoid the necessity of pumping tailings.

Sites were also considered to the north, in the drainage of Starichnaya Creek, which drains into the Malyi Anyui River about 25 km north. The Malyi Anyui River flows northward and ultimately drains into the East Siberian Sea. No acceptable sites could be identified to the north of the plant site outside of the Kaiemraveem Creek drainage. Valleys to the north are generally characterized by subdued topography, with no favorable damsite locations, and also the runoff catchment area at any tailings impoundment site in the main valley of Starichnaya Creek would be large and require major diversions. 5.3.1 Environmental Impacts Socioeconomically, the two variations are equal. The cost for constructing Alternative #2 is less than for #1. The real difference is in the environmental impacts. Alternative #2 offers the following advantages:


1. Reduced footprint – by combining the wasterock facility and the tailings impoundment, the total disturbed area will be significantly reduced;

2. Reduced ancillary facilities – Alternative #2 is significantly closer to the mill than Alternative #1. Based on this, all utility corridors will be shorter and have less of a chance for a spill.

5.4 Site Power The anticipated power at the Kupol project is approximately 15.5 MW of power. Two options are available for the supply: a diesel generating plant constructed on the mine site, or an approximately 230-km long, high-voltage transmission line connecting the mine to the nuclear power plant in Bilibino. It does not appear that the nuclear power plant in Bilibino has the capacity to supply the mine site. The Bilibino nuclear power plant has four reactors installed that are rated 12 MW each with 50% electricity and 50% steam heat. Allowing for one unit down for maintenance at all times, the facility could produce a total of 18 MW of electricity. Deducting for safety factors, and it appears that they cannot supply the power needed for the mine. Nonetheless, the administration has asked CMGC to review the potential using power supplied for the facility. 5.4.1 Socioeconomic Impacts Using the existing nuclear power facility would continue jobs in Bilibino at the nuclear power facility. The facility is the largest employer of the area (approximately 750 employees). It also would provide a much needed power consumer and provide some relief from the on-going decrease in demand due to the shrinking economy and the population migration out of the area. Additionally, if the nuclear power plant were used, a 230 km power line would have to be constructed from Bilibino to the minesite. Construction would provide additional jobs and would potentially bring power to the 300 person village of Illirny along the route. 5.4.2 Environmental Impacts The nuclear reactors were put into operations in 1973 and have been approved to operate another 15 years before they are permanently shut down. The technology used in these reactors is outdated and long beyond the proposed operational life. By operating them at near capacity, the risk for an accident is increased. Additionally, more than 200 kilometers of line would have to be built to reach the mine. All of the construction would have to be in permafrost, requiring extensive permafrost protection measures to ensure that the permafrost is not permanent impacted. An additional line would have to be built to ensure that redundancy is built into the system.


5.4.3 Justification for choosing on-site power The region does not have the capacity to provide power to the minesite. This is especially true if CMGC rejects the notion of using an outdated nuclear power facility. The impacts from constructing a power line to the mine far outweigh the short term air quality impacts associated with an on-site, diesel power plant.


6.0 Impacts The environmental impact assessment included in this document covers the following main types of impacts:

• Air quality; • Terrain; • Soils; • Permafrost and hydrogeology; • Groundwater quality • Surface water hydrology; • Surface water quality; • Vegetation; • Fauna; • Aesthetics and visual resources; • Noise and vibration; • Socio-economics; • Archeology; and, • Wildlife preserves.

These impacts are evaluated for processes that occur during construction and operations. Mitigation and monitoring measures are presented in the Environmental Action Plan (EAP) that accompanies this EIA. The EAP will be updated and disclosed as needed to reflect changes in project design, mitigation measures, or monitoring.

6.1 Air Quality Impacts The nearest community (Illirny) is located more than 90 km from the site. This community does not have any major industrial emissions other than a small power plant. Therefore, for modeling of air quality impacts from the proposed facility, background concentrations of pollutants are assumed to be zero at the site. Construction and operation of the Kupol mine has been designed to minimize all possible air quality impacts. Air quality can be affected by three types of sources:

1. Stationary point sources – these are sources that are stationary and can reasonably be emitted through a stack or vent. For the Kupol mine site, the only significant stationary point source emissions are the diesel generators (25 MW total);

2. Mobile sources – these are sources that are mobile on the site. They include all haul trucks and personnel vehicles. Haul trucks will be a major source of dust emission at the site. Emissions from mobile sources will be intermittent and dispersed. Emissions from mobiles sources are not included in the impact assessment; and,

3. Fugitive sources – these are sources that are stationary and cannot be reasonably emitted through a stack or vent. These sources include roads, stockpiles, and waste rock piles.


6.1.1 Stationary Sources During exploration and construction the only stationary sources of pollution include:

• Diesel power station; • Lubricant storage; • Welding and metal cutting; and, • Underground mine ventilation.

With the exception of the diesel power plant, all stationary sources are assumed to be insignificant (emissions < 1.0 tonne/year). Based on dispersion modeling for operating conditions, the diesel power station operated during is not anticipated to significantly impact ambient air quality. During construction and operation the following stationary sources of pollution will exist:

• Diesel power station; • Crushing plant and transfer point wet scrubber; • Lubricant storage; • Welding and metal cutting; and, • Underground mine ventilation.

With the exception of the diesel power plant, all stationary sources are assumed to be insignificant (emissions < 1.0 tonne/year). Total emissions from the diesel power station are provided in Table 26 (based on an average consumption of 15.5 MW).

Table 26. Emissions from diesel generators Average

Throughput SO2 factor PM Factor CO Factor NOX Factor

MMBtu (lb/MMBtu) (lb/MMBtu) (lb/MMBtu) (lb/MMBtu) 51.6 0.51 0.09 0.85 3.2

Emissions SO2 PM CO NOX Generated (tonne/yr) (tonne/yr) (tonne/yr) (tonne/yr) 104.8 18.5 175 390 Notes:

(1) Emission Factors from USEPA AP-42, fifth edition (2) Based on Sulfur content of the fuel of 0.5% (3) Assumes 8760 hours of operation

6.1.2 Fugitive Sources Significant dust can be generated during mechanical disturbance of granular material exposed to the air. Dust generated from these open sources is deemed to be fugitive because it is not discharged to the atmosphere in a confined flow space. Fugitive sources include the open pit, site roads, stockpiles, tailings pond and waste rock piles.


The potential drift distance of particles is governed by the initial injection height of the particle, the terminal settling velocity of the particle and the degree of atmospheric turbulence. Theoretical drift distance, as a function of particle diameter and mean wind speed, has been computed for fugitive dust emissions. Results indicate that for a typical mean wind speed of 16 km/hr, particles larger than 100 µm are likely to settle out within 6 to 9 meters from the edge of the road or other point of emission. Particles that are 30 to 100 µm are likely to settle out within a few hundred feed from the road. The site road emissions (both active and inactive) will be distributed over the entire length of the site and will not disperse far from the road based on the horizontal velocity of the dust. Therefore, it is assumed that the emissions will not have any significant, long-term environmental impacts. The stockpiles are designed to be transient in nature. Because of the small size of the stockpiles and the transient nature, it is assumed that fugitive dust will be negligible over the life of the project. The tailings dam can also be a source of fugitive particulate dust under windy conditions. However, the tailings facility is predicted to quickly freeze after cessation of operations. Additionally, the waste rock pile will be located within the tailings facility and will be saturated and frozen to prevent acid rock generation and, as a parallel benefit, prevent dust mobilization. Control techniques for fugitive dust sources generally involve watering, chemical stabilization, or reduction of surface wind speed with windbreaks or source enclosures. The Kupol project will primarily employ watering methods at the site to reduce fugitive dust. It is anticipated that this will reduce fugitive dust emissions in excess of 50%. It should be noted that the site is covered in snow during 9 months of the year. Furthermore, during the summer months, there are several periods of high precipitation.

6.2 Topography and Land Disturbance The impacts associated with land disturbance include both direct and indirect impacts. Direct impacts include:

Changes in topography; Removal of vegetation and soil cover; Reduction in surface water quality; and, Changes to hydrobiological indicators.

These changes will last the duration of the project, approximately 7 years. Indirect impacts include impacts that can occur when direct impacts are not mitigated. They include:

• Wind and water erosion; and, • Chemical leaching.


The zone of impacts will be confined to the watershed basins directly at the site (primarily Kaiemraveem and Starichnaya) and can last in excess of 20 years without proper mitigation measures. Impacts to topography and land disturbance can also impact flora, fauna and surface water quality. These impacts will be addressed later in the Impacts Section. 6.2.1 Existing impacts Since 1998, the project license area has undergone significant impacts due to exploration. This includes the following land disturbances:

Table 27. Land disturbed during exploration Facility Disturbance (ha)

Mancamp Area 6.9 Helicopter pad and fuel farm 1.3 Explosive Storage 4.3 Internal Roads 20.0 Exploration Area 21.4 Mill/plant/new mancamp site 56.5 Total 110.4

Additionally, during the 2005 construction season, the land for the airport construction will be disturbed (70 ha). 6.2.2 Construction impacts Project goals for development of the deposit are based upon maximum use of previously-developed land in the vicinity of proposed sites for construction of the mine, processing mill, tailings facility, man camp, auxiliary facilities, and access road. Total new disturbance that will occur as a result of construction (including existing disturbance) are listed in Table 28.

Table 28. New disturbance occurring during construction Facility Disturbance (ha)

Mine Complex 28.6 Mill Complex 56.5 Tailings Facility 103.0 Old Mancamp/Reagent Storage 8.2 Airport 70.0 Roads and Service Corridors 24.0 Explosives Storage 8.6 Landfill 3.0

6.2.3 Operational surface disturbance Surface disturbance impacts during operations will be reduced based on the Reclamation Plan. The plan calls for reclamation of any construction area that is not used during operations or any surface disturbance that cannot be shown to be used during operations. Additionally, Bema will be aggressive about locating and reclaiming disturbed areas during operations.


6.3 Soil Impacts Impacts to physical characteristics of soil during construction, salvaging and stockpiling will include compaction, soil structure loss, increased core fragment content due to mixing. Soil loss may be induced through transportation and erosion. Increased soil runoff may occur due to decreased infiltration and reductions in permeability. 6.3.1 Acid Rock Drainage Mineralization at the Kupol Project is a quartz-adularia type, low-sulfidation gold system localized along a high angle vein complex and hosted in andesite flows and pyroclastic rocks. The vein records repeated episodes of dilation and mineralization and is composed of up to five different banded quartz and quartz breccia assemblages. Au and Ag mineralization is associated with the earliest dilational event and includes electrum, Ag-sulfides and Ag-sulfosalts. The andesites, which are Cretaceous in age, include a basal unit dominated by flows (85 to 90%) interbedded with discrete andesite crystal tuffs and an upper unit of tuffs (50%) interbedded with flows. The andesite package is overlain by younger (Tertiary age) rhyolite and basalt flows. Post-mineralization dikes, believed the feeders for the Tertiary rhyolite and basalt flows, parallel and, particularly in the south, crosscut the vein in a narrow 100 to 200 m wide corridor. The andesites along the vein have been altered and include: 1) a propylitic assemblage characterized by the presence of carbonate, chlorite, and, epidote; 2) a carbonate dominant assemblage including calcium carbonate, iron carbonate, and dolomite; 3) a texturally destructive argillic alteration characterized by strong clay alteration; 4) a sericite assemblage; 5) an adularia assemblage and; 6) silicification with quartz dominant veining and pervasive silica flooding of the adjacent country rock. The alteration assemblages are often superimposed upon one another because of the multiple and overlapping mineralization events. A total of 276 representative samples of waste rock representing each of the lithology, alteration, and mineralization assemblages and 95 composite samples of ore (from the metallurgical test program) were collected from splits of core from a) exploration and geotechnical drill holes of the 2003 and 2004 BGC drilling campaign and b) historic Russian exploration drill holes. These samples were analyzed in the geochemical investigation at Kupol following a staged sequence of: a) static testing designed to characterize the range and variability in the geochemistry of the different material types and; b) kinetic tests designed to characterize the dynamic performance and reactivity of the waste rock at Kupol. All of the samples (276 waste rock and 95 ore samples) underwent acid base accounting tests (ABA) and strong acid digestion with metals analysis (ICP) to characterize the sulfur speciation, acid-generating potential, acid neutralizing potential, and gross whole rock chemistry. One in three samples (78 plus 1 ore sample) from the 2003 drilling, covering the range in ABA and metals concentrations, were selected for short term leaching tests (Synthetic Precipitation Leaching Procedure, EPA-1312) to identify readily solubilized constituents. One in ten samples (23 plus 1 ore sample) from the 2003 drilling, covering the range of ABA, metal concentrations, and metal solubility, were selected for Net Acid-generating tests. A total of 32 samples, including 24 from the 2003 drilling (1 in ten) and 8 from the 2004 drilling were selected for humidity cell kinetic tests. Humidity cell tests are designed to quantify, under


laboratory conditions, ARD and ML reaction rates and to confirm ARD/ML behavior of sample classes estimated by the static tests. Additional test work included X-ray diffraction analyses (XRD) and grain size analyses of the humidity cell samples. The results of the geochemical test work for the waste rock lithology, alteration, and mineralization assemblages have been grouped into three ARD codes for use in the resource model: non-acid generating (NAG); acid-generating (AG); and potentially acid-generating (PAG). The ARD codes are based on simplified logging codes for pyrite mineralization and carbonate alteration from the detailed geologic drill hole logs that define four categories: high pyrite and low carbonate – AG; low pyrite and high carbonate – NAG; and high pyrite and high carbonate or low pyrite and low carbonate – PAG. Of the 276 waste rock samples, 51 (18%) are AG, 71 (26%) are NAG, and 154 (56%) are PAG. Results of the ABA tests have shown that all of the categories of waste rock at Kupol are sulfur-bearing and the dominant sulfur species in the waste rock is sulfide sulfur (Spy). Total sulfur (Stot) and Spy range from below the detection limit to 6 weight percent. The Stot and Spy concentrations decrease from AG through NAG waste rock. At the same time, all of the waste rock categories in the resource model contain some degree of acid neutralizing potential in the form of either carbonate veins or pervasive carbonate alteration. The neutralizing potential (NP) ranges from values less then 0 (negative NP indicates the presence of natural acidity in the rock) to almost 200 kg CaCO3/tonne. The NP increases from the AG to the NAG waste rock. Using the neutralizing potential:acid-generating potential (NP:AGP) ratio and the standard industry criteria for ARD potential: two thirds of the AG waste rock is acid-generating; over two thirds of NAG material is acid neutralizing; and half of the PAG material is acid neutralizing. The proportions of acid-generating, acid neutralizing, and uncertain categories for each of the ARD codes are shown in Table 29. The “uncertain” category is generally considered potentially acid-generating awaiting the results of kinetic testing to confirm the ARD potential.

Table 29. Proportions of acid-generating, uncertain, and acid-neutralizing waste rock in the ARD domains from waste rock samples of the ARD database

NP:AGP (1) >3 <3 and >1 <1 ARD domain Count Acid neutralizing Uncertain Acid-generating AG 51 21.6% 9.8% 68.6% PAG 154 53.9% 11.7% 34.4% NAG 71 71.8% 11.3% 16.9% (1) after SRK, 1992 Table 30 shows the model predicted tonnes of the waste rock to be mined from the open pit and underground using 1) the tonnes of each of the ARD domains from the resource model and 2) the proportions of acid neutralizing, uncertain, and acid-generating waste within each of the domains shown in the Table 29 above. The NP:AGP ratio of the waste rock mined will be measured in the field by Leco furnace sulfur and carbonate analyses of blast hole cuttings as part of the mine operations program.


Table 30. Model predicted tonnes of waste rock mined from the open pit and underground and

not used for back fill. NP:AGP 2006 2007 2008 2009 2010 2011 2012 2013 2014 Total Non-acid generating >3 1231000 2058000 1874000 1549000 1243000 0 0 0 0 7955000 Potentially acid-generating <3 and >1 276312 487774 477794 389000 303000 0 0 0 0 1933880 Acid-generating waste <1 886321 1646496 1821627 1589517 862544 0 0 0 0 6806504 Unclassified Tonnes 43147 60607 22987 0 0 0 0 0 0 126741 Total Waste Tonnes 2436780 4252877 4196407 3527517 2408544 0 0 0 0 16822125

Similar estimates of the acid-generating potential of the model NAG, PAG, and AG domains in waste rock types were found using the NAGpH and “NAG” values from the Net Acid Generation test work. Using the median values of NAGpH and “NAG”, AG is acid-generating (NAGpH less than 4.5, and “NAG” greater than 5 kg H2SO4/tonne) and NAG and PAG are non-acid generating (NAGpH greater than 4.5, and “NAG” generally less than 5 kg H2SO4/tonne). The ore samples have lower Stot (below detection to 4 weight percent) than portions of the waste rock and roughly equal proportions of sulfate sulfur (SSO4) and Spy. NP in the ore samples ranges from below zero to 115 kg CaCO3/tonne and the samples fall within the field of acid-generating material (NP:AGP ratio less than 1). The ICP test results provide an indication of metals that are present in anomalously high concentrations compared to similar, unmineralized rocks. Comparison of the results of the strong acid digestion and ICP analysis of the waste rock samples indicates that silver, arsenic, gold, bismuth, mercury, antimony, selenium, and thallium occur in AG, PAG, and NAG samples at concentrations above normal basalts. Silver, arsenic, gold, bismuth, cadmium, copper, mercury, molybdenum, lead, antimony, selenium, thallium, and zinc are anomalously high in the ore samples. The SPLP test provides an indication of elemental solubility under standard laboratory conditions in the waste rock using a 24 hour leaching procedure. The SPLP leachate ranges in pH from 3 to almost 10 and shows in increase an pH from AG samples (average 4 and median of 6) to NAG samples (average 5 and median 9). Alkalinity in the SPLP leachate also increases from 10 mg CaCO3/L in the leachate from AG samples to over 20 mg CaO3/L in the leachate from the NAG samples. The chemistry of all of the SPLP leachate from AG, PAG, and NAG samples is dominated by sulfate, which ranges in concentration from 1 to 2000 mg/L. Generally, the metal concentrations in the SPLP leachate are low. Using the maximum concentrations of soluble metals in the SPLP, leachate: aluminum, arsenic, beryllium, calcium, cobalt, copper, iron, magnesium, manganese, nickel, silicon, strontium, vanadium, and zinc all appear to be mobile in slightly acid solution. The humidity cell tests provide the key parameters to understanding the geochemical behavior of the waste rock at Kupol, namely the long-term (20 weeks) leaching characteristics of the different material types. Results indicate that the pH of the leachate derived from: a) AG waste rock samples is typically low, below pH 4; b) NAG waste rock


samples is mixed but dominantly alkaline, above pH 7 and; c) PAG samples is variable but generally above pH 7. Humidity cell leachate from AG samples typically has high sulfate, zero alkalinity, and high metal concentrations, although the sulfate and metals concentrations typically decline with time. The humidity cell leachate from NAG samples typically has low sulfate, moderate alkalinity, and moderate metal concentrations. Humidity cell leachate from the PAG samples is variable, with low alkalinity, generally low sulfate, and generally low metals concentrations. The full ARD report can be found in Appendix B. 6.3.2 Potential ARD Impacts The runoff from either unsaturated or saturated NAG rock is not expected to exceed direct discharge limits for any parameter. Thus, prior treatment of NAG rock runoff is not required before discharge to receiving waters provided sufficient dilution is available prior to the edge of the mixing zone, or reaching the ambient point-of-compliance, such that Russian Class 1 fisheries standard limits are met. Concentrations in runoff from AG waste rock: During 2006 and 2007, AG waste rock will be stockpiled in the catchment upstream of the tailings dam prior to starter dam completion in late 2007. Runoff from this stockpile will be minimized, collected, and treated as discussed in Appendix F – AMEC Water Management Plan. Note that due to the late start of pit construction in 2006 (about October), there will be negligible runoff from the AG waste rock stockpile during this year, and a treatment system will not be required for this source until spring 2007, and will only operate during 2007. AG rock will also be exposed in the pit walls during 2006 and 2007 and will yield runoff from direct precipitation on the open pit and thawback inflows from the pit walls as discussed in Appendix E, AMEC Site Water Balance. Pit inflows will be minimized, collected, as treated as discussed in the Water Management Plan (AMEC, 2005a). The late start of pit development in 2007 will result in negligible pit inflows requiring management during 2006. Thus, the pit inflow treatment system need only operate for 2007. From late 2007 through the duration of the project’s operational phase (about 2014), all pit inflows will be sent to the tailings pond, and any AG waste rock stockpile runoff will be contained behind the tailings dam. Thus, a water treatment system will not be required from late 2007 through 2014.

6.4 Permafrost A schematic diagram of typical thermal conditions in a permafrost zone is shown in Figure 11. An active layer that undergoes a seasonal freeze/thaw cycle defines the upper zone. In the simplest case, the ground remains permanently frozen below the active layer to the depth of the 0oC isotherm, which is defined as the base of the permafrost. The temperature variation within the permafrost is dependent on the geothermal gradient. Figure 13


represents the average case, free from any complexities of surface slope, surface cover and seasonal temperature variations. To preserve the integrity and stability of an underground excavation the thermal equilibrium must be maintained. This is extremely important when driving through highly fractured rock masses. The heat generated by diesel engines, fresh water and electrical power must be controlled. The impacts to permafrost layers and aquifer structure resulting from project construction are expected to be minimal. Beyond the disturbance resulting from the mine operations, all impacts associated with project development are expected to be superficial in nature and negligible upon surface resources. When construction is completed, all permafrost layers will return to original depths. Operational impacts are also expected to be minimal. Based upon tailings deposition freezing models, there are no long-term impacts to permafrost predicted. The mine is both open pit and underground. The underground mine is not expected to impact permafrost and surface structures have been designed to minimize impacts to permafrost. There will be some nominal thawback in the pit and underground but it is unlikely to create any long-term impacts on the permafrost.


Figure 13. Typical Permafrost Zone


6.5 Hydrogeological Impacts The entire facility is located in a permafrost environment. No direct impacts are anticipated to groundwater with the exception of the required consumption of potable and process water from the well located south of the facility. It is anticipated that 50,000 – 60,000 m3 will be required annually from this well. The aquifer has been preliminarily calculated to have as much as a 6 million m3 capacity. No long-term impacts are anticipated for this facility.

6.6 Surface Water Quality Impacts The Kupol Gold Project will include activities such as open pit mining, underground mining, ore stockpiling and processing, tailings and waste management, and camp operation that will generate volumes and qualities of water requiring management. The feasibility mine plan shows 7.1 million tonnes of material to be processed through the mill during a seven year mine life. The criteria used in assessing surface water quality impacts were to develop a water management plan, site water balance, and conduct surface water quality modeling efforts consistent with a feasibility level of analysis and two alternative tailings storage scenarios. The full report evaluating surface water quality impacts is provided in Appendix G, AMEC Surface Water Quality Modeling Report. Concentrations of all relevant water quality parameters were compared against the World Bank direct discharge criteria for open pit (World Bank, 1999a) and underground (World Bank, 1999b) mines, and against the Russian Class 1 fisheries standards (and drinking water/recreational use standards where no fisheries standard was available) (VNIRO, 1999). The following three general project phases were considered: (1) the initial year of site construction in 2006; (2) the first full year of pit development and prior to construction of the starter tailings dam in 2007; and (3) the operations phase of the project from 2008 through 2014 after the closed-circuit process plant-tailings pond water balance and process mill start-up has occurred. Two tailings alternatives were also considered separately in the modeling approach. The following four distinct source profiles were used in the coupled mass balance-geochemical ambient surface water quality model: (1) background concentrations in the Kaiemraveem River; (2) background concentrations in contributing catchments of the Kaiemraveem River and from natural surface runoff; (3) concentrations in runoff from NAG waste/quarried rock; and (4) concentrations in runoff from treated and untreated AG waste rock. Background concentrations were obtained from the baseline water quality and hydrology program conducted by Bema (2005). Concentrations in NAG and AG runoff were estimated based on knowledge of baseline and mine site hydrology, and the comprehensive 2003-2004 geochemical testing program (Atkin, 2005). Modeling of secondary geochemical reactions, and the potential alterations to estimated runoff water quality from NAG and AG rock surfaces due to particulate/colloidal phase retention in pore spaces and settling ponds, were investigated using an equilibrium based approach.


Dissolved phase concentrations from unsaturated and saturated NAG rock, and unsaturated AG rock, were then input to a series of linked mass balance calculations coupled to the site water balance model. Using this approach, and assuming complete mixing of all water sources prior to the ambient regulatory point-of-compliance on the Kaiemraveem River at a location 500 m downstream of the last potential input from project site runoff, the resulting concentrations of the various water quality parameters of interest were calculated. During 2006, runoff will occur from the following site facilities: plant and camp site, haul and access roads, the airstrip, and the explosives storage facility. No direct discharge exceedences into either the Kaiemraveem or Starichnaya Rivers, or their contributing catchments, under either of the tailings alternatives are expected from site runoff in 2006 under the World Bank criteria (World Bank, 1999a and 1999b). In addition, no Russian Class 1 fisheries standard exceedences are expected from site runoff into either the Kaiemraveem or Starichnaya Rivers under either of the tailings alternatives at the ambient regulatory points-of-compliance located 500 m downstream of the last project discharge into either of these rivers during 2006 under the Russian Class 1 fisheries standards (VNIRO, 1999). During 2007, runoff will occur from the following site facilities: plant and camp site, haul and access roads, the airstrip, the explosives storage facility, the AG waste rock stockpile, the ore stockpile, and pit inflows. Runoff from the AG waste rock stockpile, the ore stockpile, and pit inflows will be treated with lime neutralization prior to release into receiving waters. No direct discharge exceedences into either the Kaiemraveem or Starichnaya Rivers, or their contributing catchments, under either of the tailings alternatives are expected from site runoff in 2007 under the World Bank criteria (World Bank, 1999a and 1999b). In addition, with the exception of zinc, no Russian Class 1 fisheries standard exceedences are expected from site runoff into either the Kaiemraveem or Starichnaya Rivers under either of the tailings alternatives at the ambient regulatory points-of-compliance located 500 m downstream of the last project discharge into either of these rivers under the Russian Class 1 fisheries standards (VNIRO, 1999). During 2007, zinc concentrations at the ambient point-of-compliance in the Kaiemraveem River are expected to exceed the Russian Class 1 fisheries standards (VNIRO, 1999) by about 10% under both tailings alternatives. The minimal exceedence factor for this parameter, as well as its limited spatial and temporal duration (≤4-5 months over a 14 year project life), combined with generally conservative regulatory levels for zinc in freshwaters suggests that the short-term elevated zinc level in the Kaiemraveem River will not pose a significant threat to aquatic life. Over the period from 2008 through 2014, runoff will occur from the following site facilities: plant and camp site, haul and access roads, the airstrip, the explosives storage facility, and the downstream crest of the tailings dam. No direct discharge exceedences into either the Kaiemraveem or Starichnaya Rivers, or their contributing catchments, under either of the tailings alternatives are expected from site runoff from 2008 through 2014 under the World Bank criteria (World Bank, 1999a and 1999b). In addition, no Russian Class 1 fisheries standard exceedences are expected from site runoff into either the Kaiemraveem or


Starichnaya Rivers under either of the tailings alternatives at the ambient regulatory points-of-compliance located 500 m downstream of the last project discharge into either of these rivers from 2008 through 2014 under the Russian Class 1 fisheries standards (VNIRO, 1999). The current water and waste management planning for the project, as indicated by the results of the current modeling exercise, suggest there will be no significant water quality impacts on the Kaiemraveem and Starichnaya Rivers, including their contributing catchments, from development of the Kupol Gold Project between 2006 and 2014.

6.7 Vegetation Impacts Vegetation plays an important role in controlling erosion, providing wildlife habitat, and maintaining biological diversity. Disturbance to vegetation has the potential to adversely affect these functions. Because the proposed design maximizes use of areas that were previously cleared, existing disturbances make up the majority of the vegetation impacts.

6.8 Wildlife Impacts The existing disturbance and human presence associated with the man-camp and exploration have reduced the numbers of many species native to this area. The potential disturbance to wildlife populations in the area related to proposed mining activities include:

• Direct habitat removal; • Increased human presence; and, • Possible wildlife exposure to chemical and industrial reagents.

To date, more than 250 ha have been disturbed. Additionally, more than 300 persons are on site (expanding to more than 800 during peak construction). It is likely that any animals that lived in the area have moved to a different area. A firm policy regarding hunting, fishing, and poaching will be implemented by the project. The policy will serve to protect populations of large game species from poaching by the workers in the area. No firearms will be allowed in the License Area or at the mancamp except for security purposes. However, repair and maintenance of the winter road will improve access to an area rich in game, providing opportunity for commercial and amateur hunting. These potential impacts are not inherent to use of the road by the company and will occur only if game wardens fail to enforce regulations regarding hunting and fishing in the area. The company is not responsible for poaching along the road by anyone other than its own employees while they are on duty or in transport to and from the site.


6.9 Aquatic Biology Impacts Surface water quality impacts are expected to be negligible (see Section 6.5). The major impacts have already occurred during exploration and construction of the man camp. Impacts included increased concentrations of nitrogen and phosphates in the streams due to extensive blasting. Additionally, high suspended solid concentrations will be observed for a short time period during construction due to erosion and surface run-off. During operations suspended solids should be reduced due to erosion control measures, reclamation, and proper environmental management.

6.10 Aesthetics/Visual Resources The Kupol area is located more than 90 kilometers from the nearest settlement. The entire license area is isolated and accessible primarily by the winter road. Impacts have been isolated to the valleys of Kaiemraveem and Starichnaya. These impacts include visual disturbance to the area as well as a reduction in aesthetics due to the man made facilities. During operations, Bema will minimize additional aesthetic and visual impacts by minimizing new disturbance to the areas around the mine site. Additionally, during operations, Bema will aggressively pursue reclamation of previously impacted areas to reduce aesthetic and visual resources impacts. The ultimate goal is to return the environment to conditions prior to exploration. The closure plan will return most of the land to native grades and vegetation. However, the pit area will have a small lake and the walls will not be regarded to match existing topography.

6.11 Noise Impacts In conjunction with development of the deposit, the primary stationary and mobile sources of noise pollution include the following technological processes and mechanisms:

• Drilling of blasting holes (using hand-held drill); • Ore extraction (by means of blasting); • Transport of crushed ore (loading-delivery machinery); • Ventilation of underground mine workings (adit ventilators); • Transport and dumping of ore (conveyor belts, etc.); • Primary crushing of ore (cone crusher); • Grinding of ore (ball crushers and autogenous grinders); • Electricity generation (diesel generators); and, • Transport of supplies and personnel (automobiles).

The listed sources of technological noise result in various impacts on the environment depending upon their location (at the surface or in underground mine workings), sound-


insulating properties of enclosure structures and construction materials, levels of noise intensity and the type of sound (broad-band or impulsive), and duration of impact (continuous or periodic). Approximate levels of sound intensity in octave bands of geometric mean frequencies are given in Table 31.

Table 31. Noise Levels from various sources of Noise pollution Level of Sound Intensity dBA

Frequency hz Type of Equipment

63 125 250 500 1000 2000 3000 4000 8000 Mining Extraction Equipment and Mining Transport Equipment

Hand-held drills 99 106 107 107 100 108 109 108 114 Telescope drills 104 108 113 112 107 110 106 103 115 Drilling machinery 104 106 107 108 107 108 107 105 115 Loading machinery 89 93 93 97 96 97 94 91 106 Dumping machinery 99 102 101 98 99 98 96 95 105 Localized ventilation equipment 93 96 98 106 98 95 86 78 103 Primary ventilation equipment 106 113 117 116 109 100 92 82 116 Automobiles 100 98 93 87 85 77 68 59 93

Diesel-Electrical Equipment Diesel generator (Caterpillar) 119 129 119 110 104 99 97 93 114 Diesel exhaust 103 104 111 109 107 107 106 104 114

Processing Equipment Conveyor belt 105 106 107 99 96 92 89 85 <=4 hBall crusher 101 103 104 107 110 109 104 95 <=4 hBall crusher 99 115 117 123 123 121 117 107 <=4 hCone crusher 100 103 102 97 90 88 85 79

Practically all of the given noise sources capable of having an appreciable impact on the environment are located either in the underground workings, in the pit, or at the surface in the vicinity of the processing mill or the surface mining complex. The only negative impacts will be to wildlife in the area. Expected impacts from noise sources will be within World Bank Group guidelines. If a source is found to be in non-compliance with the guidelines, appropriate hearing protection measures will be implemented for affected employees. Noise will not affect any neighboring communities.

6.12 Socioeconomic Impacts This section evaluates potential social and economic impacts of the proposed Kupol Gold project. Evaluation of the impacts associated with the proposed project must consider the current social and economic environment of the local area. The project-related impacts, both temporary and permanent, must also be related to changes in the overall economic picture of the area, including continued mining exploration, potential expansion, development, and lay-offs. Cumulative impacts may compound or offset one another, and these impacts may vary


through different phases of development. Future changes in employment and phasing of other projects may result in changes to the impacts presented. 6.12.1 Population and Demographics Populations of towns and villages in the Bilibinskii, Chaunskii, and Anadyrskii Regions have been steadily declining since 1988, particularly in the largest towns (Bilibino, Pevek, and Anadyr) over the course of the past four years. By reducing unemployment and promoting growth in the mining, construction, and the supporting services sectors, the Kupol project would be a factor in curbing the steady out-migration trend. In addition to providing job openings in the mining and construction industries, the project would also encourage growth of supporting services. These increases include transportation, shipping, communication, and retail. 6.12.2 Housing The Kupol project is located far from all population centers. Workers will all have to be transported in from supporting communities (Bilibino, Magadan, and Anadyr). The greater part of the work force is expected to be made up of people already living in the Magadan Oblast and the Chukotka Okrug. As noted in Section 6.12.1 (Population and Demographics), the area has suffered a severe decline in population. It is anticipated that any demand for additional housing for workers moving in from other areas should be easily satisfied. 6.12.3 Economic Impact The project will generate more than US $150 million in taxes (after VAT rebate) that will be split among federal, regional and local levels over the course of 7 years. Additionally, based on the license agreement between Bema Gold Corporation and the government of Chukotka, Bema has paid in excess of US $30 million directly to the region. Within 90 days of a positive feasibility study, Bema must also pay US $5.00 per ounce for 75% of the gold identified in the proven and probable reserves (approximately 29 million). In addition to the positive social aspects of the projects mentioned above, CMGC is establishing a Charitable Foundation that will support developments in health care, training and education, indigenous peoples’ traditional activities, and sustainable small to medium (SME) enterprises within the ChAO. This fund will become operational after the 1st full year of operations (2009) and will be funded by CMGC in the amount of US $3,000,000 over the life of the project. This includes US $1,000,000 invested in the foundation in the 1st quarter 2009 and US $250,000 per year beginning in 2010 and lasting 8 years. Additionally, CMGC will solicit other organizations to join the foundation and contribute to the development of the region.


6.12.4 Indigenous Population There are no direct adverse impacts to indigenous peoples. Extensive public consultation will be conducted throughout the life of the project. During initial PCDP, the IP’s in the closest village (Illirny) have asked CMGC to consider training and employment programs for IP’s wishing to work at the Kupol mine site. In particular, they have requested that IP women be considered for training and employment. Overall, reindeer pasture that will be lost to use by indigenous people is very small when compared to the available reindeer pasture as determined by the Land Use Committee in Omsukchan. Additionally, the Anadyr/Bilibino/Chaunsk Land Use Committees have determined that the land immediately around the Kupol project is ill-suited for reindeer pasture with low productivity (2.5 reindeer*day/hectare). In any case, once the final land allotment is issued for the Kupol mine, Bema will be required to pay the Anadyr Committee on the Environment (Oblkomekologiya) for reindeer pasture that has been lost due to development of this project.


7.0 Bibliography

1. AMEC, Site Water Management Plan: Kupol Project – Chukotka Autonomous Okrug, Far East Russia, Vancouver, BC, Canada, 2005

2. AMEC, Surface Water Quality Modeling Report: Kupol Project – Chukotka

Autonomous Okrug, Far East Russia, Vancouver, BC, Canada, 2005

3. AMEC, Tailings Management Report: Kupol Project – Chukotka Autonomous Okrug, Far East Russia, Vancouver, BC, Canada, 2005

4. AMEC, Water Management Plan: Kupol Project – Chukotka Autonomous Okrug, Far

East Russia, Vancouver, BC, Canada, 2005

5. AMEC, Wasterock Storage Facility: Kupol Project – Chukotka Autonomous Okrug, Far East Russia, Vancouver, BC, Canada, 2005

6. Atkins, Steve, Geochemical Characterization of Waste Rock and Ore Materials at the

Kupol Project, Russian Far East, Golden, Colorado, USA, 2005

7. Bema Gold Corporation, Preliminary Economic Assessment: Kupol Gold Project, Far East Russia, Bema Gold Corporation, Vancouver, BC, Canada, 2004

8. CIA - The World Factbook – Russia, 6/14/05,

http://www.odci.gov/cia/publications/factbook/geos/rs.html, 6/17/05 9. Committee for Statistics, RF, Социально-экономическое Положение Анадырского

Района в Январе-Остябре 2003 Года [Socioeconomical Situation of the Anadyr Region in January – October, 2003], Anadyr, RF, 2003

10. Flint, V.E., A Field Guide to the Birds of Russia and Adjacent Territories, Princeton

University Press, Princeton, New Jersey, 1984 11. Gorbatcheva, Valentina, The Peoples of the Great North, Art and Civilization of

Siberia, Parkstone Press LTD, Singapore, 2000 12. Goskomstat, RF, Краткий статистический сборник, Анадырский Район, 1996-

2002 гг. [A Short Statistical Compendium, Anadyr Region, 1996-2002], Andyr, RF, 2003

13. Goskomstat, RF, Краткий статистический сборник, Билибинский Район, 1996-

2002 гг. [A Short Statistical Compendium, Bilibino Region, 1996-2002], Andyr, RF, 2003


http://www.odci.gov/cia/publications/factbook/geos/rs.html

14. Goskomstat, RF, Паспорт Социально-экономического положения г.Анадырь, 1997-2001 [Passport for the Socioeconomical Situation for the Town of Anadyr, 1997- 2001], Anadyr, RF, 2002

15. Goskomstat, RF, Паспорт Социально-экономического положения г. Билибино,

1997-2001 [Passport for the Socioeconomical Situation for the Town of Bilibino, 1997- 2001], Anadyr, RF, 2001

16. Newell, Josh, The Russian Far East, A Reference Guide for Conservation and

Development, Daniel and Daniel Publishers, Inc., McKinleyville, California, 2004

17. Russia Administrative and Territorial Divisions, http:www.photius.com/countries/russia.htm, 2/14/2005

18. Russia Overview of the Economy, http://www.worldwide-

tax.com/russia/rus_over_economy.asp, 6/14/05

19. Russia’s Chukotka Autonomous Region Overview, 1998, http://www.users.qwest.net/~kryopak/ChukotkaHomePage.htm, 2/8/2005

20. Tichosky, John, EDITED Social Economic Development of Chukotka, personal

communication, 2005

21. Tichosky, John, Summary Information about the Social-Economic Development of the Chukotka Autonomous District (Okrug), 2004

22. USEPA, AP-42, Emission Factors, Chapter 11 – Emission Factors for Metallic

Minerals Processing, fifth edition, reformatted 1/95

23. VNII-1, Baseline Studies of the Kupol Deposit (in Russian), 2003-2004, Personal Correspondence

24. VNIRO, Maximum Allowable Concentrations for Rivers and Waterways, Fishery

Standards, Moscow, RF, 1999

25. World Bank, World Bank Environment, Health and Safety Guidelines: Open Pit Mines, The World Bank Group, Washington, DC, USA, 1999

26. World Bank, World Bank Environment, Health, and Safety Guidelines: Underground

Mines, The World Bank Group, Washington, DC, USA, 1999

27. World Bank, Pollution Prevention and Abatement Handbook, 1998, Towards Cleaner Production, The World Bank Group, Washington, DC, USA, 1999


http://www.worldwide-tax.com/russia/rus_over_economy.asp

http://www.worldwide-tax.com/russia/rus_over_economy.asp

http://www.users.qwest.net/%7Ekryopak/ChukotkaHomePage.htm

Appendix A Baseline Data

Kupol Environmental Impact Assessment

Appendix B Geochemical Characterization of Waste Rock and

Ore Materials


Geochemical characterization of waste rock and ore materials at the Kupol Project, Russian Far East

June 2005

Prepared for:

Bema Gold Corporation Suite 3100, Three Bentall Centre

595 Burard Street Vancouver, BC, V7X 1J1

Canada

Prepared by:

Steven Atkin, Ph.D. 111 Red Rocks Vista Drive Morrison, Colorado 80465

U.S.A.

Bema –geochemical characterization

pre-reas-II_1June05_final S. Atkin

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Table of Contents Summary (ii)-(v) 1.0 Introduction 1 1.1 Background 1 1.2 Objectives 1 1.3 Approach 2 2.0 Project setting 4 2.1 Location 4 2.2 Topography 4 2.3 Climate 4 2.4 Hydrology 5 2.5 Geology 6 2.6 Environmental chemistry 11 3.0 Mine plan 15 3.1 ARD potential 15 3.2 Model predicted tonnes by ARD potential 16 4.0 Test program 20 4.1 Sample selection 20 4.2 Test program 21 4.3 Comparison with industry standards 25 5.0 Test results 27 5.1 Acid Base Accounting (ABA) 27 5.2 Net Acid Generating (“NAG”) test 33 5.3 Strong acid digestion and multi-element analysis by ICP 35 5.4 Synthetic Precipitation Leaching Procedure (SPLP, EPA 1312) 40 5.5 Humidity cell test 45 5.6 Mineralogy 54 5.7 Grain size 57 5.8 QA/QC 57 6.0 Conclusions 59 6.1 Geochemical test program 59 6.2 Test results 60 6.3 Recommendations for future test work 62 6.4 Mitigation options 63 Bibliography 66 Appendix A – Tables and figures from the text Appendix B – Raw sample and geochemical data Appendix C – Test methods Appendix D – Carbonate and pyrite code maps from the geologic model Appendix E – Reviewer comments



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Summary In August 2004, Bema Gold Corporation (BGC) commissioned a geochemical investigation to characterize the potential for acid rock drainage (ARD) and metals leaching (ML) from waste rock at their Kupol Project, located in Chukotka, Far East Russia. The objectives of the investigation were: a) to identify the geochemically different types of ore and waste rock in the deposit; b) to characterize, using an industry standard test program, the geochemistry of each type of material including all of the different lithology, alteration, and mineralization assemblages of the waste rock; and c) to identify issues for potential ARD and ML and control options for the facilities at the planned mining operation. Mineralization at the Kupol Project is a quartz-adularia type, low-sulfidation gold system localized along a high angle vein complex and hosted in andesite flows and pyroclastic rocks. The vein records repeated episodes of dilation and mineralization and is composed of up to five different banded quartz and quartz breccia assemblages. Au and Ag mineralization is associated with the earliest dilational event and includes electrum, Ag-sulfides and Ag-sulfosalts. The andesites, which are Cretaceous in age, include a basal unit dominated by flows (85 to 90%) interbedded with discrete andesite crystal tuffs and an upper unit of tuffs (50%) interbedded with flows. The andesite package is overlain by younger (Tertiary age) rhyolite and basalt flows. Post-mineralization dikes, believed the feeders for the Tertiary rhyolite and basalt flows, parallel and, particularly in the south, crosscut the vein in a narrow 100 to 200 m wide corridor. The andesites along the vein have been altered and include: 1) a propylitic assemblage characterized by the presence of carbonate, chlorite, and, epidote; 2) a carbonate dominant assemblage including calcium carbonate, iron carbonate, and dolomite; 3) a texturally destructive argillic alteration characterized by strong clay alteration; 4) a sericite assemblage; 5) an adularia assemblage and; 6) silicification with quartz dominant veining and pervasive silica flooding of the adjacent country rock. The alteration assemblages are often superimposed upon one another because of the multiple and overlapping mineralization events. A total of 276 representative samples of waste rock representing each of the lithology, alteration, and mineralization assemblages and 95 composite samples of ore (from the metallurgical test program) were collected from splits of core from a) exploration and geotechnical drill holes of the 2003 and 2004 BGC drilling campaign and b) historic Russian exploration drill holes. These samples were analyzed in the geochemical investigation at Kupol following a staged sequence of: a) static testing designed to characterize the range and variability in the geochemistry of the different material types and; b) kinetic tests designed to characterize the dynamic performance and reactivity of the waste rock at Kupol. All of the samples (276 waste rock and 95 ore samples) underwent acid base accounting tests (ABA) and strong acid digestion with metals analysis (ICP) to characterize the sulfur speciation, acid generating potential, acid neutralizing potential, and gross whole rock chemistry. One in three samples (78 plus 1 ore sample) from the 2003 drilling, covering the range in ABA and metals concentrations, were selected for short term leaching tests (Synthetic Precipitation



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Leaching Procedure, EPA-1312) to identify readily solubilized constituents. One in ten samples (23 plus 1 ore sample) from the 2003 drilling, covering the range of ABA, metal concentrations, and metal solubility, were selected for Net Acid Generating (NAG) tests. A total of 32 samples, including 24 from the 2003 drilling (1 in ten) and 8 from the 2004 drilling were selected for humidity cell kinetic tests. Humidity cell tests are designed to quantify, under laboratory conditions, ARD and ML reaction rates and to confirm ARD/ML behavior of sample classes estimated by the static tests. Additional test work included X-ray diffraction analyses (XRD) and grain size analyses of the humidity cell samples. The results of the geochemical test work for the waste rock lithology, alteration, and mineralization assemblages have been grouped into three ARD codes for use in the resource model: non-acid generating (NAG); acid generating (AG); and potentially acid generating (PAG). The ARD codes are based on simplified logging codes for pyrite mineralization and carbonate alteration from the detailed geologic drill hole logs that define four categories: high pyrite and low carbonate – AG; low pyrite and high carbonate – NAG; and high pyrite and high carbonate or low pyrite and low carbonate – PAG. Of the 276 waste rock samples, 51 (18%) are AG, 71 (26%) are NAG, and 154 (56%) are PAG. Results of the ABA tests have shown that all of the categories of waste rock at Kupol are sulfur-bearing and the dominant sulfur species in the waste rock is sulfide sulfur (Spy). Total sulfur (Stot) and Spy range from below the detection limit to 6 weight percent. The Stot and Spy concentrations decrease from AG through NAG waste rock. At the same time, all of the waste rock categories in the resource model contain some degree of acid neutralizing potential in the form of either carbonate veins or pervasive carbonate alteration. The neutralizing potential (NP) ranges from values less then 0 (negative NP indicates the presence of natural acidity in the rock) to almost 200 kg CaCO3/tonne. The NP increases from the AG to the NAG waste rock. Using the neutralizing potential:acid generating potential (NP:AGP) ratio and the standard industry criteria for ARD potential: two thirds of the AG waste rock is acid generating; over two thirds of NAG material is acid neutralizing; and half of the PAG material is acid neutralizing. The proportions of acid generating, acid neutralizing, and uncertain categories for each of the ARD codes are shown in Table 1 below. The “uncertain” category is generally considered potentially acid generating awaiting the results of kinetic testing to confirm the ARD potential. Table 1 Proportions of acid generating, uncertain, and acid neutralizing waste rock in the ARD domains from waste rock samples of the ARD database NP:AGP (1) >3 <3 and >1 <1 ARD domain Count Acid neutralizing Uncertain Acid generating AG 51 21.6% 9.8% 68.6% PAG 154 53.9% 11.7% 34.4% NAG 71 71.8% 11.3% 16.9% (1) after SRK, 1992



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Table 2 shows the model predicted tonnes of the waste rock to be mined from the open pit and underground using 1) the tonnes of each of the ARD domains from the resource model and 2) the proportions of acid neutralizing, uncertain, and acid generating waste within each of the domains shown in the Table 1 above. The NP:AGP ratio of the waste rock mined will be measured in the field by Leco furnace sulfur and carbonate analyses of blast hole cuttings as part of the mine operations program.

Table 2 Model predicted tonnes of waste rock mined from the open pit and underground and not used for back fill. NP:AGP 2006 2007 2008 2009 2010 2011 2012 2013 2014 Total

Non-acid generating >3 1231000 2058000 1874000 1549000 1243000 0 0 0 0 7955000

Potentially acid generating <3 and >1 276312 487774 477794 389000 303000 0 0 0 0 1933880

Acid generating waste <1 886321 1646496 1821627 1589517 862544 0 0 0 0 6806504

Unclassified Tonnes 43147 60607 22987 0 0 0 0 0 0 126741

Total Waste Tonnes 2436780 4252877 4196407 3527517 2408544 0 0 0 0 16822125

Similar estimates of the acid generating potential of the model NAG, PAG, and AG domains in waste rock types were found using the NAGpH and “NAG” values from the Net Acid Generation test work. Using the median values of NAGpH and “NAG”, AG is acid generating (NAGpH less than 4.5, and “NAG” greater than 5 kg H2SO4/tonne) and NAG and PAG are non-acid generating (NAGpH greater than 4.5, and “NAG” generally less than 5 kg H2SO4/tonne). The ore samples have lower Stot (below detection to 4 weight percent) than portions of the waste rock and roughly equal proportions of sulfate sulfur (SSO4) and Spy. NP in the ore samples ranges from below zero to 115 kg CaCO3/tonne and the samples fall within the field of acid generating material (NP:AGP ratio less than 1). The ICP test results provide an indication of metals that are present in anomalously high concentrations compared to similar, unmineralized rocks. Comparison of the results of the strong acid digestion and ICP analysis of the waste rock samples indicates that silver, arsenic, gold, bismuth, mercury, antimony, selenium, and thallium occur in AG, PAG, and NAG samples at concentrations above normal basalts. Silver, arsenic, gold, bismuth, cadmium, copper, mercury, molybdenum, lead, antimony, selenium, thallium, and zinc are anomalously high in the ore samples. The SPLP test provides an indication of elemental solubility under standard laboratory conditions in the waste rock using a 24 hour leaching procedure. The SPLP leachate ranges in pH from 3 to almost 10 and shows in increase an pH from AG samples (average 4 and median of 6) to NAG samples (average 5 and median 9). Alkalinity in the SPLP leachate also increases from 10 mg CaCO3/L in the leachate from AG samples to over 20 mg CaO3/L in the leachate from the NAG samples. The chemistry of all of the SPLP leachate from AG, PAG, and NAG samples is dominated by sulfate, which ranges in concentration from 1 to 2000 mg/L. Generally, the metal concentrations in the SPLP leachate are low. Using the maximum concentrations of soluble metals in the SPLP,



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leachate: aluminum, arsenic, beryllium, calcium, cobalt, copper, iron, magnesium, manganese, nickel, silicon, strontium, vanadium, and zinc all appear to be mobile in slightly acid solution. The humidity cell tests provide the key parameters to understanding the geochemical behavior of the waste rock at Kupol, namely the long term (20 weeks) leaching characteristics of the different material types. Results indicate that the pH of the leachate derived from: a) AG waste rock samples is typically low, below pH 4; b) NAG waste rock samples is mixed but dominantly alkaline, above pH 7 and; c) PAG samples is variable but generally above pH 7. Humidity cell leachate from AG samples typically has high sulfate, zero alkalinity, and high metal concentrations, although the sulfate and metals concentrations typically decline with time. The humidity cell leachate from NAG samples typically has low sulfate, moderate alkalinity, and moderate metal concentrations. Humidity cell leachate from the PAG samples is variable, with low alkalinity, generally low sulfate, and generally low metals concentrations. Recommendations for mine design, operations, and closure include:

• Design of the mine facilities holding waste rock in such a way as to minimize runon and, where necessary, maximize runoff.

• Utilize “Thin Lift” technology in facility construction to maximize freezing and compartmentalization of the waste.

• Monitor waste rock material types (AG, PAG, and NAG) during operations to identify potentially ARD (and ML) generating waste for special handling. Results of the ABA and “NAG” tests show that total sulfur and carbonate analyses or NAGpH tests provide possible field and laboratory tests that can quickly identify ARD generating waste during mining.

• Use selective placement of acid-generating waste rock to: a) minimize sulfide oxidation, b) minimize contact of infiltration and seepage from the acid generating waste, and; c) maximize contact of facility runoff or discharge with acid neutralizing waste rock.

• Incorporate a monitoring program to track facility discharge water chemistry in order to identify potential sources of ARD in the facility.



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1.0 Introduction 1.1 Background The Kupol Project is a world-class epithermal vein gold and silver deposit located in Chukotka, Far East Russia. Current mine plans assume the resource will be initially developed by open pit mining methods with the waste rock delivered to the tailing facility. The open pit development is scheduled to commence in 2006 and underground development is scheduled to begin sometime around 2007 (Wardrop, 2005). The mill is planned to being processing ore in 2007. The open pit is scheduled to provide the first several years of ore production to the mill, both the open pit and underground will supply ore feed to the mill in the mid years, and the underground will provide ore to the mill in the later stages of operations. Current development plans call for mining to provide mill feed for an average plant throughput rate of 3000 tonnes per day declining late in the mine life. 1.2 Objectives The waste rock mined during operations and delivered to the tailing facility or exposed in the mine void walls at Kupol may be potential sources of acid rock drainage (ARD) and metals leaching (ML). A capacity to predict the ARD and ML potential of the waste rock at Kupol is essential to facilitate proactive mitigation planning for the periods encompassing the mine planning and development phase, the operational phase, and closure phase of the project life. In August 2003, Bema Gold Corporation (BGC) commissioned a geochemical investigation by Dr. Steven Atkin to design and implement a sampling and analysis plan (SAP) to characterize the geochemistry and ARD/ML potential of the waste rock at the Kupol Project. The objectives of the test program were multifaceted and included:

• Identification of the ore and waste rock types that will potentially be delivered to the tailing facility, and exposed in the pit walls as a result of mining at Kupol. Although samples of ore have been included in the investigation, it is anticipated that the spent ore will be contained in the zero discharge tailing facility and will have minimum impact on the environment down gradient of the facility.

• Characterization of the geochemistry of each of the ore and waste rock types. The

test program was designed to characterize for each lithology, alteration, and mineralization assemblage the bulk composition, the ARD potential, and the ML rates.

In addition, the geochemical investigation was to propose criteria to identify:

• potential sources of ARD/ML within the resource model, • potential sources of ARD/ML for special handling during operations,



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• suggestions for placement options and tailing facility design alternatives during tailing facility construction, and

• possible tailing facility closure design alternatives following cessation of mining activities at the site.

Staff geologists with the Exploration Department of Bema Gold Corporation provided insight into the geology of the Kupol gold deposit and support during sample selection. The mitigation alternative review was based upon the “Report on AMEX, 2005, Report of Waste Rock Management Feasibility Design, Kupol Project, Chukotka Autonomous Okrug, Far East Russia, AMEC Earth & Environmental May 2005. Waste Rock Management Feasibility Design” (AMEC, 2005) and the design specifications for the tailing facility in the “Tailings Facility Feasibility Design Report” (AMEC, 2005) prepared by AMEC Earth and Environmental (AMEC), Vancouver, BC, Canada. The tonnages of waste rock materials to be delivered to the tailing facility during mine life were based upon the mine plans prepared by Wardrop Engineering, Inc. (Wardrop), Vancouver, BC, Canada. In addition, BGC commissioned an external third party technical review by Mr. Mark Logdson, President, Geochimica Inc., Aptos, California, USA, of the SAP for the Kupol Project and the results of the geochemical test work. In addition, Mr. Logsdon participated in discussions of mitigation alternatives based upon his broad understanding of the geochemistry of ARD/ML and his worldwide experience in mining operations and closure design, particularly in high arctic environments. Mr. Logdon’s review comments are included in Appendix E. 1.3 Approach The methodology employed to characterize the ARD/ML potential for the Kupol Project used best international industry practices and included:

1) Review of the site geology to identify the lithology, alteration, and mineralization assemblages within the waste rock present at the site and the relative abundance of each of the lithology, alteration, and mineralization assemblages.

2) Sample selection, with the cooperation of the site geologists, based upon the expected tonnage of waste rock mined, the relative proportions of each lithology, alteration, and mineralization assemblage in the drill hole database, and selection of samples to include all geographic and depth ranges within the deposit.

3) Testing to include a broad range of industry standard static and kinetic test protocols designed to characterize the ARD and ML properties of each of the waste rock lithology, alteration, and mineralization assemblages. The geochemical test program included a range of ore samples to characterize the ARD and ML properties of ore prior to beneficiation.

4) Compilation of the geochemical test results and comparison with industry standard criteria to identify the ARD/ML potential of the waste rock at the Kupol site.



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5) Identification of possible ARD/ML mitigation options that could be applied during mine planning, operations and at closure to reduce the potential for ARD/ML within the tailing facility.



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2.0 Project setting 2.1 Location The Kupol Project is located in the high Arctic of the Chukotka Autonomous Okrug, of the Russian Far East, Russian Federation. The site is roughly 430 air km west of Anadyr, the capitol of Chukotka. The project is located above the Artic Circle at 66.7 degrees north latitude at an elevation of roughly 650 m above mean sea level. 2.2 Topography The area consists of tundra and low mountains and is characterized by prevailing low, rounded hills with occasional flat, midland areas 100 to 200 meters above the intervening riverbeds (BGC, draft environmental report, 2004). The tops are divided by wide (200 to 300 m) but shallow (20 to 30 m) saddles. The watershed slopes are not likely to exceed 15-20 degrees; their foothills and riverbeds are overlain with young talus-solifluction and solifluction sediments. The absolute elevations of the hills surrounding the property do not exceed 700-1050 m (1034.2 m for Malakhai Mountain and 815.0 m for Kupol Mountain). 2.3 Climate The climate of the region around the Kupol site belongs to the continental climatic region of the subarctic climate belt with extremely severe weather consisting of long and cold winters (8-8.5 months), overcast weather, and short summer periods (2.5 months) (BGC, draft environmental report, 2004). The average annual air temperature at the Kupol site, with only minor variances, is near -13 °C. The absolute minimum average monthly temperatures occur in January and February (-58 °C). During the warmest months (June, July, and August), the average air temperatures are 8.3; 11.3; and 10 °C; respectively. Using the closest weather stations “Ilerynei” and “Ostrovnoye”, the total annual precipitation is less than 277 mm annually. Winter snow fall is approximately 116 mm or roughly 46% of the annual precipitation. The relative humidity at the Kupol site reaches 71%. Surface water evaporation is 280 mm during summer months resulting in a slightly negative net environmental water balance on an average-annual basis. The average annual wind speed is 2.1-2.6 m/s with a maximum wind speed of 20 m/s. The maximum wind speed recorded at the closest weather stations is 24 m/s. The maximum wind speed recorded by the weather station installed at site (2003) is 30 m/s. Permafrost is distributed throughout the Kupol Property area. Depending on geomorphology, thickness of permafrost layer goes down from surface to 200 to 320 meters and reaches its maximum deep under riverbeds (based on literature reviews).



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Thickness of seasonal melting varies from 0.02-1.5 meters in river valley terrains to 12.4 meters on watersheds. Temperature gradient within permafrost rock is 0.023˚C/m. Seasonal thaw in the deposit area begins in early June and runs through September. In the bald peak areas, the thaw depth largely depends on slope exposure and is 0.8-2.0 m; in talus patches, the thaw depth is limited by thickness of talus. In Arctic tundra patches, the thaw depth depends on mechanical structure of sediments and varies between 0.8 and 1.8 m. The seasonal thaw is 0.3-0.4 m in typical hummocky and hillocky (mound) tundra areas and 2.5 m in local well-drained areas. 2.4 Hydrology The territory around the Kupol site belongs to the watersheds of the Srednyi Kaiemraveem and Malyi Anyui Rivers (BGC, draft environmental report, 2004). The Srednyi Kaiemraveem drains into the Mechkereva River. The Mechkereva River is a right tributary of the Anadyr River. The Anadyr River is one of the largest waterways on the Chukotka peninsula, and the waters flow from west to east through the middle part of Chukotka and drains into the Bering Sea. The Malyi Anyui River is a right tributary of the Kolyma River. The Kolyma is also one of the largest waterways of the far northeastern part of Russia and flows from south to north to the Eastern Siberian Sea in the Arctic Ocean. The site of the ore body and the proposed operations will be located within the northern part of water catchment basin of the Srednyi Kaiemraveem. The three small creeks in the proposed mine area that feed the Kaiemraveem run only during summer months and are a product of precipitation runoff and snowmelt. Pervaya Creek is located north of the deposit. It has a watershed catchment of 4.0 km2 and a straight creek length of 2.7 km (with an overall length of 5.2 km). Vtoraya Creek is located within the catchment area with a watershed catchment area of 1.7 km2and a straight length of 1.3 km (with an overall length of 2.0 km). Tretaya Creek has a watershed catchment area of 5.0 km2 and a straight length of 3.9 km (with an overall length of 9.6 km). Flow measurements at the site have not been taken on any regular basis and are presented here based on an analogous river with similar characteristics located within the region. During a runoff event, it is likely (BGC, draft environmental report, 2004) that the rivers will have the following volumetric flow rates: Kaiemraveem River: 58.8 m3/s; Pervyi Creek: 13.9 m3/s; Vtoraya Creek: 11.0 m3/s; and, Tretyi Creek: 12.4 m3/s.



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2.5 Geology 2.5.1 Regional geology The Kupol deposit is located in Cretaceous age volcanics of the Okhotsk-Chukotka volcanogenic belt (OCVB) (Rhys, 2004). The OCVB unconformably overlies Jurassic and older age folded rocks and is in turn overlain by basalts of probable Tertiary age. The mineralization at Kupol is hosted in a sequence of andesitic volcanics within a probable caldera complex exposed to the west of the Kaiemraveem River valley. East of the deposit and east of the Kaiemraveem River valley, the Cretaceous age igneous rocks consist of a sequence of shallow dipping rhyolite, dacite, and basalt flows and tuffs and probable subvolcanic intrusives (Rhys, 2004). Since the rocks at Kupol do not appear to extend east of the Kaiemraveem River valley, the correlation between the two Cretaceous sequences is uncertain. Rhys (2004) suggested that the lack of correlation across the valley may be the results of:

• an angular unconformity in which the rocks east of the Kaiemraveem River valley are younger than and overly the Kupol sequence or

• a major fault zone along the Kaiemraveem River valley has juxtaposed the two sequences, in which case their relative ages are unknown.

The deposit lies within the Kaiemraveem River lineament which consists of a sequence of north south trending lineaments, components of a regional fault system (Rhys, 2004). It has been suggested that owing to the lack of correlation of Cretaceous rock across the Kaiemraveem River, the lineament represents a “significant” fault that may have influenced or controlled the faults that localized the vein mineralization at Kupol (Rhys, 2004, after Garagan and MacKinnon, 2003). Rhys suggested, however, that additional mapping may be required to better define the timing, displacement sense and magnitude of regional faults for a complete understanding of nature of the Kaiemraveem River lineament. 2.5.2 District geology Mineralization at the Kupol Project is a quartz-adularia type, low sulfidation gold system localized along a high angle vein complex (Vartanyan, et al., 2001; personal communication, 2003). Mineralization occurs along a strike length of over 2 km and is open at both ends as well as at depth. Plate 2.1 is a geologic map of the project site showing the drill hole and sample locations. Plates 2.2a and 2.2b are east west cross sections through the deposit on the 90972 and 91090 northings, respectively, showing geology, drill hole traces, and sample locations.



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Geologic evidence indicates that the hydrothermal event that brought gold to the fracture systems at Kupol was long-lived and included multiple pulses resulting in:

• repeated dilation and vein mineral precipitation and • superimposed hydrothermal alteration events within the wall rock adjacent to the

vein.

Geology Rhys (2004) summarized the work of Garagan and MacKinnon (2003) who defined the principal lithologies at Kupol. The sequence currently envisioned at Kupol consists of a basal porphyritic andesite that hosts most of the mineralization, capped by a sequence of andesite tuffs. The andesites are in turn overlain by younger rocks of basalt and rhyolite composition. Basal porphyritic unit The basal porphyritic unit consists of over 700 m of massive, plagioclase porphyritic and phyric andesite flows and probable crystal tuffs (Rhys, 2004). The tuff unit comprises about 10 to 15 % of the main andesitic sequence but includes discrete ash/lapilli tuff marker beds which have been used to determine stratigraphic position within the deposit and the location and displacement of faults. The markers indicate the stratigraphy at Kupol forms a shallow east dipping domal structure centered on the Central Zone with the beds dipping to the northeast in the North and Northern Extension and to the southeast in the Southern zone. Rhys (2004) noted variations in vein thickness and grade where the markers intersect the vein suggesting the markers may have influenced permeability, dilatancy and/or chemical reactivity in the wall rock during the mineralization event. The best defined tuff in the hanging wall of the deposit is the Main Marker, a 5-30 m thick ash and lapilli tuff. A second marker occurs 200 m below the Main Marker. The Main Marker occurs at increasing depths to the north ranging from 30 m in the northern South Zone and southern Big Bend zone to 40 to 80 m in the area of the central pit. The Main Marker is absent and presumably eroded to the south in the Southern Extension and southern South Zone. Only a single marker, occurring at a depth of 470 m, is traceable over any length in the footwall of the deposit. Upper tuff unit The upper tuff unit is exposed in the northern limits of the deposit in the Northern Extension, in the down thrown block north of two significant cross faults, the Premola and Northern faults (Rhys, 2004). The unit is absent and believed eroded to the south, but is assumed to occur roughly 80-100 m stratigraphically above the Main Marker. The upper tuff is composed of several thick tuff units up to 100 m thick interbedded with plagioclase porphyry andesite. The unit is approximately 150 m thick and is composted



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of 50% tuff. Rhys postulated that the upper tuff acted as an aquitard, reducing fluid flow along the vein resulting in overpressure and enhanced vein dilation. The upper tuff unit is unconformably overlain in the Northern Extension by a mafic unit composed of basalt and basaltic andesite. A unit composed of rhyolite flows and tuffs occurs in drill holes above the basalt and is believed to unconformably overly the mafic unit. The rhyolite units in the Northern Extension are thought to be correlative with a rhyolite flow dome complex in the South Zone and both are viewed as extrusive equivalents of post-mineralization rhyolite dikes encountered within the vein complex at depth. Dikes Rhyolite dikes occur within a 100 to 200 m wide corridor surrounding and cross cutting the Kupol vein. The dikes, which may be over 50m thick and make up between 20% and 50% of the rocks within the corridor are either a) oriented parallel to the vein system striking north to northeast and, dipping steeply to the east or subvertical or b) splay off the vein parallel set trending north to northeast with steep westerly dips. The dike swarm passes though the vein in the south and diverges from the vein in the Central and North Zones. The dikes a) do not appear to offset the vein but rather exploit syn-dike dilational features and b) are unmineralized where dikes truncate mineralization indicating the dikes post-date the mineralization. Post-mineralization mafic dikes consisting of amygdaloidal plagioclase porphyritic basalt have also been encountered in drill holes in the Big Bend area. Alteration The rocks adjacent to the vein and in the hanging wall above the vein have been hydrothermally altered. Hydrothermal alteration also occurs in the footwall rocks beneath the vein but the alteration envelope appears to be much thinner in the footwall than in the hanging wall. The hydrothermal alteration includes 1) a propylitic assemblage characterized by the presence of carbonate, chlorite, and epidote within the andesitic volcanic rocks, 2) a carbonate dominant assemblage including calcite, iron carbonate, and dolomite, 3) a texturally destructive argillic alteration characterized by strong clay alteration, 4) a sericite assemblage 5) an adularia assemblage and 6) silicification with quartz dominant veining and pervasive silica flooding of the adjacent country rock. A seventh alteration assemblage is characterized by abundant gypsum mineralization. Although a single alteration type may locally dominate the mineral assemblage, the alteration assemblages often over lap such that a clay dominant system may include carbonate veining and a carbonate assemblage may include minor clay alteration. Mineralization Vartanyan, et al. (2001) identified five mineral complexes and mineral associations within each complex in the ore veins from Kupol. The complexes and mineral associates are shown in paragenetic order in Table 2.1 based upon cross cutting relationships within



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the vein complex. Appendix A includes a list of the minerals identified by the authors and the chemical compositions of each of the minerals. The earliest complex, the pyrite-adularia-quartz assemblage, preceded the main gold-bearing hydrothermal event. The arsenopyrite-pyrite-adularia-quartz and gold-stefanite-pyrargyrite-adularia-quartz complexes are the primary gold-bearing assemblages. Emplacement of the antimonite-marcasite-quartz complex followed the main gold-bearing events. The acanthite-jarosite complex is a supergene weathering assemblage superimposed on the original hydrothermal mineral assemblages. Included in Table 2.1 are the temperatures of formation of several of the mineral associations within the gold-stefanite-pyrargyrite-adularia-quartz complex from fluid inclusion studies by Vartnyan, et al (2001). Based upon the results of the geothermometry and textural evidence, the authors suggest that the mineralization at the Kupol site formed at shallow depths (less then 200 bars pressure) and over temperatures of between 160 and 260º C. Early mineralization was associated with elevated concentrations of sulfur, arsenic, and antimony while later stages of the hydrothermal event were characterized with elevated concentrations of selenium, antimony and sulfur. Table 2.1 Mineral complexes and associations (after Vartanyan et al, 2001) Assemblage Dominant mineralogy Temperature Pyrite-adularia-quartz Pyrite-adularia-quartz Arsenopyrite-pyrite-adularia-quartz Arsenopyrite-pyrite Tennantite-pyrite Amethyst-quartz Gold-stefanite-pyrargyrite-adularia-quartz Gold-perceite-chalcopyrite 260-240º C Gold-freibergite-stephanite-pyrargyrite 220-200º C Gold-aquilarite-Se-pyrargyrite 180-160º C Antimonite-marcasite-quartz Pyrite-marcasite Bertierite-antimonite Gypsum-anhydrite-chlorite Acanthite-jarosite Acanthite-covellite Acanthite-jarosite Iron hydroxide Current mapping of the vein has identified five separate and sequential dilational events recorded by unique textural and mineralogical assemblages (Rhys, 2004) including:

1. Early crustiform quartz-adularia stage – consisting of crustiform and colloform banded quartz +/- yellow adularia + sulfosalts/Ag sulfides/marcasite with Au and Ag grades proportional to sulfosalt abundance. Mineralogically the stages includes: friebergite, stephanite, pyrargyrite, tetrahedrite, acanthite, electrum, marcasite, chalcopyrite, and sphalerite.

2. Quartz breccia stage – consisting of cream to yellow colored quartz +/- jarosite breccia containing crustiform banded vein (Stage 1) fragments. Au and Ag grades in this stage are believed to be derived from contained fragments of the crustiform stage (Stage 1).



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3. Grey-green matrix breccia – consisting of a younger quartz matrix breccia localized within the Big Bend Zone. This stage is believed to post-date Au and Ag mineralization as it contains no evidence of “new” sulfosalt mineralization.

4. Massive white quartz fill – consisting of a massive quartz overprinting of the quartz breccia (Stage 2) in the Big Bend Zone that takes the form of massive opaque white quartz in discrete bands up to 4 m thick through the vein.

5. Late amethyst phase – consisting of a volumetrically minor assemblage of lenses, irregular veins and vug fillings of purple amethyst.

Structure The vein system at Kupol occupies a pre- or syn-mineral structure termed the Kupol structure (Garagan and MacKinnon, 2003) that acted as the conduit to localize vein mineralization. If one assumes an east side down motion, Rhys (2004) estimated a displacement along the fault of over 100 m. High angle shear and breccia zones strike generally north south and dip steeply to the east parallel to the vein system. The faults show evidence of syn- as well as post-mineralization movement such as alteration within the fault system and mineralized clasts within the fault breccia. The most significant post-mineralization faults are the northwest trending Permola and Northern faults located between the North Zone and the Northern Extension. The Permola fault is exposed at the surface and consists of a south to southeast dipping, 0.5 to 1 m wide breccia zone. The Northern fault is not exposed but is inferred to strike parallel to the Permola fault and dip to the north. The faults also acted as conduits for percolation of groundwater through the mineralized host rocks and include evidence of weathering and oxidation within the fault zones. Soils and weathering Soil development at the Kupol site is limited to the valley bottoms. Cold climate weathering is characterized by mechanical freeze-thaw action, thermal stress and wind erosion. As a result, the upper slopes of the hills are typically composed of broken rubble and aprons of talus on the lower slopes. Where soils are developed in tundra, they are characterized by a brown organic top layer underlain by a gray to yellowish mineral layer. Chemical weathering of the surface outcrops in the mineralized zone has locally developed a thick clay horizon as exposed in many of the trenches and surface disturbances in the area below the camp. Weathering along the fracture system into the mineralized zones has resulted in development of secondary minerals including iron hydroxides and secondary sulfate minerals along facture surfaces and along some of the contacts at depth.



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2.6 Environmental geochemistry 2.6.1 Acid generation ARD and ML are natural byproducts of oxidation of sulfide minerals that are exposed to air and water. The reactions yield low pH or acidic water that has the potential of mobilizing heavy metals that are present in minerals in the host rock. The reactions can be both chemically and biologically driven. Although often associated with mining activities, ARD/ML is also found where sulfide-bearing rocks are exposed to the atmosphere naturally, at construction sites, or other areas disturbed by human activity. ARD/ML generally occurs under the following conditions (after Hutchison et al, 1992):

• The mined material contains sulfide minerals that react either chemically or biologically to produce acid at a rate faster than the acid can be neutralized by alkalinity present within the rock.

• The facility containing the rock material allows air and water to circulate through

the pile in sufficient quantities to support the chemical reactions.

• As a result of runon, rainfall falling on the pile, or snow melt, water is allowed to infiltrate through the pile to dissolve the products of the chemical and biological reactions and transport the acidic drainage into the environment.

Oxidation of pyrite (FeS2) and marcasite (FeS2) is the major source of acid generation in mined material. Reactions of other sulfide minerals and other forms of sulfur also release acidity and heavy metals to the environment. Table 2.2 lists some of the heavy metals released from oxidation of sulfide mineral reported at Kupol (refer to Appendix A). Table 2.2 Minerals reported at Kupol and their oxidation products Mineral Composition Aqueous end products of

oxidation reactions Pyrite FeS2 Fe3+, SO4

2-, H+ Marcasite FeS2 Fe3+, SO4

2-, H+ Arsenopyrite FeAsS Fe3+, AsO4

3-, SO42-, H+

Chalcopyrite CuFeS2 Cu2+, Fe3+, SO42-, H+

Tetrahedrite-tennenite Cu12(Sb,As)4S13 Cu2+, SbO3-, AsO43-, SO4

2-, H+ Galena PbS Pb2+, SO4

2-, H+ Jarosite KFe3(SO4)2(OH)6 K+, Fe3+, SO4

2-, H+



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The following reactions characterize the three stages of pyrite oxidation (after Hutchison et al, 1992):

(1) 2 FeS2 (pyrite) + 7 O2 + 2 H2O = 2 Fe2+ + 4 SO42- + 4 H+

(2) 4 Fe2+ + 10 H2O + O2 = 4 Fe(OH)3 + 8 H+

(3) 2 Fe2+ + O2 + 2 H+ = 2 Fe3+ + H2O

(4) FeS2 (pyrite) + 14 Fe3+ + 8 H2O = 15 Fe2+ + 2 SO4

2- + 16 H+ Stage 1 proceeds by Reactions (1) and (2) and is characterized by generally moderate pH (above 4.5), high sulfate concentrations, low dissolved iron concentrations, and low acidity. The reactions are typically abiological. Stage 2 also proceeds by reactions (1) and (2) however, ferrous iron oxidation (Reaction 2) is driven by the activity of Thiobacillus ferro-oxidans (T. Ferro-oxidans). The pH range for Stage 2 is generally between 2.5 and 4.5, and the discharge is characterized by high sulfate concentrations, high acidity, high dissolved iron concentrations, and low Fe3+/Fe2+. The final stage in the progression of pyrite oxidation, Stage 3, proceeds by Reactions (3) and (4), and here, reaction rates are totally determined by the activity of T. ferro-oxidans. The pH of the discharge during Stage 3 pyrite oxidation may be below 2.5 and characterized by high sulfate, acidity, and total iron concentrations and high Fe3+/Fe2+ ratios. One can see in the progression from abiological Reactions (1) and (2) to biologically catalyzed Reactions (3) and (4) a rapid increase in acidity released for each mole of pyrite consumed. 2.6.2 Acid neutralization Although acid generation takes place through sulfide mineral oxidation, the release of acidity from Reactions (1), (2), and (4) is often moderated by the reactions with acid neutralizing minerals. Carbonate, hydroxide, and aluminosilicate minerals all neutralize acidity although to different buffer pH values. Table 2.4 shows the buffered pH ranges for some of the common acid neutralizing minerals. Table 2.4 Range of equilibrium pH for neutralizing minerals in natural water (Morin, 1990) Mineral suite Range of equilibrium pH Calcium-based carbonates 5.5 to 6.9 Iron-based carbonates 5.1 to 6.0 Aluminum-hydroxide-bearing minerals 4.3 to 5.0 Iron-hydroxide-bearing minerals 3.0 to 3.7 Aluminosilicate minerals including clay Around 3



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Reactions between calcium carbonate and atmospheric carbon dioxide can produce alkaline solutions with pH values above 8. The following reactions are common buffering reactions that occur between ARD and gangue minerals in hydrothermal alteration assemblages.

(5) 2 H+ + CaCO3 (calcite) = Ca2+ + CO2 (gas) + H2O

(6) 2 H+ + FeCO2 (siderite) = Fe2+ + CO2 (gas) + H2O

(7) 2 H+ + 2 KAlSi3O8 (potassium feldspar) + H2O = Al2Si2O5(OH)4 (clay) + 4 SiO2 (quartz) + 2 K+

2.5.3 Acid generation and neutralization estimates The ARD potential of rock material is estimated based upon the results of Acid Base Accounting (ABA) tests (refer to Section 4.2.1). The total sulfur (Stot), and sulfur species, sulfide sulfur (Spy), sulfate sulfur (SSO4), and refractory sulfur are measured either by pyrolysis or wet chemical analyses and reported in units of weight percent (wt%). Either the Stot or the Spy is converted to acid generating potential (AGP) which is reported in units of kg CaCO3/t of rock by multiplying the wt% sulfur by 31.25. The conversion factor is based upon the theoretical acidity produced by one mole of pyrite being neutralized by one mole of CaCO3. The neutralizing potential (NP) is measured by mixing the samples with a known excess of a standardized hydrochloric acid and titrating the solution with a standardized sodium hydroxide to pH 7. The NP is reported in units of kg CaCO3/t of rock. Criteria to predict the ARD generating potential of rock material have been established over the years using the AGP and NP values based upon empirical observations at many mining properties. The Net Neutralizing Potential (NNP) is the difference between the NP and AGP. The NP:AGP ratio is an alternative criterion. NNP is a traditional part of the acid base accounting and is often reported by laboratories (refer to Appendix B). The NP:AGP ratio provides a measure of the proportions of NP to AGP and the criteria can be used over a wide range of AGP values (Price, 1997). Because of the greater AGP range for which the NP:AGP ratio can be applied, Price (1997) recommended that the NP:AGP ratio not the NNP be used for:

• Identifying and separating potentially acid generating material during operations and

• As a screening criteria in the early stages of an ARD prediction program. The categorization of NP/AGP ratio after SRK (1992) has been adopted as the ABA criteria for characterizing the waste rock at Kupol. Tables 2.5 summarize the criteria using NP/AGP to characterize the ARD potential of mine rock.



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Table 2.5 Criteria for acid generating potential based upon the NNP and NP/AGP

Criteria NP/AGP (SRK, 1992) NP/AGP (Price 1997) Acid consuming > 3 >2

Uncertain acid generating < 3 and > 1 1 to 2 Acid generating < 1 <1

Using the SRK (1992) criteria for NP:AGP ratio, rock with NP greater than 3 times the AGP is considered acid consuming while rock with ratios less than 1 to 1 are considered acid generating. Rock NP/AGP between 3 and 1 are neither clearly acid generating nor acid consuming because factors other than the simple sulfide and carbonate concentrations may influence the final ARD discharge (e.g. material grain size or heterogeneity of the waste rock). The SRK authors concluded that additional test work, typically in the form of kinetic testing, is required to evaluate the ARD potential of the “uncertain” category. Price (1997) uses a slightly different set of NP:AGP ratios for his categories of ARD potential. The author suggests that ratios greater than 2 have either a low or no potential to generate ARD and ratios less than 1 are acid generating. The range between 2:1 and 1:1 is classified a “possibly” acid generating. Price cautioned that acid base accounting test work on its own provides only an estimate of the potential for ARD and suggested that additional, more accurate testing is required to develop “more refined materials handling criteria.”



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3.0 Mine Plan The current mine plan prepared by Wardrop (2005) includes an estimated 18.6 million tonnes of waste rock that will be mined from the proposed open pit and underground during the mine life. Roughly 2 million tonnes of waste from the underground will be used as backfill within the underground workings. The balance of the waste rock, approximately 17 million tonnes, will be delivered to the tailing facility during operations for either dam construction or storage and subsequent submergence. The waste rock in the Wardrop mine plan (05.06.01 ARD Schedule, 2 June 05) has been categorized based upon the ARD potential. The model includes acid generating waste (AG), potentially acid generating waste (PAG), and non-acid generating waste (NAG). 3.1 ARD potential The ARD potential is assigned to each of the waste rock blocks in the resource model based upon the pyrite and carbonate codes recorded in the detailed geologic drill hole logs (refer to Appendix A). The ARD potential code for the resource model uses the logging codes rather than using the results of laboratory analyses of selected samples because:

• Every meter of drill core has been carefully logged and assigned a pyrite and carbonate code, and the large volume of data can be directly incorporated into the geologic model and used in the resource model, and

• the geochemical database consists of 276 samples that, although representative of the different lithologic, mineralization, and alteration assemblages at Kupol, is limited in size compared to the size of the orebody and limited to the lithologic, mineralization, and alteration assemblages that can be incorporated into the geologic model.

The results of the geochemical analyses have been used to more accurately apply the detailed logging codes to characterize the range of ARD potential and metal leaching properties of the waste rock at Kupol. During geologic logging of the split core, a numeric value for the pyrite and carbonate concentrations for each lithologic interval is assigned by the geologist based upon visual pyrite abundance estimate and the reaction of the core to hydrochloric acid, respectively. Calcite (CaCO3), the dominant carbonate mineral in the Kupol deposit and the dominant source of acid neutralizing potential, reacts with hydrochloric acid (refer to Reaction (5), Section 2.5.2), and the rigor of the reaction is roughly proportional to the abundance of calcite in the sample. Table 3.1 shows the detailed logging codes for pyrite and carbonate concentrations and the respective criteria.



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Table 3.1 Detailed geologic logging codes

Code Description Code Description Pyrite codes Carbonate codes

0 < 1 volume % pyrite 0 Absent 1 1 – 3 volume % pyrite 1 Weak 2 3 – 7 volume % pyrite 2 Moderate 3 > 7 volume % pyrite 3 Strong

Comparison of the detailed logging codes and weight percent total sulfur and NP for the 276 waste rock samples analyzed as part of the ARD program (refer to Section 5.1) indicates that at higher concentrations of pyrite (measured as Stot) and carbonate (measured as NP), the logging codes may differ significantly from the laboratory measurements. This discrepancy arises because of possible sampling errors or estimation errors. Although the samples selected for the geochemical analyses are representative of the entire interval of core, only a small portion of the interval may have been collected for analyses and a still smaller portion was analyzed for ARD, leading to a potential sampling error. In addition, estimating mineral abundance and reaction strength are often very subjective observations, leading to a potential error in logging. In order to minimize the discrepancies, the detailed log codes are grouped together into two categories of low and high pyrite and low and high carbonate. Table 3.2 summarizes the criteria for each category. Table 3.2 Pyrite and carbonate categories

Model categories

Log code Criteria and explanation

Low sulfur (LoPy) 0 < 1 wt% pyrite, Roughly 75% of the log intervals coded as

“0” has Stot less than 1 wt% High sulfur

(HiPy) 1+2+3 > 1 wt% pyrite, Roughly 58% of the log intervals coded as 1, 2, or 3 has Stot greater than 1 wt%

Low carbonate (LoCarb) 0

< 40 tonnes CaCO3/1000 tonnes (1), Roughly 73% of the log interval coded as “0” has NP values less than 40 tonnes

CaCO3/1000 tonnes (1)

High carbonate (HiCarb) 1+2+3

> 40 tonnes CaCO3/1000 tonnes (1), Roughly 62% of the log intervals coded as 1, 2, or 3 has NP greater than 40 tonnes

CaCO3/1000 tonnes (1) (1) 40 tonnes CaCO3/1000 tonnes was used as an estimate of neutralizing capacity to neutralize the median wt% pyrite (0.24 wt% Stot ) for pyrite Code “0”. The resource model used the intensity of pyrite mineralization and carbonate alteration from the logging codes of the diamond drill holes to assess the ARD potential of the waste tonnages (Kupol Technical Report, March 2005). Three-D block models were first constructed that incorporated the codes 0, 1, 2, and 3 for pyrite mineralization and 0, 1, 2, and 3 for carbonate alteration (refer to Table 3.1 for definition of each code). A code of



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11 was applied to blocks over 100 m beyond the nearest data point and assigned a pyrite code of 1 and a carbonate code of zero. The 3-D models then used a nearest neighbor algorithm to assign codes to each block in the model. Since the pyrite mineralization and carbonate alteration appeared to show both strong lithologic and stratigraphic controls; interpolation routines were incorporated in the models that included lithologic units (vein, stockwork, dyke, fault and basalt), orientation domains used in the grade model, and overall orientation of the volcanic stratigraphy. The pyrite and carbonate codes for each block of the 3-D models were then incorporated in the resource model. Figures D-1a and D-1b, D-2a and D-2b, D-3a and D-3b, and D-4a and D-4b are level plans showing the distribution of blocks in the model with carbonate and pyrite codes; grey = default, yellow = very weak, orange = weak, green = moderate, and blue = strong for the 400, 450, 500 and 550 m elevations (refer to Appendix D). Figures D-5a and D-5b and D-6a and D-6b are east-west cross sections showing carbonate and pyrite codes along the 90972 and 91090 northings, respectively. Results of the ARD analyses (refer to Section 2.5.3) were used along with the combination of pyrite and carbonate codes to assign an ARD code (AG, PAG, and NAG) for each waste block of the resource model. There are four possible combinations of pyrite and carbonate codes: 1) LoPy-LoCarb, 2) LoPy –HiCarb , 3) HiPy-LoCarb, and 4) HiPy-HiCarb. The NP/AGP ratios for each of the four pyrite and carbonate codes were calculated using the median NP and AGP values for each code. Stot concentrations were used to calculate the NP/AGP ratios rather the Spy because the higher Stot concentrations were more conservative. Table 3.3 summarizes the results of the NP:AGP ratio calculations and list the number of samples used for each estimation. Table 3.3 The median NP:AGP ratios of the 276 waste rock samples

LoCarb HiCarb

LoPy 7.7

n = 80

21.2 n = 71

HiPy 0.4

n = 51

2.7 n = 74

n = the number of samples used to calculate the median NP/AGP values Table 3.3 shows that the LoPy –HiCarb combination has a NP:AGP ratio greater than 3, which falls in the field of non-acid generating (refer to Table 2.5). The combination of HiPy-LoCarb has a NP:AGP ratio of less than 1, which is within the field of acid generating waste rock. The combinations of LoPy-LoCarb and HiPy-LoCarb have ratios that are either non-acid generating or in the uncertain acid generating field. The four combinations shown in Table 3.3 were assigned ARD codes for the resource model as shown in Table 3.4.



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Table 3.3 ARD model matrix

LoCarb HiCarb < 40 kg CaCO3/t NP > 40 kg CaCO3/ t NP

LoPy <1 wt % Stot PAG NAG HiPy >1 wt% Stot AG PAG

Waste rock in the resource model that is classified as AG and NAG are dealt with as if they are acid generating and acid neutralizing, respectively. Waste rock classified as PAG is tentatively handled as non-acid generating. The classification of PAG material as non-acid generating is based upon:

• The results of geochemical test work (refer to Section 5.0) which shows that generally the PAG material, although lacking significant alkalinity, remains near neutral through time.

• The assumption that prior to operations a waste rock management strategy (AMEC, 2005) will be in place to address waste rock classification and identification methods, material handling objectives and methods, waste rock storage, facility closure plan objectives and design, and a detailed facility monitoring plan (refer to Section 6.4).

• As part of the waste rock management strategy, a program will be in place during operations that will incorporate:

• test methods to identify acid generating material within the PAG (as well

as AG and NAG) waste rock at the time the material is blasted and assayed prior to mining and

• handling methods for the acid generating waste material will be in place to isolate the acid generating material and to mitigate potential acid generation.

3.2 Model predicted tonnes by ARD potential Results of the resource model predictions of waste tonnage from the open pit and underground are shown in Table 3.5. The tonnages have been subdivided into 1) NAG, PAG, and AG based upon the resource model defined tonnages and 2) non-acid generating, potentially acid generating, and acid generating waste categories based upon the proportions of each category in the ARD sample database (refer to Section 5.1 and Table 5.2). The waste tonnages shown in Table 3.5 include only waste rock that will be available for surface disposal. Roughly 2 million tonnes of underground waste will be used for backfill in the underground workings.



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Table 3.5 Model predicted tonnes of waste rock mined from the open pit and underground and not used for back fill. NP:AGP 2006 2007 2008 2009 2010 2011 2012 2013 2014 Total






Roughly half of the waste rock at Kupol (47%) falls within the range of non-acid generating waste, with the balance of the waste rock being composed of acid generating waste (41%) and potentially acid generating waste (12%). It is anticipated that the waste rock will be tested during mining to identify, prior to shipment to the tailing facility, the ARD potential of the material.



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4.0 Test program Waste rock and ore samples have been collected from splits of whole core from exploration holes drilled by BGC during the 2003 drilling campaign, historic Russian exploration holes at the Kupol site, and geotechnical diamond drill holes of the 2004 drilling campaign. All of the geochemical test work has been conducted at VIZONSCITEC (formerly BC Research, Inc.), 3650 Wesbrook Mall, Vancouver, BC, V6S 2L2, Canada. The XRD analyses were conducted, under a subcontract from VIZONSCITEC, by Mati Raudsepp, Ph.D. and Elisabeth Pani, Ph.D., Dept. of Earth & Ocean Sciences, 6339 Stores Road, The University of British Columbia, Vancouver, BC V6T 1Z4 4.1 Sample selection The sample selection process and the number of samples collected of waste rock were based upon standard industry practices to obtain high quality, representative samples of all geologic units related to the mine plan (refer to Section 4.3, SRK, 1989). Sample selection followed a complete review of the site geology and was conducted in cooperation with the site geologists in order to:

• characterize the variability in lithology, alteration, and mineralization within the known geologic units.

• characterize the spatial variability along strike, at depth, and laterally away from

the vein system. A total of 234 waste rock samples were collected during fall 2003 from core recovered during the BGC drilling program and from the historic Russian exploration holes. An additional 42 samples were collected from oriented core holes that were drilled in the footwall of the deposit as part of the 2004 geotechnical investigation for pit design. Table 4.1 summarizes the number of samples of each major rock type, alteration type, and mineralization type identified in the detailed drill hole logs. Note that the number of samples with each of the different alteration assemblages present does not sum to the total number of samples for each rock type because of the superposition of overlapping alteration assemblages.



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Table 4.1 Geochemical sample matrix by lithology and alteration assemblage (1) Major COUNT Lithologies containing some degree of each alteration assemblage Total APROP AADUL ASIL ASER AARG ACARB MPY Andesite flows 73 36 1 19 19 10 50 32

Andesite tuff 127 68 1 39 51 41 65 78

Sediments 7 4 0 2 4 0 6 3 Rhyolite flows 7 3 0 3 1 1 3 2

Rhyolite dikes 45 7 0 5 8 8 13 2

Mafic intrusives 3 2 1 1 1 1 3 2

Faults 9 4 0 2 1 6 3 3 Surface 2 0 0 0 0 2 1 0 Veins 3 1 0 2 1 2 0 2

(1) Alteration and mineralization is identified in each lithology type when it has a logging code of 1 or greater, i.e. weak to strong alteration or mineralization. Abbreviations include: APROP = propylitic alteration, AADUL = adularia alteration, ASIL = silicification alteration, ASER = sericitic alteration, AARG = argillic alteration, ACARB = carbonate alteration, MPY = pyrite mineralization The complete list of waste rock samples is shown in Appendix B. Table B.1 in Appendix B includes the sample number (ARD-__), as well as the sampling and assay intervals, drill hole number, drill hole northing, easting, and elevation, and also includes the lithology, alteration, and mineralization codes. Plate 2.1 is a map showing the location of the different geochemical samples. Plates 2.2a and 2.2b are east west cross sections through the deposit that illustrates the distribution of samples across the orebody. Ninety five samples of ore material were included in the ARD characterization testing. The ore samples are derived from metallurgical composite samples and are shown in Appendix B, Table B.2. 4.2 Test program The test program follows an industry standard, staged progression of static and kinetic testing (refer to SRK, 1989; Price, 1997). Static tests are analyses designed to measure the quality and quantity of different constituents in a sample at one point in time (Price, 1997). Static tests include elemental analyses, sulfur species and neutralization potential analyses, pH measurements, particle size analyses and mineralogical analyses. Kinetic tests provide a measure of the dynamic performance or reactivity of a sample and include laboratory procedures such as column tests or humidity cell tests and field studies such as on-site rock pile tests. Table 4.2 lists the tests conducted on samples of waste and ore from the Kupol Project and the number samples of waste rock and ore used in each test format. A brief description of each test is included in Section 4.2 and a more detailed test description is included in Appendix C.



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Table 4.2 Static and kinetic testing of the waste rock and ore samples Description Waste rock samples Ore samples Acid base accounting tests (ABA) 276 95 Net acid generating tests (NAG) 23 Strong acid digestion with metals analyses (ICP) 276 90 Synthetic precipitation leaching procedure (SPLP, EPA-1312) 78 1 Humidity cell tests 31 1 X-ray diffraction (XRD) 10 QA/QC checks 32 4.2.1 Acid Base Accounting tests (ABA) The combined measurements of sulfur species, neutralization potential (NP), and pH with the calculation of the acid generating potential (AGP), net neutralizing potential (NNP) and the neutralizing potential and acid generating potential ratio (NP/AGP) are collectively referred to as Acid Base Accounting (ABA)(after Price, 1997). The application of the ABA results is discussed in detail in Section 2.5.3. Total sulfur and sulfur species have been included in the ABA analytical suite in order to identify the different sources of soluble sulfur. Sulfate has been included in the analytical suite because both gypsum and jarosite have been reported in the deposit, and jarosite is a potential source of mineral acidity. Sulfide sulfur has been included in the test suite because pyrite as well as other iron sulfide minerals are the primary sources of acidity during weathering (refer to Section 2.5.1). Refractory sulfur, including organic sulfur, is estimated by difference: total sulfur minus (sulfate sulfur plus sulfide sulfur). Total sulfur (Stot) in the Kupol samples has been measured by Leco Induction Furnace and Automatic Sulfur Analyzer. Sulfate sulfur (SSO4) is analyzed by first digesting the sample in 1 N hydrochloric (HCl) acid and the extract analyzed for sulfate by turbidimetric method. Sulfide sulfur (Spy) is measured in the washed residue. The residue is subjected to aqua regia digestion to oxidize the sulfide present in the sample and the extract analyzed for sulfate again using the turbidimetric method. Stot , SSO4 , and Spy are reported in units of weight percent (wt%). The NP is measured in the Kupol samples by the Sobek method (Sobek, 1978). The sample is first tested by application of a few drops of 1:3 HCl to determine a “fizz” rating which is an indication of the abundance of CaCO3 in the sample. Based upon the “fizz” rating, either 0.1 or 0.5 N HCl is added to the sample and heated until the reaction is complete. The solution is then titrated with either 0.1 or 0.5 N sodium hydroxide to a pH of 7. The results are reported in units of kg CaCO3/tonne of rock. The paste pH is measured by applying distilled water in a 1:1 or 2:1 mass ratio to a sample of < 250 mesh material and the pH measured in the solution. Although a qualitative measure of the reactivity of the sample, paste pH above 7 is assumed to be indicative of the presence of carbonate neutralizing capacity (or drilling fluids). Paste pH



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below 4 suggests the presence of readily available acidity from prior acid generation or oxidation reactions. 4.2.2 Net Acid Generation (NAG) Test The “NAG” test (after Miller, 1998) is a field test designed to measure the acid generated as a result of oxidation of sulfide minerals in a sample by reaction with hydrogen peroxide (H2O2). The “NAG” test is an industry standard test, particularly in Australia and the South Pacific, to identify ARD source rock. Two measurements are derived from the NAG test, 1) the NAGpH which is measured from the H2O2 solution after it has been allowed to react with the sample and 2) the “NAG”, which is calculated after the H2O2 solution is titrated with NaOH to pH 4.5 (Miller, 1998). The “NAG” is reported in units of kg H2SO4/tonne. The NAGpH and “NAG” can be used to identify acid generating material at a mine site, particularly after the test results are calibrated against standard ABA results. As a rule, the following criteria have been used to identify different categories of ARD potential (refer to Table 4.3). Table 4.3 Criteria for acid generating rocks using the “NAG” test

NAGpH “NAG” (kg H2SO4/tonne)

Non-acid generating 4.5 0 Potentially acid generating < 4.5 ≥ 5

4.2.3 Strong acid digestion multi-element analysis by ICP The objective of the multi-element analysis is to provide a measure of the solid phase concentrations of various metals that may be of environmental concern. Results of the analyses are compared to results found in normal rock and soils to identify elements in elevated concentrations. During the analysis, a known volume of aqua regia solution (HCl:HNO3 – 3:1) is applied to a small amount of sample and allowed to react for 2 hours. An aliquot of sample solution is then recovered and analyzed using an Inductively Coupled Plasma (ICP) instrument. The results are reported in either ppm for trace elements or weight percent for major elements. 4.2.4 Synthetic Precipitation Leaching Procedure (SPLP, EPA 1312) The Synthetic Precipitation Leaching Procedure (SPLP) test is a test method designed to determine the mobility of organic and inorganic analytes present in a soil or rock sample. SPLP test results are useful for comparison to water quality standards in order to identify potentially mobile elements of concern. Because of the high water to rock ratio (20 to 1



24

by weight), the measured concentrations derived from the test may not accurately reflect the actual water to rock ratios at the site and therefore the expected leachate water chemistry. The extracted liquid for the SPLP tests of Kupol samples were analyzed for a) metals by ICP-MS: aluminum, antimony, arsenic, barium, beryllium, bismuth, boron, cadmium, calcium, chromium, cobalt, copper, iron, lead, lithium, magnesium, manganese, mercury, molybdenum, nickel, phosphorus, potassium, selenium, silicon (SiO2 ), silver, sodium, strontium, tellurium, thallium, tin, uranium, titanium, thorium, vanadium, zinc, and zirconium and b) general parameters: pH, alkalinity, acidity, sulfate, total dissolved solids (TDS), bicarbonate, fluoride, chloride, nitrate, and nitrite. 4.2.5 Humidity cell tests Humidity cell tests are the primary, industry wide test format for predicting reaction rates under aerobic conditions. The resulting data provides a measure of the rates of metals release, acid generation, and acid neutralization under the geochemical conditions encountered in the test. The results of the humidity cell tests will be used in the future to model the behavior of waste rock in the tailing facility and in the pit walls. The humidity cell format used for the Kupol samples is based upon the MEND method. The humidity cell test is begun by placing a 1 kg (dry weight) sample of 80% minus ¼ inch rock sample in a humidity cell chamber and applying 750 mL of demineralized water to the sample. The sample is allowed to soak for 2 hours and the lixiviant is recovered and filtered and preserved for dissolved analyses. This is defined as Week 0 and the sample is analyzed for the same suite of analytes as the SPLP sample. In the week following the initial flushing, dry air is passed over and through the sample for three days. On the morning of the fourth day, the moist air cycle begins and moist air is passed over and through the sample for the next three days. On the seventh day, the second flushing cycle begins when 500 mL of demineralized water is applied to the sample and allowed to soak for 2 hours. The Week 1 leachate is then recovered and analyzed for a short list of analytes. The seven day cycle is repeated for a total of 20 weeks. Table 4.2 shows the analysis schedule for the humidity cell tests at Kupol. Table 4.2 Humidity cell analysis schedule

Analyses Weeks Short list (input volume, output

volume, pH, conductivity, sulfate) 1, 2, 3, 4, 5, 6, 7, 8, 9, 10,11, 12, 13, 14, 15, 16,

17, 18, 19, and 20 Full analysis 0, 4, 8, 12, 16, and 20

4.2.6 Mineralogical analyses (X-ray diffraction – XRD) Mineralogy plays a large role in determining the rates of metals release, acid generation, and acid neutralization (Price, 1997). The mineralogical analyses at Kupol include X-ray diffraction analyses of the humidity cell samples and petrographic analyses of chip samples of vein and wall rock material.



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4.2.7 Grain size analyses Reaction rates for oxidation of waste rock in the tailing facility, in the pit walls, and in the humidity cell test chambers are proportional to the reactive surface area of the material. The grain size analysis of the humidity cell test material provides the data to quantify the surface area of the material in the cell. The results of the humidity cell tests are scaled to the fit the waste rock in the tailing facility and pits wall by converting the concentrations from the monthly chemical analyses reported from the laboratory into a rate function for generation of a mobile chemical load. The equation for the conversion takes the form:

(1) C (mg/L) * R (L/kg) * A (kg/m2) * T (1/week) = mg/m2/wk

C = the concentration of dissolved ion in the monthly flush sample R = application rate of lixiviant in the monthly flush of the humidity cell A = reactive surface area of the rock in the humidity cell from the grain size analyses T = sampling frequency

4.2.8 QA/QC The quality control program for the geochemical test work at Kupol includes both the laboratory QA/QC program and 10% duplicate analyses of all leachate chemical analyses. The leachate duplicate analyses include:

• nine analysis or 1 in 10 of the 95 SPLP tests leachate samples, • three sets of 6 duplicates of the monthly full chemical analyses from the humidity

cell tests for weeks 12, 16 and 20 of the 2003 sample ARD test program, and • one duplicate from each of the monthly full chemcial analyses from the humidity

cell tests for weeks 0, 4, 8, 12, and 16 of the 2004 sample ARD test program . 4.3 Comparison with industry standards. The ARD geochemical sampling program follows industry standard format in which 1) the number of samples is tied to the proposed tonnage of material expected to be mined and 2) the individual sample selection is comprehensive and based upon the understanding of the geology at the time of sampling. The number of samples collected is roughly proportional to the abundance of each of the lithology, alteration, and mineralization assemblages identified within the deposit. The test program likewise follows the recommended test sequence including a range of static tests (ABA, SPLP, mineralogy, and whole rock chemical analyses) to characterize the composition and variability in geochemistry of the various lithology, alteration, and mineralization assemblages and kinetic testing using the humidity cell test format to characterize the



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long term leaching characteristics and rates of each of the lithology, alteration, and mineralization assemblages . The current sampling and testing program is limited in the number of samples collected and tested from within or near the current open pit. When initial samples were collected in 2003, the lateral extent and the distribution of gold mineralization in the orebody were only beginning to be understood. The first suite of samples therefore was designed to characterize the waste rock within the entire ore zone with emphasis on the waste rock lying along the possible limits of a mine void for a large open pit. With time and further definition of the orebody, the proportions of reserves mined by open pit versus underground methods has changed resulting in a smaller open pit and greater underground production. The sampling program still remains representative of the waste rock of the deposit as a whole, however only about 1 in 10 of the samples is from the proposed open pit. In order to improve the characterization of the waste rock in the proposed open pit:

• Additional samples may be collected and tested from exploration holes that penetrate the pit limits to better aid the model prediction of the abundance of AG, PAG, and NAG material types within the proposed mine void.

• The operational monitoring ARD program (refer to Section 6.4) will be conducted

during mining to identify waste rock material types for special handling.



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5.0 Test results The test results for ABA, strong acid digestion ICP analyses, NAG, and SPLP tests are discussed in the context of the resource model categories of acid generating (AG), potentially acid generating (PAG), and non-acid generating (NAG) waste rock. Each of the three categories is identified in the resource model based upon detailed drill hole geologic logging codes (refer to Section 3.1 for a discussion of the resource model). The table in Appendix A lists the sample ID, the lithology and lithology codes, the carbonate alteration and pyrite codes (0 to 3), and the category of material (LoPy, HiPy, LoCarb, and HiCarb). Fifty-one (18%) of the waste rock samples in the geochemical data set are AG, while 71 (26 %) are NAG. The balance of the geochemical samples (154 or 56%) are PAG. This compares with the resource model estimates for the entire planned pit of 27% AG, 10% NAG, and 64% PAG waste rock. 5.1 Acid Base Accounting (ABA) ABA testing is based upon several simplifying assumptions: a) that all of the sulfide material is pyrite or other iron sulfide, the iron sulfide is available to react, and the iron sulfide is oxidized instantly and completely, and b) all of the neutralizing capacity is carbonate neutralizing capacity, the carbonate is available to react, and the acid consuming reactions are instantaneous and run to completion. ABA testing does not consider chemical kinetics or reaction rates and as a consequence neither predicts the ARD conditions at an intermediate time period nor the affects of other sources of acid generation or neutralization (refer to Section 2.5). To account for the uncertainty in the ARD predictions using ABA test results, the criteria for acid generation and neutralization include a broad range of NNP and NP/AGP values that are classified as uncertain or potentially acid generating (refer to Table 2.5). Results of the ABA test work on 276 waste rock samples and 95 ore samples are shown in Appendix B and summarized in Table 5.1. 5.1.1 Waste rock characterization Paste pH Paste pH is an indicator of the presence of readily solubilized mineral acidity from prior oxidation reactions or the presence of natural neutralizing capacity (refer to Section 4.2.1). The range in paste pH of 2 to 10 is similar for the AG, PAG, and NAG samples suggesting that that waste rock in each of the categories has been at least locally oxidized and contains some ambient neutralizing capacity. This is consistent with our understanding of the mineralization and alteration at Kupol. The geologic model shows that 1) the alteration envelopes are tightly controlled by the vein geometry and 2) the alteration types overlap. The difference between the median and the mean suggests that the distribution of pH values in each population is not normally distributed and the



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subsequent discussion will be based upon the median values (median is the middle of a population, i.e. one half the population with values greater than and one half of the population with values less then the median value). The paste pH increases from a median of 7.00 for AG to a median of 8.3 for NAG samples, which is as expected. Sulfur species In 1978 the United States Environmental Protection Agency (EPA) identified as “potentially problematic” material with NNP of as little as -5 kg CaCO3/tonne (Sobek, 1978). In the absence of any neutralizing potential, -5 kg CaCO3/tonne is less then 0.2 wt% Stot or Spy or less than 0.1 volume percent pyrite; suggesting that even a very small amount of pyrite can be acid generating. Fortunately, the rocks at Kupol contain varying amounts of carbonate neutralizing potential. The range in Stot concentration is 0.01 to 6 wt% for AG, PAG, and NAG waste rock types. The median Stot concentrations range from 1.6 and 0.11 wt% and decrease from AG to NAG classifications, which is consistent with the assumption that the source of acid generation is the sulfide-bearing minerals. The dominant sulfur species is Spy. The median SSO4 is 0.01 to 0.02 wt% which is at or near the detection limit. Spy ranges from below the detection limit to nearly 6 wt%. The median Spy concentration decreases from 1.5 wt% for AG to 0.01 wt% for NG, paralleling the decrease in Stot concentration. The median AGP, which is derived from the sulfide sulfur concentrations (refer to Section 4.2.1), decreases from 47 kg CaCO3/tonne in AG to 1.8 kg CaCO3/tonne in NAG waste rocks. Neutralizing potential Carbonate minerals occur both in discrete veins and as disseminated grains in the altered waste rock throughout the deposit however the NP increases from the AG to NAG waste rock types. The range in NP for the AG samples is between -23 and 114 kg CaCO3/tonne and -15 to 164 kg CaCO3/tonne for the NAG samples. The negative values indicate the presence of natural acidity in the sample at the time of analysis. The high NP values, values greater then a few kg CaCO3/tonne, indicate that carbonate alteration is widely distributed throughout the deposit, even within zones of strong sulfide mineralization. The median NP values increase from 6 kg CaCO3/tonne in the AG samples to almost 70 kg CaCO3/tonne in NAG samples.



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Table 5.1 Summary of ABA data from ore samples and waste rock samples for AG, PAG, and NAG categories

Sample Paste Total Sulfate Sulfide Acid Generating Neutralization NP/AGP pH Sulfur (1) Sulfur (1) Sulfur (1) Potential (2) Potential (Wt.%) (Wt.%) (Wt.%) (Kg CaCO3/Tonne) (Kg CaCO3/Tonne)

Acid generating - AG Count 51 51 51 51 51 51 51 Minimum 2.00 0.01 0.01 0.01 0.31 -23 -0.21 Maximum 9.20 5.67 1.78 5.04 158 114 59.9 Average 3.58 1.82 0.18 1.69 52.8 25.6 4.65 Median 7.00 1.57 0.02 1.50 46.9 6 0.44

Potentially acid generating - PAG Count 154 154 154 154 154 154 154 Minimum 2.70 0.01 0.01 0.01 0.31 -22 -11.5 Maximum 10.00 6.01 2.29 5.94 186 172.1 404 Average 4.46 1.13 0.13 1.00 31.2 44.8 29.4 Median 8.00 0.555 0.01 0.28 8.6 30.5 3.85

Non-acid generating - NAG Count 71 71 71 71 71 71 71 Minimum 3.00 0.01 0.01 0.01 0.31 -14.6 -11.2 Maximum 9.50 5.11 2.74 4.55 142 164 524 Average 4.64 0.76 0.18 0.56 17.55 60.3 85.3 Median 8.30 0.11 0.01 0.06 1.77 67.3 21.2

Ore samples Count 95 95 95 95 95 95 95 Minimum 2.90 0.02 0.01 0.01 0.31 -9.00 -2.56 Maximum 8.50 3.82 1.22 2.91 91.1 115 22.7 Average 3.71 0.80 0.30 0.39 12.32 2.69 0.58 Median 3.90 0.61 0.27 0.20 6.14 -0.70 -0.11

(1) One half the detection limit is used for values reported as below the detection limit. (2) The acid generating potential (AGP) is based upon the sulfide sulfur concentration. (3) The average paste pH is derived from the average H+ concentration. NP:AGP ratio criteria The NP:AGP ratio is a well established criteria to identify ARD potential in mined material (refer to Section 3.1). Because of the range of sulfur species and carbonate concentrations in the AG, PAG, and NAG material types, the ARD codes do not lend themselves to simple categories of acid neutralizing, acid generating, or uncertain ARD potential. The mean and median values of NP:AGP ratios for AG waste rock are 4.6 and 0.4 respectively. The mean ratio would suggest that the AG material type may be non-acid generating, however two thirds of AG samples fall in the acid generating field (refer to Table 5.2). The mean and median NP:AGP ratios for the NAG material type are all above 3, suggesting the NAG material is acid neutralizing. The PAG material type has a broad range of NP:AGP ratios however, the mean and median values are both within the



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field of acid neutralizing (29.4 and 3.85, respectively) and over half of the samples fall within the field of acid neutralizing waste. Table 5.2 Summary of proportions of the AG, PAG, and NAG material in each ARD category

ARD code NP:GP < 1 1 < NP:AGP < 3 NP:AGP > 3 Acid generating Uncertain Acid neutralizing

AG 69% 10% 22% PAG 34% 12% 54% NAG 17% 11% 72%

Figure 5.1 is a plot of all of the AGP and NP data for the waste rock samples identified by code, AG (●), PAG (∆), and NAG (▀). The figure graphically shows the broad range of NP and AGP for each of the code types. However, as noted in Table 5.2, each code (AG, PAG, and NAG):

• includes samples that are acid generating and acid neutralizing and • includes a majority of samples that are either acid generating (AG) or acid

neutralizing (NAG and PAG). Figure 5.1 NP versus AGP

0

20

40

60

80

100

120

140

160

180

200

0 20 40 60 80 100 120 140 160 180 200

Neutralizng potential (NP) (kg CaCO3/tonne)

Aci

d ge

nera

ting

pote

ntia

l (A

GP)

(kg

CaC

O3/

tonn

e)

AG NAG PAG

3:1

1:1Acid generating waste

Potentially acid generating waste

Non-acidgenerating waste



31

ABA Criteria for waste handling From an operational perspective, it is ideal to have a simple set of criteria to identify potentially acid generating materials for special handling (refer to Section 6.4). One mechanism to identify acid generating waste is “on site” analyses of the sulfide content and neutralizing potential of the waste rock. Sulfur and carbon analyses can be done using a Leco CR-12 carbon and sulfur analyzer according to ASTM method D-4239. Figure 5.2 is a plot of total sulfur versus the NP:AGP ratio for all of the 276 waste rock samples. The figure shows that:

• 100% of the waste rock with a total sulfur concentration of 3 wt% or greater is acid generating

• Roughly 60% of the samples have total sulfur less that 1.0 wt% and 15% have total sulfur concentrations less than the detection limit of 0.02 wt%.

• Roughly 80% of the waste rock with a total sulfur concentration of 1% or less have NP:AGP ratio of over 3 (i.e. the field of acid neutralizing)

• Roughly 20% of the waste rock samples have NP:AGP ratios of less than 3 suggesting that a significant portion of the low sulfur-bearing waste rock is potentially acid generating and neutralizing potential laboratory tests may be required to prevent co-mingling of acid generating and acid neutralizing waste rock. and 5% with NP/AGP between 1 and 3 (i.e. the field of uncertain acid generation).

Figure 5.2 Total sulfur versus NP/AGP for all waste rock samples

0.01

0.1

1

10

100

1000

0 1 2 3 4 5 6 7

Total sulfur (wt%)

NP/

AG

P ra

tio

AG NAG PAG

NP:AGP = 3:1

NP:AGP = 1:1

100% of the samples with Stot > 3.0 wt% have NP/AGP ratios < 1 60% of all samples have Stot < 1.0 wt% and 15% of all samples have Stot below the detection limitof 0.02 wt%

Roughly 80% of the sampleswith Stot < 1.0 wt% have NP/AGP ratios > 3



32

5.1.2 Ore samples Paste pH The paste pH of the ore samples is lower than for the waste rock samples (median 3.9) (refer to Table 5.1). The lower paste pH may be the result of either a) supergene weathering of the hydrothermal sulfide minerals by preferential surface water flow through the fractured vein system or b) hydrothermal oxidation of primary sulfides to secondary sulfate minerals such as jarosite during the collapse of the near surface hydrothermal system. There is textural evidence that suggests the jarosite of the vein system is the result of collapse of the hydrothermal system (Shein, personal communication, 2004). Sulfide species The ore samples have lower Stot and sulfur species then the waste rock samples. The ore samples were collected for metallurgical testing and are composites of selected high grade ore vein intercepts. The vein samples are typically silica-rich often with little sulfide mineralization or sulfides encapsulated in quartz with little or no wall rock material. Stot ranges from 0.02 to 3.8 wt% compared to 0.01 to 6 wt% for the waste rock samples, with a median value of 0.6 wt%. Unlike the waste rock samples, SSO4 and Spy are in roughly equal amounts in the ore samples. The range of SSO4 in the ore samples is 0.01 to 1.2 wt% while the range in Spy is 0.01 to 2.9. The median SSO4 and Spy in the ore samples are identical, 0.27 and 0.2, respectively. Neutralizing potential Although dominated by silica, the ore samples contain varying amounts of carbonate. The NP varies from -9 to 115 kg CaCO3/tonne which is similar to NP values of the AG samples. The median value of NP is negative, -0.7 suggesting that on the whole the ore samples lack significant natural NP. During processing, alkali will be added to the ore material as part of the cyanidization process, and so the spent ore will contain a limited amount of neutralizing potential when the material is delivered to the tailing facility or dry stack. NP/AGP ratio criteria The ore samples all fall into the range of acid generating material. The mean and median NP/AGP are less then 1 (0.58 and -0.1, respectively). Current plans call for the spent ore to be stored in a dry stack. The potential for the spent ore to generate ARD will depend upon several factors (refer to Section 2.5) including:



33

• The availability of oxygen to gain access to the sulfides in the tailing to drive the sulfide oxidation reactions.

• The potential for water to infiltrate into the stack during operations and post-closure, react with the contained tailing material, and discharge from the stack.

• The residual lime contained in the spent ore from the alkalinity amendment applied during processing to neutralize acid generated by oxidation in the stack.

5.2 Net Acid Generation (“NAG”) Test The “NAG” test is designed as a field test procedure to identify material types during mining by simply measuring the pH of a solution derived from mixing crushed rock and hydrogen peroxide (the NAGpH) (refer to Section 4.2.2). The “NAG” can be measured in the assay laboratory at the mine site by a more sophisticated titration procedure. The “NAG” tests were conducted using 23 waste rock samples, including 5 AG samples, 10 PAG samples and 8 NAG samples. The results of the net acid generating tests are summarized in Table 5.3. A complete list of “NAG” test results is shown in Appendix B. Table 5.3 Summary of Net Acid Generation test from waste rock samples for AG, PAG, and NAG categories

NAGpH NAG (1) su (kg H2SO4/tonne)

Acid generating - AG Count 5 5 Minimum 2.32 0.00 Maximum 6.12 76.4 Average 2.54 29.0 Median 2.43 26.6

Potentially acid generating – PAG Count 10 10 Minimum 2.3 0 Maximum 7.89 36.2 Average 3.20 4.40 Median 5.17 0

Non-acid generating – NAG Count 8 8 Minimum 2.44 0 Maximum 7.07 33.1 Average 3.16 6.15 Median 6.12 0

(1) Based on titration to pH 4.5 NAGpH The NAGpH is the pH derived from oxidation of iron sulfide minerals in the sample. The NAGpH ranges from 2.3 to almost 8 in the AG, PAG, and NAG samples indicating



34

all three of the categories of waste in the resource model contained both pyrite mineralization and carbonate alteration. The median NAGpH increases from 2.4 in the AG samples to over 6 in the NAG samples. NAG The “NAG”, like the NP, measures the amount of acidity available by titration with sodium hydroxide. The “NAG” range decreases from 0 to 76 Kg H2SO4/tonne in the AG samples to 0 to 33 kg H2SO4/tonne in the NAG samples. The median “NAG” decreases from 26.6 in the AG samples to 0 kg H2SO4/tonne in the NAG samples. NAGpH and “NAG” criteria Miller (1998) indicated that NAGpH values greater than 4.5 and “NAG” values less then 5 are non-acid generating. Based upon these criteria and using the median NAGpH and “NAG”, the AG category of waste is acid generating and the PAG and NAG categories are non-acid generating. NAGpH Criteria for waste handling An alternative criteria to total sulfur analyses for identification of acid generating waste rock during operations is the NAGpH (refer to Section 5.1.1). Figure 5.3 is a plot of the NAGpH versus the NP/AGP. The figure shows that:

• 100% of the waste rock with a NAGpH of 3 or less is acid generating and • 100% of the waste rock with a NAGpH of 5 or above fall outside of the field of

acid generation, including 92% that have NP/AGP over 3 (i.e. the field of acid neutralizing).



35

Figure 5.3 NAGpH versus NP:AGP ratio for all of the 23 waste rock samples

0.01

0.1

1

10

100

1000

0.00 1.00 2.00 3.00 4.00 5.00 6.00 7.00 8.00 9.00

NAGpH

NP:

\/AG

P ra

tio

100% acid generating waste rock

92% acid neutralinzing waste rock

5.3 Strong acid digestion and multi-element analysis by ICP The discussion of the tests results from the ABA and “NAG” has addressed the potential for the waste rock at Kupol to generate acidic seepage. However, of equal significance is the chemical load of the seepage, i.e metals of concern for regulatory authorities in either surface or groundwater. The strong acid digestion and multi-element analyses is a test procedure designed to identify those elements that are available to leach. The procedure is based upon two assumptions a) that concentrations of dissolved species in groundwater in contact with “normal” rock are below a reference standard such as drinking water standard and 2) rock with elemental concentrations significantly above the concentrations in “normal” rock have the potential to leach and increase the groundwater concentrations of anomalous elements to levels above the reference standard. The test does not measure the leaching characteristics of the rock. Tables 5.4a and 5.4b summarize the results of the ICP analyses of 276 waste rock samples. The waste rock samples used in the ICP analyses included 51 AG samples, 154 PAG samples, and 71 NAG samples. Table 5.5 summarizes the results of the ICP analyses for 90 ore samples. The complete test results for the waste rock and ore samples are shown in Appendix B. Tables 5.4a, 5.4b, and 5.5 include the elemental concentrations in a “normal” basalt as a reference standard (after Price, 1997). “Normal” basalt has been used for comparison purposes because the host rocks at Kupol include



36

andesites and basaltic andesite and the table referenced by Price (after Turkenian and Wedepohl, 1961) includes only the broad categories of “normal” ultrabasic, “normal” basalt or “normal” granite. Using the maximum concentration and a criterion of 10 times the concentration of the element in the “normal” basalt, the concentrations in the waste rock samples of silver, arsenic, gold, mercury, and antimony are anomalously high in the AG, PAG, and NAG categories. Selenium and bismuth are high compared to “normal” basalt, however 80% of the analyses of selenium are below the detection limit and 60% of the bismuth analyses are below the detection limit. Thallium concentrations are anomalous in the AG and PAG samples. Generally the maximum, mean, and median concentrations decrease from the AG to the NAG categories. Using the same criteria, the concentrations of silver, arsenic, gold, cadmium, copper, mercury, molybdenum, lead, antimony, selenium, thallium and zinc are anomalous in the ore samples. The concentration of bismuth is also high compared to “normal” basalt however 60% of the bismuth analyses are below the detection limit. The elements that occur in elevated concentrations in the ore samples are the most common pathfinder elements in virtually all epithermal Au systems.



37

Table 5.4a Summary of ICP data from waste rock samples for AG, PAG, and NAG categories (1) Ag Al As Au B Ba Bi Ca Cd Co Cr Cu Fe Ga Hg K La Mg ppm % ppm ppb ppm ppm ppm % ppm ppm ppm ppm % ppm ppm % ppm % avg basalt 0.1 7.80 2.0 4.0 5 330 0.007 7.60 0.2 48.0 170.0 87.0 8.65 17 0.09 0.80 15 4.60 Detection limit <0.1 <0.5 <1 <0.1 <0.1 <1 <0.01 % below detect 19% 0% 0% 19% 17% 0% 57% 0% 34% 0% 0% 0% 0% 1% 9% 0% 0% 0%

Acid generating - AG Count 51 51 51 51 51 51 51 51 51 51 51 51 51 51 51 51 51 51 Minimum 0.05 0.05 3 0.25 0.5 4 0.05 0.02 0.05 0.5 10.8 0.6 0.77 0.5 0.005 0.03 1 0.01 Maximum 100 1.88 4343.4 11500.3 72 67 2.7 3.07 0.8 31.5 239.2 54.6 5.11 9 3.87 0.52 36 1.9 Average 2.93 0.88 420.53 312.52 3.28 13.94 0.24 0.87 0.13 14.05 44.78 21.04 3.36 3.60 0.55 0.12 14.75 0.59 Median 0.5 0.81 193 25.4 2 10 0.05 0.49 0.1 14.2 36.5 19.7 3.57 3 0.28 0.11 15 0.45

Potentially acid generating - PAG Count 154 154 154 154 154 154 154 154 154 154 154 154 154 154 154 154 154 154 Minimum 0.05 0.13 0.8 0.25 0.5 4 0.05 0.03 0.05 0.2 7.5 0.3 0.26 0.5 0.005 0.02 4 0.02 Maximum 87.9 4.85 4278.4 5408.4 8 328 2.1 4.19 0.5 24.7 124.8 64.5 5.65 11 5.41 0.58 27 2.43 Average 1.05 1.22 209.96 62.45 1.91 22.95 0.15 1.35 0.10 11.91 35.14 16.95 3.03 4.67 0.26 0.11 16.08 0.84 Median 0.1 1.27 36.9 2.9 2 13 0.05 1.13 0.1 13.9 29.7 18.3 3.48 4.5 0.06 0.1 16 0.72

Non-acid generating - NAG Count 71 71 71 71 71 71 71 71 71 71 71 71 71 71 71 71 71 71 Minimum 0.05 0.19 0.6 0.25 0.5 5 0.05 0.09 0.05 0.2 1.2 0.5 0.53 1 0.005 0.03 9 0.03 Maximum 4 2.39 1041.9 239.7 14 173 1.7 5.82 0.4 29.1 98 33.4 5.65 11 1.69 0.22 36 2.11 Average 0.23 1.31 61.80 10.84 2.52 26.34 0.14 1.98 0.10 12.51 32.56 16.35 3.19 5.13 0.11 0.09 15.38 0.99 Median 0.1 1.43 20.5 1.9 2 15 0.05 1.71 0.1 14 29 17.7 3.7 5 0.04 0.09 15 1.15 (1) One half the detection limit is used for values reported as below the detection limit.



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Table 5.4b Summary of ICP data from waste rock samples for AG, PAG, and NAG categories (1) Mn Mo Na Ni P Pb Sb Sc Se Sr Th Ti Tl U V W Zn ppm ppm % ppm % ppm ppm ppm ppm ppm ppm % ppm ppm ppm ppm ppm avg basalt 1500 1.5 1.8 130.0 11 6.0 0.2 30.0 0.05 465 4.0 1.380 0.2 1.0 250 0.7 105 Detection limit <0.1 <0.5 <0.001 <0.1 <1 <0.1 % below detect 0% 0% 0% 3% 0% 0% 0% 0% 81% 0% 0% 23% 32% 0% 6% 34% 0%

Acid generating - AG Count 51 51 51 51 51 51 51 51 51 51 51 51 51 51 51 51 51 Minimum 20 1.2 0.004 0.3 0.003 2.5 0.4 0.6 0.25 5 0.2 0.0005 0.05 0.1 0.5 0.05 3 Maximum 5147 18.4 0.174 14.1 0.094 71.9 107.2 10.2 2.7 136 8.5 0.206 6.7 1.4 129 0.5 153 Average 908.51 4.34 0.03 6.52 0.06 9.69 7.76 4.10 0.56 34.65 1.76 0.01 0.55 0.45 41.75 0.13 71.43 Median 724 3.3 0.014 6.6 0.068 7.6 3.4 3.8 0.25 24 1.3 0.001 0.2 0.4 39 0.1 66

Potentially acid generating - PAG Count 154 154 154 154 154 154 154 154 154 154 154 154 154 154 154 154 154 Minimum 22 0.5 0.004 0.05 0.002 1.6 0.1 0.4 0.25 6 0.5 0.0005 0.05 0.1 0.5 0.05 2 Maximum 2554 12.4 0.576 22.8 0.108 32.5 51.9 15.5 6.1 1362 10.2 0.25 7.7 1.8 188 2.1 153 Average 787.16 2.70 0.05 5.28 0.06 7.66 3.21 4.43 0.46 67.83 2.89 0.02 0.33 0.63 47.52 0.16 57.77 Median 778.5 2.05 0.03 5.8 0.0695 6.8 1.7 3.95 0.25 45 1.9 0.002 0.1 0.5 43 0.1 58

Non-acid generating - NAG Count 71 71 71 71 71 71 71 71 71 71 71 71 71 71 71 71 71 Minimum 78 0.4 0.004 0.05 0.003 1.2 0.1 0.7 0.25 11 0.5 0.0005 0.05 0.1 0.5 0.05 17 Maximum 2658 5.1 0.35 11.2 0.104 25.6 10.6 19.4 1.3 558 14.7 0.302 1.1 5.2 213 2.6 129 Average 939.01 2.05 0.07 4.89 0.06 6.40 1.94 5.56 0.33 81.37 2.83 0.03 0.12 0.64 64.72 0.14 58.73 Median 865 1.9 0.038 5.6 0.069 6 1.2 5 0.25 58 1.6 0.004 0.1 0.5 64 0.1 58 (1) One half the detection limit is used for values reported as below the detection limit.



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Table 5.5 Summary of ICP data from ore samples (1) Ag Al As Au B Ba Bi Ca Cd Co Cr Cu Fe Ga Hg K La Mg ppm % ppm ppb ppm ppm ppm % ppm ppm ppm ppm % ppm ppm % ppm % avg basalt 0.1 7.80 2.0 4.0 5 330 0.007 7.60 0.2 48.0 170.0 87.0 8.65 17 0.09 0.80 15 4.60 Detection limit <1 <0.1 <0.1 <1 <0.01 <1 <0.01 % below detect 0% 0% 0% 0% 11% 0% 64% 0% 6% 0% 0% 0% 0% 11% 0% 1% 3% 2%

Ore samples count 28 90 89 90 90 90 90 90 90 90 90 90 90 90 90 90 90 90 In 10.2 0.05 45 1403 0.5 1 0.05 0.01 0.05 0.5 48.5 14 0.39 0.5 0.05 0.005 0.5 0.005 Max >100 0.86 8845.8 99999 6 42 1.2 2.36 17.5 21.4 249.6 1747.5 5.01 4 14.38 0.54 13 1.04 average 0.21 1784 17250 2.32 12.9 0.11 0.22 1.08 3.75 106 101 1.55 1.40 1.69 0.11 3.31 0.07 median 0.175 1293.7 10943 2 10 0.05 0.08 0.3 2.4 99.55 52.05 1.3 1 0.89 0.1 3 0.025 Mn Mo Na Ni P Pb Sb Sc Se Sr Th Ti Tl U V W Zn ppm ppm % ppm % ppm ppm ppm Ppm ppm ppm % ppm ppm ppm ppm ppm avg basalt 1500 1.5 1.8 130.0 11 6.0 0.2 30.0 0.05 465 4.0 1.380 0.2 1.0 250 0.7 105 Detection limit <0.1 <0.001 <0.1 <0.1 <1 <0.1 % below detect 0% 0% 0% 0% 0% 0% 0% 0% 0% 0% 11% 69% 1% 6% 1% 17% 0%

Ore samples count 90 90 90 90 90 90 90 90 90 90 90 90 90 90 90 90 90 In 12 1.7 0.001 2.1 0.001 7.1 7.2 0.2 0.6 4 0.05 0.0005 0.05 0.05 0.5 0.05 7 Max 2956 33.2 0.083 11.7 0.11 1705.7 506.3 6.2 52.1 320 1.5 0.008 15.1 0.6 36 1.7 1700 average 316 9.33 0.01 5.05 0.02 128 82.3 1.44 8.75 42.3 0.32 0.001 2.18 0.20 10.1 0.16 122 median 134.5 6.45 0.008 4.7 0.0185 57.75 56.75 1.25 4.35 23 0.2 0.0005 1.1 0.2 7 0.1 47 (1) One half the detection limit is used for values reported as below the detection limit.



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The ICP analyses of waste rock and ore samples highlight the concentrations of metals that are anomalous by comparison with chemically similar and unmineralized rocks. Table 5.6 summarizes the metals that occur in elevated concentrations in the majority of samples. Table 5.6 Metals occurring in elevated concentrations in ore and waste ARD categories Ore AG PAG NAG

Elements Ag, As, Cd, Cu, Hg, Mo, Pb, Sb,

Se, Tl, Zn

Ag, As, Au, Hg, Sb, Tl

Ag, As, Au, Hg, Sb, Tl

Ag, As, Au, Hg, Sb,

The elevated gold and silver concentrations in the veins and the adjacent wall rock are explained by the presence of gold and silver mineralization. Vartnyan, et al (2001) concluded that the early mineralization at Kupol is associated with elevated concentrations of sulfur, arsenic, and antimony and the later stages of the hydrothermal event have elevated concentrations of selenium, antimony and sulfur. The other elements such as the base metals lead, copper, and zinc are common trace elements in many hydrothermal systems as are bismuth, mercury, molybdenum, selenium, and thallium. 5.4 Synthetic Precipitation Leaching Procedure (SPLP, EPA 1312) The Synthetic Precipitation Leaching Procedure (SPLP) is a test designed to characterize the leaching characteristics of readily solubilized mineral acidity and salts. Due to the short time period of the test (24 hours), the SPLP test is a valid test format to assess the leaching characteristics of non-sulfide-bearing rock or completely oxidized rocks. The short duration of the tests does not allow time for iron-sulfide mineral oxidation. The SPLP test results are useful in identifying metals that can be leached from waste rock rather than as an estimate of absolute concentrations of individual elements in the leachate derived from the tailing facility or pit walls. The chemistry of the discharge from a tailing facility for example depends upon many factors not addressed in the SPLP tests including the water to rock ratio, reaction kinetics, mineral precipitation and sorption within the pile, and the affects of downstream sorption and dilution. Seventy eight waste rock samples and one ore sample were used in SPLP testing. Seventeen samples for the SPLP tests are from the AG category of the resource model, 38 within the PAG category, and 23 within the NAG category. The results of the SPLP tests are summarized in Tables 5.6a, 5.6b, and 5.6c. The complete set of SPLP test results is shown in Appendix B The pH of the leachate from the SPLP tests ranges from roughly 3 to over 9 for the three waste rock categories which, is similar to the ranges in paste pH. Figure 5.4 is a plot of paste pH versus SPLP leachate pH. The SPLP leachate pH is slightly higher than the equivalent paste pH which may result from longer contact time of the SPLP lixiviant with the sample material. The range in SPLP leachate pH indicates the presence of both



41

soluble mineral acidity and natural neutralizing potential in all of the different waste rock types although the median pH increases from 6 to 9 in the AG and NAG categories. The alkalinity and acidity concentrations in the SPLP leachate are only estimates since these parameters were not measure in all samples. Alkalinity was not measured in samples with low pH, while acidity was not measured in samples with elevated pH values. The number of samples with measured alkalinity and acidity are shown in the count. Sulfate is the dominant anion in the analyses. Figure 5.5 is a plot of the sulfate ion concentration versus total dissolved solids (TDS). Sulfate concentrations range from 1 to almost 2000 mg/L for all waste rock types although the median declines from AG samples to NAG samples. TDS ranges from 12 to almost 3000 mg/L and shows a similar decline in median values from 133 to 60 mg/L from AG to NAG samples. Chloride, fluoride, nitrate and nitrite concentrations are generally low for all of the samples. Figure 5.4 Paste pH and SPLP leachate pH

0.0

2.0

4.0

6.0

8.0

10.0

12.0

0 2 4 6 8 10 12

Paste pH (su)

SPLP

leac

hate

pH

(su)



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Table 5.6a Summary of SPLP data from ore and waste rock samples for AG, PAG, and NAG categories (1) pH Alkalinity ACID 4.5 ACID 8.3 SO4 TDS HCO3 Cl F NO3-N NO2-N Fe Al Sb As Ba

su mgCaCO3/L mgCaCO3/L mgCaCO3/L mg/L mg/L mgCaCO3/L mg/L mg/L mg N/L mg N/L mg/L mg/L mg/L mg/L mg/L

Detection limit <0.001 <0.001 <0.001

% below detect 0 46% 15% 33%

Acid generating - AG

Count 17 16 6 14 17 17 16 17 17 17 17 17 17 17 17 17

Minimum 2.9 0.5 4 2 1.4 14 0.5 0.1 0.05 0.01 0.001 0.025 0.008 0.0005 0.0005 0.0005

Maximum 9.0 31 572 1120 1790 2935 30.9 23 2 0.058 0.004 190 110 0.041 1.13 0.019

Average 3.8 7.8 151 158 372 596 7.46 3.05 0.23 0.03 0.00 25.7 11.7 0.0054 0.1664 0.0048

Median 5.9 1.5 13.5 9 73 133 1.5 0.53 0.05 0.01 0.001 0.51 0.21 0.002 0.031 0.002

Potentially acid generating - PAG

Count 38 35 6 20 38 38 35 38 38 38 38 38 38 38 38 38

Minimum 3.30 0.5 1.0 1 1 12 0.5 0.1 0.1 0.01 0.001 0.025 0.005 0.0005 0.0005 0.0005

Maximum 9.7 55.0 336 1140 1970 2856 45.47 7.4 0.8 0.172 0.005 297 78.5 0.009 0.9 0.037

Average 4.5 16.2 61 76 212 320 14.32 1.41 0.13 0.05 0.00 9.6 2.7 0.0019 0.0704 0.0060

Median 7.8 16.0 6.0 12 36.7 96 15.9 0.605 0.1 0.0515 0.001 0.185 0.23 0.00075 0.0085 0.001

Non-acid generating - NAG

Count 23 22 2 7 23 23 22 23 23 23 23 23 23 23 23 23

Minimum 3.80 1.0 2.0 1 1.2 22 1.0 0.1 0.1 0.01 0.001 0.025 0.014 0.0005 0.0005 0.0005

Maximum 9.70 44.0 116 776 1910 2683 30.7 4.4 0.8 0.105 0.007 257 48.2 0.07 2.73 0.031

Average 5.06 22.7 59 120 275 434 18.68 0.71 0.16 0.04 0.00 11.7 2.6 0.0047 0.1364 0.0048

Median 9.00 22.0 59.0 8 5.6 60 18.03 0.33 0.1 0.052 0.001 0.08 0.44 0.001 0.004 0.002

Ore samples

NSVZ 6.1 3 N/A 2 94 110 2.999016 0.62 0.05 0.01 0.001 0.62 0.013 0.001 0.003 0.0005 (1) One half the detection limit is used for values reported as below the detection limit.



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Table 5.6b Summary of SPLP data from ore and waste rock samples for AG, PAG, and NAG categories (1) Be Bi B Cd Ca Cr Co Cu Fe Pb Li Mg Mn Hg Mo Ni

mg/L mg/L mg/L mg/L mg/L mg/L mg/L mg/L mg/L mg/L mg/L mg/L mg/L ug/L mg/L mg/L

Detection limit <0.001 <0.001 <0.05 <0.0002 <0.001 <0.001 <0.001 <0.05 <0.001 <0.05 <0.02 <0.0005 <0.001

% below detect 84% 100% 95% 90% 0% 94% 62% 72% 35% 92% 0% 3% 0% 96% 72% 65%

Acid generating – AG

Count 17 17 17 17 17 17 17 17 17 17 17 17 17 17 17 17

Minimum 0.0005 0.0005 0.025 0.0001 1.15 0.0005 0.0005 0.0005 0.025 0.0005 0.002 0.22 0.001 0.01 0.00025 0.0005

Maximum 0.012 0.0005 0.025 0.0056 563 0.016 0.3 0.25 173 0.004 0.041 119 32.2 0.06 0.0058 0.14

Average 0.0020 0.0005 0.0250 0.0007 72.9 0.0017 0.0535 0.0334 24.3 0.0010 0.014 12.37 2.58 0.0129 0.0016 0.0239

Median 0.0005 0.0005 0.025 0.0001 13.3 0.0005 0.009 0.001 0.53 0.0005 0.01 2.39 0.4 0.01 0.00025 0.004

Potentially acid generating – PAG

Count 38 38 38 38 38 38 38 38 38 38 38 38 38 38 38 38

Minimum 0.0005 0.0005 0.025 0.0001 0.1 0.0005 0.0005 0.0005 0.025 0.0005 0.001 0.025 0.001 0.01 0.00025 0.0005

Maximum 0.005 0.0005 0.08 0.0008 575 0.011 0.36 0.18 289 0.002 0.13 98.3 11.8 0.03 0.0096 0.11

Average 0.0008 0.0005 0.0288 0.0001 49.1 0.0008 0.0202 0.0093 9.3772 0.0006 0.019 9.63 1.16 0.0105 0.0009 0.0075

Median 0.0005 0.0005 0.025 0.0001 8.255 0.0005 0.0005 0.0005 0.145 0.0005 0.0085 2.155 0.0165 0.01 0.00025 0.0005

Non-acid generating – NAG

Count 23 23 23 23 23 23 23 23 23 23 23 23 23 23 23 23

Minimum 0.0005 0.0005 0.025 0.0001 1.53 0.0005 0.0005 0.0005 0.025 0.0005 0.002 0.15 0.001 0.01 0.00025 0.0005

Maximum 0.011 0.0005 0.06 0.0014 573 0.003 0.24 0.22 246 0.0005 0.028 96.4 11.2 0.02 0.0062 0.11

Average 0.0011 0.0005 0.0265 0.0002 73.3 0.0006 0.0134 0.0103 11.2 0.0005 0.010 11.86 0.74 0.0104 0.0008 0.0062

Median 0.0005 0.0005 0.025 0.0001 5.87 0.0005 0.0005 0.0005 0.025 0.0005 0.006 0.89 0.005 0.01 0.00025 0.0005

Ore samples

NSVZ 0.0005 0.0005 0.025 0.0004 22.1 0.0005 0.011 0.001 0.72 0.0005 0.007 6.68 2.5 0.01 0.00025 0.007 (1) One half the detection limit is used for values reported as below the detection limit.



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Table 5.6c Summary of SPLP data from ore and waste rock samples for AG, PAG, and NAG categories (1) PO4 K Se SiO2 Ag Na Sr Te Tl Th Sn Ti U V Zn Zr

mg/L mg/L mg/L mg/L mg/L mg/L mg/L mg/L mg/L mg/L mg/L mg/L mg/L mg/L mg/L mg/L

Detection limit <0.15 <0.001 <0.00025 <0.001 <0.001 <0.0001 <0.0005 <0.001 <0.001 <0.0005 <0.001 <0.005 <0.01

% below detect 92% 0% 97% 0% 99% 0% 1% 100% 91% 99% 100% 87% 82% 53% 61% 100%

Acid generating - AG

Count 17 17 17 17 17 17 17 17 17 17 17 17 17 17 17 17

Minimum 0.075 0.6 0.0005 0.7 0.000125 1.12 0.006 0.0005 0.00005 0.00025 0.0005 0.0005 0.00025 0.0005 0.0025 0.005

Maximum 0.8 13.2 0.004 3.3 0.0006 22.7 0.95 0.0005 0.0054 0.0012 0.0005 0.0005 0.0071 0.13 2.04 0.005

Average 0.1618 3.09 0.0008 1.71 0.000153 5.66 0.273 0.0005 0.000662 0.000306 0.0005 0.0005 0.000938 0.012353 0.208235 0.005

Median 0.075 2.6 0.0005 1.7 0.000125 3.07 0.11 0.0005 0.00005 0.00025 0.0005 0.0005 0.00025 0.0005 0.012 0.005

Potentially acid generating - PAG

Count 38 38 38 38 38 38 38 38 38 38 38 38 38 38 38 38

Minimum 0.075 0.2 0.0005 0.8 0.000125 1.61 0.0005 0.0005 0.00005 0.00025 0.0005 0.0005 0.00025 0.0005 0.0025 0.005

Maximum 0.7 14.3 0.0005 7.4 0.000125 32.1 4.89 0.0005 0.0043 0.00025 0.0005 0.004 0.0029 0.063 0.52 0.005

Average 0.0947 2.17 0.0005 2.65 0.000125 8.31 0.322 0.0005 0.000262 0.00025 0.0005 0.000842 0.000442 0.004553 0.055013 0.005

Median 0.075 1.55 0.0005 2.3 0.000125 6.415 0.0835 0.0005 0.00005 0.00025 0.0005 0.0005 0.00025 0.0005 0.0025 0.005

Non-acid generating - NAG

Count 23 23 23 23 23 23 23 23 23 23 23 23 23 23 23 23

Minimum 0.075 0.4 0.0005 0.3 0.000125 0.15 0.026 0.0005 0.00005 0.00025 0.0005 0.0005 0.00025 0.0005 0.0025 0.005

Maximum 0.2 3.3 0.0005 8.4 0.000125 17.8 1.93 0.0005 0.00005 0.00025 0.0005 0.006 0.0066 0.011 1.93 0.005

Average 0.0804 1.52 0.0005 3.08 0.000125 7.57 0.251 0.0005 0.00005 0.00025 0.0005 0.000848 0.00065 0.003326 0.102717 0.005

Median 0.075 1.4 0.0005 3.2 0.000125 6.19 0.085 0.0005 0.00005 0.00025 0.0005 0.0005 0.00025 0.003 0.0025 0.005

Ore samples

NSVZ 0.075 1.6 0.0005 0.8 0.000125 1.06 0.017 0.0005 0.00005 0.00025 0.0005 0.0005 0.00025 0.0005 0.026 0.005 (1) One half the detection limit is used for values reported as below the detection limit.



45

Figure 5.5 Sulfate and TDS concentrations

0.0

500.0

1000.0

1500.0

2000.0

2500.0

3000.0

3500.0

0 500 1000 1500 2000 2500

TDS (mg/L)

Sulfa

te (m

g/L)

Based upon the SPLP leaching test, metals are generally not very mobile in most of the waste rock samples from Kupol. Between 90% and 100% of the analyses of bismuth, boron, cadmium, chromium, lead, mercury, phosphate, selenium, silver, tellurium, thallium, thorium, tin, and zirconium are below their respective detection limits and 80% of the analyses of beryllium, titanium and uranium are below their respective detection limits. The metals concentrations in the SPLP leachate do not directly compare to water quality standards for various reasons, e.g. the SPLP samples have been crushed and have a very high surface area, the conditions in the test chamber do not accurately reproduce conditions on site, etc. Using the maximum SPLP concentrations for elements that have been mobilized in more than 10% of the analyses, however the following elements are mobile in one or more of the resource model categories of waste rock AG, PAG, or NAG: aluminum, arsenic, beryllium, calcium, cobalt, copper, iron, magnesium, manganese, nickel, silicon, strontium, vanadium, and zinc. 5.5 Humidity cell test Humidity cell tests measure time dependent changes in acidity and metals release rates. Typically two sets of reactions are recorded by humidity cell tests as shown in Figure 5.6. The early chemical load in the humidity cell leachate results from dissolution reactions of readily solubilized salts and mineral acidity. Iron sulfide oxidation does not contribute to the early chemical load due to reaction kinetics. The reservoir of soluble elements that



46

form the early chemical load is typically depleted within the first few pore volumes of flushing. The long term chemical load is derived from oxidation reactions of primary sulfide minerals. Once initiated, iron sulfide oxidation, in the presence of excess oxygen, progresses through the three stages discussed in Section 2.5. Carbonate neutralizing potential, when present, buffers the pH to near neutral until either a) all of the carbonate is consumed (i.e. AGP > NP) or b) all of the acidity is consumed (i.e. NP > AGP). The dissolution reactions and oxidation reactions continue regardless of the presence of neutralizing potential however the high pH reduces the sulfide oxidation rates and promotes precipitation of secondary oxyhydroxide minerals and metal sorption. Figure 5.6 Typical profiles of chemical loading in humidity cell leachate over time

Two sets of humidity cell tests have been conducted, the first tests with samples collected from the 2003 drilling campaign were initiated on 3 March 2004 and the second set of tests with samples collected from the 2004 drilling campaign were begun on 24 December 04.. The humidity cell tests include 32 samples of which 9 are AG samples, 8 are NAG samples, 13 are PAG samples, and 2 are an ore sample. Table 5.7 summarizes the results of the 2003 samples for the week 20 weekly analyses and the results for selected elements of the full chemical analyses. The table also lists the pyritic sulfur concentration for comparison with the measured sulfate concentration in the weekly analyses and the NP:AGP ratios for comparison with the pH and alkalinity measurements from the leachate samples. All of the data is included in Appendix B.



47

Table 5.7 humidity cell test results summary for week 20 ARD code Sample Spy NP NP:AGP pH SO4 alk Fe Al As Sb be mn se Tl cu pb zn hg Wt% T CaCO3/kT su mg/L mg/L mg/L mg/L Mg/L mg/L mg/L mg/L mg/L mg/L mg/L mg/L mg/L mg/L AG HiPyLoCarb ARD - 48 5.04 3 0.02 2.71 586 0 166 23.2 0.16 0.011 0.002 3.44 0.002 0.002 0.11 0.0005 0.98 0.0000HiPyLoCarb ARD - 62 1.86 4.2 0.07 3.50 174 0 13.2 0.85 0.32 0.003 0.003 5.92 0.0005 0.0011 0.047 0.0005 1.01 0.0000HiPyLoCarb ARD - 81 2.00 -2.2 -0.04 3.44 648 0 3.95 2.2 0.0005 0.0005 0.002 26.9 0.001 0.0004 0.01 0.0005 0.23 0.0000HiPyLoCarb ARD - 152 0.06 2 1.07 4.06 51 0 0.025 0.14 0.0005 0.0005 0.0005 0.23 0.0005 0.00005 0.016 0.0005 0.021 0.0000HiPyLoCarb ARD - 129 2.63 -17.2 -0.21 2.34 850 0 214 16.4 0.22 0.0005 0.0005 0.067 0.01 0.0009 0.14 0.0005 0.015 0.0000HiPyLoCarb ARD - 153 4.05 -23 -0.18 2.19 1290 0 297 18.9 0.68 0.0005 0.001 0.075 0.005 0.0002 0.16 0.0005 0.14 0.0000HiPyLoCarb ARD - 157 1.33 -3.2 -0.08 2.89 1155 0 16.9 25.6 0.006 0.0005 0.009 0.76 0.004 0.00005 0.16 0.0005 0.32 0.0000HiPyLoCarb ARD-240 0.1 18 5.8 8.43 1 24 0.025 0.18 0.003 0.002 0.0005 0.005 0.0005 0.00005 0.0005 0.0005 0.0025 0.0000HiPyLoCarb ARD-242 0.02 -1.3 -2.1 3.01 512 #N/A 1.2 21.1 0.033 0.0005 0.007 2.9 0.0005 0.0004 0.1 0.0005 1.01 0.0000NAG LoPyHiCarb ARD - 38 1.30 1.7 0.04 3.53 378 0 0.68 3.68 0.0005 0.0005 0.004 1.27 0.002 0.00005 0.005 0.0005 0.068 0.0000LoPyHiCarb ARD - 77 0.49 134.8 8.81 7.45 4 27 0.025 0.029 0.004 0.012 0.0005 0.002 0.0005 0.00005 0.0005 0.0005 0.0025 0.0000LoPyHiCarb ARD - 92 0.01 133.5 427.20 8.55 3 49 0.025 0.036 0.008 0.006 0.0005 0.004 0.0005 0.00005 0.0005 0.0005 0.0025 0.0000LoPyHiCarb ARD - 128 0.01 54.2 173.44 6.63 6 25 0.025 0.035 0.01 0.01 0.0005 0.003 0.0005 0.00005 0.0005 0.0005 0.0025 0.0000LoPyHiCarb ARD - 163 3.65 -14.6 -0.13 2.73 808 0 73.5 7.25 0.048 0.0005 0.002 1.16 0.001 0.00005 0.066 0.0005 0.12 0.0000LoPyHiCarb ARD - 200 2.16 91.9 1.36 7.41 11 18 0.025 0.03 0.001 0.002 0.0005 0.003 0.0005 0.00005 0.0005 0.0005 0.0025 0.0000LoPyHiCarb ARD - 211 3.18 101.7 1.02 7.80 10 41 0.025 0.023 0.64 0.049 0.0005 0.011 0.0005 0.00005 0.0005 0.0005 0.0025 0.0000LoPyHiCarb ARD-256 0.03 28.7 30.6 8.39 3 18 0.025 0.1 0.001 0.002 0.0005 0.007 0.0005 0.00005 0.0005 0.0005 0.073 0.0000PAG HiPyHiCarb ARD - 96 0.01 45.8 146.56 8.86 1 38 0.025 0.16 0.003 0.0005 0.0005 0.001 0.0005 0.00005 0.0005 0.0005 0.0025 0.0000HiPyHiCarb ARD - 106 0.46 112.9 7.86 7.51 14 36 0.025 0.031 0.0005 0.0005 0.0005 0.062 0.0005 0.00005 0.0005 0.0005 0.0025 0.0000HiPyHiCarb ARD - 205 2.00 83.3 1.33 7.81 18 26 0.025 0.047 0.011 0.006 0.0005 0.003 0.0005 0.00005 0.0005 0.0005 0.0025 0.0000HiPyHiCarb ARD - 208 1.24 134.8 3.48 8.02 20 23 0.025 0.029 0.01 0.003 0.0005 0.007 0.0005 0.00005 0.0005 0.0005 0.0025 0.0000HiPyHiCarb ARD - 213 1.96 5.7 0.09 4.85 96 1 0.36 0.045 0.001 0.0005 0.0005 3.57 0.0005 0.00005 0.004 0.0005 0.052 0.0000HiPyHiCarb ARD-251 1.89 23.5 0.4 9.01 1 47 0.025 0.16 0.004 0.002 0.0005 0.002 0.0005 0.00005 0.001 0.0005 0.0025 0.0000HiPyHiCarb ARD-239 0.14 51.3 11.7 8.49 4 22 0.12 0.21 0.002 0.0005 0.0005 0.012 0.0005 0.00005 0.001 0.0005 0.006 0.0000LoPyLoCarb ARD - 5 0.15 -1.2 -0.25 3.12 292 0 0.36 7.75 0.009 0.001 0.002 0.11 0.0005 0.0002 0.98 0.0005 0.059 0.0000LoPyLoCarb ARD - 10 0.25 113.1 14.69 7.56 73 19 0.025 0.018 0.004 0.0005 0.0005 0.004 0.0005 0.00005 0.0005 0.0005 0.0025 0.0000LoPyLoCarb ARD - 187 0.01 -3.6 -11.52 4.05 33 0 0.025 0.15 0.0005 0.0005 0.0005 0.93 0.0005 0.00005 0.003 0.0005 0.089 0.0000LoPyLoCarb ARD - 198 0.01 7.9 25.28 7.50 2 6 0.025 0.074 0.006 0.0005 0.0005 0.013 0.0005 0.00005 0.0005 0.0005 0.0025 0.0000LoPyLoCarb ARD-244 0.2 16.1 2.6 8.05 1 19 0.07 0.14 0.002 0.0005 0.0005 0.005 0.0005 0.00005 0.0005 0.0005 0.017 0.0000LoPyLoCarb ARD-259 0.9 14.6 0.5 7.78 15 21 0.025 0.04 0.002 0.001 0.0005 0.038 0.0005 0.00005 0.0005 0.0005 0.018 0.0000ORE ore NSVZ 0.07 3.5 1.60 7.76 24 6 0.025 0.0025 0.004 0.0005 0.0005 0.48 0.0005 0.00005 0.0005 0.0005 0.0025 0.0000ore ARD-269 0.18 -1.8 -0.3 2.61 1532 #N/A 17.5 14.1 0.14 0.0005 0.009 22.8 0.003 0.00005 0.17 0.002 1.08 0.0000



48

Weekly data Weekly data includes sulfate, pH and conductivity analyses. The data for the AG, PAG, and NAG samples are shown in Figures 5.7, 5.8, and 5.9 and include pH and sulfate analyses. The pH of the AG samples (refer to Figure 5.7) is low, less than 5. The one exception is sample ARD-242 which remains circum-neutral. Since the pH of the flushing solution is generally around 5, this suggests the lack any neutralizing potential in the AG samples and the presence of some soluble acidity. The pH remains nearly constant or slightly declining because of the lack of neutralizing potential to raise the pH. The sulfate shows a decline from a maximum of 10,000 mg/L for several of the samples with the lowest pH (ARD-153 and ARD-129) to around 1000 mg/L. Sample ARD-242 has correspondingly low sulfate concentrations in the leachate. The presence of non-acid generating AG material attests to the heterogeneity of the ARD categories. The source of the elevated sulfate and low pH in the early analyses may be iron sulfate minerals such as jarosite. The pH of the NAG samples (refer to Figure 5.8) is mixed, although the majority of the samples have a pH around 8. Two samples (ARD-38 and ARD-163) have pH values below 4. Sample ARD 128 shows a sudden decline in pH in week 20, however the alkalinity in the leachate remains at 25 mg/L (refer to Table 5.7) suggesting that the drop in pH is a short term fluctuation rather than a long term trend. The presence of acidic samples within the NAG category is consistent with the heterogeneity of the resource model categories (refer to 5.1.1). The pH of the alkaline samples remains near neutral, and two of the samples with initially low pH (ARD-128 and ARD-200) show a pronounced increase in pH with time. Since the NAG samples are the primary source of neutralizing potential in the tailing facility, it is important that these samples have remained near neutral during the duration of the 20 week test. The sulfate concentrations in the NAG samples are low (generally less than 100 mg/L) and declining. The sulfate concentrations of the two acidic NAG samples are high, due to the presence of sulfide minerals or acid generating sulfate minerals, and declining. The pH of the PAG samples is mixed and often erratic (refer to Figure 5.9). Samples ARD-5 and ARD-187 are acidic (pH below 4), while the pH of samples ARD-96, ARD-205, ARD-208, ARD-106, ARD-239, ARD-251, and ARD-244 is alkaline (pH above 8). The pH of samples ARD-10, ARD-198, and ARD-213 is more erratic, suggesting disequilibrium within the sample between the slightly acidic flushing solution and carbonate neutralizing potential. Sulfate concentrations range from over 1000 to less than 10 mg/L, and all decline with time. The two ore samples (KP03-NSVZ and ARD-269) show distinctly different trends in both pH and sulfate concentrations. KP03-NSVZ, which consists of silica-rich sheeted veins, shows an increasing pH and decreasing sulfate concentration with time. As shown in Table B.3 in Appendix B, this is a low sulfide sample (0.07 wt% Spy). Sample ARD-269, a small vein in the footwall of the ore zone, shows a low pH (around 3) and sulfate



49

concentrations around 1000 mg/L. ARD-269 has slightly higher sulfide sulfur than KP03-NSVZ but a negative NP suggesting 1) that it has undergone some level of oxidation which would have dissolved any neutralizing capacity and 2) will continue to be a source of acidity over time. The long term kinetic tests are consistent with the ARD predictions of the static tests (refer to Table 5.7). Samples with NP:AGP ratios less than 1 are generally acidic with pH values that are below 4 or 5 but can range as high as 8. Samples with NP:AGP ratios greater than 3 are all alkaline with pH between 6.6 and 8.9. The samples with NP:AGP ratios in the uncertain field (between 1 and 3) are all alkaline (pH between 7.4 and 8) except AG sample ARD 152 with pH of 4. The high pH of the samples with NP:AGP ratios between 1 and 3 supports the suggestion that the NP:AGP ratios greater than 1 are, if not acid neutralizing, at least not acid generating. The results of the kinetic test also support the assumption about the acid generating potential of the three ARD categories AG, PAG, and NAG (refer to Section 3.1 and 5.1.1 and Table 5.7). AG samples identified as HiPyLoCarb have pH values generally less than 4 and sulfate concentrations of around 1000 mg/L. NAG samples (LoPyHiCarb) have pH values generally between 7 and 8 and sulfate concentrations generally less than 10 mg/L. The PAG samples including LoPyLoCarb and HiPyHiCarb have near neutral pH and sulfate concentrations generally between 10 and 100 mg/L. Both LoPyLoCarb and HiPyHiCarb include the range of pH and sulfate concentrations of the PAG samples. There is no evidence to suggest that either LoPyLoCarb or HiPyHiCarb are more likely to have either low pH (high sulfate) or high pH (low sulfate).



50

Figure 5.7 Weekly pH and sulfate analyses for AG samples

0.00

2.00

4.00

6.00

8.00

10.00

12.00

0 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20

Weeks

pH (s

u)

ARD-48ARD-62ARD - 81ARD-152ARD-129ARD-153ARD-157ARD-240ARD-242

1

10

100

1000

10000

100000

0 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20

Weeks

Sulfa

te(m

g/L)

ARD-48ARD-62ARD-81ARD-152ARD-129ARD-153ARD-157ARD-240ARD-242



51

Figure 5.8 Weekly pH and sulfate analyses for NAG samples

0.00

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8.00

10.00

12.00

0 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20

Weeks

pH (s

u)

ARD-38ARD-77ARD-92ARD-128ARD-163ARD-200ARD-211ARD-256

1

10

100

1000

10000

100000

0 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20

Weeks

Sulfa

te (

mg/

L)

ARD-38ARD-77ARD-92ARD-128ARD-163ARD-200ARD-211ARD-256



52

Figure 5.9 Weekly pH and sulfate analyses for PAG samples

0.00

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6.00

8.00

10.00

12.00

0 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20

Weeks

pH (s

u)

ARD-96ARD-106ARD-205ARD-208ARD-213ARD-239ARD-251ARD-5ARD-10ARD-187ARD-198ARD-244ARD-259KP03-NSVZARD-269

1

10

100

1000

10000

100000

0 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20

Weeks

Sulfa

te (m

g/L)

ARD-96ARD-106ARD-205ARD-208ARD-213ARD-239ARD-251ARD-5ARD-10ARD-187ARD-198ARD-244ARD-259KP03-NSVZARD-269



53

Full chemical analyses Early-time data from the humidity cell tests is generally used to characterize the potential discharge chemistry during the operational phase of a project. The long term, steady state data is used to characterize the potential discharge water chemistry from a facility after closure. Table 5.7 summaries the result of the week 20 and week 16 analyses for major elements (alkalinity, iron, aluminum, and manganese), trace elements (arsenic, antimony, beryllium, selenium, thallium, and mercury), and base metals (copper, lead, and zinc). Time series plots of concentrations for these elements are also included in Appendix A. The figures include the relevant groundwater and drinking water standards (SanPiN) and in-stream surface water standards (MAC) for reference purposes only. The actual chemistry of any discharge from the tailing facility and pit walls will depend upon many factors not addressed in the humidity cell tests including the water to rock ratio, reaction rates, mineral precipitation, and metal adsorption, among other factors. Carbonate alkalinity (HCO3

-, and CO32-) is present only in leachate with pH above 6.4.

Below 6, the dominant carbonate is H2CO3. Leachate samples for the AG and low pH PAG and NAG categories have zero alkalinity. Iron, aluminum, and manganese concentrations are high (and declining) in the AG samples and generally lower and declining to near or below the reference standards in the neutral pH PAG and NAG samples (refer to Table 5.7). The elevated metals concentrations result from the low pH leachate either reacting with primary sulfides or secondary metal-rich salts. The solubility of each of the metals is pH dependent. At near neutral pH, iron, aluminum, and manganese form oxides or oxyhydroxides that effectively remove these metals from solution. An added benefit of iron and manganese precipitation is that these metal oxides and oxyhydroxides sorb metals such as arsenic, antimony, copper, lead, and zinc from solution. Figure 5.10 is a plot showing the sorptive edges for arsenic, lead, copper, zinc and other metals. It is advantages in a tailing facility to maintain a near neutral pH in order to reduce dissolution but also to promote metals precipitation and sorption. Figure 5.10 Sorptive edges of metals in the presence of a ferrihydrite substrate



54

The trace metal concentrations generally mirror the major cation concentrations and are elevated in the AG samples and other samples with low pH. Minerals such as tetrahedrite-tennenite ((Cu,Fe,Zn,Ag)12(As,Sb)4S13) and Ag-sulfosalts (proustite (Ag3AsS3) or pyrargyrite(Ar3SbS3)) are sources of both arsenic and antimony and the time series leaching patterns are very similar for both metals. The concentrations of beryllium, selenium, and thallium are elevated in the AG samples but decline over time. The concentrations of selenium and thallium decline to near or below the respective reference standards. The reference standards for beryllium are below the detection limits of the analyses. Mercury concentrations are very low, generally around 0.00001 mg/L. The concentrations of base metals copper, lead, and zinc are elevated but declining in AG samples and other samples with low pH. The concentrations of copper and zinc generally fall between the SanPiN and MAC standards while the concentration of lead is below both standards. 5.6 Mineralogy Two sets of mineralogical data are available to characterize the composition of the different lithologic, mineralization, and alteration assemblages in the rocks at Kupol:

• Petrographic analyses of drill hole samples conducted by the exploration group and

• XRD analyses of samples from the ARD test program. Petrographic analyses The petrographic analyses show that the rocks at Kupol range from unaltered and mildly altered (deuteric overprint) andesite, andesitic basalt, and rhyolite to near monomineralic vein assemblages composed of 90% quartz. Table A.3 in Appendix A summaries the results of petrographic analyses of 34 rock samples from different lithologic and alteration assemblages. The mineral names in bold are minerals found in the rock in abundances greater than 5% of the modes from polished sections. The same minerals appear in different “alteration” assemblages, and the only differences arise from the relative abundance of each mineral. Quartz, K-feldspar (either primary igneous orthoclase or hydrothermal adularia), and plagioclase (either primary igneous andesine or hydrothermal albite and anorthoclase) are present in all of the samples, except the quartz veins where plagioclase is absent. Relict mafic minerals are present in the unaltered rhyolite and andesite and are altered to combinations of chlorite, carbonate, quartz, sericite, and other accessory minerals in the altered equivalent rock. Sericite, sericite-clay, carbonate, and pyrite are common alteration or mineralization phases to almost all of the different assemblages. Sulfides and sulfosalts are restricted to the vein assemblage and include: chalcocite (Cu2S), acanthite-argentite (Ag2S), covelite (CuS), Ag-sulfosalts (proustite (Ag3AsS3) or pyrargyrite(Ag3SbS3)), sphalerite ((Zn,Fe)S), and chalcopyrite (CuFeS2).



55

XRD analyses The XRD analyses are qualitative to semi-quantitative estimates of the mineralogy of the sample material and do not indicate the relative abundance of each mineral identified but rather an indication of the presence of each phase. Table 5.8 lists the different minerals identified in the 10 samples analyzed by XRD as well as several parameters from the ABA data set and the results of the humidity cell tests. The mineralogical data has been broken down into several categories:

• The relict primary or secondary silicate minerals, quartz, potassium feldspar, and plagioclase feldspar

• The sulfide and sulfate minerals that are sources of iron and sulfate: pyrite, jarosite, and gypsum

• The carbonate minerals, sources of alkalinity: calcite, dolomite, ankerite, and siderite

• The aluminum silicate minerals, sources of soluble aluminum: chlorite, muscovite (sericite), illite, montmorillinite, and kaolinite

• Miscellaneous minerals: rutile, goethite, magnetite, and gottardiite (?). High Spy (around 2 wt% or greater) is typically associated with measured pyrite. High sulfate concentrations are generally associated with the presence of pyrite, however very high sulfate (greater than 300 mg/L) in the week 20 leachate is also associated with the presence of sulfates (jarosite or gypsum with or without pyrite, refer to ARD-5). Measurable alkalinity in the week 20 leachate is associated with one or more of the carbonate minerals, calcite, dolomite, siderite or ankerite although the highest alkalinity is typically associated with calcite. High leachate iron concentrations are typically associated with low pH and pyrite although jarosite may be an additional source of iron (refer to ARD-153). High aluminum in the leachate is associated with low pH and the likely source of aluminum (in the absence of alunite) is the clays (sericite, illite, montmorillinite, or kaolinite). For pH values around 3 or less, illite appears to be a major source of aluminum in solution, . .



56

Table 1. Results of qualitative phase analysis

Sample ID ARD ARD ARD ARD ARD ARD ARD ARD ARD KPO3

62 153 92 163 211 5 96 205 213 NSVZ

ARD AG AG NAG NAG NAG PAG PAG PAG PAG ORE

Chemistry (week 20)

Pyrite (wt%) 1.86 4.05 0.01 3.65 3.18 0.15 0.01 2 1.96 0.07 pH 3.5 2.09 8.55 2.73 7.8 3.12 8.86 7.81 4.85 7.5 Alk 0 0 49 0 41 0 38 26 1 6 SO4 174 1290 3 808 10 292 1 18 96 2 Fe 3.2 297 0.03 73.5 0.03 0.36 0.03 0.03 0.36 0.025 Al 0.85 18.9 0.04 7.25 0.02 7.75 0.02 0.05 0.05 0.074

Mineralogy Quartz SiO2 X X X X X X X X X

K-feldspar KAlSi3O8 X X X X X

Plagioclase NaAlSi3O8 – CaAl2Si2O8 X X X X X

Pyrite FeS2 X X X X X X

Jarosite K2Fe63+(SO4)4(OH)12 X X X

Gypsum CaSO4·2H2O X X X X

Calcite CaCO3 X X X

Dolomite CaMg(CO3)2 X

Siderite Fe2+CO3 X X

Ankerite Ca(Fe2+,Mg,Mn)(CO3)2 X

Chlorite (Mg,Fe2+)5Al(Si3Al)O10(OH)8 X X X X X

Muscovite KAl2AlSi3O10(OH)2 X X X X X

Illite K0.65Al2Al0.65Si3.35O10(OH)2 X X X X X X

Montmorillonite (Na,Ca)0.3(Al,Mg)2Si4O10(OH)2·nH2O X

Kaolinite Al2Si2O5(OH)4 X X

Rutile TiO2 X Goethite α –Fe3+O(OH) X

Magnetite Fe3O4 X

Gottardiite Na3Mg3Ca5[Al19Si117O272]·93H2O ?



57

5.7 Grain size analyses The grain size analyses of the humidity cell test material are used to scale up the results of the test to fit each of the facilities modeled (refer to Section 4.2.7). Figure 5.11 is a figure showing the percent passing for a range of screen sizes for test material used in the different humidity cells. The MEND humidity cell specifications require that 80% of the sample material be less than ¼ inch (6.35 mm). Figure 5.11 suggest that the test material ranges from between 64 and 97 % less than 6.30 mm. Sample ARD-NSVZ was crushed for metallurgical testing and is significantly finer grained than the other humidity cell charges (98% less than 6.3 mm). The grain size analyses shown by dashed lines are for samples collected from the 2004 drilling campaign. The grain size analyses are included in Appendix B. Figure 5.11 Grain size analyses for the humidity cell test material

0.00

20.00

40.00

60.00

80.00

100.00

120.00

<0.25 0.250 0.500 1.00 2.00 4.00 6.30 9.50

Screen size (mm)

Perc

ent p

assi

ng

ARD-5ARD-10ARD-38ARD - 48ARD-62ARD-77ARD-81ARD-92ARD-96ARD-106ARD-128ARD-129ARD-152ARD-153ARD-157ARD-163ARD-187ARD-198ARD-200ARD-205ARD-208ARD-211ARD-213KP03-NSVZARD - 239ARD - 240ARD - 242ARD - 244ARD - 251ARD - 256ARD 259

5.8 QA/QC The quality control program includes standard laboratory duplicates and blanks as well as duplicates of SPLP and humidity cell leachate samples. Results of the ABA analyses are shown graphically in Appendix A, which are plots of neutralizing capacity and sulfur species versus the respective duplicate analyses. The sulfate and total sulfur analyses and



58

duplicate analyses match very closely. The sulfide sulfur and neutralizing potential are slightly higher in the duplicate analyses when compared to the original analyses but within the error range of ±20%. Results of the ICP metals analyses also show good reproducibility with duplicate analyses falling within the range of ±20% error. Results of the SPLP test generally meet the QA/QC criteria established by Standard Methods for anion-cation balance. The exceptions are limited to analyses that are very dilute with TDS values generally less than 100 mg/L (anion charge less the 3 meq/L) where minor analytical errors can result in a disproportionately higher charge imbalance. The duplicate analyses show good reproducibility and generally fall within ±20%, except for manganese which shows greater variation. Duplicate analyses for the humidity cell tests were conducted using a) six leachate water samples from the monthly analyses for two monthly sampling suites (weeks 16 and 20) for the 2003 samples and b) one leachate water sample from the five monthly analyses (weeks 0, 4, 8, 12, and 16). Plots of metals concentrations from the original analyses and the duplicate analyses are shown in Appendix A. The agreement between duplicate analyses is very good.



59

6.0 Conclusions Preliminary results of the geochemical test program at Kupol indicate that the waste rock to be delivered to the tailing facility and exposed in the pit walls is sulfide-bearing and a potential source of ARD/ML. A capacity to predict the ARD and ML potential of the waste rock at Kupol is essential to facilitate pro-active mitigation planning for the periods encompassing the mine planning and development phase, the operational phase, and closure phase of the project life. This section addresses:

• The validity and applicability of the sampling and testing program • The results of the tests and application to the current mine plan • Recommendations for future work • Mitigation measures within a geochemical framework

An evaluation of the test program by Mr. Logsdon, the third party reviewer, is included in Appendix E. 6.1 Geochemical test program The sampling and geochemical testing program at the Kupol Project adequately addresses site-specific issues for ARD and ML based upon the current understanding of the mineralization and altered country rock adjacent to the vein system. The sampling plan follows best industry practices in the number and criteria for sample selection. Waste rock sampling intervals were identified from each of the rock, alteration, and mineralization types used in the exploration drill hole geological logs and representative samples from each of the identified zones were collected by site geologists. Ore samples used in the geochemical test work were collected by site geologist and metallurgists for metallurgical test work. Within the limits of availability of waste rock material to sample, the sample selection process included samples from all areas of the South, Big Bend, Central, and Northern exploration zones and from all depths. The test program followed the best industry practices for sulfide-bearing rock materials including a screening level suite of standard static testing and progressing through to more long term kinetic test work. All of the 276 waste rock samples from both the 2003 and 2004 drilling campaigns and the majority of the ore samples underwent ABA and partial digestion and multi-element ICP analysis for initial screening. Results of the screening level investigations were used to select samples from the 2003 drilling campaign with a representative range of acid generating potential and whole rock chemistry for more detailed analyses including one in three samples for SPLP testing and one in ten samples for NAG testing. The results of the static tests were used to identify a range of representative samples for kinetic testing. The tests provide critical information concerning the acid generating potential of the waste rock at Kupol, the metal availability for leaching from the waste rock, and the leaching characteristics of the different waste rock types under laboratory conditions. The kinetic tests in conjunction with the tailing facility engineering design and results of



60

the tailing facility hydrological and thermal investigations will provide the basis for prediction of the chemistry of any water discharging from the facility. The kinetic tests results in conjunction with the final pit design and the results of a site-wide hydrological investigation, will be used to evaluate pit wall runoff chemistry and the chemistry of any standing water within the mine void following cessation of surface mining operations. 6.2 Test results Results of the geochemical test work have been applied to the three waste rock categories identified in the resource model: acid generating (AG), potentially acid generating (PAG), and non-acid generating (NAG). The categories of the resource model are based upon boundaries in the geologic model constructed using pyrite mineralization codes and carbonate alteration codes from the detailed geologic drill hole logs. The selection of the individual codes used for the boundaries was based upon the results of the ABA test work, where weight percent total sulfur (Stot) and NP were applied to the individual code designations. Results of the ABA tests have shown that all of the categories of waste rock at Kupol are sulfur-bearing and the dominant sulfur species in the waste rock is sulfide sulfur (Spy). Stot and Spy range from below the detection limit to 6 weight percent. The Stot and Spy concentrations decrease from AG through NAG waste rock. At the same time, all of the waste rock categories in the resource model contain some degree of acid neutralizing potential in the form of either carbonate veins or pervasive carbonate alteration. The NP ranges from values less then 0 (negative NP indicates the presence of natural acidity in the rock) to almost 200 kg CaCO3/tonne. The NP increases from the AG to the NAP waste rock. Using industry standard criteria of NP:AGP ratio for identification for the ARD potential based upon results the ABA test work indicates:

• AG is acid generating (NP/AGP is generally less than 1) • NAG is acid neutralizing (NP/AGP greater than 3) • PAG is within the range of acid neutralizing to uncertain and potentially acid

generating. Owing to the uncertainty in logging codes, each of the ARD material types contains some portion of acid generating and acid neutralizing material types. . Similar results are shown using the NAGpH and “NAG” values from the Net Acid Generation test work.

• AG is acid generating (NAGpH less than 4.5, and NAG greater than 5 kg H2SO4/tonne)

• NAG is non-acid generating (NAGpH greater than 4.5, and NAG generally less than 5 kg H2SO4/tonne)



61

• PAG is within the range of non-acid generating and potentially acid generating. The ore samples have lower Stot (below detection to 4 weight percent) than some of the waste rock types and roughly equal proportions of sulfate sulfur (SSO4) and Spy. NP in the ore samplers ranges from below zero to 115 kg CaCO3/tonne and the samples fall within the range of field of acid generating material (NP/AGP less than 1). The ICP test results provide an indication of metals that are present in anomalous concentrations. Comparison of the results of the strong acid digestion and ICP analysis of the waste rock samples indicates that silver, arsenic, gold, bismuth, mercury, antimony, selenium, and thallium occur in AG, PAG, and NAG samples in concentrations above normal basalts. Silver, arsenic, gold, bismuth, cadmium, copper, mercury, molybdenum, lead, antimony, selenium, thallium, and zinc are anomalously high in the ore samples. The SPLP test provides an indication of elemental solubility in the waste rock under a set of standard laboratory conditions. The SPLP leachate has a range in pH of 3 to almost 10 and shows in increase in pH from AG samples (average 4 and median of 6) to NAG samples (average 5 and median 9). Alkalinity in the SPLP leachate also increases from an average of less than 10 mg CaCO3/L in the leachate from AG samples to over 20 mg CaO3/L in the leachate from the NAG samples. The chemistry of all of the SPLP leachate AG, PAG, and NAG samples is dominated by sulfate, which ranges from 1 to 2000 mg/L. Generally the metal concentrations in the SPLP leachate are low. Using the maximum concentrations of soluble metals (i.e. greater than 10% of the analyses above the respective detection limits) in the SPLP leachate: aluminum, arsenic, beryllium, calcium, cobalt, copper, iron, magnesium, manganese, nickel, silicon, strontium, vanadium, and zinc appear to be mobile in slightly acidic solution. The humidity cell tests provide the key parameters to the understanding of the geochemical behavior of the waste rock at Kupol, namely the long term leaching characteristics of the different material types. The critical issue for the humidity cell tests is the durability of the carbonate alkalinity in the NAG and PAG samples over time. Results indicate that the pH of the leachate derived from a) AG waste rock samples is typically low, below 4, b) NAG waste rock samples is mixed but dominantly alkaline, above 7, and c) PAG samples is variable but above 7. Humidity cell leachate from AG samples typically has high sulfate, zero alkalinity, and high metal concentrations, although the sulfate and metals concentrations typically decline with time. The humidity cell leachate from NAG samples typically has low sulfate, moderate alkalinity, and low metal concentrations. Humidity cell leachate from the PAG samples is variable with low alkalinity, generally low sulfate, and low metals concentrations. Fifty-one or 18%, of the waste rock samples in the geochemical data set are AG while 71, or 26 %, are NAG (refer to Table 6.1). The balance of the samples (154, or 56%, of the samples) is PAG. The majority of the AG and NAG tonnage lies in the hanging wall, where a large plume of pyrite mineralization and carbonate alteration extends away for the vein. Table 6.1 shows the number of samples by NAG, PAG, and AG domains and



62

the proportions of non-acid generating, potentially acid generating or uncertain, and acid generating material in each of the NAG, PAG, and AG domains. Table 6.1 Proportions of acid generating, uncertain, and acid neutralizing waste rock in the ARD domains from waste rock samples of the ARD database NP:AGP (1) >3 <3 and >1 <1 ARD domain Count Acid neutralizing Uncertain Acid generating AG 51 21.6% 9.8% 68.6% PAG 154 53.9% 11.7% 34.4% NAG 71 71.8% 11.3% 16.9% (1) after SRK, 1992 Table 6.2 shows the model predicted tonnes of the waste rock to be mined from the open pit and underground using 1) the tonnes of each of the ARD domains from the resource model and 2) the proportions of acid neutralizing, potentially acid generating or uncertain, and acid generating waste within each of the domains shown in the Table 6.1 above.

Table 2 Model predicted tonnes of waste rock mined from the open pit and underground and not used for back fill. NP:AGP 2006 2007 2008 2009 2010 2011 2012 2013 2014 Total






The locations of the high pyrite mineralization and high carbonate alteration zones for the 400, 450, 500, and 500 m elevations in the hanging wall at Kupol are shown in Appendix D. Neither zone is exposed at the surface. The figures show the absence of significant pyrite mineralization and carbonate alteration in the footwall. 6.3 Recommendations for future test work The current sampling and test programs are comprehensive and adequate for this level of investigation. As noted in Section 4.3, additional sources of information about the distribution of the different ARD model codes within the mine void will become available from:

• Additional samples may be collected and tested from exploration drill holes that penetrate the pit to better aid the model prediction of the abundance of AG, PAG, NAG material types within the proposed mine void.

• The operational monitoring ARD program (refer to Section 6.4) will be conducted

during mining to identify waste rock material types for special handling.



63

• The geochemical database will be kept current and will be compiled into a readily

accessible database format (e.g. Microsoft Access). 6.4 Mitigation options for waste rock disposal based upon geochemical considerations. Based upon the current understanding of the hydrogeologic conditions at Kupol, it is likely that the bedrock mined during operations will be from the zone of permafrost and will be saturated and frozen (refer to Section 2.2). It is also likely that material mined and used in construction or placed in the tailings facility, particularly during winter operations, will remain below 0 ºC after placement. It is also likely, however, that the waste rock in the pile will not be saturated, particularly in the lower portions of the early waste rock piles, at least during operations and prior to submergence. Based upon the last assumption, it is likely that during operations and into the early post closure period:

• The access to air circulating through the unsaturated waste rock will expose sulfide minerals to oxidation, although the expected low temperatures within the pile will reduce significantly the oxidation reaction rates.

• Percolation of infiltrating water through the unsaturated rock will have the potential to dissolve and carry ARD and dissolved metals, but the flux of liquid water through the entire facility may be limited by the low T of the waste rock pile.

The objectives of the ARD/ML mitigation are:

• To minimize ARD production within the tailing facility and to neutralize any potential ARD using NAG waste rock.

• To minimize infiltration of water into and through the waste rock and air flow through the pile.

• Monitor the performance of the facility during operation and into the early post-closure time period.

The following recommendations should be incorporated in the Waste Rock Management Feasibility Design report (AMEC, 2005) that should address the overall environmental objectives, facility engineering design criteria, waste rock classification and identification methods, material handling objectives and methods, closure plan objectives and design, and a detailed facility monitoring plan. 1) Design and construct diversion structures above the tailing facility to prevent uncontrolled upgradient runon from reaching the tailing facility 2) Design the facility for maximum runoff after closure. 3) Use “Thin Lift” construction techniques for where waste rock is used in construction. Recent research has shown that low temperature significantly reduces chemical reaction rates. “Thin lift” construction, whereby 2 to 3 m lifts are used, promotes freezing within



64

the pile, and whether or not saturated, will significantly reduce oxidation reaction rates. If the waste rock is dozed during construction, incorporating fine layers or caps within the pile may reduce communication through the pile. Fine grain material (e.g. silt size or finer) often has higher pore water contents and may reach saturation, which, upon freezing, may provide internal barriers to air and water flow through the facility. 4) Monitor waste rock material types (AG, PAG, and NAG) during operations for special handling and placement. This includes sampling and analyzing blast hole cuttings for indicators of potential ARD generation at the same time the cuttings are assayed for gold and silver grades. The results of the ARD tests must be available with the assay grades for identification of waste rock ARD potential for flagging after the blast and prior to mining. Criteria typically used in operating mines in identify ARD potential of waste rock include:

• Total sulfur and carbonate analyses based upon Leco furnace carbon and sulfur analyzer test results. Stot and NP analyses for the 276 samples have shown that total sulfur and carbonate concentrations combined are practical measures of ARD potential in waste rock from Kupol. Based upon the NP:AGP ratio, the current data shows that:

o 100% of the material above a cutoff of 3 wt% Stot have NP:AGP ratios less than 1, within the field of acid generating waste rock while

o 80% of the waste rock below a cutoff of 1 wt% Stot have NP:AGP ratios above 3, within the field of acid neutralizing waste. Combining the sulfur and carbonate concentrations measurements would optimize selection of non-acid generating and acid generating waste during operations.

• “NAG” values based upon Net Acid Generation Tests either in the field or in the assay laboratory. NAGpH values from the current ARD database suggest similar cutoff values to identify acid generating and acid neutralizing waste in the field. Using the NAGpH values and NP:AGP ratios from the ARD samples shows that:

o 100% of the samples with NAGpH values less than 3 are acid generating while

o 100% of the samples with NAGpH values greater than 5 are acid neutralizing.

5) Use selective placement of a) acid generating and b) non-acid generating waste rock material types. Several suggestions have been presented for waste rock placement to minimize potential environmental impacts depending upon the final disposal plan:

• Selective placement of acid generating material away from surface water infiltration, air circulation, and within the coldest portion of any waste rock pile to minimize sulfide oxidation. This design uses non-acid generating material for cover during construction and at closure. Non-acid generating material on the surface of the pile will minimize ARD runoff during and after construction. The non-acid generating cover needs to be 2 times the local active-layer thickness to ensure adequate performance after settlement and freeze-thaw cycles have occurred. Empirical evidence from natural well drained tundra (refer to Section



65

2.2) and using talus patches as an in-situ analogue for waste rock, one would probably estimate the active zone at 2.5 m or perhaps a bit more.

• Subaqueous placement of AG and acid generating and potentially acid generating PAG material in the tailing facility (AMEC, personal communication, 2005). The advantages of the subaqueous placement design are:

o Sequestering the acid generating waste in a zero discharge facility. o Submerging the sulfide-bearing waste to prevent subaerial oxidation of sulfides

and, over the long term, allowing for the potential of further isolation of the sulfide-bearing waste should the water-saturated tailings freeze creating an impermeable ice mass.

6) Incorporate a regular monitoring program beginning during construction and following through operations to identify potential sources of ARD within the tailing facility:

• to control any potential discharges from the facility • to identify possible problems and, when possible, recommend changes to the

operational procedures • to identify any changes required in the facility design or closure plan.

Monitoring devices may include down gradient surface sites and monitor wells as well as thermisters within the waste rock and within the tailing dam. In addition, a test pile of waste rock constructed early in the construction phase of the mine with regular monitoring and analyses of discharge may provide empirical data for predicting the long term discharge water chemistry.



66

Bibliography AMEX, 2005, Report of Waste Rock Management Feasibility Design, Kupol Project, Chukotka Autonomous Okrug, Far East Russia, AMEC Earth & Environmental May 2005. AMEX, 2005, Tailings Facility Feasibility Design Report, Kupol Project, Chukotka Autonomous Okrug, Far East Russia, AMEC Earth & Environmental May 2005. Garagan, T. and MacKinnon, H, 2003, Technical Report, Kupol Project, Chukotka A.O. Russian Federation, Bema Gold Corporation, submitted to Sedar. Hutchison, I. P. G. and Ellison, R. D., 1992, Mine Ware Management, prepared for California Mining Association, Lewis Publishers. Miller, S. D., 1998, Static Net Acid Generation Test Procedure (NAG Test), unpublished manuscript Morin, K. A., 1990, Problems and Proposed Solutions in Predicting Acid Drainage with Acid-Base Accounting, in Acid Mine Drainage Design for Closure, BAC/MAC Joint Annual Meeting, Vancouver. Rhys, D.A., 2004, Structural Study of the Kupol Deposit, Chukotka Autonomous Region, Eastern Russia, unpublished company report. Rice, W. A., 1997, Draft Guidelines and Recommended Methods for the Prediction of Metal Leaching and Acid Rock Drainage at Minesites in British Columbia, prepared for the British Columbia Reclamation Section, Energy and Minerals Division, Ministry of Employment and Investment. Sobek, A.A., Schuller, W.A., Freeman, J.R. and Smith, R.M. (1978), Field and Laboratory Methods Applicable to Overburden and Minesoils, Report EPA-600/2-78-054, U.S. National Technical Information Service Report PB-280 495. Steffen Robertson and Kirsten (SRK), 1989, Draft Acid Rock Drainage Technical Guide, prepared for the British Columbia Acid Mine Drainage Task Force Report. Steffen Robertson and Kirsten (SRK), 1992, Mine Rock Guidelines Design and Control of Drainage Water Quality, Report NO. 93301, prepared for Saskatchewan Environment and Public Safety, Mines Pollution Control Branch, Prince Albert, Saskatchewan, Canada Vartanyan et. al., 2001, unpublished company report.



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BEMA GOLD CORP.KUPOL PROJECT

Vertical cross section-looking North2004 Geological Interpretation

ARD Sample locations

Section 90972.00 N

Scale 1:1000 Date: 11/05/05 User ID: T. McK.

7700

0 E

77000 E

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77050 E

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4-15

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KP

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64

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KP

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76

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6

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TP90

970

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Drillhole- Lithocode (Column and Left Text)

0 10 20 30 40 501:1000

BEMA GOLD CORP.KUPOL PROJECT

Vertical cross section-looking North2004 Geological Interpretation

ARD Sample locations

Section 91090.00 N

Scale 1:1000 Date: 11/05/05 User ID: T. McK.

7705

0 E

77050 E

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7725

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ARD-36

ARD-37

ARD-67

ARD-68

ARD-69

ARD-70

ARD-71

7705

0 E

77050 E

7710

0 E

77100 E

7715

0 E

77150 E

7720

0 E

77200 E

7725

0 E

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TP91090.1 TP91090.1TP91090.2 TP91090.2

TP91100. TP91100.2

BASALTDYKEFAULTSTOCKVEIN

Geology SlicesAndesite (10-13) Basalt (14)Int-Maf Volcanic (15-19) Int-Maf Fragmental (20-29)Clastic Sediments (30-39) Felsic Volcanic (40-44)Felsic Fragmental (45-49) Intermediate Dyke (52)Porphyritic Dyke (53) Dacite Dyke (54)Basalt Dyke (55) Dyke Margin (56)Altered Dyke (59) Intrusive (60-69)Fault (70-74) Surficial (80)Vein (90) Vein-Banded (91)Vein-Breccia (92-93) Stockwork (94-95)Wallrock Breccia (96) Silica Breccia (97)

Drillhole- Lithocode (Column and Left Text)

0 10 20 30 40 501:1000

Appendix C AMEC Tailings Report


AMEC Earth & Environmental a division of AMEC Americas Limited 2227 Douglas Road, Burnaby, BC Canada V5C 5A9 Tel +1 (604) 294-3811 Fax +1 (604) 294-4664 www.amec.com

Tailings Facility Feasibility Design Report

Kupol Gold Project Chukotka Autonomous Okrug, Far East Russia

Submitted to:

Bema Gold Corporation Vancouver, BC

Submitted by:

AMEC Earth & Environmental, a division of AMEC Americas Limited

Burnaby, BC

03 June 2005

AMEC File: VM00329

Bema Gold Corporation Tailings Facility Feasibility Design Report Kupol Gold Project, Far East Russia 03 June 2005

AMEC File: VM00329 S:\PROJECTS\VM00329 - Kupol Gold Tailings & Geotech\Tailings Dam Design Task 230\Report\FINAL REPORT\Tailings Facility Feasibility Design Report_No Figures.docPage i

TABLE OF CONTENTS

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SUMMARY .................................................................................................................................1

1.0 INTRODUCTION................................................................................................................8

2.0 SITE DESCRIPTION..........................................................................................................9 2.1 General......................................................................................................................9 2.2 Climate ......................................................................................................................9 2.3 Baseline Hydrology..................................................................................................11

2.3.1 General ........................................................................................................11 2.3.2 Precipitation .................................................................................................12 2.3.3 Temperature ................................................................................................13 2.3.4 Evaporation..................................................................................................13 2.3.5 Streamflows .................................................................................................14

2.4 Surficial Geology .....................................................................................................14 2.5 Permafrost Setting ...................................................................................................15 2.6 Seismicity ................................................................................................................15 2.7 Alternative Technologies..........................................................................................16 2.8 Alternative Sites.......................................................................................................17

3.0 SITE INVESTIGATIONS ..................................................................................................19 3.1 Geotechnical Drilling................................................................................................19

3.1.1 Program/Findings.........................................................................................19 3.2 Mapping...................................................................................................................19

4.0 DESIGN CRITERIA..........................................................................................................20 4.1 Tailings Material ......................................................................................................20 4.2 Tailings Production Schedule ..................................................................................21 4.3 Tailings Density and Volume ...................................................................................22 4.4 Waste Rock .............................................................................................................23

5.0 DESIGN OF TAILINGS FACILITIES.................................................................................25 5.1 Design Criteria.........................................................................................................25 5.2 Dam Design.............................................................................................................26

5.2.1 Design Section.............................................................................................26 5.2.2 Foundation Preparation................................................................................27 5.2.3 Cutoff Trench Preparation/Seepage Control.................................................27

5.3 Water and Materials Balance...................................................................................28 5.4 Tailings Management ..............................................................................................29 5.5 Waste Rock Management........................................................................................31 5.6 Seepage Interception and Recovery........................................................................31 5.7 Diversion Ditches.....................................................................................................32 5.8 Spillway ...................................................................................................................32

6.0 ALTERNATIVE TAILINGS SITES ....................................................................................34 6.1 Alternative #1...........................................................................................................34


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TABLE OF CONTENTS

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6.1.1 Dam Siting ...................................................................................................34 6.1.2 Volume-Elevation Relationship.....................................................................34 6.1.3 Water and Materials Balance .......................................................................35 6.1.4 Diversion Ditches .........................................................................................36 6.1.5 Design Analyses ..........................................................................................37

6.1.5.1 Thermal Analyses ..........................................................................37 6.1.5.2 Limit Equilibrium Stability Analyses................................................39 6.1.5.3 Seismic Loading Conditions...........................................................40 6.1.5.4 Seepage Analyses.........................................................................41

6.2 Alternative #2...........................................................................................................41 6.2.1 Dam Siting ...................................................................................................41 6.2.2 Volume-Elevation Relationship.....................................................................41 6.2.3 Water and Materials Balance .......................................................................42 6.2.4 Diversion Ditches .........................................................................................43 6.2.5 Design Analyses ..........................................................................................43

6.2.5.1 Thermal Analyses ..........................................................................43 6.2.5.2 Limit Equilibrium Stability Analyses................................................44 6.2.5.3 Seepage Analyses.........................................................................45

7.0 TAILINGS DAM SCHEDULING AND CONSTRUCTION..................................................46 7.1 Starter Dam .............................................................................................................46

7.1.1 Schedule......................................................................................................46 7.1.2 Foundation Preparation................................................................................48 7.1.3 Temporary Collection Pond..........................................................................48 7.1.4 Construction Diversion .................................................................................48 7.1.5 Fill Materials – Borrow Sources, Specification and Placement .....................48

7.1.5.1 Bituminous Geomembrane ............................................................49 7.1.5.2 Zone 1 – Rockfill Embankment ......................................................49 7.1.5.3 Zone 2 – Transition Zone...............................................................49 7.1.5.4 Zone 3 – Bedding Layer.................................................................49 7.1.5.5 Zone 4 – General Fine Grained Fill ................................................50

7.2 Dam Raising ............................................................................................................50 7.3 Construction Quality Control and Quality Assurance................................................50

8.0 MONITORING AND INSTRUMENTATION.......................................................................53

9.0 COST ESTIMATE ............................................................................................................54

10.0 CLOSURE PLAN..............................................................................................................57 10.1 Closure Objectives...................................................................................................57 10.2 Closure Strategy......................................................................................................58 10.3 Closure Monitoring...................................................................................................60

11.0 FURTHER SITE INVESTIGATIONS REQUIRED.............................................................61

12.0 LIMITATIONS AND CLOSURE ........................................................................................62


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TABLE OF CONTENTS

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REFERENCES .........................................................................................................................63

LIST OF FIGURES

Figure 2.1 - Regional Climate Stations......................................................................................10 Figure 2.2 - Comparison of Precipitation ...................................................................................12 Figure 2.3 - Comparison of Temperature ..................................................................................13 Figure 6.1 - Alternative #1 Tailings Valley .................................................................................34 Figure 6.2 - Tailings Alternative #1 – Storage Elevation Curve..................................................35 Figure 6.3 - Starter Dam Thermal Status June Year – 1 (2007) ................................................38 Figure 6.4 - Starter Dam Thermal Status October – Year – 1(2007)..........................................38 Figure 6.5 - Tailings Dam Thermal Status October – Year – 1(2007) ........................................39 Figure 6.6 - Stability analysis section – Alternative #1 – Tailings Dam ......................................40 Figure 6.7 - Alternative # 2 Tailings Valley ................................................................................41 Figure 6.8 - Tailings Alternative #2 – Storage Elevation Curve..................................................42 Figure 7.1 - Preliminary Construction Schedule – Tailings Impoundment..................................47 Figure 9.1 - 3 Year Starter Dam Quantity and Cost Estimate Alternative #1..............................55 Figure 9.2 - 3 Year Starter Dam Quantity and Cost Estimate Alternative #2..............................56

LIST OF TABLES

Table 2.1 - Monthly Runoff Coefficients Kupol...........................................................................14 Table 4.1 - Typical Kupol Tailings Gradation Analysis...............................................................20 Table 4.2: - Kupol Gold Project Mine Production Schedule (tonnage) .......................................22 Table 4.3 - Kupol Gold Project Mine Production Schedule (volume) .........................................23 Table 4.4 - Total Waste Rock Tonnages ...................................................................................23 Table 4.5 - Waste Rock Tonnages and Quantities by ARD Classification .................................24 Table 6.1 - Alternative #1 Required Dam Crest Elevation .........................................................34 Table 6.2: - Diversion Ditch Dimensions....................................................................................36 Table 6.3 - Summary of material parameters used in the stability analysis................................40 Table 6.4 - Summary of stability analysis results for the Main Dam...........................................40 Table 6.5 - Alternative #2 Required Dam Crest Elevation .........................................................41 Table 6.6 - Diversion Ditch Dimensions.....................................................................................43 Table 7.1 - 3 Year Starter Dam Rockfill Quantities Alternative #1 and #2..................................50 Table 7.2 - QA/QC Requirements .............................................................................................51

LIST OF APPENDICES

Appendix A Technical Information Sheet ES3 Bituminous Liner Appendix B Water Balance – Alternatives #1 and #2 Appendix C Thermal Analysis Report


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IMPORTANT NOTICE

This report was prepared exclusively for Bema Gold Corporation by AMEC Earth & Environmental, a wholly owned subsidiary of AMEC Americas Limited. The quality of information, conclusions and estimates contained herein is consistent with the level of effort involved in AMEC services and based on: i) information available at the time of preparation, ii) data supplied by outside sources, and iii) the assumptions, conditions and qualifications set forth in this report. This report is intended to be used by Bema Gold Corporation only, subject to the terms and conditions of its contract with AMEC. Any other use of, or reliance on, this report by any third party is at that party’s sole risk.


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SUMMARY

This report presents the feasibility design of tailings management facility for the Kupol Gold Project, in Chukotka Autonomous Okrug, Far East Russia. The design is presented to a level to support a project feasibility study being prepared by Bema Gold Corporation (Bema). Two alternative methodologies were considered for tailings disposal for the Kupol Project:

• A filtered, dry stack tailings storage system • A conventional slurry tailings system with tailings impounded by a tailings dam

Design and economic analyses have been completed by Bema that indicates the dry stack system to be overly costly, in terms of both capital and operating cost. As a result, the dry stack system is no longer being considered for the project. A slurry tailings system has been selected as the most appropriate for the Kupol Project. In addition to tailings storage, the impoundment will also provide permanent submerged storage of mine waste rock that is classified as acid generating. Site selection studies were undertaken in 2003 and 2004 to identify potential tailings impoundment sites. This report presents feasibility design for two alternative sites that have been identified. Alternative #1 is located southwest of the plant site in the valley of a right tributary of the Middle Kaiemraveem River. Alternative #2 is located immediately south of the plant site on the southern part of the Granichnaya Mountain, in the valley of the Tretyi Creek. Additional exploration drilling has to be completed to confirm that the Alternative #2 tailings site does not conflict with economic mineralization. Hence, both alternative sites are being carried through the feasibility stage until the condemnation drilling is completed to confirm that Alternative #2 can be used for tailings storage. The Kupol Mine Site is located slightly north of the Arctic Circle, in a very cold environment. The climate is characterized by extremely severe weather consisting of long, cold winters (8-8.5 months), overcast weather, and short summer periods (2.5 months). The site is underlain by permafrost to depths in excess of 300 m, with the depth of the active permafrost layer about 1.2 m. Hydrologic analyses have been carried out based on data from two years of site meteorological monitoring, correlated to data from weather stations at Ilirney and Ostrovnoje, provided by Bema Gold, and data from Markovo Station obtained form the World Meteorological Observatory. Ilirney is the closest station to the site, located approximately 80 km west of the Kupol mine site, and is considered to be representative of the site. Data from Markovo was found to provide a better correlation to the site data and has been extrapolated for use as the design parameters for this report. The Markovo data indicates that the annual amount of precipitation at the site is in the order of 380 mm, while the Ilirney data shows about 278 mm/year. The Markovo data was analyzed statistically to derive peak flow rates for design of diversions and other water handling structures. In 2004 JSC Dalstroyizyskania (Dalstroy) carried out a geotechnical site investigation at the Kupol site. Investigations included 37 boreholes for Alternative #1, 9 boreholes for the



proposed Dry Stack site, and 3 boreholes for Alternative #2, for a total of 49 boreholes. The drilling was carried out by dry coring. The tailings will be typical, finely-ground gold mine tailings, with 100% finer than 150 microns. Gold will be recovered using a cyanide leach. Following gold recovery, the tailings will pass through a cyanide destruction circuit to reduce the total cyanide concentration to a typical level of about 5 mg/L. Tailings will be transported to the pond in a tailings pipeline as slurry, with a slurry density of 50% solids by weight. The feasibility study shows 7.1 million tonnes of material to be processed through the mill during a seven year mine life. The criteria provided by Bema was to develop a facility capable of holding 12 million tonnes of waste. This extra capacity will be used to store material not included in the feasibility study (inferred) or material discovered in the future. The Kupol Gold Project is scheduled to have a nominal plant throughput of 3000 tonnes of tailings per day (tpd) during Years 1 through 7 (open pit phase years 1 through 3 and 4 years of underground operations). The rate of processing beyond the feasibility study is assumed to continue underground at a rate of 2000 tpd to the final 12 million tonnes for a mine life of about 13.5 years. For design purposes, the in-situ dry density of deposited tailings is estimated at 1.15 tonnes/m3 based on the experience at Bema’s Julietta Mine. This value of in-situ density is somewhat lower than typical gold tailings values, but is selected as being representative of tailings that will be deposited entirely sub-aqueously and hence will have relatively low density and high void ratio. The design tailings volume based on this density is approximately 10,500,000 m3. The open pit and underground mining phases are scheduled to produce approximately 10 million tonnes of acid-generating (AG) and potentially acid-generating (PAG) waste. All of the AG waste will be stored in the tailings impoundment where it will be permanently submerged. The PAG waste will be used for construction of the tailings dam, stored in the tailings impoundment or used for underground backfill. At an assumed density of 2 tonnes/m3 the waste rock will occupy about 3.5 M m3 of storage in the tailings impoundment. The total amount of solids to be stored in the tailings impoundment is expected to be about 14 million m3. The following design criteria have been adopted for design of the tailings dam:

• The dam should have sufficient freeboard at all times to store the projected tailings volume and waste rock, plus 4 m depth of water (at least 2 m of free water and up to 2 m of ice in winter);

• A minimum of 1.5 m freeboard should be provided above the resulting tailings

and water level, comprising 1.0 m for storage of a Probable Maximum Flood (PMF) Event, plus 0.5 m for waves above that level;

• The tailings starter dam will be constructed of non-acid generating (NAG)

blended with some PAG waste rock and be sized to provide storage for the first three years of tailings production, AG waste rock and remaining PAG waste rock plus the design water storage and freeboard requirements;



• The tailings slurry water will contain total cyanide, at initial concentrations in the order of 5 mg/L. Total cyanide within tailings pore spaces may persist at these concentrations. The impoundment dam will be constructed so that there will be minimal seepage through the dam and foundations. It is assumed that there will be no seepage via deep groundwater flow, as the tailings basin will be rendered essentially impervious by permafrost, and will remain frozen. The dam face will be fully lined with a bituminous geomembrane to minimize seepage loss through the dam and foundations. The membrane will be tied to bedrock at the upstream toe to form a seepage cutoff.

The dam design selected for the Kupol tailings facility is a rockfill embankment with an impervious geomembrane on the upstream slope. The selection of this dam section is dictated by the climate and availability of materials. A rockfill embankment can be constructed in cold weather conditions, so the dam construction season will not be limited by the short summer season. The dam will be constructed of NAG and PAG rock from mine waste or quarried. It is planned that the starter dam shell will be constructed with NAG and PAG material from pre-production open pit development, or if this cannot be scheduled, then quarried NAG rock would be used. The dam will be raised in the downstream direction. It is anticipated that the downstream dam shell will be raised near to its projected ultimate crest elevation during the first three years of mine operation, when the open pit will be in production and waste rock will be available. The storage capacity will be increased above the starter dam level in later years by extending the geomembrane up the slope and also extending the upstream foundation seepage cutoff up the dam abutments. The seepage barrier will consist of a bituminous geomembrane liner placed on the upstream face and anchored to bedrock at the upstream toe of the dam. The geomembrane will be extended up the dam face as the dam is raised. Bituminous membranes are mainly composed of needle-punched, non-woven, polyester geotextile (300 gm/m2) impregnated and faced with bituminous mastic (mixture of bitumen and filler) creating a liner product with a typical thickness of about 4.8 mm. Bituminous membranes are considered to have much greater permanence than plastic liners. The material has high durability under extremely low temperatures. Bituminous geomembrane can also tolerate pH values in the range of 2 to 11 and can be installed in temperatures as low as -25º C. A technical information sheet for the bituminous membrane (type ES3) is presented in Appendix A. Tailings, waste rock (all AG and the portions of PAG not used in dam construction) and water management will be carried out using procedures that will allow effective operation in the cold climate at Kupol. The primary strategy will be to maintain a water cover over the tailings surface during winter. The impounded waste rock will mostly be placed above the active pond for first several years of production and ultimately be covered by tailings. Tailings will be discharged underwater, beneath the ice cover, to avoid freezing on exposed beaches, which could lead to large losses of water. The water cover will also allow for process water reclaim during winter.



During winter, there will be no input of runoff, and water will be depleted through losses to storage in tailings voids. The minimum depth of water will be maintained throughout the winter by either:

a) providing enough excess water in storage at the beginning of winter, through storage of runoff during the summer months, or

b) adding fresh makeup water from the process water wells during winter. To avoid water surplus in the pond, diversions will be installed to divert most of the catchment area upstream of the tailings facility. For Alternative #1, a total of 5150 m of diversion ditches would be required. For Alternative # 2, the total ditch length would be 4200 m. The tailings impoundment is designed to be a zero discharge facility. Regular monitoring of the tailings pond water depth and volume will be required to ensure that the operating criteria are being met. The water balance will be updated on a monthly basis throughout the year. Decisions on water additions will be made based on the ongoing water balance monitoring. During the spring freshet (the main spring runoff event), side discharges on diversion ditches may be opened to allow discharge into the pond on a controlled basis, in order to achieve target water depths at the beginning of winter, or conversely water balance observations may indicate that maximum diversion should be achieved to prevent water surpluses in the pond. Start-up of milling is scheduled for June 2008. Makeup water will be stored prior to start-up to allow the reclaim pump barge to begin operation, and also to provide water in storage for the following winter. Construction of the impervious liner and foundation cutoff are scheduled for completion in the fall of 2007. This construction schedule will allow runoff from the freshet in May and/or June of 2008 to be collected in the pond. Management of the diversions will be required to collect the target water volume. Tailings will be distributed by discharging at varying positions around the impoundment. The objectives of tailings distribution will be to:

a) make full use of the tailings storage capacity by filling all areas of the pond; b) provide additional seepage protection at the upstream toe and abutments of the dam by

discharging tailings preferentially along the crest of the dam, and c) provide adequate depth of water under the reclaim pump barge at all times.

For Alternative #1, the tailings line would be laid along the dam crest for initial operations and tailings discharged through a series of spigots. In later years, a tailings line will be raised with the dam crest and a separate line constructed to the back of the pond to allow for periodic discharge from the rear of the impoundment. For Alternative #2, the tailings line would be initially laid across the crest of the starter dam and discharged through a series of spigots, and would be extended in later years along the east side of the pond.



Geotechnical investigations for Alternative #1 indicate that the valley bottom is underlain by alluvial and colluvial deposited silty sands and gravels, with depths up to 19 m. The dam location was selected to minimize the depth of cutoff at the upstream toe. The cutoff depth is expected to be about 5 m. A preliminary geotechnical investigation at Alternative #2 indicates that bedrock is near surface over most of the dam footprint with the exception of the in-filled stream channels in the valleys where deposits with thicknesses up to 10 m have been encountered. The estimated construction quantities and costs favour Alternative #1 as the tailings dam fill quantities are significantly lower for the same storage volume. The 3 Year starter dam for Alternative #1 would require about 650,000 m3 of rockfill compared to about 1,500,000 m3 for Alternative #2. However, the actual construction cost will depend on the cost of rockfill. If the dams are constructed of open pit NAG and PAG waste rock, the cost of the higher volume would not be significant, and the cost difference of the two alternatives would narrow. Otherwise, Alternative #2 would have the following advantages:

• Reduced environmental footprint for the overall project; • Closer to the millsite, with a shorter tailings line; • No tailings pumping, Alternative #1 requires some pumping to lift tailings around the

saddle west of the millsite; and • Smaller catchment, lesser diversion work required.

At the end of mine life, a closure plan will be implemented for the tailings impoundment, with the following objectives:

• Leave the impoundment in a stable condition, so that the tailings will be permanently contained;

• Prevent acid generation of sulphide tailings and waste rock; • Provide a permanent, stable stream channel and spillway for conveyance of surface

runoff through the tailings impoundment; and • Ensure that all water discharges from the tailings impoundment will meet discharge

standards for protection of fish-bearing waters downstream. Following is the strategy that will be implemented for closure of the tailings pond.

1. At the time of closure, the tailings pond water quality will be analyzed to determine whether it is suitable for release. If the water meets all parameters for discharge, excess water will be released from the pond. Water release will take place over a period of time, through a planned and regulated process. Water releases would be timed to coincide with peak flow periods in Kaiemraveem Creek, at discharge rates which would be a defined percentage of the flow of Kaiemraveem Creek.

2. Excess tailings pond water would not be released upon cessation of mining and milling

activities if it did not meet effluent discharge standards. The water would be held or treated until suitable for release. Improvements in water quality would be achieved through natural degradation and, if necessary, through treatment.



3. The tailings pond water release will likely take place over one or more years following cessation of mine operations. The objective of the release will be to draw down the pond water to a minimum level, to create a small water pond. A permanent discharge spillway will be constructed to maintain the pond at this low level for closure. For Alternative #1 tailings pond, the spillway would be constructed along the west side of the pond, adjacent to the right abutment of the dam. For Alternative #2, the discharge spillway would be cut through the bedrock ridge on the northeast side of the pond.

4. The tailings pond surface will be monitored following shutdown to assess the rate of

ongoing consolidation settlement. Settlements may occur for several years, but it is anticipated that consolidation will not run its full course as the tailings mass will be frozen before complete consolidation occurs. The process of freezing and consolidation will be monitored through a combination of surface settlement gauges, piezometers and thermistors. The purpose of monitoring the consolidation and settlement process will be to assess the final surface contours of the tailings pond for design of drainage courses and closure cover

5. During and after the ongoing process of consolidation and freezing, a closure cover will

be constructed over the tailings surface. It is anticipated that the cover will consist of about 2 m thickness of quarried, non-reactive rockfill, to serve the following functions:

• To provide protection from surface erosion of the tailings. Additional erosion

protection will be placed in the main drainage courses on the tailings surface; and

• To provide a thermal barrier to prevent seasonal thawing of the tailings, for prevention of oxidation of sulphides.

6. On completion of the closure cover and construction of erosion protection on main

drainages through the pond, the runoff diversions will be taken out of service and the main drainages routed through the tailings pond and spillway

7. After a period of monitoring to confirm that the tailings surface cover and all drainage

courses are stable, the site access roads and pipeline corridors will be re-graded to match the natural topography, and reclaimed with native vegetation to minimize the potential erosion that could occur from exposed surfaces in such permafrost terrain.

The above closure plan is considered preliminary and conceptual. Details of the plan can best be finalized during mine operations and through closure trials. However, the plan is considered to be robust. There is little doubt that the tailings deposits will become totally frozen within a number of years, so there will be no ongoing concerns with respect to quality of pore water discharges or oxidation of sulphide tailings. Spillways will be constructed in rock so should be permanent and durable, preventing any long term erosion of tailings.



Additional site investigations will be required in 2005, with the following key tasks to be carried out:

• Exploration drilling to confirm that Alternative #2 site would not sterilize any significant ore reserves;

• Boreholes in the footprint of the dam for Alternative #1 tailings dam, as the 2004 investigation was carried out upstream of the final alignment;

• Cored boreholes along the alignment of Alternative #2 dam site; • Geological mapping of the damsite and ditch alignments; • Installation of thermistors to take advantage of the 2005 drill holes; and • Exploration drilling to determine suitable NAG quarries, with acid-base accounting (ABA)

testing.



1.0 INTRODUCTION This report presents the design of tailings storage facilities for the Kupol Gold Project, in Chukotka Autonomous Okrug, Far East Russia. The design is presented to a level to support a project feasibility study being prepared by Bema. The proposed Kupol Gold Project entails development of a high-grade gold and silver mine. The project site is located in the Chukotka Autonomous Okrug, at latitude 66°47’ slightly above the Arctic Circle, and longitude 169°19’. The project is located in a very cold region of continuous permafrost. Access to the site is via helicopter year-round and by winter road in the months of December through March from Bilibino, a city about 200 km northwest from the site. Major equipment and supplies for the project will be transported by ship during the summer to the East Siberian Sea port of Pevek, which is about 325 km northeast of Bilibino. An airstrip is planned to be constructed as part of the project to allow year-round access by fixed-wing aircraft. The Kupol Mine will be developed initially as a small open pit to access near-surface ore, with underground mine development to also start at an early stage of the project, so the mine will transition in its early years to underground mining. A gold milling and beneficiation plant will be constructed with a design throughput of 3000 tonnes per day. The plant will include:

• Crushing • Grinding • Whole ore cyanidation • Gold refinery • Cyanide destruction circuit to reduce residual cyanide in tailings to low levels

This report presents feasibility designs of two tailings sites: Alternative #1, located southwest of the plant site in the valley of a right tributary of the Middle Kaiemraveem River, and Alternative #2, located immediately south of the plant site on the southern part of the Granichnaya Mountain, in the valley of the Tretyi Creek. Alternative #2 is the favoured site due to a number of advantages that are presented under Section 3 below. However, additional exploration drilling has to be completed to confirm that the tailings site does not conflict with economic mineralization, and geotechnical investigations have to be carried out to confirm adequate foundation conditions.



2.0 SITE DESCRIPTION 2.1 General The Kupol area is formed by rounded bedrock hills, incised by streams that generally flow only in summer months (May to October). Bedrock outcrops over much of the site. Where bedrock outcrops it is generally highly shattered in the upper 2 m to 3 m from frost action. The bedrock is covered on gentler slopes by organic-rich colluvium. There are no trees, other than some willow in valley bottoms. The dominant vegetation on organic-rich slopes is “tussocks”, a low grassy plant, while areas of shallow bedrock are covered by lichens and low shrubs. There appear to be no significant deposits of low-permeability soils. Granular deposits exist in stream valleys to depths up to 19 m. 2.2 Climate The proposed Kupol Mine Site is located within the Chukotka Autonomous Okrug (Region) in Eastern Russia, slightly above the Arctic Circle, in a very cold environment. The climate in this region is characterized by long cold winters (8 to 8.5 months), short summers (2.5 months), and even shorter springs and falls. Kupol is located at the separation of the watersheds of the Bering and Chukchi Sea within the Andyrsky Mountains. The Andyrsky Mountains run east-west and have elevations over 2000 m. Some key parameters from this data analysis, provided by Bema, are summarized as follows:

• The estimated average annual air temperature at the Kupol site, with only minor variances, is near -13 °C;

• Weather data from Ilirney shows the annual precipitation to be about 278 mm. For the design basis, the average total amount of precipitation is estimated from analysis of data at the Markovo station to be 380 mm. The total number of days with average daily temperature above zero does not exceed 50 days. Positive average daily temperatures are first noted in the first 10 days of June. The transition from positive average daily temperatures to negative average daily temperatures occurs in the middle 10 days of September;

• The approximate annual amount of evaporation from surface water sources is 280 mm; and

• Snow begins falling in the middle of September and achieves a maximum depth in March. The average depth of snow cover is observed to be 38-45 cm. As a result of heavy winds, the valleys are filled with snowdrifts and the tops of ridges and hills and steeps slopes are blown bare. The average snow density reaches 160 kg/m3 with water equivalent content of 107 mm. During the winter months, approximately 174 mm of water equivalent precipitation is expected. Snow is approximately 46% of the annual precipitation.



The atmospheric circulation over Kupol moves north during winter and south during summer. This north flow brings moist air from the Bering Sea to the site, suggesting that climate stations located along the north-south path would provide a better climate correlation to the mine site. Three regional climate stations were considered for the extension of the mine site data - Ilirney, Ostrovnoje and Markovo and are shown on Figure 2.1.

Figure 2.1 - Regional Climate Stations



Ilirney is the closet station to the site located approximately 150 km west of the Kupol mine site. Markovo is located about 660 km south and Ostrovnoje is located about 690 km west of the site. All three stations have over 20 years of data. Ilirney is the closest climate station to the site and can be utilized to extend the minesite climate station. However, the periods of record at the minesite and Ilirney do not overlap and therefore a direct correlation between this regional station and the mine site could not be adequately established. This holds true for the Ostrovnoje station as well. Markovo is the only climate station with recorded data that coincides with the two summer seasons of data recorded at the mine in 2003 and 2004. Markovo is better suited to establish a direct correlation with the site data. Other reasons for employing Markovo (discussed in detail in the following paragraphs) include the north-south movement of wind currents, long-term monthly patterns and conservatism. As discussed above, the wind currents that accumulate and deposit moisture from the Bering Sea circulate in a north-south direction. The Markovo station is located along this path south of the mine site. Ilirney is located west of the mine site and on the other side of the drainage catchment divide between the Bering Sea and Chukchi Sea to the north, and may therefore be influenced by different moisture patterns. A plot of long term period of record monthly precipitation between Ilirney, Ostrovnoje, Markovo and the site’s two years of data over the summer period from June to September suggests that Ilirney’s peak month occurs in July, while Markovo and the mine site experiences the peak in August, as illustrated in Figure 2.2. This suggests Ilirney experiences a different rainfall pattern in the summer than Markovo and the mine site. Finally, it is prudent to err on the conservative side given that the tailings impoundment for is a closed circuit system with no discharge permitted. Ilirney’s long term average annual precipitation from 1945 to 1992 is estimated to be 278 mm, while Markovo’s estimate is 380 mm over the same period. It would be more conservative to assume a greater annual inflow volume at the feasibility stage to ensure design flows and volumes will not overtop the tailings dam. 2.3 Baseline Hydrology 2.3.1 General A correlation is drawn between the Markovo climate station and the site-recorded data for two summers in 2003 and 2004. The comparison with seven months (July-September 2003 and June-September 2004) of rainfall data produced a good relationship between the two stations, with the site experiencing an estimated 15% more precipitation than at Markovo. This difference in precipitation could be explained by the orographic influence caused by an elevation difference between Markovo and the mine site. The Markovo station is at an elevation of 33 m and the mine site is at an elevation of 650 m. This correlation should evolve over time as more data are collected from the mine site.



2.3.2 Precipitation Employing the correlation between Markovo station and the site to the 72 years (1895-1999) of precipitation data at Markovo station, a derived average annual precipitation at the mine of 380 mm is estimated. Applying this same relationship to annual maximum daily precipitation at Markovo station, the 24-hour, 200-year storm event is estimated to be 63.5 mm, and the 24-hour, Probable Maximum Precipitation (PMP) is estimated to be 206 mm. It should be noted that short-duration storm events will experience slightly higher peaking ratios when compared with a monthly period. Lack of storm tracking information and erratic behaviour at the daily resolution level between the site and Markovo station data prevent establishing a daily peak ratio, so a monthly volumetric ratio was employed. As more data become available at the site, a better peaking factor will be established.

Figure 2.2 - Comparison of Precipitation



2.3.3 Temperature A comparison of the average monthly temperatures at Ostrovnoje, Ilirney, Markovo and the mine (Figure 2.3) indicates that the summer temperatures are approximately the same, but winter temperatures at the minesite and Markovo are slightly higher than at Ostrovnoje and Ilirney. This difference in temperature is likely due to the effect of atmospheric circulation. This observation should be verified as more mine site data are collected. Based on the two years of site data the mean annual temperature is -10°C. The average temperature in winter (October to May) is -17.8°C and the average temperature in summer (June to September) is 8.2°C.

Figure 2.3 - Comparison of Temperature

2.3.4 Evaporation The only regional evaporation data available is from the Ostrovnoje climate station. Both Ostrovnoje and the mine are located above the Arctic Circle, are within the same latitude region, and experience similar solar radiation exposure. It is therefore assumed that the evaporation at Ostrovnoje (annual amount of evaporation from surface water sources of 280 mm) would provide a reasonable approximation of the evaporation at the mine.



2.3.5 Streamflows The hydrologic characteristics of the watersheds in the vicinity of Kupol mine are typical of arctic regions. Extreme flood events are caused by a large rainfall event combined with the spring snowmelt. The watersheds that will be influenced by the work at Kupol are Kaiemraveem Creek and Starichnaya River. Kaiemraveem Creek flows south to join the Anadyr River, which flows into the Bering Sea. Starichnaya River, the headwaters of which are a few kilometres north of the mine, flows north into drainages that discharge to the Arctic Ocean. In the vicinity of the mine, Kaiemraveem Creek has steeper valleys than Starichnaya River. With high winds experienced at the site, snow will likely be blown into the valley. High runoff coefficients are expected during the spring freshet expected to occur mostly in May, as the ground will still be frozen. Flow data has been collected at two recording sites along an unnamed creek, which is an upper catchment tributary of Kaiemraveem Creek, west of the mine camp. The unnamed creek flows through the site for Alternative #1. The first station is located immediately upstream of the proposed dam location, while the second is located towards the upper limit of the impoundment. The maximum-recorded stream discharge at the first station is 0.54 m³/s, while the maximum-recorded flow at the second station is 0.24 m³/s. Runoff coefficients are calculated using the recorded flow and rainfall data collected at the mine site. Initially, negligible groundwater effects and only storm events are considered in computing the runoff coefficients. However, the spring freshet in May or early June increases the runoff coefficient substantially for the month with drainage occurring over frozen ground. Additionally, recorded hydrograph data suggest that saturated ground surface condition and subsequent groundwater releases increase the runoff from rainfall events in August and September. The resulting coefficient estimates therefore include these assumptions and are summarized in Table 2.1 below.

Table 2.1 - Monthly Runoff Coefficients Kupol

MONTH COEFFICIENT June 0.80 July 0.60 Aug 0.60 Sept 0.90

These coefficients will be reviewed and revised as more site-specific stream flow and rainfall data are collected. 2.4 Surficial Geology The Kupol site is characterized by rounded bedrock hills incised by streams, with colluvium on lower slopes of the hills and fluvial deposits in valley bottoms. The surfaces of the bedrock hills are mantled by 1.5 m to 2 m of weathered and frost-shattered bedrock. Flatter slopes are covered with organic-rich colluvium. The organic-rich soil slopes are ice-rich, and generally



covered by tussocks. Stream channels contain alluvial deposits of sand, gravel, silt and boulders. 2.5 Permafrost Setting The Kupol site lies at approximately 66o47’ North Latitude in the far east of Russia in a continuous permafrost zone, i.e. permafrost is laterally continuous (see Figure 329-001). The area is underlain by deep, cold permafrost. The depth of the active permafrost zone is noted on Russian topographic maps to be 1.2 m, and this was confirmed by excavation of test pits during the 2003 site visit. From thermistors installed in the mine area, the depth of permafrost was extrapolated to depths of in excess of 300 m below surface, and likely to as deep as 450 m. A thermistor in the valley of Alternative #1 confirmed that the dam foundation soils are in permafrost with a temperature of about -6°C below the active permafrost zone. 2.6 Seismicity The level of regional seismic activity was assessed by AMEC’s Moscow office, on the basis of The Construction Norms and Rules (SNiP) II-7-81* “Construction in seismic regions” (current edition 2000) for development of adequate technical solutions. This SNiP specifies tolerance of construction to seismic impact (in our case it means resistance of the dam structure). The seismicity was assessed for the Kupol site at N66° 47’; E169° 20’.

According to SNiP 2.06.01-86 (Appendix 2), a rockfill tailings dam with height 49.5 m (for Alternative #2, foundation on bedrock case) and dam with height 53.5 m (for Alternative #1, foundation on frozen soils) are categorized as Class II of Hydraulic Engineering Construction. These dams correspond to level II of responsibility (p. 1.3* of SNiP II-7-81* “Construction in seismic regions” and SNiP 2.06.01-86 “Hydraulic Engineering Construction”); meaning map OSR-97-B can be used for seismic zoning. From seismic zoning map OSR-97-B, the seismic activity level of this region is classified as 6 points for 5% probability of exceedence in 50 years. This seismic level applies to the towns of Anadyr, Bilibino and Ilirney, which are all located in the same region as Kupol. According to SNiP II-7-81*, for this level of regional seismic activity, seismicity does not need to be taken into consideration in design of these structures. The earthquake by intensity grade 6 pertains to geological processes hazard category “Hazardous” (SNiP 22-01-95). Therefore, assessment of hazard of geological factors complex must be executed on base of their research and results of forecast of modifying geological conditions due to interaction of construction object with geological environment (SNiP 22-01-95). However, according to SNiP 22-01-95 (current edition 1996) “Geophysics of hazardous natural impacts”, natural hazards must be assessed. The interaction of geological processes, including probable seismic actions, frost actions and other exogenous geological processes with a planned facility must be investigated and researched. SNiP 22-01-95 classifies an intensity grade 6 earthquake as “Hazardous” with regard to natural hazards. Therefore, at the design



stage, geological hazards such as solifluction, scree slopes or kurums (formation of fields of large rock blocks as a result of frost jacking of fractured bedrock out of the active layer, which are slowly moving on the surface) and other permafrost-related phenomena should be assessed in relation to safety of the tailings impoundment. 2.7 Alternative Technologies Alternatives analysis for the Kupol tailings system included assessment of alternative tailings management technologies and alternative tailings storage locations. Alternative tailings technologies considered for Kupol included:

• Filtered, dry stack tailings • Conventional slurry tailings disposal

Filtered, dry stack tailings Filtered tailings systems are based on use of belt or pressure filter equipment to dewater tailings material to the point that it can be transported mechanically, by trucking or conveying, to a tailings disposal site. The material is placed and spread on a dry stack. A filtered, dry stack system was considered for the following:

• Avoids cold weather water management problems • Eliminates need for impoundments dams and thereby substantially reduces geotechnical

stability issues; • Conserves water as less water is trapped as ice or in voids of tailings; • Allows concurrent reclamation of dry stack slopes; and • Eliminates seepage.

Filtered, dry stack tailings system was discounted for the following:

• Capital costs, filtration plants have a high capital cost; • Operating cost, power draw for filtration plant is high, filter tailings have to be hauled to

the designated dry stack dump area; • Uncertainty of ore consistency, the potential for clay inclusion with the ore will cause

upsets of the filtration process, and • Dust, the windy environment at Kupol can results in a severe dust problem from the dry

stack dump (the dust could contain residual cyanide from the mill process). As a result of the estimated high capital and operating costs of the filter and dry stack system, as well as the risk that the system may not perform adequately in all ore types, this technology was considered inappropriate for the Kupol Project and was eliminated from further consideration.



Conventional slurry tailings disposal systems Conventional slurry tailings disposal systems in cold climates such as at Kupol can be challenging to operate, due to freezing problems. However, there has been considerable experience with slurry tailings disposal systems in arctic climates and methods have been developed for overcoming slurry and water handling challenges. Some of these methods include:

• Maintaining a depth of water in the pond to allow tailings to be discharged under water during winter to prevent freezing that would occur on tailings beaches;

• Use of bubbler systems to maintain open water around reclaim pump barges; and • Heating and insulating pipelines to avoid freezing.

Following site selection studies in which sites were identified that would be suitable for construction of a tailings dam to create a storage impoundment, it was decided that a conventional slurry system would be most suitable, based on both lower cost and lower risks than for a filtered, dry stack system. It should also be noted that Bema has operational experience with cold climate conventional slurry tailings disposal at the Julietta Mine located in the Magadan Region, Far East Russia. 2.8 Alternative Sites During a site visit to Kupol by Bema and AMEC in August 2003, a site selection study was undertaken to identify key components of the project, including the plant site, the open pit waste dump site, the accommodation camp, potential water supply sources and potential tailings storage sites. Site selection was carried out using 1:50,000 scale topographic maps to locate potential sites. The potential sites were then viewed by helicopter flyover and walking inspections. The key criteria for site selection were as follows:

• In a valley with a relatively level valley bottom gradient, and a confining point for dam construction, to achieve an acceptable ratio of tailings storage volume to dam fill volume;

• In a valley with modest drainage area, to avoid the necessity of handling large diversion flows; and

• Located relatively near the plantsite to minimize the length of tailings pipelines, and located at a lower elevation to minimize or avoid the necessity of pumping tailings;

Figure 329-002 presents the two alternative sites that were identified for tailings storage. Both sites are south of the plantsite, in the catchment of Kaiemraveem Creek, and ultimately in the drainage basin of the Anadyr River into which Kaiemraveem Creek discharges some 60 km to the south. Anadyr River flows eastward to discharge into the Pacific Ocean and is an important salmon river.



Sites were also considered to the north, in the drainage of Starichnaya Creek, which drains into the Mayli Anjuj River about 25 km north. The Mayli Anjuj River flows northward and ultimately drains into the East Siberian Sea. No acceptable sites could be identified to the north of the plantsite outside of the Kaiemraveem Creek drainage. Valleys to the north are generally characterized by subdued topography, with no favourable damsite locations, and also the runoff catchment area at any tailings impoundment site in the main valley of Starichnaya Creek would be large and require major diversions. The first tailings impoundment site investigated (Alternative #1) shown on Figure 329-003 is located southwest of the plant site in the valley of a right tributary of the Middle Kaiemraveem River to the southwest of the Kupol lease area. The site has good topographic characteristics. The gradient of the stream in the valley bottom is gentle, estimated at about 2.5% slope. There is a favourable damsite location, controlled by bedrock abutments where the valley narrows, and upstream the valley widens to provide storage. The deep permafrost in the region would contribute to forming a reliable seepage barrier for the impoundment. Some pumping would be required to lift tailings over the topographic saddle to the west of the millsite, but from that point tailings would flow by gravity. The second tailings impoundment investigated (Alternative #2) shown on Figure 329-004 is located immediately south of the plant site on the southern part of Granichnaya Mountain, in the valley of Tretyi Creek. Alternative #2 was presented as the Waste Dump in the Preliminary Economic Assessment. Optimization of the pit design has reduced the volume of mine waste rock such that a separate waste rock dump is no longer necessary and the site is now considered as tailings storage site Alternative #2. The site is in a broad, gently-sloping valley to the southwest of the proposed open pit. The site is favourable from the standpoint of geotechnical stability, as bedrock is essentially at surface over most of the site, with a thin mantle of weathered material overlying the bedrock with thicker overburden in the valley bottoms. Alternative #2 has the following advantages over Alternative #1:

• A reduced environmental footprint; • Closer to the millsite, with a shorter tailings line; • No tailings pumping (Alternative #1 requires some pumping to lift tailings around the

saddle west of the millsite); • Lower dam raising costs (utilizing waste rock, short haul from the open pit); • Smaller catchment, lesser diversion work required; and • Good geotechnical foundation conditions.

Based on the above considerations, Alternative #2 is being considered as the primary tailings storage option. However, Bema has to complete exploration work at this site to confirm that the tailings impoundment would not conflict with any resource that may be present. Therefore, for purposes of this study both Alternative #1 and #2 are being designed to feasibility level.



3.0 SITE INVESTIGATIONS 3.1 Geotechnical Drilling In 2004 JSC Dalstroyizyskania (Dalstroy) carried out a site investigation at the Kupol site. Investigations included 37 boreholes for Alternative #1, 9 boreholes for the proposed Dry Stack site, and 3 boreholes for Alternative #2, for a total of 49 boreholes. Alternative #1 borehole locations are shown on Figures 329-005 and 006. The drilling was carried out by dry coring. A detailed Geotechnical drilling program is currently being carried out for the Alternative #2 site, which will be incorporated in the detailed design of the facility. 3.1.1 Program/Findings During drilling, samples were laid out on the ground for logging and sampling. Borehole depths varied from 2.2 m to 32 m, with bedrock encountered in all but three boreholes (#14a – 3 m depth, #17 – 7 m depth, and #11 – 14 m depth). Translated logs for all the boreholes were provided by Bema Gold. To verify the borehole translations the original Russian logs were translated by A. Tchekchovski, AMEC geological engineer and permafrost specialist. AMEC engineering designs presented herein are based on the AMEC translation of the Dalstroy logs. Temperature profiles were determined in some of the holes. Bema installed a thermistor in one of the boreholes, BH-4, in Alternative #1. 3.2 Mapping A base map for the Kupol Project has been compiled, with the majority of the site represented by 5 m contours. Detailed surveying of the original Alternative #1 was carried out providing 1 m contour accuracy. A detailed survey is absent for portions of the upper impoundment limit and for a portion of impoundment covering off the diversion alignment. In the area of Alternative #2, 5 m contour accuracy has been provided. The current surveys are considered sufficient for the feasibility level development.



4.0 DESIGN CRITERIA 4.1 Tailings Material SGS Lakefield Research Limited carried out a suite of tests on typical tailings material, identified for this program as Mixed Iron Cyanide Tailings Fresh Solids. This material is considered to be typical tailing from the Big Bend Ore. The tailing material will be the residue from cyanide leaching of the ore, followed by cyanide destruction. The results of testing typical tailings from bench scale metallurgical tests are presented in SGS Lakefield (2005a) and SGS Lakefield (2005b). The tailings will be typical, finely-ground gold mine tailings, with 100% finer than 0.15 mm. The material gradation is shown on Table 4.1 below.

Table 4.1 - Typical Kupol Tailings Gradation Analysis

Sieve Size (US Stan.)

Particle Size (mm)

Weight Passing %

1” 25.400 100.0 3/4” 19.050 100.0 3/8” 9.500 100.0 #10 2.000 100.0 #12 1.700 100.0 #40 0.425 100.0 #60 0.250 100.0 #100 0.150 100.0 #200 0.075 97.4

- 0.040 78.5 - 0.029 68.4 - 0.022 59.3 - 0.016 49.3 0.012 42.2 - 0.009 33.2 0.006 28.2 0.005 21.1 0.001 8.0

Other parameters derived for the tailings in the testing program included:

• Specific Gravity = 2.64 • Atterberg Limits – could not determine Liquid or Plastic limits, indicating that the material

is non-plastic • Compaction Characteristics as defined by Standard Proctor test (ASTM D-698)

o Maximum dry density = 1.748 o Optimum moisture content = 16.6%



Gold will be dissolved from the ore in a cyanide leach process. Following leaching, the tailings will pass through a cyanide destruction circuit, and will also be treated to reduce thiocyanates. Design residual levels in the tailings following treatment, prior to discharge to the tailings impoundment, will be 5 mg/L total cyanide and 10 mg/L total thiocyanate. SGS Lakefield (2005b) presents the results of a Mineralogical Examination of Mixed Iron Cyanide Tailings Fresh Solids from Kupol Project. This work concluded that, mineralogically, the material was composed mainly of silicate minerals (>95%). Sulphur is present primarily as pyrite (FeS2), with a modal abundance of 2.2% in the sample examined. The pyrite predominantly occurs as liberated grains (90%) while the remaining 10% occurs as attached or locked particles associated with quartz. The results of a modified ABA tests completed on the fresh, Mixed Iron CND Tailings show the sample had moderate sulphide sulphur (S=) and sulphate (SO4) contents (0.56% and 0.67% respectively). A negative net neutralization potential (Net NP) of -9.5 t CaCO3/1000 t, and an NP/ acid potential (AP) ratio of 0.45 indicate that the fresh Mixed Iron CND Tailings Fresh Solids have the potential to generate acidic drainage. In comparison, after aging 90 days, the Mixed Iron CND tailings had a higher potential to generate acidic drainage, with a lower negative Net NP of -19.9 t CaCO3/1000 t, and a smaller NP/AP ratio of 0.19. Thus, the tailings material would be classified as acid generating. 4.2 Tailings Production Schedule The feasibility study shows 7.1 million tonnes of material to be processed through the mill during a seven year mine life. The criteria provided by Bema was to develop a facility capable of holding 12 million tonnes of waste. This extra capacity will be used to store material not included in the feasibility study (inferred Years 8 thru 13.5) or material discovered in the future. The Kupol Gold Project is scheduled to have a nominal plant throughput of 3000 tonnes of tailings per day (tpd) during Years 1 through 7 (open pit phase years 1 through 3 and 4 years of underground operations). The rate of processing beyond the feasibility study is assumed to continue underground with a processing rate of 2000 tpd to the final 12 million tonnes for a mine life of about 13.5 years. The complete mine production schedule for the life of the Kupol Gold Project is summarized in Table 4.2 (Bema provided an upper bound for the ore reserve of 12 Mt or about 13.5 years).



Table 4.2: - Kupol Gold Project Mine Production Schedule (tonnage)

Year Annual Tonnage (000's)

Cumulative Tonnage (000,s)

1 (June – December) 537 537 2 1,095 1,632 3 1,095 2,727 4 1,095 3,822 5 1,095 4,917 6 1,095 6,012 7 1,074 7,086 8 730 7,816 9 730 8,546 10 730 9,276 11 730 10,006 12 730 10,736 13 730 11,466 14 534 12,000

4.3 Tailings Density and Volume For design purposes, the in-situ dry density of deposited tailings is estimated at 1.15 tonnes/m3

based on the experience at Bema’s Julietta Mine. This value of in-situ density is somewhat lower than typical gold tailings values, but is selected as being representative of tailings that will be deposited entirely sub-aqueously and hence will have relatively low density and high void ratio. A void ratio (e, which is the ratio of the volume of voids to the volume of solids within a unit volume of the tailings in-place) of e = 1.26 was calculated assuming the specific gravity of the rock of 2.6 (from Bema). Thus, the total volume of tails to be deposited in the tailings impoundment over the life of the mine is

• 12 M tonnes / 1.15 tonnes/m3 = 10.5 Mm3 Bema should undertake annual tailings beach surveys and water pond soundings to reconcile actual tailings tonnage discharge with surveyed volumes to determine the actual average in place dry density being achieved within the impoundment. Based on the above value of in-situ density, the annual tailings volume are as shown on Table 4.3.



Table 4.3 - Kupol Gold Project Mine Production Schedule (volume)

Year Annual Volume (Mm3)

Cumulative Volume (Mm3)

1 467 467 2 952 1,419 3 952 2,371 4 952 3,323 5 952 4,276 6 952 5,228 7 934 6,162 8 635 6,797 9 635 7,431 10 635 8,066 11 635 8,701 12 635 9,336 13 635 9,970 14 530 10,500

4.4 Waste Rock The geochemical make-up of the waste rock to be produced from the open pit and underground development was initially based on a visual classification system developed by Dr. Steve Atkin. The classification scheme identified the waste rock as non-acid generating (NAG), potentially acid generating (PAG) and acid generating (AG) based on field observations during exploration drilling. The details of the classification scheme are discussed in the geochemical report issued by Atkin (Atkin 2005). At start-up a program will be implemented to sample every blast hole for total sulphur to better define the waste rock classification into NAG, PAG and AG. For feasibility design a schedule of waste was provided by Wardrop and Bema. Table 4.4 presents a summary of the waste rock tonnages.

Table 4.4 - Total Waste Rock Tonnages

Open Pit Underground

17 M t 1.5 M t The resulting estimated waste rock tonnages and volumes specific handling requirements are as shown on Table 4.5 below.



Table 4.5 - Waste Rock Tonnages and Quantities by ARD Classification

Quantity AG PAG NAG

Tonnes (x106) 7.6 2.0 8.7

Volume (m3) (based on 2 t/m3)

3.8 1.0 4.4

Placement/Usage 90% placed in the tailings

pond, 10% used as

underground backfill

100% mixed with NAG rock and used as

dam fill (any excess PAG will be placed

with AG in the tailings pond)

100% for dam and site

infrastructure construction



5.0 DESIGN OF TAILINGS FACILITIES 5.1 Design Criteria The following design criteria have been adopted for design of the tailings dam:

• The dam should have sufficient freeboard at all times to store the projected tailings volume, plus 4 m depth of water (at least 2 m of free water and up to 2 m of ice in winter). The purpose of the 2 m minimum free water depth is to maintain sufficient depth beneath the pump barge to minimize the potential for tailings entrainment in the pumped reclaim water

• A minimum of 1.5 m freeboard should be provided above the resulting tailings

and water level, comprising 1.0 m for storage of a Probable Maximum Flood Event, plus 0.5 m for waves above that level. In the first few years it will be necessary to store more than 4 m depth of water to provide sufficient reclaim water in the winter months due to the initially smaller area of the tailings basin

• The impoundment should be constructed so that there will be minimal seepage

through the dam and foundations. It is assumed that there will be no seepage via deep groundwater flow, as the tailings basin will be rendered essentially impervious by permafrost. This assumption is verified by thermal modeling

• The bulk of the starter dam will be constructed in the winter months in order to

‘lock-in’ the permafrost. Thermal modelling shows that the permafrost in the foundation material is preserved by the placement of rockfill in the winter minimizing the potential for foundation settlement

• The tailings starter dam will be constructed to provide storage for the first years

of tailings production, AG (all) and PAG (portions of) waste rock and the design water storage and freeboard requirements

• The tailings are sulphide-bearing and are therefore potentially acid generating.

As such, it is necessary to inhibit oxidation of the tailings, and/or prevent release of any water from the tailings pond that is low in pH and laden with metals. During operation, with continuous discharge of slightly basic pH (target pH of 8 to 8.5) slurry tailings into the impoundment, oxidation will not occur given the saturated condition of the tailings. The liner system, which will minimize seepage from the impoundment, will assist in maintaining the tailings in a saturated condition. At closure, the tailings will freeze and be covered, as described in more detail in Section 10.0

• Acid generating waste rock will be placed within the ultimate pond limit to ensure

that it will become submerged within the impoundment at closure. Runoff from the waste rock will be collected in the tailings pond

• The tailings slurry water will contain cyanide, at initial concentrations as high as 5

mg/L (total cyanide). Cyanide in the pond will undergo natural degradation, but



cyanide within tailings pore spaces, and within groundwater, may persist at relatively high concentrations. As such, and as described previously the dam face will be lined with a bituminous geomembrane liner to minimize seepage loss.

5.2 Dam Design 5.2.1 Design Section The dam design selected for the Kupol tailings facility is a rockfill embankment with an impervious geomembrane liner on the upstream slope. The selection of this dam section is dictated by the climate and availability of materials. A rockfill shell can be constructed in cold weather conditions, so the dam construction season will not be limited by the short construction season. Due to lack of impervious soils in the area, a geomembrane liner is the best option for creating an impervious zone in the dam. This design will be applied to both alternative locations, with foundation preparation to suit each site. The upstream liner is intended to reduce the possibility of seepage out of the pond during operations. In the longer term, that is closure, it is predicted that the entire tailings body will become frozen and therefore seepage is not a design issue. The upstream liner is potentially subject to loading mechanisms from the floating ice sheet either due to possible adhesion of the ice cover to the liner and then jacking by pond level changes, or due to moving ice during spring thaw. These mechanisms require additional consideration, but at this stage of design it is planned that the liner will be covered by a granular layer to assist in decoupling the liner from any ice effects. The bituminous liner may not require a cover material; this has been identified as a potential cost savings and requires further investigation. The dam will be initially constructed as a starter dam providing storage for the first 3 years of production. It is planned that the starter dam shell will be constructed with NAG and PAG material from pre-production open pit development, or if this cannot be scheduled, then quarried NAG rock would be used. The dam will be raised in the downstream direction. The downstream dam embankment will be raised during the mine life as required. The storage capacity would be increased above the starter dam level in later years by extending the geomembrane liner up the slope and also extending the upstream foundation seepage cutoff up the dam abutments. The starter dam through ultimate dam is designed with a crest width of 10 m and upstream and downstream slopes of 2.5H:1V and 2.1H:1V respectively (the downstream slope conforms to the Russian regulations for the overall closure slope for waste rock dumps). The mine waste rock or quarried rock fill in the dam shell will be placed and compacted by dozer and haul truck traffic. Transition and bedding layers for protection of the geomembrane will be produced by processing quarried or pre-stripped mine NAG rock. The transition and bedding layers will be placed in uniform lifts and compacted to achieve the required density specifications as presented in Section 8.1.4.



5.2.2 Foundation Preparation The footprint of the dam will be cleared and grubbed of all organic and ice-rich materials. The remaining sands and gravels are frozen and are considered suitable foundation materials if maintained in a frozen state. The dam abutments will also be cleaned of any organic material prior to rockfill placement. 5.2.3 Cutoff Trench Preparation/Seepage Control The principal seepage barrier will consist of a bituminous geomembrane liner placed on the dam upstream face and anchored at the dam crest and sealed to bedrock at the upstream toe of the dam. The geomembrane liner will be extended up the dam face as the dam is raised. Bituminous membranes are mainly composed of needle-punched, non-woven, polyester geotextile (300 gm/m2) impregnated and faced with bituminous mastic (mixture of bitumen and filler). The geomembrane liner should be anchored into a seepage cutoff trench with minimum width of 6.0 m and should extend through the foundation soils and weathered bedrock into competent intact bedrock as shown on Figure 329-009. The core contact area should be excavated to bedrock and then cleaned by hand and air jetting to develop a clean surface that will provide a suitable base for application of concrete treatment. If the bedrock is heavily fractured and jointed with ice or sand and gravel infilling, this material shall be removed to allow the shotcrete to penetrate the joints/fractures by 50 or 60 mm. All exposed bedrock of the seepage cutoff/anchor trench shall be inspected to determine if such treatment is required. The bituminous geomembrane liner (discussed in detail below) can be adhered directly to the concrete surface. Propane torches are used to heat the bituminous geomembrane liner making the surface ‘sticky’. The sticky surface is essentially welded to the concrete surface when pressure is applied by rollers. Since the threads of the geotextile component of the bituminous liner are completely coated with bituminous mastic, they are protected against outside physio-chemical attack. The physio-chemical resistance of the geomembrane liner is thus the same as that of bitumen. Bituminous mastics are known to last for millennia and have fully protected items that had been encased in similar mastics as far back as Mesopotamian times. Of particular note is that geomembranes including bitumen as a key element are gaining increasing acceptance for containment of hazardous radioactive wastes where longevity is a critical requirement. In terms of the durability of the material under extremely low temperatures, it is understood in discussions with the supplier that the geomembrane can be installed in temperatures as low as -25º C and undergoes minimal expansion/contraction with variances in temperature. Bituminous geomembrane can also tolerate pH values in the range of 2 to 11, well beyond the anticipated operating ranges. Another advantage of bituminous geomembrane liners is that they display excellent resistance to puncturing by aggregates, illustrated by the fact that they have been used directly under railroad ballast. In this regard they are superior to high and low density polyethylene liners (HDPE and LDPE), which require finer-grained bedding (at Kupol this would have to be



processed material) and/or geotextiles (a geotextile-clay composite liner (GCL) would be required with an HDPE or LDPE liner at Kupol). Bituminous geomembrane liners can be installed on steeper slopes (2H:1V) than traditional HDPE or LDPE liners (maximum 2.5H:1V) resulting in a savings on construction rockfill unless a liner cover material is required which requires a flatter 2.5H:1V slope. Installation of the bituminous geomembrane is straightforward, the overlapping zone of geomembrane is propane torched to allow the bitumen to become sticky, and then rolled to create a sealed joint. Quality control testing will be carried out to ensure the competency of the welded seams. 5.3 Water and Materials Balance The water and materials balance for Alternatives #1 and #2 were based on the following tailings inflow and deposition characteristics:

• Nominal dry solids inflow of 3,000 metric tonnes per day (MTPD) for the first 7 years • Nominal dry solids inflow of 2,000 MTPD for the balance of the 13.5-year inferred mine

life • Solids specific gravity of 2.60 • Percent solids by weight of 50% • Tailings water deposition rate of 125 m³/hr at the nominal 3,000 MTPD • Tailings water deposition rate of 88.3 m³/hr at the nominal 2,000 MTPD • Fresh water make-up rate of 40 m³/hr • Reclaim rate of 83 m³/hr at the nominal 3,000 MTPD • Reclaim rate of 55.3 m³/hr at the nominal 2,000 MTPD • Tailings void ratio of 1.26. • AG waste rock, open pit = 7,000,000 tonnes, underground = 600,000 tonnes. • Waste rock dry density 2 tonnes/m3 • Contaminated site runoff (ore stockpile) = 9,000 m3/yr (average annual) • Domestic waste water from the treatment plant (average annual) = 55,000 m3/yr • Pit dewatering average annual inflows = 59,000 m3/yr • Pit wall thawback = varies from peak of 42,000 m3/yr in Year 1 of production, decreasing

to about 9,000 m3/yr in year 5, assumed negligible after Year 5. The tailings impoundment is designed as a closed-circuit process with no discharge, so the impoundment must have enough capacity to store average annual net water inflows, a 200-year return period wet year, and a Probable Maximum Flood (PMF) event with adequate freeboard during the mine life. The probability of a 200-year wet year occurring over the mine life is 5 to 10%. To allow for water to be added to the pond during spring runoff the west diversion ditch of Alternative # 1 or east diversion ditch of Alternative #2 will be constructed with a side-discharge release structure immediately upstream of the dam crest’s abutment. Releases through this facility will be conducted during the spring/summer months (May to September) on an annual or as needed basis. Water levels and climate forecasts will be continuously monitored to determine the volume of water released into the tailings impoundment and water management guidelines for the tailings facility will be developed during the mine’s operation. The planned releases into the pond will provide buffer storage for winter month water withdrawals while



maintaining the minimum 4-m depth over the tailings level. Additional make up water will be pumped from process water wells as needed. The average annual runoff coefficient is estimated to be 0.70 over the spring/summer months. The monthly distribution over this period is based on average monthly runoff estimates computed from site rainfall-runoff records. Average annual temperatures from October to April are below freezing so snowmelt runoff generated from accumulated precipitation over this period occurs in May, based on site recorded and regional station data. Direct precipitation on the pond will accumulate over the winter and thaw along with catchment runoff during the month of May. The annual precipitation estimate of 380 mm is based on site and the Markovo regional station data. Evapotranspiration and sublimation are accounted for in the catchment runoff with preliminary estimates of 120 mm and 70 mm, respectively. These values are based on previous arctic region studies conducted in northern Canada and the regions in the Siberia. Annual open water evaporation of 280 mm and its monthly distribution for the open pond area during the summer months are estimated from the Ostrovnoje regional station data in the absence of site-specific information. This station is expected to experience similar regional solar radiation exposure and humidity levels as the site. Seepage rates are assumed to be zero, as the dam embankment is lined with a bituminous geomembrane and the foundation soils are shown to freeze and to maintain its frozen state during summer months, based on thermal modeling (discussed in Section 6.1.5.1 and 6.2.5.1). A 15-year, monthly water/mass balance model was developed to simulate the hydrologic behaviour of the tailings impoundment. The detailed water balance spreadsheets for Alternatives #1 and #2 are presented in Appendix B. The baseline scenario of the model employs average annual precipitation, runoff and evaporation for the duration of the mine life. The mine life was derived from the two processing rates mentioned above and a total mined quantity of 12 million metric tonnes. The starter dam is assumed to be completed by October of the year prior to mine start-up, and precipitation runoff is expected to accumulate henceforth. Milling startup is scheduled for the following June. Inflow sources are from tailings solids and water, process water wells, catchment runoff, direct pond precipitation, ore stockpile drainage, open pit runoff and thawback inflows, and sewage treatment plant effluent discharge. Outflows consist of evapotranspiration, sublimation, evaporation, void losses and reclaim. It should be noted that the evapotranspiration and sublimation outflows are included in the net catchment runoff yield computations. 5.4 Tailings Management Tailings and water management will be carried out using procedures that will allow effective operation in the cold climate at Kupol. The primary strategy will be to maintain a water cover over the tailings surface during winter. Tailings will be discharged underwater, beneath the ice cover, to avoid freezing on exposed beaches. Freezing of tailings discharged on beaches could lead to large losses of water through freezing of slurry on the beaches. The water cover will also allow for process water reclaim during winter. This strategy of providing a water cover will be the primary criterion for tailings and water balance management. Regular monitoring of the water depth will be required to ensure that the operating criterion is being met. The water balance will be updated on a monthly basis throughout the year.



Decisions on water additions will be made based on the ongoing water balance monitoring. During the spring freshet, side discharges on diversion ditches may be opened to allow discharge into the pond on a controlled basis, in order to achieve target water depths at the beginning of winter, or conversely water balance observations may indicate that maximum diversion should be achieved to prevent water surpluses in the pond. Start-up of milling is scheduled for June 2008. Makeup water will be stored prior to start-up to allow the reclaim pump barge to begin operation, and also to provide water in storage for the following winter. Construction of the impervious liner and foundation cutoff are scheduled for completion in the fall of 2007. This construction schedule will allow runoff from the spring runoff in May and June of 2008 to be collected in the pond. Management of the diversions will be required to collect the target water volume. Tailings will be distributed by discharging at varying positions around the impoundment. The objectives of tailings distribution will be to:

• make full use of the tailings storage capacity by filling all areas of the pond; • provide additional seepage protection at the upstream toe and abutments of the dam by

discharging tailings preferentially along the crest of the dam, and • provide adequate depth at the reclaim pump barge at all times.

The alignment of tailings lines to achieve the above objectives is shown on Figures 329-003 and 004. For Alternative #1, the tailings line will be laid along the dam crest for initial operations and tailings will be discharged through a series of spigots. The spigot pipes will extend down the dam face to discharge directly into the water pond. In the second year of operation, a tailings line will be constructed to the back of the pond. Tailings will be discharged from a single point from this tailings line, rather than spigotting. This rear discharge point will be used primarily during summer months. In subsequent years of operation, the rear tailings line will be extended as the pond extends further up the valley. The reclaim barge will be initially positioned as indicated on Figures 329-003, and as the pond rises will retreat up the relatively steep slope on the north side of the tailings valley. For Alternative #2, the tailings line will also be positioned initially on the crest of the dam and tailings discharged by spigotting. In future years, the line will be extended to the far end of the dam and along the east side of the impoundment. The reclaim barge will be positioned at the northwest corner of the impoundment, and will retreat up the slope as shown on the Figure 329-004.



5.5 Waste Rock Management NAG and PAG waste rock mined from the open pit and underground will be used as dam fill. PAG used for the tailings starter dam construction will be placed in layers so that it is well incorporated with the NAG rock, so that available carbonates in the NAG will be available for buffering any acidity that may develop. All other waste rock classified as PAG or AG rock (based on blasthole assays) will be placed within the ultimate tailings impoundment. Placing the AG waste rock in the impoundment will ensure that any leachate generated from the waste rock is contained within the zero discharge tailings facility. Conceptual designs have been developed that will allow the waste rock to be placed in the tailings pond without significant initial capital cost impact, and minimal impact on pond operations or water balance. The conceptual plan for the disposal of AG waste rock in the tailings pond is to minimize any impact on starter dam volumes and the on-going operations of the tailings pond. For Alternative #1 the rock will be placed in the northeastern corner of the impoundment starting in pre-production. Once the dam is raised by mine production waste rock and sufficient storage capacity is available the AG and PAG waste rock will be placed into the pond. For Alternative #2, beginning in the pre-production period, a waste rock pile will be developed in the eastern arm of the Y-shaped valley in which the impoundment is sited. Small storage ponds will be developed to collect runoff from the waste rock. For both alternatives the maximum waste rock height will be 2 to 3 meters below the final tailings elevation, to ensure that the waste rock is completely encapsulated within the tailings and ultimately frozen in a saturated state. 5.6 Seepage Interception and Recovery The tailings impoundment for both alternatives is designed to be water-tight. Deep permafrost underlying the tailings sites is expected to provide an effective barrier to seepage from the impoundments. The dam will be constructed with an impervious liner, with the liner sealed to bedrock to form an effective barrier to seepage through or beneath the dam. It is expected there will be little or no seepage through or beneath the dam. However, there is some possibility that there may be some minor seepage, either through shallow bedrock beneath the seepage cutoff or through minor imperfections in the liner. To detect and monitor any potential seepage, a sump will be constructed downstream of the dam toe. The sump will be positioned to collect any water flowing downstream of the dam, either surface water from the downstream side of the dam or from valley slopes. The sump will be constructed with a cutoff to bedrock, or into permafrost below the active layer, to ensure complete interception of groundwater flow. Water in the sump will be sampled and tested regularly, as part of the site water quality monitoring program. If the water does not meet criteria for discharge to Class 1 fisheries streams, a pump will be installed in the sump and all water will be pumped back to the tailings impoundment during the summer months.



5.7 Diversion Ditches Diversions will be required to prevent catchment runoff from entering the tailings basins, under either tailings site alternative. Diversion ditches will be sized to handle statistically projected 200-year return period, 24 hour flood peak flows. The diversions will form part of the water management system and will be operated throughout the mine life. Side-discharge structures will be constructed in the ditches to allow release to the tailings impoundment during the freshet to provide makeup water for the following year. These water releases must be planned and monitored, to provide the amount of water indicated by pond level readings, pond soundings, and water balance modelling. The ditches will be constructed over two typical terrain types, and are designed for these terrain types as follows:

• For ditches in organic soils overlying ice-rich colluvium, the ditches will be constructed as shown on Typical Ditch Section A on Figure 329-010. In the winter prior to ditch construction (winter 2005-2006), ditch access roads will be constructed of general rockfill materials. The rockfill will have a minimum thickness of 2 m, and will be placed over frozen ground and is expected to provide sufficient cover to maintain freezing conditions beneath the fill. In the following summer of 2006, the ditch will be excavated upslope of the access road, using excavators working on the access road. Ditch spoil may be cast to the downslope side of the access road, and may be accessed at a later date for reclamation of the ditches for mine closure. The ditches should be over-excavated about 0.6 m beyond the final ditch dimensions to allow placement of rockfill on the ditch bottom and sides, to provide erosion protection and some thermal insulation for the ditch. In ice-rich soils the ditches may require some on-going maintenance as dictated by performance until a new geothermal equilibrium is achieved.

• For ditches in bedrock, the typical section is shown as Typical Ditch Section B on Figure

329-010. Ditches in rock will be excavated to form a trapezoidal or triangular section, as shown, and the excavated materials used to construct the ditch access road. It is expected that much of the excavation will be able to be achieved by ripping with a bulldozer, although some drilling and blasting will also be likely be needed.

5.8 Spillway The tailings impoundment is designed as a closed-circuit system and is capable of accommodating all net inflows including average annual rainfall and catchment runoff, a 200-year wet year, and a PMF event with minimum freeboard for wave run-up, without spilling during the operational mine life of 15 years.



At closure, a spillway designed to pass the PMF with adequate freeboard for wave run-up will be necessary to provide emergency releases should the net inflows accumulate enough over time such that the water level approaches the dam crest elevation, potentially overtopping the dam. As a preliminary and conservative estimate, the facility may be sized as a trapezoidal section along the right abutment, excavated in competent rock with a base width of 10 m, a side slope of 0.5H:1V and a minimum depth of 2.5 m.



6.0 ALTERNATIVE TAILINGS SITES 6.1 Alternative #1 6.1.1 Dam Siting Alternative #1 tailings site looking up-valley from the proposed dam location is shown on Figure 6.1.

Figure 6.1 - Alternative #1 Tailings Valley The dam alignment was optimized based on the geotechnical investigation results. Geotechnical conditions in the valley bottom generally consist of 1 to 2 m of ice-rich peat and silt, underlain by as much as 17 m of frozen granular (sand and gravel) fill. The upstream toe of the dam was shifted downstream of its initially selected location, to coincide with the location where the minimum depth of overburden is present over bedrock in the valley bottom. At the selected location, the depth of overburden, as determined by geotechnical drilling, is about 5 m. Abutment conditions consist of similar stratigraphy as the valley bottom with considerably reduced thicknesses of sand and gravel, silt and organics. Bedrock is exposed, or covered by a thin veneer of weathered bedrock, in some locations. The dam is designed with an upstream seepage cutoff trench anchored in bedrock. A starter dam will be positioned perpendicular to the valley, with successive dam raises that would angle the dam slightly up-valley. Figure 329-005 presents the geotechnical cross section of the valley at the dam centreline. The successive raises are angled to keep the dam from day-lighting over the right abutment. 6.1.2 Volume-Elevation Relationship Table 6.1 presents the required crest elevations for the 3-Year Starter Dam and the 12 M tonne Ultimate Dam.

Table 6.1 - Alternative #1 Required Dam Crest Elevation

Dam Tailings Elevation (masl) Dam Crest (masl)

3-Year Starter Dam 539 545

12 M tonne tailings plus waste rock - Ultimate Dam

556 561.5

Figure 6.2 presents a summary of the required tailings dam storage and the corresponding dam crest elevations. The storage elevation curve presented in Figure 6.2 incorporates the 3.5 Mm3 of AG waste rock produced during the open pit phase of mining and the ultimate dam crest is represented by the total volume stored of approximately 10,500,000 m3 of tailings plus 600,000 m3 of AG rock produced from the underground operation.



Figure 6.2 - Tailings Alternative #1 – Storage Elevation Curve Based on the design criteria for tailings, waste and water storage, and freeboard, the Alternative #1 3-year starter dam will be constructed to crest El. 545 m, for an embankment height of 33 m, as shown on Figure 329-006. The ultimate dam, designed to store 12 M tonnes of tailings, will have its crest at El. 561.5 m, for a maximum embankment height of 49.5 m. 6.1.3 Water and Materials Balance The most effective alternative for diverting and retaining net inflows into the tailings impoundment to adequately maintain a balance between required reclaim water volume and sufficient storage capacity is to maintain diversion ditches constructed around the perimeter, upslope of the tailings impoundment inundation level. The total diverted catchment area is approximately 6.4 km² out of a total 7.6 km². Two ditches are proposed for diverting the catchment runoff – west and east diversion ditches, as illustrated in Figure 329–003. The west diversion ditch will collect runoff from the west slope drainage area that comprises 4.7 km², while the east diversion ditch will collect runoff from the east slope tributary area of 1.7 km². Each ditch is designed to convey 24-hour, 200-year return period storm event flows around the respective abutments, upslope of the dam embankment and into the main stem of the creek at an appropriate distance downstream of the dam toe. The simulated results from the 14-year, monthly water/mass balance model (discussed in Section 5.3) suggest that annual releases into the pond from the west diversion ditch facility will be required for a few months a year between May and September. The inflow volumes will be required to initially accumulate sufficient start-up water capacity for the mill process, and to provide sufficient water depth for reclaim during the winter months. The model suggests that minimal to no external makeup water is needed beyond releases into the pond from the diversion ditches, but an external water supply should be available in the event that an extreme dry year occur, or should reclaim water be unavailable due to cold weather operating problems. The 3-year tailings and water level is estimated to be 539 m and 553.5 m, respectively. For the final year of operations, tailings and water levels are estimated to be 556 m and 560 m, respectively (the AG rock final elevation is predicted to be about El. 553 m at closure, completely buried by tailings). The dam crest staging should exceed these levels by at least 1.5 m to provide adequate freeboard to store a PMF storm event and to accommodate wind-generated wave run-up. Two additional scenarios are modeled with a 200-year wet year and a 200-year dry year to independently occur during the operational life of the mine. These extreme-year scenarios are selected to test the tailing’s impoundment’s ability to meet operation water needs and to determine its capacity to retain excess water volume with adequate freeboard. For the 200-year wet year, the annual precipitation was estimated to be 664 mm. Simulation results from the model suggest that water levels in the impoundment will rise between 0.4 m to 3.8 m over an



average annual year, depending on the maturity of operation. These water level increases can be mitigated with reduced diversion releases from the ditches during the year For the 200-year dry year scenario, the annual precipitation was estimated to be 91 mm. Simulation results from this model suggest that the water level will drop between 0.5 m to 3.5 m over an average annual year, depending on the operational stage. The reduction in water level can be mitigated with increased diversion releases over the spring/summer months and, at lower water levels, supplemented by make-up water from process water wells with up to 550 USGPM. 6.1.4 Diversion Ditches Alternative #1 has a catchment area of about 7.6 km2 upstream of the dam location. Diversions will be installed all around both sides of the tailings impoundment, at an elevation just above the projected ultimate storage level, to reduce the catchment area to about 1.2 km2. Both diversion ditches will be extended downstream past the limit of the tailings dam toe to discharge to the existing drainage course. The ditches will be constructed with a design gradient along the ditch bottom of 0.5%. These ditches will be designed to convey the 24-hour, 200-year storm event with a minimum 0.3 m of freeboard. HEC-HMS, a drainage basin runoff routing model, was used to route the rainfall event to develop peak design discharges for each tributary catchment at each ditch location. To be conservative, an SCS type II storm distribution was used to produce the largest peak for a given rainfall event and time of concentration. The input parameters for this model will be adjusted as more site stream flow, rainfall and catchment characteristic data are collected. For Alternative #1, ditches on the east side of the tailings pond (draining 1.7 km²) and on the west side (draining 4.7 km²) of the tailings pond will be constructed to collect catchment runoff and discharge it downstream of the dam toe, back into the unnamed creek. Based on the HEC-HMS model, the resulting ditch design flows for a 24-hour, 200-year storm event are:

• 3.2 m³/s for the west ditch • 1.8 m³/s for the east ditch

The resulting ditch cross-sectional configuration is trapezoidal with the dimensions presented in Table 6.3:

Table 6.2 - Diversion Ditch Dimensions

Ditch Location Base Width (m) Side-Slope Channel Depth (m)

West 1 m 1.5H:1V 1.3 m (Includes 0.3 m freeboard)

East 1 m 1.5H:1V 1.1 m (Includes 0.3 m freeboard)

A side-discharge release structure, based on either a gated or stop-log system, will be constructed in the west diversion ditch, just upstream of the dam crest to admit summer flows into the tailings impoundment to help maintain a water balance reserve for winter withdrawal



needs. The discharge facility will be sized to accommodate average annual spring thaw flows and will typically be operated on an as-needed basis to supplement mine operations water supply needs. 6.1.5 Design Analyses 6.1.5.1 Thermal Analyses Thermal analysis was carried out for a typical tailings dam section with the less favourable foundation conditions of Alternative #1. The dam section was modeled in stages to best represent the developing tailings facility. The geothermal modeling program SIMPTEMP, 2D version, (developed in-house by AMEC) was used to analyze geothermal regimes for three stages of the dam. The simulator uses the finite element method to compute a numerical solution of the heat transfer problem. Published physical/mathematical algorithms are used in the SIMTEMP model, and the simulation process has been verified: against well-known analytical solutions of the heat transfer problem, and compared with numerical solutions produced by other commercial/non-commercial geothermal software. AMEC has successfully used the SIMPTEMP program for a variety of geothermal applications over a period of approximately ten years. The complete thermal analysis report is attached as Appendix C.



The geothermal calculations are base on heat conduction in response to the in-situ ground temperatures and seasonally varying air temperatures. As the Kupol Dam section is constructed out of rockfill, it is possible that heat transfer from convection of cold (denser) air may slowly percolate through the downstream section. Experience with the rockfill dams in Russia and in northern Canada indicate that additional, and advantageous, cooling may result from this effect. The starter dam is scheduled to be constructed in the winter months to preserve the permafrost in the dam foundation soils. Figure 6.3 presents the starter dam thermal conditions at the end of winter construction.

Figure 6.3 - Starter Dam Thermal Status June Year – 1 (2007)

Figure 6.4 below presents the thermal state of the starter dam at the end of the first summer showing the impact of a summer’s warming. Figure 6.4 indicates that the seepage cutoff will remain in a frozen state after the first exposed summer thaw.

Figure 6.4 - Starter Dam Thermal Status October – Year – 1(2007)

Figure 6.5 presents the tailings dam thermal status after the first summer of tailings deposition, further illustrating that the seepage cutoff and the dam foundation will remain in a frozen condition during operations.

Seepage Cutoff Trench




Figure 6.5 - Tailings Dam Thermal Status October – Year – 1(2007)

The thermal analysis demonstrates that placement of cold rockfill in the winter will preserve the dam foundation in a frozen state. Maintaining foundation permafrost will prevent any thaw settlements in the foundation and provide an effective barrier for any seepage that passes the liner containment system. During the final design, recommendations for a suitable array of temperature measurement devices to be deployed in the dam and the foundation to validate the thermal predictions will be made. 6.1.5.2 Limit Equilibrium Stability Analyses Based on the data yielded by the thermistor approximately 120 m upstream of the Alternative #1 dam location, and confirmed by borehole sampling, the dam foundation soils are frozen beneath the shallow depth of active permafrost. The foundation soils underlying the downstream extension of the dam are expected to remain frozen (they will be buried more deeply due to fill placement and hence be further insulated against surficial seasonal thaw). Limit equilibrium stability analyses were carried out on the proposed dam raise geometry to confirm adequate factors of safety during operation and beyond, assuming unfrozen conditions in the dam foundation materials. Two-dimensional analyses were completed with the limit equilibrium slope stability program SlopeW™ (Version 5.02, Geo-Slope International Ltd). The analysis method used was the Morgenstern-Price method of slices with a half sine side function to calculate the inter-slice forces satisfying both moment and force equilibrium. The stability of the proposed dam raise to crest elevation 561.5 m was evaluated for the maximum reasonable embankment phreatic level (see dashed line on Figure 6.6). As the raised dam will be completely lined along the upstream face, the phreatic level within the embankment fill would not be expected to reach significant levels. Figure 6.6 below illustrates the dam geometry used in the analyses. This cross section is consistent with Section B-B’ (Figure 329-005) and represents the maximum height section through the dam.


Dam raise



The dam foundations are generally low ice content soils that are expected to stay frozen. The “long term” strength of frozen soil is frictional. In order to account for the unlikely possibility of minor thaw in the foundation a modest pore pressure parameter of Ru = 0.2 was assigned to the stability analyses. However, the critical failure surface did not engage this possible weak layer, and the critical factor or lowest factor of safety is as shown on Figure 6.3. There is a modest possibility of minor movements induced by creep in the ice poor foundation soils during the operational life and into the closure phase. These movements are not considered to be a significant design issue.

Figure 6.6 - Stability analysis section – Alternative #1 – Tailings Dam Table 6.4 gives a summary of the soil strength parameters used in the assessment. All materials were assumed to be Mohr-Coulomb frictional materials.

Table 6.3 - Summary of material parameters used in the stability analysis

Material Bulk Unit Weight

γγγγb (kN/m3) Effective Friction

Angle - φφφφ’ (degrees) Ru

Tailings 11 0 0 Rockfill 18 38 0

Foundation 18 32 0.2 The stability analyses results suggest that the embankment has an ample factor of safety under static and static conditions with the current design as shown in Table 6.5 below:

Table 6.4 - Summary of stability analysis results for the Main Dam

Analysis condition Factor of Safety

Static 1.62 Pseudostatic n/a – see below

6.1.5.3 Seismic Loading Conditions As described in Section 2.6, the regional seismicity of the site under Russian classification system is 6 points. For this level of regional seismic activity, seismic risk, SNiP II-7-81* states that detailed seismic design analyses are not required for design of the dams. Regardless of the seismicity classification, experience worldwide has shown that rockfill dams have high resistance to earthquake shaking. Because of the high shear strength of rockfill embankments, and the fact that they have no mechanism for loss of strength, rockfill dams have no risk of failure under earthquake loading. The only effect of earthquake may be some small deformations; well within the tolerance of liner systems should any such deformations occur.



6.1.5.4 Seepage Analyses The thermal modelling carried out for the tailings dam illustrates that the foundation soils at depth will remain in a frozen condition during construction and through operations, and will thereby preclude any deep seepage paths from the tailings impoundment. The critical components of the facility for seepage control will be the liner and liner anchor trench. The integrity of these will rely on stringent quality control during construction. Any seepage which may occur through liner defects or past the liner anchor trench is expected to be prevented from continuing past the dam by the frozen foundation and embankment. 6.2 Alternative #2 6.2.1 Dam Siting Tailings Dam Alternative #2 considers the construction of a tailings facility in the upper portion of the valley as shown on Figure 329-003. The location selected ties in to rock abutments on both ends of the dam. An on-going site investigation for Alternative #2 has shown that for the majority of the dam footprint, bedrock is exposed or is relatively shallow. Several holes drilled in the valley in the right of the proposed impoundment have encountered overburden up to 10 meters thick. The Alternative #2 tailings dam site, looking west along the dam alignment is shown on Figure 6.7.

Figure 6.7 - Alternative # 2 Tailings Valley 6.2.2 Volume-Elevation Relationship Table 6.5 presents the required crest elevations for the 3-year Starter Dam and the 12 M tonne Ultimate Dam.

Table 6.5 - Alternative #2 Required Dam Crest Elevation

Dam Tailings Elevation (masl) Dam Crest (masl)

3-year Starter Dam 534 539.5 12 M tonne Ultimate Dam 555 560.5 Figure 6.8 presents a summary of the required tailings dam storage and the corresponding dam crest elevations. The storage elevation curve presented in Figure 6.8 incorporates the 3.5 Mm3 of AG waste rock produced during the open pit phase of mining and the ultimate dam crest is represented by the total volume stored of approximately 10,500,000 m3 of tailings plus 600,000 m3 of AG rock produced from the underground operation.



Figure 6.8 - Tailings Alternative #2 – Storage Elevation Curve

Based on the design criteria established for water storage and freeboard, the tailings 3-year starter dam should be constructed to crest El. 540 m for an embankment height of 33 m as shown on Figure 329-007. The ultimate dam, designed to store 12 M tonnes of tailings at El. 555 m for a maximum embankment height of 53.5 m. 6.2.3 Water and Materials Balance The most effective alternative for diverting and retaining net inflows into the impoundment and adequately maintain a balance between required reclaim water volume and sufficient storage capacity is to construct perimeter diversion ditches upslope of the tailings impoundment inundation level. The total diverted catchment area is approximately 2.6 km² out of a total 3.5 km² drainage area. Two ditches are proposed for diverting the catchment runoff – west and east diversion ditches. Each ditch is designed to convey 24-hour, 200-year return period storm event flows. The west diversion ditch will collect runoff from the west slope drainage area that comprises 0.33 km². Because of the steep terrain on the west abutment, a smaller ditch is proposed to minimize extensive slope cuts. The ditch will drain south and discharge downstream of the dam toe. The east diversion ditch will collect runoff from the north and east slope tributary areas of 2.2 km². The ditch will drain eastward and discharge over a saddle point on the east catchment boundary upstream of the dam embankment. A side discharge facility will be constructed at the downstream end of this ditch to admit make-up water collected from catchment runoff into the impoundment. The simulated results from the 14-year, monthly water/mass balance model (discussed in Section 5.3) suggest that annual releases into the pond from the west diversion ditch facility will be required for a few months a year between May and September. The inflow volumes will be required to initially accumulate sufficient start-up water capacity for the mill process, and to provide sufficient water depth for reclaim during the winter months. The model suggests that minimal to no external makeup water is needed beyond releases into the pond from the diversion ditches, but an external water supply should be available in the event that an extreme dry year occur, or should reclaim water be unavailable due to cold weather operating problems. The 3-year tailings and water levels are estimated to be 534 m and 538.5 m, respectively. For the final year of operations, tailings and water levels are estimated to be 555 m and 559 m, respectively. The dam crest staging program should exceed these levels by at least 1.5 m to provide adequate freeboard to store a PMF storm event and to accommodate wind-generated wave run-up. Two additional scenarios are modeled with a 200-year wet year and a 200-year dry year to independently occur during the operational life of the mine. These extreme-year scenarios are selected to test the tailings impoundment’s ability to meet operation water needs and to determine its capacity to retain excess water volume with adequate freeboard. For the 200-year



wet year, the annual precipitation was estimated to be 664 mm. Simulation results from this model suggest that water levels in the tailings impoundment will rise between 0.5 m to 2.8 m over an average annual year, depending on the maturity of operation. These water level increases can be mitigated with no May or summer diversion releases during the year depending on the water level stage. I For the 200-year dry year scenario, the annual precipitation was estimated to be 91 mm. Simulation results from this model suggest that the water level will drop between 0.3 m to 3.6 m over an average annual year, depending on the operational stage. The reduction in water level can be mitigated with diversion releases over the entire spring/summer months and make-up water from process water wells up to 450 USGPM. 6.2.4 Diversion Ditches Alternative #2 will have two diversion ditches, one on the east side of the tailings pond (draining 2.2 km²) and one on the west side (draining 0.33 km²) of the tailings pond. The diversion ditches will be constructed to collect catchment runoff and discharge downstream of the dam toe. Based on the HEC-HMS model, the resulting ditch design flows for a 24-hour, 200-year storm event are:

• 0.68 m³/s for the west ditch • 2.4 m³/s for the north ditch

The resulting ditch cross-sectional configuration is trapezoidal with dimensions presented in Table 6.6:

Table 6.6 - Diversion Ditch Dimensions


West 1 m 1.5H:1V 0.8m (Includes 0.3 m freeboard)


A side-discharge release facility based on either a gated or stop-log system, will be constructed in the east diversion ditch, just upstream of the dam crest to admit summer flows into the impoundment to help maintain a water balance reserve for winter withdrawal needs. The discharge facility will be sized to accommodate average annual spring thaw flows and will typically be operated on an as-needed basis to supplement mine operations water supply needs. 6.2.5 Design Analyses 6.2.5.1 Thermal Analyses Thermal analysis was carried out on a typical dam section from Alternative #1, as discussed in Section 6.1.5.1. The results from that analysis are applied to the Alternative #2 site as the design criteria are similar. Preliminary investigation indicates that the seepage cutoff/anchor



trench for the Alt #2 dam may not be excavated as deep as what is required at the Alt #1 site. As such, it is recommended that a cover (1 to 2 m) of fine grained material be placed over the seepage cutoff anchor trench to help preserve the cutoff trench in a frozen condition through the first filling season. During the final design, recommendations for a suitable array of temperature measurement devices to be deployed in the dam and the foundation to validate the thermal predictions will be made. 6.2.5.2 Limit Equilibrium Stability Analyses Stability analyses were not carried out for Alternative #2. Alternative #1 was demonstrated to be stable on a soil foundation. Alternative # 2, to be constructed on a stronger, bedrock foundation, is certain to also be stable.



6.2.5.3 Seepage Analyses The thermal modelling carried out for the tailings dam illustrates that the foundation soils at depth will remain in a frozen condition during construction and through operations, and will thereby preclude any deep seepage paths from the tailings impoundment. The critical components of the facility for seepage control will be the liner and liner anchor trench. The integrity of these will rely on stringent quality control during construction. Any seepage which may occur through liner defects or past the liner anchor trench is expected to be prevented from continuing past the dam by the frozen foundation and embankment.



7.0 TAILINGS DAM SCHEDULING AND CONSTRUCTION This section applies to the construction of either of the considered alternatives. 7.1 Starter Dam The starter dam is designed to store approximately 3 years of tailings production, AG and portions of the PAG waste rock and provide storage for the required amount of mill process water. 7.1.1 Schedule A preliminary construction schedule for the tailings impoundment is shown on Figure 7.1. The schedule is based on an assumed start-up of milling operations in June 2008. Critical aspects of the schedule are summarized below:

• Access to the tailings dam site will be developed in the winter of 2005-2006 with the construction of a rockfill road of sufficient width and depth to maintain the road foundation in a frozen condition Construction access roads for the diversions will also be constructed in winter 2005-2006;

• In the summer of 2006 the dam foundation will be prepared and a rockfill quarry established. Diversion ditches will need to be in place in summer 2006 to divert runoff around the foundation construction operations in 2007;

• Starter dam rockfill embankment construction will be carried out in the winter of 2006-2007 so as to ‘lock-in’ the permafrost beneath the dam. The embankment must be completed prior to summer to allow for the summer activities described below;

• Production and placement of the geomembrane transition and bedding layers will be carried out in the summer of 2007, followed by placement of the bituminous geomembrane. The seepage cutoff trench and any associated bedrock cleaning, concreting and grouting will also be completed in summer 2007. The liner will be sealed into the cutoff trench and the trench backfilled. The dam must be complete in fall 2007, ready to store runoff during the spring thaw of 2008;

• All tailings and reclaim piping, the reclaim pump barge, powerlines, monitoring and instrumentation and all other components of the tailings system will be in place prior to start-up; and

• First filling of the tailings impoundment will commence in fall of 2007 with the majority of the water being captured during the spring 2008 spring thaw.


AMEC File: VM00329 Task 230 S:\PROJECTS\VM00329 - Kupol Gold Tailings & Geotech\Tailings Dam Design Task 230\Report\FINAL REPORT\Tailings Facility Feasibility Design Report_No Figures.doc Page 47

Figure 7.1 - Preliminary Construction Schedule – Tailings Impoundment



7.1.2 Foundation Preparation Foundation preparation for Alternative #1 will consist of removal of all ice-rich and/or organic soils. The upper 1 to 3 meters of bedrock has likely been affected by frost penetration and may require removal. All exposed foundation rock (abutments, seepage cutoff trench) is required to be inspected by the site Engineer to determine the rock quality. Rock deemed unacceptable will be removed and if the foundation conditions are irregular such to affect the effectiveness of the contact of the geomembrane, the rock may have to be treated with dental concrete or slush grout. 7.1.3 Temporary Collection Pond About 1,250,000 m3 of AG rock will be produced prior to the completion of the tailings dam. Runoff from the waste rock will likely be unacceptable for discharge and must be collected. For Alternative #2 the waste rock will be placed in the eastern valley of the impoundment, a small water retaining dam will be required to capture and store the runoff from the waste rock. For Alternative #1 the AG waste rock will be placed in the bottom of the impoundment, runoff from the waste rock will be collected behind a temporary waste pile runoff collection structure discussed in the following section. 7.1.4 Construction Diversion For Alternative #2 a temporary, field-fit, construction diversion ditch will be constructed upstream of the dam foundation area to divert flows from the remaining western catchment downstream of the permanent diversion ditches. For Alternative #1 a small diversion dam or sump and pump system will be constructed to collect the runoff from the impoundment. A geomembrane lined cofferdam will be required to divert stream flows into the diversion ditch. The ditch and cofferdam system or sump and pump system will be designed to the same risk level of an annual exceedence probability (1%) of an equivalent 100-year return storm prorated to the operation period of the diversion facility. The coffer dam crest will be constructed to maintain a 1-m freeboard during its operation to accommodate excessive stream flows beyond the design flow. This temporary diversion system will be in place at each respective alternative site during the starter dam seepage cut-off wall construction. Once the liner installation and cut-off wall construction are completed, the cofferdam will be breached or the sump abandoned and the stream discharge will be permitted to return to its original creek alignment. The permanent diversions will also be diverted into the impoundment to allow collection of water for start-up. 7.1.5 Fill Materials – Borrow Sources, Specification and Placement The dam fill zones are identified on Figures 329-006 and 009. Material and placement procedures are summarized below: 7.1.5.1 Bituminous Geomembrane A bituminous geomembrane will be placed on the upstream face of the dam to provide the impervious element. Bituminous membranes are manufactured by impregnating geotextile with



bitumen. The material has been shown to have superior resistance to puncture and tear as compared to plastic geomembranes, is superior in cold climates as it maintains flexibility to very low temperatures (-70°C) and can be placed and seamed in low temperatures (to -25°C), and it has almost indefinite longevity. Bitumen-saturated fabrics have been found intact in archaeological sites in Mesopotamia, indicating the material will last for millennia. Bituminous liners do not require a high degree of skill to place and seam, as the sealing process consists of simply heating the surfaces with a propane torch. The material specified for the Kupol tailings dam is Coletanche ES-3, material that has a thickness of 4.84 mm and is used in applications requiring toughness and durability. The material can be sealed directly onto rock or concrete surfaces to form an effective seal. The liner for the Kupol dams will be laid directly on the bedding layer and be covered will a soil layer to protect it from being damaged by the winter ice sheet. 7.1.5.2 Zone 1 – Rockfill Embankment The general dam fill will consist of open pit waste rock. The general rockfill should have a maximum particle size of 1000 mm. The material will be placed in horizontal lifts, with Lift thicknesses a maximum of 1.5 m, and will be compacted by travel of the hauling equipment. Dam fill will consist either of NAG or PAG rock. Blasthole assays will be used to control geochemical quality of rockfill. No AG rock will be placed in the dam The starter dam fill quantities for Alternatives #1 and #2 utilizing a bituminous geomembrane are presented in Table 7.1. 7.1.5.3 Zone 2 – Transition Zone Below the bedding layer a transition zone (Zone 2) is required to provide an intermediate size between the bedding material (Zone 3) and embankment rockfill (Zone 1). The transition zone will consist of visually selected or screened blasted NAG rock with a maximum particle size of 200 mm. The material will be placed in a single 1.5 m lift parallel to the slope. 7.1.5.4 Zone 3 – Bedding Layer The bedding zone (Zone 3) will consisting of granular fill with a maximum particle size of 75 mm. The material may be processed from quarried NAG rock, or alternatively sourced from alluvial deposits in Kaiemraveem Creek. The material will be placed in a single, 500 mm lift parallel to slope. Following placement the surface will be smoothed to prepare for liner placement.



Table 7.1 - 3 Year Starter Dam Rockfill Quantities Alternative #1 and #2

Starter Dam Rockfill Volume (m3) Transition (m3) Bedding (m3)

Alternative #1 645,000 41,000 17,000 Alternative #2 1,525,000 81,000 32,000 7.1.5.5 Zone 4 – General Fine Grained Fill The seepage cutoff/anchor trench will be backfilled with general fine grained fill (Zone 4). This general fill may consist of dam foundation strippings or other fine-grained material. The material should be placed without damaging the liner, with the objective being to backfill the cutoff trench to provide an extra line of seepage control upstream of the cutoff trench. 7.2 Dam Raising The current mine plan shows that the open pit will be completed at the end of Year 3. NAG and selectively placed PAG waste rock produced from the open pit will be used as dam fill and the general dam fill (Zone 4) may be completed to the projected final elevation at the end of Year 3. Zones 2 and 3 can be processed and placed and the geomembrane extended as required by the rising pond level. 7.3 Construction Quality Control and Quality Assurance The tailings dam and associated facilities must be constructed with a high level of quality control and quality assurance (QA/QC) to ensure that the facility performs satisfactorily. It is anticipated that all critical aspects of the work will be carried out under monitoring by a representative of AMEC, as the designer of record. The QA/QC work for the various components of the impoundment is summarized in Table 7.2 below.



Table 7.2 - QA/QC Requirements

Testing Component Critical aspects Fulltime monitoring

by designer? Test Type Frequency

Foundation preparation

Remove all organics to the level of dense soil or bedrock

Yes Visual observation Continuous

Seepage cutoff trench

Excavate to the level of intact bedrock, thoroughly clean all loose materials

Yes Visual observation, photographic record and mapping,

Continuous

Rockfill Zone 1 Place in lifts not exceeding 1.5 m, remove any material exceeding 1 m diameter. Compaction with haulage equipment.

No Lecos Furnace testing to confirm geochemical classification, visual observation of lifts and compaction, photographic record

Transition Zone 2 Production

Blasted rock, maximum particle size 200 mm

No Visual observation Daily

Transition Zone 2 Placement

Place in single 1.5 m lift parallel to slope

No Control thickness with fill stakes

1 stake every 1000 m2

Bedding Layer Zone 3 Production

75 mm minus No Gradation analysis 1 test per 1000 m

Bedding Layer Zone 3 Placement

Place in single 500 mm lift parallel to slope. Compact surface with roller to prepare for liner

Yes Control thickness with fill stakes

1 stake every 1000 m2

Seepage cutoff grouting (if required)

Detailed program to be designed if required

Yes Detailed program to be developed

Continuous

Seepage cutoff dental concrete

Create smooth and level surface for liner attachment. Place concrete on cleaned bedrock surface

Yes

Bituminous liner Liner to be laid with minimum 500 mm overlap, overlap to be sealed by “torching”, liner surface to be free

Yes Visual observation and photographic record

Continuous



Testing Component Critical aspects Fulltime monitoring

by designer? Test Type Frequency

of defects Liner seal to cutoff trench

Liner to be “torched” to smooth bedrock or concrete surface over full design width

Yes Visual observation and photographic record

Continuous



8.0 MONITORING AND INSTRUMENTATION The following monitoring will be carried out during mine operations:

• Thermal monitoring. Thermistors will be installed in the tailings dam and foundations to monitor temperatures, to provide comparison to the thermal performance predicted by modeling. A minimum of 5 thermistors will be installed in drillholes through the crest of the dam, extending a minimum of 10 m into the dam foundations. These will be read on a monthly basis for the pre-production period, and for the first 3 years of operation, after which the monitoring frequency will be reviewed.

• Movement monitoring. Survey monuments will be installed along the downstream crest of the tailings dam. The monuments will consist of a steel rod cast in the top of concrete monuments that are placed to a depth of at least 2 m in the rockfill. The monuments will be surveyed for vertical and horizontal movements on a monthly basis for the first year of operation, following which the surveys will be done quarterly.

• Seepage monitoring. If any seepage is detected downstream of the dam, flow measuring weirs will be installed to measure flows on a weekly basis during the summer season. Water quality samples will also be taken weekly and analysed for critical constituents. If the water quality is not acceptable for release, it will be captured in the seepage catchment sump and pumped into the tailings pond.

• Tailings pond water levels. The tailings pond water level will be recorded on at least a monthly basis.

• Water balance. All inputs and outflows contributing to the water balance will be continuously monitored, and the tailings pond water balance will be updated monthly. The water balance will be critical in meeting water management objectives.

• Tailings pond water quality. Samples of tailings pond water will be taken weekly and analysed for total and dissolved solids, cyanide, thiocyanates, nitrates, pH, and other critical constituents.

• Tailings beach surveys and bathymetry. Tailings surface surveys will be undertaken annually to determine the deposited volume and allow calibration of the actual deposited tailings density, and also to calibrate and re-set the water balance annually.



9.0 COST ESTIMATE The cost estimates presented in this report are based on current assumptions:

• All costs associated with earthworks are to be provided by Bema • Costs presented are current at the time of this report issue.

The purpose of these estimates is to provide a baseline estimate for comparison purposes should changes be made through the course of the project. Capital cost estimates were prepared for Alternatives #1 and #2. Quantity takeoffs and cost estimates were prepared for liners, instrumentation and other civil components, with some exclusions. Earthworks, pipelines, pumps, plant facilities, water well requirements are to be costed by others with access and haul roads requiring further clarification. All costs are expressed in $US. Summarized capital cost estimates with certain identifiable indirect costs are presented below for the 3 year dams with bituminous geomembrane liner. The cost estimates are based on typical unit costs for northern mining construction from AMEC’s experience in North America with adjustments made for the lower Russian labour costs. The unit cost estimates are preliminary and represent a best guess for anticipated costs. Unit costs should be further developed in conjunction with BEMA. Quantities and capital cost estimates for Alternatives #1 and #2 are presented in Figures 9.1 and 9.2, respectively.



Figure 9.1 - 3 Year Starter Dam Quantity and Cost Estimate Alternative #1



Figure 9.2 - 3 Year Starter Dam Quantity and Cost Estimate Alternative #2



10.0 CLOSURE PLAN 10.1 Closure Objectives At the end of mine life, a closure plan will be implemented for the tailings impoundment, with the following objectives:

• To leave the impoundment in a stable condition, so that the tailings and AG waste rock will be permanently contained;

• To prevent acid generation of sulphide tailings and waste rock; • To provide a permanent, stable stream channel and spillway for conveyance of surface

runoff through the tailings impoundment; and • To ensure that all water discharges from the tailings impoundment will meet discharge

standards for protection of fish-bearing waters downstream. The anticipated conditions at the end of mine operation will be as follows:

• The tailings dam will have a height of about 49.5 m (Alternative #1) or 53.5 m (Alternative #2). The dam will be constructed of rockfill which will have ample geotechnical stability for the long term. The dam would also be resistant to long term erosion;

• The tailings pond will have a water cover of about 4 m, which is the planned operating

condition over the life of the mine. A cyanide destruction circuit will operate during the mine life to reduce cyanide to low levels. Similarly, a process for reduction of thiocyanate will be operated during mine life. It is possible that, in spite of treatment, there may be some low concentrations of residual cyanide and/or thiocyanates, and possibly some other chemicals such as ammonia or nitrate, remaining in the free water pond at the end of mine life;

• The deposited tailings are expected, from the results of thermal modelling presented in

Appendix C, to freeze from the top down and bottom up over a period of 20 to 30 years; • Diversions will be in place at the end of mine operation and they will have been

maintained during the operating life. The diversions are being designed to convey a flood estimated as a 1 in 200-year return period. The diversions are not considered permanent following closure and would be expected to eventually fail through lack of maintenance or exceedence of design conditions;

• The pond will have a positive water balance after operations cease, and there will be a

surplus of water that will have to be released from the pond; and • There will be a number of site access roads and pipe corridors that will have been

maintained through the operating life of the mine and require decommissioning.



10.2 Closure Strategy Following is the strategy that will be implemented for closure of the tailings pond.

1. The tailings pond water quality will be monitored to determine whether it is suitable for release. If the water meets all parameters for discharge, excess water will be released from the pond. Water release will take place over a period of time, through a planned and regulated process. Water releases would be timed to coincide with peak flow periods in Kaiemraveem Creek, at discharge rates which would be a defined percentage of the flow of Kaiemraveem Creek. The details of release rates and allowable discharge parameters would be defined at later stages of the project, to meet specific permit requirements.

2. Excess tailings pond water would not be released upon cessation of mining and milling

activities if it did not meet effluent discharge standards. The water would be held until suitable for release. Improvements in water quality would be achieved through natural degradation and, if necessary, through treatment. Experience at other northern gold mines (INAC Colomac 2004) has shown that, in spite of short summers, significant natural degradation of cyanide occurs in tailings ponds during the ice-free period. Free cyanide exists only at high pH levels, which are maintained through lime addition in gold cyanidation circuits. In an exposed tailings pond, pH levels drop naturally through the mechanism of CO2 exchange. As the pH drops the level of free cyanide also drops. This mechanism has been well demonstrated. Following permanent shutdown of the processing plant, the process of natural degradation of cyanide would be monitored. Should this natural process not be sufficiently effective, treatment would be enhanced through continued operation of the cyanide destruction and thiocyanate treatment circuits, followed by release. The treatment and release strategy would be either to treat and directly release the treated water, or following treatment the water would be stored to allow further natural degradation, and release would follow the same planned and regulated process as described in point 1 above

3. Results of the analyses completed on the aged decant liquids from the Mixed Iron CND

Tailings sample show that after ageing for 90 days, mercury reported at concentrations greater than the World Bank Guidelines for Mining and Milling Effluent. Results of the analyses completed on the aged decant liquids from the High Iron CND Tailings show that mercury, arsenic, and copper reported at concentrations above the World Bank Guidelines in the Day 3 and Day 15 samples (Day 45 analysis results are pending). Simple water treatment is expected to be able to reduce the mercury, arsenic, and copper in the supernatant tailings waters prior to discharge upon closure.

4. The tailings pond water release will likely take place over one or more years following

cessation of mine operations. The objective of the release will be to draw down the pond water to a minimum level, to create a small remaining water pond. A permanent discharge spillway will be constructed to maintain the pond at this low level for closure. For Alternative #1 tailings pond, the spillway would be constructed along the left (north) side of the pond and adjacent to the north abutment of the dam. For Alternative #2, the discharge spillway would be cut through the bedrock ridge on the northeast side of the pond and would be separate from the dam and waste rock pile.



5. The tailings pond surface will be monitored following shutdown to assess the rate of ongoing consolidation settlement. Settlements may occur for several years, but it is anticipated that consolidation will not run its full course as the tailings mass will be frozen before complete consolidation occurs. The process of freezing and consolidation will be monitored through a combination of surface settlement gauges, piezometers and thermistors. The purpose of monitoring the consolidation and settlement process will be to assess the final surface contours of the tailings pond for design of drainage courses and closure cover

6. During and after the ongoing process of consolidation and freezing, a closure cover will

be constructed over the tailings surface. The cover will serve the following functions:

• To provide protection from surface erosion of the tailings. Additional erosion protection will be provided in the main drainage courses on the tailings surface

• To provide a thermal barrier to prevent seasonal thawing of the tailings, for

prevention of oxidation of sulphides. Published data and site investigations at the site indicate that the mean depth of active permafrost on the site (the depth to which seasonal thawing and freezing occur) is about 1.2 m. The cover thickness required to provide the equivalent depth of this active permafrost zone will depend on the thermal properties of the cover material. It is anticipated that the cover material will be quarried, NAG rockfill. However, the final cover design can best be determined through field trials during the mine’s operating life. Holubec, 2004 documents four sites in Canada that depend on covers over reactive tailings in permafrost regions, all of which are warmer sites (higher mean average annual temperature) than Kupol and all of these sites are depending on development of permafrost to prevent oxidation of reactive tailings.

On completion of the closure cover and construction of erosion protection on main drainages through the pond, the runoff diversions will be taken out of service and the main drainages routed through the tailings pond and spillway

After a period of monitoring to confirm that the tailings surface cover and all drainage courses are stable, the site access roads and pipeline corridors will be regraded to mimic the natural topography, and reclaimed with native vegetation to minimize the potential erosion that could occur from exposed surfaces in such permafrost terrain. The above closure plan is considered preliminary and conceptual. Details of the plan can best be finalized during mine operations and through closure trials. However, the plan is considered to be robust. There is little doubt that the tailings deposits will become totally frozen within a short number of years, so there will be no ongoing concerns with respect to quality of pore water discharges or oxidation of sulphide tailings. Spillways will be constructed in rock so will be permanent and durable, preventing any long term erosion of tailings.



10.3 Closure Monitoring Following closure, monitoring will continue as follows:

• Thermal monitoring. Thermistors will be maintained in the tailings dam for several years after closure, until equilibrium temperatures are established. A number of thermistors will also be installed in the tailings beach and foundations to monitor the progress of freezing of the tailings. The number and locations of these thermistors will be established at the time of closure.

• Movement monitoring. Survey monuments on the tailings dam will be maintained as necessary until it is established there are no ongoing movements.

• Seepage monitoring. If any seepage is detected downstream of the dam, flow measuring weirs will be maintained to measure flows on a quarterly basis during the summer season. Water quality samples will also be taken quarterly and analysed for critical constituents.

• Tailings pond water quality. Samples of tailings pond water will be taken quarterly and analysed for total and dissolved solids, cyanide, thiocyanates, nitrates, pH, and other critical constituents, until such time as the water quality is shown to be of no concern.

• Tailings beach surveys and bathymetry. Tailings surface surveys will be undertaken annually to monitor the progress of consolidation settlement of the tailings.



11.0 FURTHER SITE INVESTIGATIONS REQUIRED It is recommended that additional site investigation be carried out in the summer if 2005 to finalize the tailings disposal alternative selection and confirm the conditions assumed in the feasibility report. It is possible that condemnation drilling has been completed prior to the summer site investigation and a decision has been made to identify the preferred alternative. The site investigation should consist of a geotechnical drilling program, installation of additional thermistors and monitoring stations, detailed ground survey, detailed geological and geomorphological mapping and geophysical survey. Geologic/Geomorphologic Mapping: AMEC proposes that a senior geological engineer map the surficial geology of the tailings dam sites and proposed diversion ditch alignments. Geotechnical Drilling: Geotechnical drilling will be required to develop a thorough understanding of the tailings dam foundation conditions for Alternative #2. Survey: Ground surveys with sufficient accuracy to produce 1 m contour intervals should be completed in the following areas as a minimum:

• Alternative #2 tailings dam location; • Alternative #2 tailings impoundment up to and including the diversion ditches;

and • Access and construction road alignments.

Geophysical Survey: Geophysical surveys should be undertaken to assist in identifying depths of overburden overlying bedrock, possible ice lenses and the fracture/competent bedrock horizon. At Julietta Mine, ground penetrating radar (GPR) was used effectively to locate ice lensing in the tailings dam area. Thermistor installations: Thermistors should be installed in drill holes in the proposed tailings dam locations. The temperature data from these thermistors will be of vital importance to verify that the thermal conditions predicted by thermal modelling exist. Hydrologic monitoring: The hydrology of the tailings basin for Alternative #2 needs to be better understood to assess whether there will be an excess flow in the tailings valley catchment. A stream gauge station should be installed upstream of the northernmost diversion ditch. A site evaporation station should also be established.



12.0 LIMITATIONS AND CLOSURE Recommendations presented herein are based on a geotechnical evaluation of the findings of the site investigation noted. If conditions other than those reported are noted during subsequent phases of the project, AMEC should be notified and be given the opportunity to review and revise the current recommendations, if necessary. Recommendations presented herein may not be valid if an adequate level of review or inspection is not provided during construction. This report has been prepared for the exclusive use of Bema Gold Corp. for specific application to the area within this report. Any use which a third party makes of this report, or any reliance on or decisions made based on it, are the responsibility of such third parties. AMEC accepts no responsibility for damages, if any, suffered by any third party as a result of decisions made or actions based on this report. It has been prepared in accordance with generally accepted soil and foundation engineering practices. No other warranty, expressed or implied, is made. Respectfully submitted, AMEC Earth & Environmental, a division of AMEC Americas Limited

Reviewed by:

Per Daryl Dufault, EIT Geotechnical Engineer Earth & Environmental

Peter C. Lighthall, P.Eng. Vice President, Mining Earth & Environmental

Ed McRoberts, Ph.D., P.Eng. Chief Technical Officer



REFERENCES

Colomac Site Remediation Plan Final Report, Indian and Northern Affairs Canada, March 2004. SGS Lakefield (2005a) – Mineralogical Examination of Mixed Iron CND Tailings Fresh Solids from the Kupol Project, prepared for Bema Gold Corporation, LR10713 – MI5010-JAN05, March 3, 2005, 9 pages SGS Lakefield (2005b) – Environmental Characterization and Stability Testing of Metallurgical Tailings, Interim Report 2; Kupol Project. SGS Lakefield Reference No. 10716-002, May 17, 2005. Holubec, Igor (2004). Covers for Reactive Tailings Located in Permafrost Regions Review, MEND Report 1.61.6, Natural Resources Canada, Minerals and Metals Sector, 111 pages, October 2004. Atkin, S. 2004. Geochemical Characterization of Waste Rock and Ore Materials at the Kupol Project, Russian Far East. Prepared for BEMA Gold Corporation: Vancouver, BC, Canada.

Appendix A

Bituminous Liner Technical Information Document

Appendix B

Water Balance Alternatives #1 and #2

Appendix C

Thermal Analysis Report

Appendix D AMEC Wasterock Management Report



Report on Waste Rock Management Feasibility Design

Kupol Project Chukotka Autonomous Okrug, Far East Russia

Submitted to:

Bema Gold Corporation Vancouver, B.C.

Submitted by:


Burnaby, BC

03 June 2005

AMEC File: VM00330

Bema Gold Corporation Kupol Project, Chukotka Autonomous Okrug, Far East Russia Report on Waste Rock Management Feasibility Design 03 June 2005

Page ii

TABLE OF CONTENTS

Page

SUMMARY.................................................................................................................................... 1

1.0 INTRODUCTION.................................................................................................................. 3

2.0 WASTE MATERIAL CHARACTERIZATION........................................................................ 42.1 General........................................................................................................................42.2 Summary of Static Testing Program ........................................................................... 4

2.2.1 Site Geology .................................................................................................... 42.2.2 Description of the Static Testing Program ....................................................... 52.2.3 Results of the Static Testing Program ............................................................. 5

2.3 Summary of Kinetic Testing Program.......................................................................... 62.4 Estimated Quantities of Waste Rock........................................................................... 7

3.0 WASTE MANAGEMENT STRATEGY ................................................................................. 93.1 General........................................................................................................................93.2 Waste Characterization and Control ........................................................................... 93.3 Waste Pile Site Preparation ........................................................................................ 9

3.3.1 Foundation Preparation/Waste Pile Stability ................................................... 93.3.2 Diversion Ditching............................................................................................ 93.3.3 Collection and Treatment of Waste Rock Effluent......................................... 10

3.4 AG Placement in Tailings Impoundment ................................................................... 103.5 NAG/PAG Placement in Tailings Dam ...................................................................... 10

4.0 MONITORING.................................................................................................................... 12

5.0 CLOSURE.......................................................................................................................... 13

6.0 LIMITATIONS..................................................................................................................... 14

REFERENCES ........................................................................................................................... 15

LIST OF FIGURES

Figure 330-001 Tailings Impoundment Alternatives Site Plan Figure 330-002 Tailings Alternative #1 Waste Rock Disposal End of Pre-Production Figure 330-003 Tailings Alternative #2 Waste Rock Disposal End of Pre-Production Figure 330-004 Tailings Alternative #1 Waste Rock Pile Closure Plan Figure 330-005 Tailings Alternative #2 Waste Rock Pile Closure Plan


Page iii

LIST OF TABLES

Table 1.1 – Waste Rock Quantities by ARD Classification...........................................................1 Table 2.1 – Proportions of Acid Generating, Uncertain, and Acid Neutralizing Potential for each

of the Resource Model ARD Codes.......................................................................................6 Table 2.2 – Waste Rock ARD Code Percentages ........................................................................7 Table 2.3 – Total Waste Rock (tonnes)Produced from Open Pit and Underground.....................8 Table 2.4 – Total Waste Rock (tonnes) Requiring Surface Handling............................................8 Table 2.6: – Waste Rock Volumes by ARD Classification ............................................................8


Page iv

IMPORTANT NOTICE

This report was prepared exclusively for Bema Gold Corporation by AMEC Earth & Environmental, a wholly owned subsidiary of AMEC Americas Limited. The quality of information, conclusions and estimates contained herein is consistent with the level of effort involved in AMEC services and based on: i) information available at the time of preparation, ii) data supplied by outside sources, and iii) the assumptions, conditions and qualifications set forth in this report. This report is intended to be used by Bema Gold Corporation only, subject to the terms and conditions of its contract with AMEC. Any other use of, or reliance on, this report by any third party is at that party’s sole risk.


AMEC File: VM00330 S:\PROJECTS\VM00330 - Kupol Hydrology\WASTE ROCK FACILITY\wr report\Kupol Feasibility Waste Rock Management Feas_Design 03_June_05_FINAL.doc Page 1

SUMMARY

This report presents the feasibility design waste rock management plan for the Kupol Gold Project. From feasibility study mine planning by Wardrop, the total waste rock volume to be produced from the Kupol open pit and underground mining is estimated to be about 17 million tonnes and 1.5 million tonnes respectively. Detailed geochemical analyses have been completed on the waste rock. Over 200 representative waste samples were taken from drill core. Acid-base accounting was carried out for initial characterization, followed by kinetic testing (humidity cells) on a number of selected samples. This geochemical program, presented in detail in a separate report by Steve Atkin, indicated that a significant amount of the open pit waste materials is likely acid generating (AG), some material will be potentially acid generating (PAG) and the balance will be non acid generating (NAG). Acid generation can lead to leaching of metals, which can have the potential to impact downstream drainages. Therefore, it is prudent that a sound waste handling strategy is developed to prevent acid rock drainage. From the block model of the feasibility mine design, Table 1.1 shows the quantities of waste rock materials by ARD classification.

Table 1.1 – Waste Rock Quantities by ARD Classification

Quantity AG Rock PAG Rock NAG Rock

Mass 7.6 M tonnes 2.0 M t 8.7 M t Volume

(assuming 2.0 t/ m3) 3.8 Mm3 1.0 M m3 4.3 Mm3

The strategy proposed for waste rock management is that all acid-generating rock (AG) will be placed within the tailings impoundment and ultimately be submerged under tailings and water. Non-acid generating rock (NAG), and the majority of the potentially acid generating (PAG) rock, will be used to construct the tailings dam. The remaining PAG rock not used in dam construction will be placed with the AG waste in the tailings impoundment. At this stage of project development, two alternative tailings disposal sites are being considered:

• Alternative #1 tailings impoundment site is located southwest of the plant site in the valley of a right tributary of the Middle Kaiemraveem River to the southwest of the Kupol lease area. The tailings valley is capable of supporting future dam crest raises to increase the impoundment storage capacity if required

• Alternative #2 tailings impoundment is located immediately south of the plant site on the

southern part of Granichnaya Mountain, in the valley of Tretyi Creek. The impoundment location is such that there is limited potential to create additional storage capacity.



The two tailings disposal alternative areas developed to feasibility level are shown on Figure 330-001. Alternative #2 is the favoured site because of a shorter waste haul and shorter tailings discharge and reclaim lines. However, condemnation drilling work is still under way to confirm there is no economic mineralization beneath this site. Conceptual designs have been developed that will allow the AG waste rock to be placed in the tailings pond, with minimal impact on pond operations or water balance. The waste rock will be placed above the tailings starter pond level initially, so that disposal of AG waste rock in the tailings pond will minimize any impact on starter dam volumes and the on-going operations of the tailings pond. At the end of the life of the tailings impoundment, the waste rock will be covered by tailings. Placing AG rock in the tailings impoundment will provide the following benefits:

• All drainage from the AG waste piles will be collected and, as necessary, treated during the pre-production period, and all drainage from the AG waste piles following completion of tailings dam construction, and throughout the operational period will report to the tailings pond and be used as process water. Thus, there will be no release of any acidic drainage from the AG waste piles

• All AG waste rock will be enveloped within the saturated tailings mass. The waste rock

will eventually freeze as permafrost moves into the system following mine closure, so the material will be completely encapsulated on closure. This will provide a sound closure system with all acid generating material saturated and frozen.

Monitoring will be implemented to confirm that the objectives of acid drainage prevention and control are being achieved. Monitoring will include the following:

• Water quality sampling and testing downstream of the AG waste rock pile until the time of completion of the tailings dam in fall 2007

• Water quality sampling downstream of the tailings dam

• Quality control sampling to check that the waste material control system is properly characterizing waste materials

• Thermistors in the AG rock pile to monitor its temperature profile

• Thermistors in the tailings dam to monitor internal temperatures

• Tracking of waste rock quantities by mine operations engineering personnel with

respect to predicted and actual AG, PAG and NAG materials, and regularly updates of the material balance to ensure that the overall plan and strategy are maintained

• Settlement gauges installed on the waste rock pile to monitor its stability



1.0 INTRODUCTION This report presents the feasibility design for the disposal of waste rock for the Kupol Gold Project, in Chukotka Autonomous Okrug, Far East Russia. The design is presented to a level to support a project feasibility study being prepared by Bema Gold. The proposed Kupol Gold Project entails development of a high-grade gold and silver mine. The project site is located in the Chukotka Autonomous Okrug, at latitude 66°47’ slightly above the Arctic Circle, and longitude 169°19’. The project is located in a very cold region of continuous permafrost. Access to the site is via helicopter year-round and by winter road in the months of December through March from Bilibino, a city about 200 km northwest from the site. Major equipment and supplies for the project will be transported by ship during the summer to the East Siberian Sea port of Pevek, which is about 325 km northeast of Bilibino. An airstrip is planned to be constructed as part of the project to allow year-round access by fixed-wing aircraft. The Kupol Mine will be developed initially as a small open pit to access near-surface ore, with underground mine development to also start at an early stage of the project, so the mine will transition in its early years to underground mining. A gold milling and beneficiation plant will be constructed with a design nominal throughput of 3000 tonnes per day. This report presents feasibility designs for the disposal of waste rock as it applies to two tailings sites: Alternative #1, located southwest of the plant site in the valley of a right tributary of the Middle Kaiemraveem River, and Alternative #2, located immediately south of the plant site on the southern part of the Granichnaya Mountain, in the valley of the Tretyi Creek. This report is based on the following information:

• Geochemical characterization of the waste rock materials completed by geochemical consultant Steve Atkin (Atkin, 2005)

• Waste rock quantities and scheduling prepared by mine design consultants Wardrop



2.0 WASTE MATERIAL CHARACTERIZATION 2.1 General The Kupol deposit is a linear vein structure that runs approximately north-south for a strike length of 2 to 3 km, dipping steeply to the east and open to depths in excess of 250 m. The deposit will be mined initially by open pit, for about the first three years of mine life. Underground mine development will proceed in parallel with open pit mining and ore production will transition to underground over the early years of mining. The largest volume of mine waste rock will be generated primarily by the open pit, with lesser amounts from the underground mining phase. There is little overburden on the open pit site, and it is not expected that any overburden will be separated or stockpiled. 2.2 Summary of Static Testing Program In August 2003, Bema Gold commissioned a geochemical investigation to characterize the potential for acid rock drainage (ARD) and metals leaching (ML) from waste rock at the project site. The objectives of the program were to (1) to identify the geochemically different types waste rock in the deposit, (2) to characterize the geochemistry of the different lithologic, alteration, mineralization, and assemblages (LAMAs) of the waste rock, and (3) to identify potential ARD and ML issues and control options for project facilities. Geochemist Steve Atkin visited the Kupol Project in 2003 to develop a sampling program for geochemical characterization of the wastes. A total of 276 drill core samples were selected for static acid generation potential and other geochemical testing. Based on the static ABA test work, representative samples were then selected for kinetic test work. The details of the 2003 and 2004 geochemical testing programs are presented in Atkin, 2005. Of the 234 samples from the 2003 investigation only 18 samples were from the footwall rock. In 2004 11 drill holes were collared in the footwall of the deposit and 42 samples gathered for static testing. The details of the 2004 sample geochemical testing program are presented in Atkin, 2005. 2.2.1 Site Geology Mineralization at the project site is a quartz-adularia, low-sulfidation gold system localized along a high angle vein complex and hosted in andesite flows and pyroclastic rocks. The vein records repeated episodes of dilation and mineralization and is composed of up to five separate banded quartz and quartz breccia assemblages. Gold and silver mineralization is associated with the earliest dilational event and includes electrum, silver-sulfides and silver-sulfosalts. The andesites, which are Cretaceous in age, include a basal unit dominated by flows (85 to 90%) interbedded with discrete andesite crystal tuffs and an upper unit of tuffs (50%) interbedded with flows. The andesite package is overlain by younger (Tertiary age) rhyolite and basalt flows. Post-mineralization dikes, believed the feeders for the Tertiary rhyolite and basalt flows, parallel and, particularly in the south, crosscut the vein in a narrow 100 to 200 m wide corridor. The andesites along the vein have been altered and include the following assemblages: (1) a propylitic assemblage characterized by the presence of carbonate, chlorite, and, epidote; (2) a carbonate dominant assemblage including calcium carbonate, iron carbonate, and dolomite; (3) a texturally destructive argillic alteration characterized by strong clay alteration; (4) a sericite



assemblage; (5) an adularia assemblage; and (6) silicification with quartz dominant veining and pervasive silica flooding of the adjacent country rock. The alteration assemblages are often superimposed upon one another because of the multiple and overlapping mineralization events. 2.2.2 Description of the Static Testing Program A total of 276 representative samples of waste rock representing each of the LAMAs were collected from splits of core from exploration and geotechnical drill holes of the 2003 drilling campaign and historic Russian exploration drill holes. All of the samples underwent acid base accounting tests (ABA) and strong acid digestion with metals analysis (ICP) to characterize the sulfur speciation, acid generating potential, acid neutralizing potential, and gross whole rock chemistry. One in three samples (for a total of 78 samples) from the 2003 drilling, covering the range in ABA parameters and metals concentrations were selected for short term leaching tests to identify readily solubilized constituents. One in ten samples (24 samples) from the 2003 drilling program, and an additional 8 samples from the 2004 drilling program, covering the range of ABA parameters, metal concentrations, and metal solubilities, were selected for net acid generating (NAG) tests and subsequent humidity cell kinetic tests (see Section 2.3 for the results of the kinetic testing program). The 2004 drilling and sampling program was carried out to better classify the footwall rocks, and produced 42 samples for ABA testing. 2.2.3 Results of the Static Testing Program The results of the static geochemical test work for the waste rock LAMAs were grouped into three ARD codes for use in the resource model: (1) non-acid generating (NAG); (2) acid generating (AG); and (3) potentially acid generating (PAG). The ARD codes were based on simplified logging codes for pyrite mineralization and carbonate alteration from the detailed geologic drill hole logs that define four categories: (1) high pyrite and low carbonate – AG; (2) low pyrite and high carbonate – NAG; and (3) high pyrite and high carbonate or low pyrite low and carbonate – PAG. Of the 276 waste rock samples analyzed via static testing methods, 51 (18%) are AG, 71 (26%) are NAG, and 154 (56%) are PAG. Results of the ABA tests indicated that all of the categories of waste rock at the project site are sulfur-bearing. The dominant sulfur species in the waste rock is sulfide sulfur. Total sulfur and sulfide sulfur range from below the detection limit to 6% by weight. Total sulfur and sulfide sulfur concentrations decrease from AG through NAG waste rock. All of the waste rock categories in the resource model contain some degree of acid neutralizing potential (NP) in the form of either carbonate veins or pervasive carbonate alteration. The NP ranges from values less then 0 (negative NP indicates the presence of natural acidity in the rock) to approximately 200 kg CaCO3/t. The NP increases in moving from the AG to the NAG waste rock. Using the neutralizing potential:acid generating potential ratio (NPR=NP/AP) and standardized criteria for classifying ARD potential (i.e., NPR<1 is acid generating, 1<NPR<3 has uncertain acid generating potential, and NPR>3 has net acid neutralizing potential), 67% of the AG waste rock is acid generating, 70% of NAG material is acid neutralizing; and 51% of the PAG material is acid neutralizing (Table 2.1). The proportions of acid generating, acid neutralizing, and uncertain categories for each of the ARD codes are shown in Table 2.1. The uncertain category



is generally considered potentially acid generating awaiting the results of kinetic testing to confirm the short- through long-term ARD potential.

Table 2.1 – Proportions of Acid Generating, Uncertain, and Acid Neutralizing Potential for each of the Resource Model ARD Codes

ARD Code NPR<1 (Acid Generating)

1<NPR<3 (Uncertain)

NPR>3 (Acid Neutralizing)

AG 69% 10% 22% PAG 34% 12% 54% NAG 17% 11% 72%

Similar results were found using the NAGpH and NAG values from the Net Acid Generation test work. The median values of NAGpH and NAG suggests that the AG waste rock is acid generating (NAGpH<4.5 and NAG>5 kg H2SO4/t) and that the NAG and PAG waste rock is non-acid generating (NAGpH>4.5 and NAG generally less than 5 kg H2SO4/t). Whole rock analyses by ICP testing indicates that gold, silver, arsenic, gold, bismuth, mercury, antimony, selenium, and thallium occur in AG, PAG, and NAG waste rock samples at concentrations above normal basalts. The short term leach testing showed a range in pH from about 3 to 10 among the AG, PAG, and NAG waste rock samples. The pH of the short term leach testing increased in moving from the AG samples (average pH of 4, with a median pH of 6) to the NAG samples (average pH of 5 and median pH of 9). Alkalinity in the leachate also increased in a similar manner from 10 mg CaCO3/L in the leachate from AG samples to over 20 mg CaO3/L in the leachate from the NAG samples. The chemistry of the AG, PAG, and NAG leachates was dominated by sulfate, which ranged in concentration from 1 to 2,000 mg/L. Using the maximum concentrations of soluble metals in the short term leach testing leachate, the following elements appear to be mobile in slightly acidic solution: aluminum, arsenic, beryllium, calcium, cobalt, copper, iron, magnesium, manganese, nickel, silicon, strontium, vanadium, and zinc. 2.3 Summary of Kinetic Testing Program The humidity cell tests showed that the pH of leachate derived from AG waste rock samples was typically below pH 4 over the duration of the 20 week testing period. Leachate from AG samples generally had high sulfate, zero alkalinity, and high metal concentrations, with sulfate and metal concentrations declining over time. Humidity cell testing on NAG waste rock samples indicated a range from acidic to neutral/alkaline pH between individual samples, with the majority of samples having effluent pH values >7 over the 20 week testing period. Leachate from NAG samples typically had low sulfate, moderate alkalinity, and moderate metal. Testing on PAG samples indicated a range from acidic to neutral/alkaline pH between individual samples, with the majority of samples having effluent pH values >6 over the 20 week testing period. Humidity cell leachate from the PAG samples was variable depending on the acidity characteristics of the effluent, with higher alkalinity and lower sulfate and metal concentrations in samples having neutral pH leachate.



The following criteria were used to assess ARD potential:

AP – Acid Potential reported as kg of CaCO3 equivalent per tonne of sample. AP is a measure of the maximum potential acidity that the sample can generate if all of the contained sulphide minerals oxidize and form acid. AP is a calculated value based on the total sulphide sulphur concentration within the sample. AP = wt% Sulphide Sulphur x 31.25. The Sulphide Sulphur content in the sample is determined by analysis. The factor 31.25 is based on the acidity generated as derived from the general stoichiometry for the complete oxidation of pyrite and the subsequent hydrolysis of the Fe3+ generated. This factor therefore relates the total acidity produced with the equivalent alkalinity as CaCO3 required for neutralization.

Sobek NP – Neutralization Potential reported as kg of CaCO3 equivalent per tonne of

sample. Sobek-NP is a measure of the neutralizing potential of the sample. NP is a measured value determined by measuring the amount of acid that the sample can neutralize under standardized laboratory conditions.

2.4 Estimated Quantities of Waste Rock Wardrop incorporated the Atkin geochemical data into the open pit mine model to allow estimation of quantities of waste material sorted by ARD potential. Table 2.2 presents the percentages of AG, PAG and NAG rock from the Atkin 2005 report.

Table 2.2 – Waste Rock ARD Code Percentages

ARD Code Percentage of Waste Rock

AG 41.6% PAG 11.3% NAG 47.1%

Table 2.3 below presents a summary of the total waste tonnages provided by Bema.



Table 2.3 – Total Waste Rock (tonnes) Produced from Open Pit and Underground Mined Waste Streams, Atkins Adjusted 2006 2007 2008 2009 2010 2011 2012 2013 2014 Total

Open Pit 1,231,000 2,058,000 1,874,000 1,549,000 1,243,000 7,955,000 Underground 17,645 98,522 89,749 71,006 130,372 249,971 66,369 2,013 725,648 Subtotal 1,248,645 2,156,522 1,963,749 1,620,006 1,373,372 249,971 66,369 2,013 - 8,680,648 Open Pit 877,000 1,585,000 1,810,000 1,594,000 1,203,000 7,069,000 Underground 9,321 74,800 85,455 77,045 100,892 162,160 49,719 3,584 562,976 Subtotal 886,321 1,659,800 1,895,455 1,671,045 1,303,892 162,160 49,719 3,584 - 7,631,976 Open Pit 273,000 466,000 456,000 389,000 303,000 1,887,000 Underground 3,312 21,774 21,794 17,953 29,106 52,830 14,424 651 161,844 Subtotal 276,312 487,774 477,794 406,953 332,106 52,830 14,424 651 - 2,048,844

Unclassified Unclassified Tonnes 43,147 60,607 22,987 101,617 11,181 3,675 0 0 243,214 Total Total Waste Generated 2,454,425 4,364,702 4,359,985 3,799,621 3,020,551 468,637 130,513 6,249 - 18,604,682

NAG

AG

PAG

A portion of the waste rock will be used for underground backfill, Table 2.4 presents the amount of waste rock that will report to the surface and require handling.

Table 2.4 – Total Waste Rock (tonnes) Requiring Surface Handling Waste Not Backfilled or used for Portal Pad 2006 2007 2008 2009 2010 2011 2012 2013 2014 Total

Open Pit 1,231,000 2,058,000 1,874,000 1,549,000 1,243,000 - - - - 7,955,000 Underground - - - - - - - - - - Subtotal 1,231,000 2,058,000 1,874,000 1,549,000 1,243,000 - - - - 7,955,000 Open Pit 877,000 1,585,000 1,810,000 1,589,517 862,544 - - - - 6,724,061 Underground 9,321 61,496 11,627 - - - - - - 82,444 Subtotal 886,321 1,646,496 1,821,627 1,589,517 862,544 - - - - 6,806,504 Open Pit 273,000 466,000 456,000 389,000 303,000 - - - - 1,887,000 Underground 3,312 21,774 21,794 - - - - - - 46,880 Subtotal 276,312 487,774 477,794 389,000 303,000 - - - - 1,933,880

Unclassified Unclassified Tonnes 43,147 60,607 22,987 - - - - - - 126,741 Total Total Unbackfilled Waste from OP & UG 2,436,780 4,252,877 4,196,407 3,527,517 2,408,544 - - - - 16,822,125

NAG

AG

PAG

Table 2.6 presents the total volumes of waste rock requiring surface disposal for the life of the Kupol project, assuming a waste rock density of 2.0 t/ m3.

Table 2.5: – Waste Rock Volumes by ARD Classification

AG Rock PAG Rock NAG Rock 3.6 Mm3 1.0 M m3 4.3 Mm3



3.0 WASTE MANAGEMENT STRATEGY 3.1 General The strategy proposed for waste rock management is that all AG waste rock will be placed in the tailings impoundment and will ultimately be submerged under tailings and water. NAG and as much PAG rock as needed to make up the dam fill volumes, will be used to construct the tailings dam. The balance of PAG rock not needed for tailings dam construction will be placed in the tailings impoundment with the AG material. Detailed design of the tailings dams is presented in AMEC (2005a). 3.2 Waste Characterization and Control Waste materials will be sorted during mining into AG, NAG and PAG categories so that each material can be placed in designated waste zones. ARD classification of the wastes will be determined by sampling and assaying blasthole cuttings in the mine. The assay will be done by index sulphur testing such as the Leco S test. Following testing, blasted waste materials will be flagged in the field for identification, and haul trucks directed to the appropriate dump location. Blasthole sampling and testing for ARD classification will be carried out every blasthole at least during initial operation. Once experience is gained in understanding the relationship between rock type and ARD classification, it is possible that the frequency of ARD sampling and testing could be reduced. 3.3 Waste Pile Site Preparation 3.3.1 Foundation Preparation/Waste Pile Stability The footprint of the waste pile will be cleared and grubbed of all organic and ice-rich materials. If foundation stripping is required the stripped material will be stockpiled for later use in reclamation. The remaining sands and gravels are frozen and are considered suitable foundation materials if maintained in a frozen state, however it is anticipated that the upper meter or two of the foundation will undergo an initial period of thaw as the pond water level rises as shown by the thermal modeling presented in AMEC 2005a, Tailings Facility Feasibility Report. The waste rock will be placed with a final slope of 1.5H:1V however, the foundation soil characteristics will need to be assessed through geotechnical site investigation work in 2005 and detailed stability analyses will be carried out during final design to ensure that the waste piles will be stable during and after placement. An advantage of encapsulating the rock pile with tailings is that the tailings will buttress and thereby enhance the stability of the waste rock pile. Settlement gauges will be installed on the waste rock pile to monitor its stability and thermistors installed to monitor it thermal condition. 3.3.2 Diversion Ditching Tailings dam diversion ditches are to be constructed prior to commencement of waste rock placement. With the diversion ditches already in place, runoff through the waste rock placement areas will be minimal. Prior to the completion of the tailings starter dam additional diversion



ditches will be constructed upslope of the waste pile to intercept and divert surface and shallow groundwater flows. 3.3.3 Collection and Treatment of Waste Rock Effluent Water quality modeling (AMEC, 2005b) indicates that runoff and/or seepage water from AG waste piles may contain elevated levels of some metals, as well as having depressed pH, as a result of flushing of existing oxidation products from the waste rock and from ongoing oxidation and metal leaching. In the summer of 2007, before the tailings dam is complete, runoff and seepage from the AG waste rock pile will be sampled and tested for water quality, and if necessary will be collected and/or treated. For either alternative tailings dam location, a collection sump or pond will be constructed downstream of the waste pile, upstream of the tailings dam, to serve as a point of water quality sampling, and if necessary where pile drainage water can be intercepted and treated prior to discharge in the pre-production period. The collection sump will be formed by constructing a small dam of rockfill with a geomembrane liner. Following completion of the tailings dam, all waste pile drainage will be captured in the tailings pond. 3.4 AG Placement in Tailings Impoundment For Alternative #1 the AG waste rock will be placed in the north-east corner of the impoundment prior to the completion of the tailings starter dam, to elevation 550 m (Figure 330-002). Once the dam is completed the AG and PAG waste rock will be placed into the pond in a ring dyke pattern around the proposed barge location (Figure 330-003) up to a maximum final elevation of 553 m. For Alternative #2 the AG waste rock pile will be developed in the eastern arm of the Y-shaped valley in the tailings impoundment up to elevation 540 m (Figure 330-004). Upon completion of the tailings starter dam the AG and PAG waste rock will be placed into the pond from the central portion of the impoundment east and south towards the tailings dam to a maximum final elevation of 552 m (Figure 330-004). Design storage volumes for tailings and AG rock are as follows:

Tailings 12 million tonnes 10.5 million m3 AG Waste Rock 7.6 million tonnes 3.8 million m3

3.5 NAG/PAG Placement in Tailings Dam The tailings dam will be constructed primarily of NAG rock, with PAG material layered and mixed within the body of the dam. The objective of this scheme of incorporating some PAG rock is to use the PAG material as part of the required dam fill volume, as there is not expected to be enough NAG to construct the body of the dam, and also to reduce the volume of material that will need to be stored in the impoundment.



NAG rock will be placed at the base of the dam in all drainages where surface water may contact fill materials. As well, a minimum thickness of 10 m of NAG rockfill will be placed on the final downstream face and crest of the dam, to provide a conservative depth of non-acid generating material that may in the long term be subject to seasonal thawing (i.e. the active permafrost zone). Within the body of the dam, PAG material will be intermixed with the NAG materials. The overall proportion of NAG to PAG is expected to be about 2:1. Thus, the NAG material will provide considerable buffering for any acidity that may develop in the PAG materials. While the PAG rock is expected to not be acid generating, there is the possibility of small inclusions of material that could generate acidity. There are a number of design features or site conditions that will mitigate the risk of acid generation due to the incorporation of PAG in the dam, as follows:

• The proportion of materials in the dam would be about 2 NAG to 1 PAG, so there will be considerable buffering material available, and a generous thermal insulation layer of NAG material (at least 10 m) will be provided on the final dam crest and downstream slope

• All materials will be controlled by assay of blasthole samples to sort into acid generating

(AG, NP:AGP<1), non-acid generating (NAG, NP:AGP>3) and potentially acid generating (PAG, 3>NP:AGP>1) rocks

• The upstream portion of the dam fill will be covered by the upstream liner, so that there

would be no moisture available for any material placed under the liner

• The rockfill would be placed in the dam in thin lifts, to create low permeability zones on the trafficked surfaces, thus reducing convection and infiltration

• Thermal modelling indicates that the dam will freeze (i.e. pore water within the fine

grained, or silt to sand sized waste, will freeze to create zones or layers of solid ice-rock mass), thus limiting flow through the dam and also inhibiting oxidation of pyrite materials

• The amount of moisture entering the dam fill will be very small. The upstream side of

the fill will be cut off by the liner, and diversions will be in place along the sides, so that the only moisture entering would be direct precipitation falling on the dam

• Seepage from the dam would be sampled and tested for water quality, and if necessary

intercepted downstream and returned to the tailings pond

• The dam will be completed early in mine life, to give a long period of observation of seepage flow rates, temperatures and water quality before mine closure

• If at the end of mine life there is any residual water quality problem, a final fallback

solution could be to extend the liner over the downstream slope to fully exclude any water from entering the dam fill.



4.0 MONITORING Monitoring will be implemented to confirm that the objectives of acid drainage prevention and control are being achieved. Monitoring will include the following:

• Water quality sampling and testing downstream of the AG waste rock pile until the time of completion of the tailings dam in fall 2007

• Water quality sampling downstream of the tailings dam throughout mine operation

and into the closure period until it is confirmed that there are no long term deleterious water quality issues

• Quality control sampling of placed NAG materials to check that the waste material

control system is excluding any PAG or AG materials

• Thermistors in the AG rock pile to monitor its temperature profile.

• Thermistors in the tailings dam to monitor internal temperatures, to confirm that freezing of the embankment is occurring as per thermal modeling presented in AMEC, 2005a, and to detect any abnormal thermal patterns resulting from geochemical activity

• Tracking of waste rock quantities by mine operations engineering personnel with

respect to predicted and actual AG, PAG and NAG materials, and regularly updates of the material balance to ensure that the overall plan and strategy are maintained

• Settlement gauges will be installed on the waste rock pile to monitor its stability.



5.0 CLOSURE Details of the tailings dam and tailings impoundment closure are presented in the tailings impoundment design report, AMEC 2005a. For both tailings site alternatives AG waste rock placed within the impoundment will be placed to a final elevation 2 to 3 m below the final tailings elevation. By completing the waste rock pile below the tailings allows for tailings to be discharged over the waste rock ensuring that the waste rock is completely encapsulated within the tailings and will ultimately be frozen in a saturated state, preventing the generation of ARD. The proposed scheme of permanently submerging all AG waste rock in the tailings impoundment, which will ultimately freeze, will ensure there will be essentially no possibility of long term acid generation from these materials. The tailings themselves will be potentially acid generating. To prevent long term acid generation from the tailings, the tailings impoundment closure plan specifies that a layer of non acid-generating rockfill will be placed over the final tailings impoundment surface to provide a thermal barrier. The active permafrost zone will remain within the rockfill cover, and the tailings will remain frozen. The tailings dam itself will be monitored to ensure that the quality of runoff is acceptable for discharge to the environment. An advantage of Kupol for closure is that the site will continue to be operate following completion of the open pit, with underground mining expected to continue for many years. This will ensure that the site will be effectively monitored and any ongoing maintenance or remedial measures put in place.



6.0 LIMITATIONS Recommendations presented herein are based on an evaluation of the findings of the site investigation noted. If conditions other than those reported are noted during subsequent phases of the project, AMEC should be notified and be given the opportunity to review and revise the current recommendations, if necessary. Recommendations presented herein may not be valid if an adequate level of review or inspection is not provided during construction. This report has been prepared for the exclusive use of Bema Gold, Inc. for specific application to the area within this report. Any use which a third party makes of this report, or any reliance on or decisions made based on it, are the responsibility of such third parties. AMEC accepts no responsibility for damages, if any, suffered by any third party as a result of decisions made or actions based on this report. It has been prepared in accordance with generally accepted engineering practices. No other warranty, expressed or implied, is made. Respectfully submitted, AMEC Earth & Environmental, a division of AMEC Americas Limited

Reviewed by:

Daryl Dufault, E.I.T. Geotechnical Engineer

Per Peter C. Lighthall, P.Eng. Vice President, Mining Earth & Environmental

Ed McRoberts, Ph.D., P.Eng. Chief Technical Officer



REFERENCES

Atkin, S. 2005. Geochemical Characterization of Waste Rock and Ore Materials at the Kupol Project, Russian Far East. Prepared for BEMA Gold Corporation: Vancouver, BC, Canada. – Atkin, S, 2004b. Memorandum to Bill Lytle, Sample selection for humidity cell test work for the 2004 ARD Samples. AMEC, 2005a. Report on Tailings Disposal, Kupol Project, AMEC Earth & Environmental, June.2005 AMEC, 2005b. Report on Water Quality, Kupol Project, AMEC Earth & Environmental, June.2005

Appendix E AMEC Site Water Balance



Site Water Balance Kupol Gold Project

Chukotka Autonomous Okrug, Far East Russia

Submitted to:

Bema Gold Corporation

Submitted by:


Burnaby, BC

July 2005

AMEC File: VM 00330

Bema Gold Corporation Site Water Balance Kupol Gold Project, Far East Russia July 5 2005

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TABLE OF CONTENTS

Page

SUMMARY .................................................................................................................................1

1 INTRODUCTION................................................................................................................2

2 OVERVIEW........................................................................................................................2 2.1 Baseline Hydrology....................................................................................................2 2.2 Climate ......................................................................................................................2

2.2.1 General ..........................................................................................................5 2.2.2 Precipitation ...................................................................................................5 2.2.3 Temperature ..................................................................................................5 2.2.4 Evaporation....................................................................................................5 2.2.5 Streamflows ...................................................................................................7

2.3 Annual Unit Runoff Quantities for Various Land Use Types .......................................7 2.4 Monthly Runoff Distributions ......................................................................................8 2.5 Unit Runoff Quantities for Various Return Periods .....................................................9

3 TAILINGS IMPOUNDMENT/PROCESS PLANT CIRCUIT ...............................................11 3.1 Alternative #1...........................................................................................................13 3.2 Alternative #2...........................................................................................................14

4 INFLOWS TO THE OPEN PIT AND UNDERGROUND WORKINGS ...............................15

5 WASTE ROCK .................................................................................................................21 5.1 Alternative #1...........................................................................................................21 5.2 Alternative #2...........................................................................................................23

6 SITE INFRASTRUCTURE................................................................................................24 6.1 Plant Site .................................................................................................................25 6.2 Ore Stockpile ...........................................................................................................25 6.3 Airstrip .....................................................................................................................26

7 DESIGN CRITERIA FOR ENGINEERING STRUCTURES...............................................26 7.1 Culverts ...................................................................................................................26 7.2 Ditches and other water conveyance facilities..........................................................27 7.3 Sedimentation Ponds...............................................................................................27

8 ENGINEERING STRUCTURES.......................................................................................27 8.1 Culverts ...................................................................................................................27 8.2 Ditches ....................................................................................................................28 8.3 TMF Diversion ditches .............................................................................................29

8.3.1 Alternative #1 ...............................................................................................29 8.3.2 Alternative #2 ...............................................................................................30

8.4 Sediment Ponds ......................................................................................................30 8.5 Closure/Post-Closure Plan.......................................................................................31


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9 DISCHARGE TO RECEIVING WATERS..........................................................................31

10 CONCLUSION .................................................................................................................32

11 CLOSURE........................................................................................................................33

REFERENCES .........................................................................................................................34

LIST OF FIGURES

Figure 2.1 - Regional Climate Stations....................................................................................... 4 Figure 2.2 - Comparison of Precipitation .................................................................................... 6 Figure 2.3 - Comparison of Temperature ................................................................................... 6

LIST OF TABLES

Table 2.1 Average Annual Unit Runoff Values for Various Land Use Types............................... 7 Table 2.2 Monthly Runoff Distributions for Average, Wet and Dry Years.................................... 8 Table 2.3 Precipitation and Sublimation for Wet, Dry and Average Years .................................. 9 Table 2.4 Monthly Runoff Distributions under Varying Return Periods for Natural Land Surfaces

........................................................................................................................................... 9 Table 2.5 Monthly Runoff Distributions under Varying Return Periods for Water Surfaces........10 Table 2.6 Monthly Runoff Distributions under Varying Return Periods for Developed Surfaces 10 Table 2.7 Monthly Runoff Distributions under Varying Return Periods for Waste Rock (Tailings

Dam) during Operations.....................................................................................................10 Table 2.8 Monthly Runoff Distributions under Varying Return Periods for Waste Rock (Tailings

Dam) Closure/Post-Closure ...............................................................................................11 Table 4.1 Monthly Distributions of Runoff in Pit from Direct Precipitation for Various Return

Periods ..............................................................................................................................15 Table 4.2 Estimated Thawback Inflows over the Project Life.....................................................17 Table 4.3 Total Pit Inflows 2007 ................................................................................................18 Table 4.4 Total Pit inflows 2008 ................................................................................................18 Table 4.5 Total Pit Inflows 2009 ................................................................................................19 Table 4.6 Total Pit Inflows 2010 ................................................................................................19 Table 4.7 Total Pit inflows 2011 ................................................................................................20 Table 4.8 Total Pit Inflows 2012 ................................................................................................20 Table 4.9 Total Pit Inflows 2013 ................................................................................................20 Table 4.10 Total Pit Inflows 2014 ..............................................................................................21 Table 5.1 Total Annual Runoff from Pre-Production Waste Pile Alternative #1 for Wet Dry and

Average Years ...................................................................................................................22 Table 5.2 Total Annual Runoff (m³) from Alternative #1 Tailings Dam (during Operations) for

Wet, Dry and Average Year ...............................................................................................22 Table 5.3 Total Annual Runoff (m³) from Alternative #1 Tailings (at Closure) for Wet, Dry and

Average Year.....................................................................................................................23


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Table 5.4 Total Annual Runoff from Pre-Production Waste Pile Alternative #2 for Wet Dry and Average Years ...................................................................................................................23

Table 5.5 Total Annual Runoff (m³) from Alternative #2 Tailings Dam (during Operations) for Wet, Dry and Average Year ...............................................................................................24

Table 5.6 Total Annual Runoff (m³) from Alternative #2 Tailings (at Closure) for Wet, Dry and Average Year.....................................................................................................................24

Table 6.1 Annual Runoff from Plant Site for Wet, Dry and Average Years ...............................25 Table 6.2 Annual Runoff from Ore Stockpile for Wet, Dry and Average Years ..........................25 Table 6.3 Annual Runoff from Direct Precipitation on Airstrip for Wet, Dry and Average Years26 Table 8.1 Airstrip Access Road Culvert Summary.....................................................................27 Table 8.2 General Ditch Dimensions.........................................................................................28 Table 8.3 - Alternative #1 - Diversion Ditch Dimensions............................................................29 Table 8.4 – Alternative #2 - Diversion Ditch Dimensions ...........................................................30 Table 8.5 Summary of Sediment Pond Sizes ...........................................................................30 Table 9.1 Volume of Treatment Water (m³) ...............................................................................31 Table 9.2 Volume of Discharge Water (m³) ..............................................................................32 Table B.11.1: Site Water Balance Summary for 2007 ......................................................... 1 Table B.11.2: Site Water Balance Summary for 2008 ......................................................... 2 Table B.11.3: Site Water Balance Summary for 2009 ......................................................... 3 Table B.11.4: Site Water Balance Summary for 2010 ......................................................... 4 Table B.11.5: Site Water Balance Summary for 2011 ......................................................... 5 Table B.11.6: Site Water Balance Summary for 2012 ......................................................... 6 Table B.11.7: Site Water Balance Summary for 2013 ......................................................... 7 Table B.11.8: Site Water Balance Summary for 2014 ......................................................... 8 Table B.11.9: Site Water Balance Summary for 2015-2021 ................................................ 9

LIST OF APPENDICES

APPENDIX A : TAILINGS WATER BALANCE – Alternatives #1 and #2 APPENDIX B : SITE WATER BALANCE ANNUAL SUMMARY

IMPORTANT NOTICE

This report was prepared exclusively for BEMA Gold Corporation by AMEC Earth & Environmental Limited, a wholly owned subsidiary of AMEC Americas Limited. The quality of information, conclusions and estimates contained herein is consistent with the level of effort involved in AMEC services and based on: i) information available at the time of preparation, ii) data supplied by outside sources, and iii) the assumptions, conditions and qualifications set forth in this report. This report is intended to be used by BEMA Gold Corporation only, subject to the terms and conditions of its contract with AMEC. Any other use of, or reliance on, this report by any third party is at that party’s sole risk.

Bema Gold Corporation Site Water Balance Kupol, Russia July, 5 2005

AMEC File: VM 00330 S:\PROJECTS\VM00330 - Kupol Gold Water Mgmt\Site Water Balance\Site Water Balance - Final 05 Jul 05.doc Page 1

SUMMARY

This report presents the hydrology, development of hydrologic factors, site water balance, tailings water balance and design of hydraulic structures for the Kupol Gold Project, in Chukotka Autonomous Okrug, Far East Russia. The quantities developed in this report will be used in the other work such as the water management plan and tailings dam report. Analysis of the climate data indicates that at the site:

• average annual temperature is -13°C, • annual precipitation is estimated at 380 mm based on Markovo climate station, and • annual evaporation is estimated at 280 mm based on Ostrovnoje climate station.

Average annual unit runoff values along with monthly distributions were developed for the various land use types at Kupol. These land use types include natural surfaces, water surfaces, developed surfaces without vegetation and the waste rock. The 1-in-200, 1-in-100, 1-in-10 wet and dry year runoffs with monthly distributions were also calculated. For both tailings alternatives, a 14-year, monthly water/mass balance model was developed to simulate the hydrologic behaviour of the Tailings Management Facility (TMF). Diversion ditches will be constructed above the impoundment limit to help maintain the minimum water depth requirement of the TMF. A side discharge facility installed in one of the ditches and located upstream of the abutment will assist with make-up water requirements. During the pre-production year (2007) the runoff from the pit, plant site and ore stockpile will be passed through a lime treatment system and discharge to receiving waters. The sewage treatment plant effluent will also be discharged to the natural receiving body. Upon completion of the starter dam runoff from the pit and ore stockpile along with effluent from the sewage treatment plant will be pumped to the TMF. The plant site runoff will continue to be treated with a lime mixing and sediment pond system and discharge to the receiving water. Diversion ditches around the TMF are design for the 200–year flood. Culverts and ditches that carry runoff to the culverts are designed for the 100–year flood. The sediment ponds at the site are designed to a 10-year precipitation event combined with average one day snowmelt. The major findings of this report are:

• On a volumetric basis the treatment of the contact water during the life of the mine is manageable through the use of a lime treatment system.

• The results of the tailings water balance for either alternative show that the TMF will be

able to handle all inflows and operate as a zero discharge facility.

• Shipping containers used as culverts will be able to adequately convey the design flows at all culvert crossings around the site.



1 INTRODUCTION The Kupol Gold Project will include activities such as open pit mining, underground mining, ore stockpiling and processing, tailings and waste management, and camp operation that will generate volumes and qualities of water requiring management. This report is one of several pertaining to water management for this project and develops the quantities used in the water management plan. A description of the baseline hydrology, derivation of hydrologic factors, design criteria and design of various hydraulic structures such as culverts, ditches and sedimentation ponds is also presented in this report. The site water balance was developed for the years 2007-2014 and 2019. The years 2007 – 2014 correspond to the period when the open pit and underground mine is being developed. Therefore, thawback inflows from the pit walls and consequently the water balance will vary until the development is complete in 2014. The starter tailings dam is scheduled to be completed by the end of 2007 therefore the site water balance for the pre-production year (2007) will differ because all contact water will be treated and released. The site water balance for 2011 will be representative of any year during the operating life of the mine. The inputs and outputs of the TMF vary annually as the facility is in-filled with tailings, waste rock and net water inflows, therefore the TMF water balance will be calculated for every year of operation. A water balance was completed for both tailings alternatives. The mine site layout showing both tailings alternatives can be found in Figure 330-001 along with delineated catchments. A more detailed view of the plant site layout is presented in Figure 330-002. The feasibility study shows 7.1 million tonnes of material to be processed through the mill during a seven year mine life. The criteria provided by Bema Gold Corp. was to develop a facility capable of holding 12 million tonnes of waste. This extra capacity will be used to store material not included in the feasibility study (inferred) or material discovered in the future. The TMF water balance is calculated past the seventh year to year 14 at a mill rate of 2000 metric tonnes per day to account for this extra capacity. The term contact water will be used within this report. For this report contact water refers to water that has come into contact with sources of potential pollutants and includes runoff from the plant site, ore stockpile and some haul roads which will be collected and diverted to the TMF. 2 OVERVIEW 2.1 Baseline Hydrology 2.2 Climate The proposed Kupol Gold Project is located within the Chukotka Autonomous Okrug (Region) in Eastern Russia, slightly above the Arctic Circle, in a very cold environment. The climate in this region is characterized by long cold winters (8 to 8.5 months), short summers (2.5 months), and even shorter springs and falls. Kupol is located at the divided of watersheds between water that flows east to the Bering Sea and Pacific Ocean and north to the East Siberian Sea and Arctic



Ocean. The Andyrsky Mountains run east-west and have elevations over 2000 m. Some key parameters from this data analysis, provided by Bema, are summarized as follows:

• The estimated average annual air temperature at the Kupol site, with only minor variances, is near -13 °C;

• Weather data from Ilirney shows the annual precipitation to be about 278 mm. For the design basis, the average total amount of precipitation at the mine is estimated from analysis of data at the Markovo station to be 380 mm. The total number of days with average daily temperature above zero is less than 50. Positive average daily temperatures are first noted in the first 10 days of June. The transition from positive average daily temperatures to negative average daily temperatures occurs in the middle 10 days of September;

• The approximate annual amount of evaporation from surface water sources is 280 mm based on Ostrovnoje climate station; and

• Snow begins falling in the middle of September and achieves a maximum depth in March. The average depth of snow cover is observed to be 38-45 cm. As a result of heavy winds, the valleys are filled with snowdrifts and the tops of ridges and hills and steeps slopes are blown bare. The average snow density reaches 160 kg/m3 with water equivalent content of 107 mm. During the winter months, approximately 174 mm of water equivalent precipitation is expected. Snow is approximately 46% of the annual precipitation.

The atmospheric circulation over Kupol blows north during winter and south during summer. This north flow brings moist air from the Bering Sea to the site, suggesting that climate stations located along the north-south path would provide the best climate correlation to the mine site. Three regional climate stations were considered for the extension of the mine site data - Ilirney, Ostrovnoje and Markovo and are shown on Figure 2.1. Ilirney is the closet station to the site located approximately 150 km west of the Kupol mine site. Markovo is located about 660 km south and Ostrovnoje is located about 690 km west of the site. All three stations have over 20 years of data. Ilirney is the closest climate station to the site and can be utilized to extend the minesite climate station. However, the periods of record at the minesite and Ilirney do not overlap and therefore a direct correlation between this regional station and the mine site could not be adequately established. This holds true for the Ostrovnoje station as well. Markovo is the only climate station with recorded data that coincides with the two summer seasons of data recorded at the mine in 2003 and 2004. Markovo is better suited to establish a direct correlation with the site data. Other reasons for employing Markovo include the north-south movement of wind currents, long-term monthly patterns and conservatism. As discussed above, the wind currents that accumulate and deposit moisture from the Bering Sea circulate in a north-south direction. The Markovo station is located along this path south of the mine site. Ilirney is located west of the mine site and on the other side of the drainage catchment divide between the Bering Sea and Chukchi Sea to the north, and is therefore influenced by different moisture patterns.



A plot of long term period of record monthly precipitation between Ilirney, Ostrovnoje, Markovo and the site’s two years of data over the summer period from June to September suggests that Ilirney’s peak month occurs in July, while Markovo and the mine site experiences the peak in August, as illustrated in Figure 2.1. This suggests Ilirney experiences a different rainfall pattern in the summer than Markovo and the mine site. It is necessary to err on the conservative side given that the Tailings Management Facility (TMF) for which the climate data are employed is a closed circuit system with no spilling permitted. The estimated precipitation at the mine based on Ilirney is estimated to be 278 mm, while Markovo’s estimate is 380 mm. It would be more conservative to assume a greater annual inflow volume at the feasibility stage to ensure design flows and volumes will not overtop the TMF dam.

Figure 2.1 - Regional Climate Stations



2.2.1 General A correlation is drawn between the Markovo climate station and the site-recorded data for two summers in 2003 and 2004. The comparison with seven months (July-September 2003 and June-September 2004) of rainfall data produced a good relationship between the two stations with the site experiencing an estimated 15% more precipitation than at Markovo. This difference in precipitation can be explained by the orographic influence caused by an elevation difference between Markovo and the mine site. The Markovo station is at an elevation of 33 m and the mine site is at an elevation of 650 m. This correlation will evolve over time as more data are collected from the mine site. 2.2.2 Precipitation Employing this relationship to the 72 years (1895-1999) of precipitation data at Markovo station, a derived average annual precipitation at the mine of 380 mm is estimated. Applying this same relationship to annual maximum daily precipitation at Markovo station, the 24-hour, 200-yr storm event is estimated to be 63.5 mm, and the 24-hour, Probable Maximum Precipitation (PMP) is estimated to be 206 mm. It should be noted that short-duration storm events will experience slightly higher peaking ratios when compared with a monthly period. Lack of storm tracking information and erratic behaviour at the daily resolution level between the site and Markovo station data prevent establishing a daily peak ratio, so a monthly volumetric one was employed. As more data become available at the site, a better peaking factor will be established. Sublimation is estimated at approximately 70 mm. This is based on knowledge of Arctic climates in Canada. 2.2.3 Temperature A comparison of the average monthly temperatures at Ostrovnoje, Ilirney, Markovo and the mine (Figure 2.3) indicates that the summer temperatures are approximately the same, but the winter temperatures at the minesite and Markovo are slightly higher than at Ostrovnoje and Ilirney. This difference in temperature is likely due to the effect of atmospheric circulation. This observation could be verified as more mine site data are collected. Based on the two years of site data the mean annual temperature is -10°C. The average temperature in winter (October to May) is -17.8°C and the average temperature in summer (June to September) is 8.2°C. 2.2.4 Evaporation The only regional evaporation data available is from the Ostrovnoje climate station. Both Ostrovnoje and the mine are located above the Arctic Circle, are within the same latitude region, and experience similar solar radiation exposure. It is therefore assumed that the evaporation at Ostrovnoje, equal to 280 mm annually, would provide a reasonable approximation of the evaporation at the mine. Evaporation data could be collected at the mine site to assist in the verification of this assumption. Evapotranspiration is estimated at 120 mm based on the calculation of evapotranspiration in Canadian Arctic climates.



0

1

2

3

4

5

6

7

8

Jan. Feb. Mar. Apr. May Jun. Jul. Aug. Sep. Oct. Nov. Dec.

Pre

cipi

tatio

n (c

m)

Ilirney, Russia

Markovo, Russia

Ostrovnoje

Minesite 2003/04

Markovo 2003/04

Regional Average

Figure 2.2 - Comparison of Precipitation

-40

-30

-20

-10

0

10

20

Jan. Feb. Mar. Apr. May Jun. Jul. Aug. Sep. Oct. Nov. Dec.

Ilirney, Russia

Markovo, Russia

Ostrovnoje

Minesite 2003/04

Markovo 2003/04

Regional Average

Figure 2.3 - Comparison of Temperature



2.2.5 Streamflows The hydrologic characteristics of the watersheds in the vicinity of Kupol mine are similar to northern Canadian watersheds. Extreme flood events are caused by a large rainfall event combined with the spring snowmelt. The watersheds that will be influenced by the work at Kupol are Kaiemraveem Creek and Starichnaya River. Kaiemraveem Creek flows south into the Bering Sea and will be most affected by the proposed mine. Starichnaya River, the headwaters of which are a few kilometres north of the mine, flows north into the Arctic Ocean. In the vicinity of the mine, Kaiemraveem Creek has steeper valleys than Starichnaya River. With high winds experienced at the site, snow will likely be blown into the valley. High runoff coefficients are expected during the spring freshet (the main spring runoff event) expected to occur mostly in May or early June, as the ground will still be frozen. Flow data has been collected at two recording sites along an unnamed creek, which is an upper catchment tributary of Kaiemraveem Creek, west of the mine camp. The unnamed creek alignment flows through the site for Alternative #1. The first station located immediately upstream of the proposed dam location, while the second is located towards the upper limit of the impoundment. The maximum-recorded stream discharge at the first station is 0.54 m³/s, while the maximum-recorded flow at the second station is 0.24 m³/s. 2.3 Annual Unit Runoff Quantities for Various Land Use Types For the purposes of runoff analysis the Kupol site can be divided into the following land use types:

• Natural Surfaces • Open Water Surfaces • Developed land without vegetation, and • Waste Rock (operations and closure)

Table 2.1 summarizes the average annual unit runoff estimated for each land use type.

Table 2.1 Average Annual Unit Runoff Values for Various Land Use Types

Land Use Type Average Annual Unit Runoff (mm)

Natural Surfaces 190 Water Surfaces 100 Developed Surfaces without vegetation (airstrip, plant site, roads, storage areas) 279

Waste Rock (Tailings Dam) Operations

Closure

130 279

The average annual precipitation of 380 mm occurs as both rain and snow. During an average winter an estimated 70 mm of sublimation occurs. Evapotranspiration, a measure of



evaporation from both open water and transpiration from vegetation, is estimated at 120 mm. These numbers yield an annual runoff of 190 mm (380-70-120) from natural surfaces for an average year. Average annual evaporation at the site is estimated at 280 mm so the annual runoff from open water surfaces during an average year is 100 mm (380-70-280). Developed surfaces without vegetation includes the airstrip, plant site, pit and any other areas of development without vegetation, and is assumed to have an average annual runoff coefficient of 0.9, resulting in a net runoff of 279 mm [0.9 x (380-70)]. The waste rock located in the tailings dams is not included as part of the developed surfaces without vegetation. Average annual runoff from the tailings dam fill (waste rock) during operations is estimated at 130 mm. Runoff from the dam was calculated assuming a 3 m active layer with a 1% net wetting of the country rock at 2650 kg/m³ rock density and 25% porosity yielding 60 mm of loss. A loss of 120 mm is also included in the runoff calculation to account for evaporation losses from ponding on the waste rock and evaporative loss from the active layer which is assumed to be equal to evapotranspiration losses from natural surfaces. The average annual runoff from the waste rock consists of the total annual precipitation (380 mm) less sublimation (70 mm), rock porosity loss (60 mm), and equivalent evapotranspiration (120 mm). The tailings dam fill (waste rock) during the closure/post-closure phase of the project will remain the same size and the active layer will only contain previously wetted rock. Therefore, the runoff from the waste rock at closure will be the same as developed land without vegetation (279 mm). 2.4 Monthly Runoff Distributions Annual runoff at Kupol occurs during the open water season from May to October. Due to annual variances in climatic conditions the start and end of the open water season will vary. The monthly runoff distributions provided by Bema are summarized in Table 2.2 and will be used for the site water balance.

Table 2.2 Monthly Runoff Distributions for Average, Wet and Dry Years

Month Dry Years Average Year Wet Years January 0% 0% 0% February 0% 0% 0% March 0% 0% 0% April 0% 0% 0% May 2% 3% 1% June 47% 47% 34% July 28% 28% 34% August 18% 17% 22% September 4% 4% 7% October 1% 1% 2% November 0% 0% 0% December 0% 0% 0%



2.5 Unit Runoff Quantities for Various Return Periods The precipitation for various return periods from the 200-year wet to the 200-year dry are presented in Table 2.3. Evaporation and evapotranspiration are assumed to be constant regardless of return period. Assuming sublimation in a 200-year dry year is the same as an average year would result in a majority of the precipitation being lost to sublimation. This is an unlikely scenario, therefore sublimation during the dry years is assumed to be the same percentage of precipitation as in an average year.

Table 2.3 Precipitation and Sublimation for Wet, Dry and Average Years

Return Period Precipitation

(mm)

Over-Winter Sublimation

Losses (mm)

1-in-200 Dry Year 92 17 1-in-100 Dry Year 120 22 1-in-10 Dry Year 237 44 Average Year 380 70 1-in-10 Wet Year 523 70 1-in-100 Wet Year 640 70 1-in-200 Wet Year 668 70

The resulting return period (extreme event) runoffs for various land use types and distributed into monthly runoff are presented in Tables 2.4 to 2.8.

Table 2.4 Monthly Runoff Distributions under Varying Return Periods for Natural Land Surfaces

Unit Runoff from Natural Land Surface (mm)

Month 1-in-200 Dry Year

1-in-100 Dry Year

1-in-10 Dry Year

Average Year

1-in-10 Wet Year

1-in-100 Wet Year

1-in-200 Wet Year

January 0 0 0 0 0 0 0 February 0 0 0 0 0 0 0 March 0 0 0 0 0 0 0 April 0 0 0 0 0 0 0 May -0.9 -0.4 1.5 5.7 3.3 5 5 June -21 -10 34 89 113 153 163 July -13 -6.2 21 53 113 153 163 August -8.1 -4.0 13 32 73 99 105 September -1.8 -0.9 2.9 7.6 23 32 33 October -0.4 -0.2 0.7 1.9 6.7 9 10 November 0 0 0 0 0 0 0 December 0 0 0 0 0 0 0 Annual -45 -22 73 190 333 450 478



Table 2.5 Monthly Runoff Distributions under Varying Return Periods for Water Surfaces

Unit Runoff from Water Surface (mm)

Month 1-in-200 dry Year

1-in-100 Dry Year

1-in-10 Dry Year

Average Year

1-in-10 Wet Year

1-in-100 Wet Year

1-in-200 Wet Year

January 0 0 0 0 0 0 0 February 0 0 0 0 0 0 0 March 0 0 0 0 0 0 0 April 0 0 0 0 0 0 0 May -3.8 -3.2 -0.9 3.0 2.4 3.6 3.9 June -88 -75 -20 47 83 122 132 July -53 -45 -12 28 83 122 132 August -34 -29 -8 17 53 79 85 September -7.5 -6.4 -1.7 4.0 17 25 27 October -1.9 -1.6 -0.4 1.0 4.9 7.2 7.8 November 0 0 0 0 0 0 0 December 0 0 0 0 0 0 0 Annual -188 -160 -43 100 243 360 388

Table 2.6 Monthly Runoff Distributions under Varying Return Periods for Developed Surfaces

Unit Runoff from Developed Surface (mm)


1-in-100 Dry Year

1-in-10 Dry Year

Average Year

1-in-10 Wet Year

1-in-100 Wet Year

1-in-200 Wet Year

January 0 0 0 0 0 0 0 February 0 0 0 0 0 0 0 March 0 0 0 0 0 0 0 April 0 0 0 0 0 0 0 May 1.4 1.8 3.5 8.4 4.1 5.1 5.4 June 31.7 41.4 82 131 139 174 183 July 18.9 24.7 49 78 139 174 183 August 12.2 15.9 31 47 90 113 118 September 2.7 3.5 7.0 11.2 29 36 38 October 0.68 0.88 1.7 2.8 8.2 10 11 November 0 0 0 0 0 0 0 December 0 0 0 0 0 0 0 Annual 68 88 174 279 408 513 538

Table 2.7 Monthly Runoff Distributions under Varying Return Periods for Waste Rock (Tailings Dam) during Operations

Unit Runoff from Waste Rock During Operations (mm)


1-in-100 Dry Year

1-in-10 Dry Year

Average Year

1-in-10 Wet Year

1-in-100 Wet Year

1-in-200 Wet Year

January 0 0 0 0 0 0 0 February 0 0 0 0 0 0 0 March 0 0 0 0 0 0 0



Unit Runoff from Waste Rock During Operations (mm)


1-in-100 Dry Year

1-in-10 Dry Year

Average Year

1-in-10 Wet Year

1-in-100 Wet Year

1-in-200 Wet Year

April 0 0 0 0 0 0 0 May -2.1 -1.6 0.27 3.9 2.7 3.9 4.2 June -49 -39 6.3 61 93 133 142 July -29 -23 3.7 36 93 133 142 August -19 -15 2.4 22 60 86 92 September -4.2 -3.3 0.53 5.2 19 27 29 October -1.0 -0.8 0.13 1.3 5.5 7.8 8.4 November 0 0 0 0 0 0 0 December 0 0 0 0 0 0 0 Annual -105 -82 13 130 273 390 418

Table 2.8 Monthly Runoff Distributions under Varying Return Periods for Waste Rock (Tailings Dam) Closure/Post-Closure

Unit Runoff from Waste Rock Post Closure (mm)


1-in-100 Dry Year

1-in-10 Dry Year

Average Year

1-in-10 Wet Year

1-in-100 Wet Year

1-in-200 Wet Year

January 0 0 0 0 0 0 0 February 0 0 0 0 0 0 0 March 0 0 0 0 0 0 0 April 0 0 0 0 0 0 0 May 1.4 1.8 3.5 8.4 4.1 5.1 5.4 June 32 41 82 131 139 174 183 July 19 25 49 78 139 174 183 August 12 16 31 47 90 113 118 September 2.7 3.5 7.0 11 29 36 38 October 0.7 0.88 1.7 2.8 8.2 10 11 November 0 0 0 0 0 0 0 December 0 0 0 0 0 0 0 Annual 68 88 174 279 408 513 538 3 TAILINGS IMPOUNDMENT/PROCESS PLANT CIRCUIT The water and materials balance for Alternatives #1 and #2 were based on the following tailings inflow and deposition characteristics:

• Nominal dry solids inflow of 2,509 metric tonnes per day (MTPD) in 2008 • Nominal dry solids inflow of 3,000 MTPD for 2009-2013 • Nominal dry solids inflow of 2,943 MTPD for 2014 • Nominal dry solids inflow of 2,000 MTPD for 2015-2021 • Solids specific gravity of 2.60 • Percent solids by weight of 50% • Tailings water deposition rate of 105 m³/h at the nominal 2,509 MTPD



• Tailings water deposition rate of 125 m³/h at the nominal 3,000 MTPD • Tailings water deposition rate of 123 m³/h at the nominal 2,943 MTPD • Tailings water deposition rate of 83.3 m³/h at the nominal 2,000 MTPD • Fresh water make-up rate of 40 m³/h • Reclaim rate of 69.4 m³/h at the nominal 2,509 MTPD • Reclaim rate of 83 m³/h at the nominal 3,000 MTPD • Reclaim rate of 81.4 m³/h at the nominal 2943 MTPD • Reclaim rate of 55.3 m³/h at the nominal 2,000 MTPD • Tailings void ratio of 1.26. • AG waste rock, open pit = 7,000,000 tonnes, underground = 600,000 tonnes. • Waste rock dry density 2 tonnes/m3 • Contaminated site runoff (ore stockpile) = 9,000 m3/yr (average annual) • Domestic waste water from the treatment plant (average annual) = 66,000 m3/yr • Pit dewatering average annual inflows = 59,000 m3/yr • Pit wall thawback = varies from peak of 42,000 m3/yr in 2007 of production, decreasing

to about 9,000 m3/yr in 2010, assumed negligible after 2010. The TMF is designed as a closed-circuit process with no allowance for spill, so the impoundment must have enough capacity to store average annual net water inflows, a 200-year return period wet year, and a Probable Maximum Flood (PMF) event with adequate freeboard during the mine life. The probability of a 200-year wet year occurring in the seven year mine life is estimated to be 3.4% and will be the level of risk exposure for the facility on an average annual basis. To maintain the design minimum water depth during net inflow deficit periods, a diversion ditch will be constructed with a side-discharge release structure immediately upstream of the dam crest’s abutment. The location of this side discharge facility will depend on Alternative selected. Releases through this facility will be conducted during the spring/summer months (May to September) on an annual or as needed basis. Water levels and trends in climate will be continuously monitored to determine the volume of water released into the TMF, and an operational rule curve for the tailings facility will be developed during the mine’s operation. The excess releases will provide buffer storage for winter month withdrawals while maintaining the minimum 4-m depth over the tailings level. Additional make up water will be pumped as needed. Direct precipitation on the pond is accumulated over the winter and melted along with catchment runoff during the months of May and June. The annual precipitation estimate of 380 mm is based on site and the Markovo regional station data. Evapotranspiration and sublimation are accounted for in the catchment runoff with preliminary estimates of 120 mm and 70 mm, respectively. These values are based on previous arctic region studies conducted in northern Canada and the regions in the Siberia. Annual evaporation of 280 mm and its monthly distribution for the open pond area during the summer months is estimated from the Ostrovnoje regional station data in the absence of site-specific information. This station is expected to experience similar atmospheric conditions related to evaporation because of its location along similar latitude with the site. Seepage rates are assumed to be zero as the dam embankment is lined with a bituminous geomembrane and the foundation soils are shown to freeze and to maintain its frozen state



during summer months, based on thermal modeling (discussed in Sections 6.1.5.1 and 6.2.5.1 of the Tailings Dam Feasibility Report (AMEC, 2005)). A 14-year, monthly water/mass balance model was developed to simulate the hydrologic behaviour of the TMF. The water balance for Alternatives #1 and #2 are presented in Appendix A. The baseline scenario of the model employs average annual precipitation, runoff and evaporation for the duration of the mine life. The mine life was derived from the processing rates mentioned above and a total mined quantity of 12 million metric tonnes. The starter dam is assumed to be completed by October of 2007 and precipitation runoff is expected to accumulate henceforth. The mine operation is scheduled to initiate the following June. Inflow sources are from tailings solids and water, process water wells, catchment runoff, direct pond precipitation, ore stockpile runoff, pit runoff, pit thawback, tailings dam runoff and sewage treatment plant effluent. Outflows incorporate evaporation, void losses, reclaim and seepage. 3.1 Alternative #1 Diversion ditches constructed around the perimeter, upslope of the TMF inundation level, were the option selected to divert and retain net inflows into the TMF. This will adequately maintain a balance between required reclaim water volume and sufficient storage capacity. The total diverted catchment area is approximately 6.4 km² out of a total 7.6 km² drainage area. Two ditches are proposed for diverting the catchment runoff – west and east diversion ditches, as illustrated in Figure 330–003. The west diversion ditch will collect runoff from the west slope drainage area that comprises 4.7 km², while the east diversion ditch will collect runoff from the east slope tributary area of 1.7 km². Each ditch is designed to convey 24-hour, 200-year return period storm event flows around the respective abutments, upslope of the dam embankment and into the main stem of the creek at an appropriate distance downstream of the dam toe. Detailed ditch design is discussed in Section 8.2. The simulated results from the 14-year, monthly water/mass balance model (discussed in Section 3) suggest that annual releases into the pond from the west diversion ditch facility will be required for some period each year in the months from May to September. The inflow volumes will be required to initially accumulate sufficient start-up water capacity for the mill process, and to provide makeup water through the winter months. The model suggests that minimal external makeup water is needed beyond releases into the pond from the diversion ditches, but an external water supply must be available should an extreme dry year occur, or should reclaim water be unavailable due to cold weather operating problems. The 3-year tailings and water levels are estimated to be 539 m and 543.3 m, respectively. For the final year of operations, tailings and water levels are estimated to be 554.4 m and 5558.6 m, respectively (the AG and PAG rock final elevation are predicted to be about El. 553 m at closure, completely buried by tailings). The dam crest staging should exceed these levels by at least 1.5 m to provide adequate freeboard to store a PMF storm event and to accommodate wind-generated wave run-up. Two additional scenarios are modeled with a 200-year wet year and a 200-year dry year to independently occur during the operational life of the mine. These extreme-year scenarios are selected to test the TMF’s ability to meet operation water needs and to determine its capacity to retain excess water volume with adequate freeboard. For the 200-year wet year, the annual



precipitation was estimated to be 664 mm. Simulation results from the model suggest that water levels in the TMF will rise between 0.4 m to 3.8 m over an average annual year, depending on the maturity of operation. These water level increases can be mitigated with reduced or eliminated diversion releases from the ditches during the year. In earlier years of operation, the suspension of diversion releases is sufficient enough to mitigate increased inflows during the year and an occurrence in latter years when releases occur translates to incrementally higher water levels that are small enough to still maintain adequate freeboard. For the 200-year dry year scenario, the annual precipitation was estimated to be 91 mm. Simulation results from this model suggest that the water level will drop between 0.5 m to 3.5 over an average annual year, depending on the operational stage. The reduction in water level can be mitigated with increased diversion releases over the spring/summer months and, at lower water levels, supplemented by make-up water from process water wells. 3.2 Alternative #2 Diversion ditches constructed around the perimeter, upslope of the TMF inundation level, were the option selected to divert and retain net inflows into the TMF. This will adequately maintain a balance between required reclaim water volume and sufficient storage capacity. The total diverted catchment area is approximately 2.6 km² out of a total 3.5 km² drainage area. Two ditches are proposed for diverting the catchment runoff – west and east diversion ditches, as shown in Figure 330-004. Each ditch is designed to convey 24-hour, 200-year return period storm event flows. The west diversion ditch will collect runoff from the west slope drainage area that comprises 0.33 km². Due to the steep terrain on the west abutment, a smaller ditch is proposed to minimize extensive slope cuts. The ditch will drain south and discharge downstream of the dam toe. The east diversion ditch will collect runoff from the north and east slope tributary areas of 2.2 km². The ditch will drain eastward and discharge over a saddle point on the east catchment boundary upstream of the dam embankment. A side discharge facility will be constructed at the downstream end of this ditch to admit make-up water collected from catchment runoff into the TMF. The simulated results from the 14-year, monthly water/mass balance model discussed in Section 3 suggests that the east ditch side discharge facility will begin operation as early as 2008. Annual releases going forward will initiate in May and continue through the summer months to accumulate sufficient water volume for start-up and to provide makeup water for operations through the winter months. Releases from the spillway into the pond beyond the first year of operation should be timed to avoid utilizing the facility during the heavy flow condition of the spring freshet. The model suggests that additional make up water from other pumped sources may be needed beyond catchment inflow released from the spillway during the first fewf years of operation. The external supply must also be available should operational problems or an extreme dry year occur. The 3-year tailings and water levels are estimated to be 534 m and 539 m, respectively. For the final year of operations, tailings and water levels are estimated to be 554.9 m and 559 m, respectively. The dam crest staging program should exceed these levels by at least 1.5 m to



provide adequate freeboard to store a PMF storm event and to accommodate wind-generated wave run-up. Two additional scenarios are modeled with a 200-year wet year and a 200-year dry year to independently occur during the operational life of the mine. These extreme-year scenarios are selected to test the TMF’s ability to meet operation water needs and to determine its capacity to retain excess water volume with adequate freeboard. For the 200-year wet year, the annual precipitation was estimated to be 664 mm. Simulation results from this model suggest that water levels in the TMF will rise between 0.5 m to 2.8 m over an average annual year, depending on the maturity of operation. These water level increases can be mitigated with no May or summer diversion releases during the year depending on the water level stage. In earlier years of operation, the omission of diversion inflows is sufficient, but an occurrence in latter years translates to incrementally higher water levels that are small enough to still maintain adequate freeboard. For the 200-year dry year scenario, the annual precipitation was estimated to be 91 mm. Simulation results from this model suggest that the water level will drop between 0.3 m to 3.6 m over an average annual year, depending on the operational stage. The reduction in water level can be mitigated with diversion releases over the entire spring/summer months and make-up water from process water wells. 4 INFLOWS TO THE OPEN PIT AND UNDERGROUND WORKINGS The runoff from the catchments adjacent to the pit (as shown in Figure 330-002) will be diverted around the pit therefore the pit inflows will be from:

• Direct precipitation on the pit, and • Groundwater inflows (majority expected to be thawback of exposed permafrost in active

layer) It is assumed the pit reaches its maximum footprint (21.2 ha) by the end of 2006. Based on that assumption and the average annual unit runoff of 279 mm for developed surfaces without vegetation, the resulting average volume of surface runoff in the open pit is 59,200 m³/yr. Table 4.1 summarizes the monthly distribution of average annual runoff as well as for the wet and dry years.

Table 4.1 Monthly Distributions of Runoff in Pit from Direct Precipitation for Various Return Periods

Total annual runoff in Pit (m³)

Month 1-in-200 dry year

1-in-100 dry year

1-in-10 dry year

Average year

1-in-10 wet year

1-in-100 wet year

1-in-200 wet year

January 0 0 0 0 0 0 0 February 0 0 0 0 0 0 0 March 0 0 0 0 0 0 0 April 0 0 0 0 0 0 0 May 287 374 739 1780 865 1090 1140 June 6740 8790 17400 27800 29400 37000 38800 July 4010 5240 10300 16600 29400 37000 38800



Total annual runoff in Pit (m³)


1-in-100 dry year

1-in-10 dry year

Average year

1-in-10 wet year

1-in-100 wet year

1-in-200 wet year

August 2580 3370 6650 10100 19000 24000 25100 September 573 748 1480 2370 6060 7620 8000 October 143 187 369 592 1730 2180 2300 November 0 0 0 0 0 0 0 December 0 0 0 0 0 0 0 Annual 14300 18700 36900 59200 86500 109000 114000

Geotechnical studies at the project site indicate the thickness of the permafrost layer varies from 190 to 320 m (Dalstroyizyskania, 2004) to a depth of up to 450 m (AMEC, 2004; SRK, 2004). Maximum permafrost depth is expected to exist on exposed elevated topographic regions, such as the Kupol Dome, with minimum permafrost depth in valley bottoms. Thus, the open pit and underground workings at Kupol are not expected to expose sub-permafrost saturated regions that could potentially result in groundwater influxes to the mining operations. Groundwater quality in sub-permafrost arctic regions of Russia and North America is known to be of high salinity (Sader et al., 2003; Borisov et al., 1995; Clark et al., 2000; Douglas et al., 2000; Gascoyne and Kamineni, 1994), posing issues regarding treatment and disposal at mine sites where receiving waters generally have low dissolved solids contents. Groundwater inflows to the open pit will thus likely be restricted to a thawback of exposed permafrost in the active layer during the summer months. This process will only occur while the open pit is increasing in size and exposing fresh permafrost (e.g., 2007 through 2011). Once the open pit reaches its maximize size (2011), and future operations concentrate on the underground mining activities, no net annual efflux of water from thawing of the pit wall active layer is expected to occur. Previous modeling of thawback inflows at Kupol assumed a simplified pit geometry with wall located primarily in competent andesite having volumetric ice content of 10%. Subsequent thermal modeling indicated a water yield of ca. 0.272 m3/m2 of freshly exposed wall face of competent andesite. Subsequent years of exposure were predicted to yield negligible thawback flows. The lithology of the pit wall is not currently included in the resource model, and thus a distribution of exposed lithologies as a function of the pit life cannot be determined. For the purposes of the current water balance, thawback inflows to the open pit were assumed to arise from competent andesite only. Based on the thermal modeling, clayey ash tuff beds are estimated to yield ca. 0.14 m3/m2, heavily weathered bedrock would yield up to ca. 0.616 m3/m2, and fracture rhyolite dykes would yield 0.12 m3/m2. Previous modeling under a prior pit design indicated that thawback inflows from competent andesite would dominate the total thawback inflow (ca. 90%) over each year of pit development. Thus, the assumption of a pit wall located exclusively in competent andesite appears reasonable for the present water balance. Based on an assumed thawback inflow rate from freshly exposed competent andesite of 0.272 m3/m2/y, with negligible thaw flows from exposed walls in subsequent years after excavation (<0.016 m3/m2/y), and the total pit wall and base areas from the current mine plan at the time of writing, the total thawback inflows over 2007 through 2011 is given in Table 4.2.



Table 4.2 Estimated Thawback Inflows over the Project Life

Year Total Pit Wall Area (m2)

Incremental Increase in Exposed Pit Wall Area (m2/y)

Thawback Inflows (m3/y)

2007 203,400 49,000 55,300 2008 261,000 57,600 15,700 2009 309,800 48,800 13,300 2010 342,100 32,300 8,800 2011-2019 342,100 0 0

Thus, using the total pit wall and base areas for each of 2007 through 2011, the estimated inflows to the open pit due to thawback in exposed permafrost are expected to decrease from 42,000 m³/y in 2006 to 8,800 m³/y in 2010. Note that no net thawback of permafrost is expected to occur during 2011 through 2023 since no fresh pit wall is being exposed during this period, and because pit walls exposed for longer than one summer season are expected to yield negligible thawback inflows. Inflows from wall thawback in permafrost will also likely occur in the underground workings. Based on limited available design information, and assuming an average annual air temperature 5°C, previous site-specific thermal modeling suggested that each 100 m of galleries in competent andesite of higher volumetric ice content (10%) would yield about 300 m3/y of inflows in the year of exposure. Note that the underground workings will not be actively heated. However, average daily temperatures are above freezing from May through September (Bema, 2004), and thus, active ventilation will raise air temperatures in the underground workings above the freezing point during this period. Continuing thawback during the summer months in subsequent years following initial exposure of the wall faces is expected to yield significantly lower flow rates at <125 m3/y per 100 m of galleries (AMEC, 2004). The current primary (1°) and secondary (2°) development of underground galleries is expected to progress as follows by year: 2006, 1,090 m of 1°; 2007, 1,520 m of 1° and 2,520 m of 2°; 2008, 1,340 m of 1° and 2,140 m of 2°; 2009, 2,040 m of 1° and 2,110 m of 2°; 2010, 470 m of 1° and 3,910 m of 2°; 2011, 2,070 m of 1° and 5,350 m of 2°; 2012, 850 m of 1° and 1,220 m of 2°; 2013, 94 m of 1° and 0 m of 2°. No underground development is expected during 2014, and the late start to development in 2006 will likely prevent any significant thaw flows in the underground during this year due to cold ambient temperatures in the ventilation system. For the purposes of the preliminary underground workings water balance estimates, It is assumed that only current primary and secondary development will be ventilated. Other areas of the underground are assumed to not be ventilated following the year of their development, and are assumed to re-freeze due to both the presence of surrounding permafrost around the workings. A thawback inflow rate of 125 m3/y per 100 m of galleries is assumed for primary and secondary development following the year of construction. For the third year after construction of new primary and secondary underground, the lack of active ventilation (especially during the summer months) is expected to result in no thaw inflows from these regions.



Based on an inflow rate of 3 m3/m/y of galleries in the year of their development, 1.25 m3/m/y from galleries in the year following their development, and 0 m3/m/y from galleries in the third and future years after their development, along with the development schedule presented above, the estimated annual thawback flows in the underground workings are as follows: 2007, 13,500 m3; 2008, 15,500 m3; 2009, 16,800 m3; 2010, 18,300 m3; 2011, 27,700 m3; 2012, 15,500 m3; 2013, 2,900; and 2014, 100 m3. The 100 m3 of thawback flows in 2014 is due to reduced inputs from the 94 m of primary development taking place during 2013. Total inflows to the open pit are the sum of inflows from net thawback of exposed permafrost and surface runoff from direct precipitation on the open pit. Total pit inflows during an average year are expected to decrease from 101,200 m³/y in 2007 to 68,000 m³/y in 2011, after which time total pit inflows in an average year will remain constant at 59,400 m³/y for 2012 through 2023. To distribute the thawback inflows on an annual basis it was assumed that thawback occurred equally over the June to September period. The monthly distribution of the average, wet and dry year runoff for 2006 to 2010 is presented in Tables 4.3 to 4.7.

Table 4.3 Total Pit Inflows 2007

Total annual Pit Inflows (m³) 2007


1-in-100 dry year

1-in-10 dry year

Average year

1-in-10 wet year

1-in-100 wet year

1-in-200 wet year

January 0 0 0 0 0 0 0 February 0 0 0 0 0 0 0 March 0 0 0 0 0 0 0 April 0 0 0 0 0 0 0 May 286 374 738 1780 865 1090 1140 June 23900 26000 34500 45000 46600 54300 56000 July 21200 22400 27500 33800 46600 54300 56000 August 19800 20600 23800 27300 36200 41200 42300 September 17800 17900 18700 19600 23300 24800 25200 October 143 187 369 592 1730 2180 2280 November 0 0 0 0 0 0 0 December 0 0 0 0 0 0 0 Annual 83100 87500 105700 128000 155300 177800 182800

Table 4.4 Total Pit inflows 2008



1-in-100 dry year

1-in-10 dry year

Average year

1-in-10 wet year

1-in-100 wet year

1-in-200 wet year






1-in-100 dry year

1-in-10 dry year

Average year

1-in-10 wet year

1-in-100 wet year

1-in-200 wet year





1-in-100 dry year

1-in-10 dry year

Average year

1-in-10 wet year

1-in-100 wet year

1-in-200 wet year





1-in-100 dry year

1-in-10 dry year

Average year

1-in-10 wet year

1-in-100 wet year

1-in-200 wet year




Table 4.7 Total Pit inflows 2011



1-in-100 dry year

1-in-10 dry year

Average year

1-in-10 wet year

1-in-100 wet year

1-in-200 wet year





1-in-100 dry year

1-in-10 dry year

Average year

1-in-10 wet year

1-in-100 wet year

1-in-200 wet year





1-in-100 dry year

1-in-10 dry year

Average year

1-in-10 wet year

1-in-100 wet year

1-in-200 wet year

January 0 0 0 0 0 0 0 February 0 0 0 0 0 0 0 March 0 0 0 0 0 0 0 April 0 0 0 0 0 0 0





1-in-100 dry year

1-in-10 dry year

Average year

1-in-10 wet year

1-in-100 wet year

1-in-200 wet year

May 286 374 738 1780 865 1090 1140 June 7400 9500 18000 28500 30100 37800 39500 July 4700 6000 11000 17300 30100 37800 39500 August 3300 4100 7400 10800 19700 24700 25800 September 1300 1500 2200 3100 6800 8400 8700 October 143 187 369 592 1730 2180 2280 November 0 0 0 0 0 0 0 December 0 0 0 0 0 0 0 Annual 17200 21600 39800 62100 89400 111900 116900




1-in-100 dry year

1-in-10 dry year

Average year

1-in-10 wet year

1-in-100 wet year

1-in-200 wet year


5 WASTE ROCK Based on the geochemical testing (Atkin 2005) approximately 40% of the waste rock is anticipated to be acid generating (AG) and therefore will be placed in the TMF. The remaining waste rock will used in the construction of the TMF dam. Therefore the runoff from the dam will be governed by the runoff from the waste rock. During the first two years of pit development prior to the completion of the starter dam (2006 and 2007) the waste rock will be placed within the inundation level of the TMF. 5.1 Alternative #1 During the pre-production phase of the mine a waste pile will be created within the inundation level of the TMF, as indicated on Figure 330-003. The pre-production waste pile will occupy an area of 13.7 ha but collect runoff from an area of 15 ha. The resulting total average annual



runoff is 46,300 m³, as summarized in Table 5.1 as well as the wet and dry years and the monthly distributions.

Table 5.1 Total Annual Runoff from Pre-Production Waste Pile Alternative #1 for Wet Dry and Average Years

Total annual runoff from Pre-Production Waste Rock Pile

Alternative #1(m³)


1-in-100 dry year

1-in-10 dry year

Average year

1-in-10 wet year

1-in-100 wet year

1-in-200 wet year

January 0 0 0 0 0 0 0 February 0 0 0 0 0 0 0 March 0 0 0 0 0 0 0 April 0 0 0 0 0 0 0 May -422 -292 256 1389 874 1210 1290 June -9920 -6860 6020 21800 29700 41100 43900 July -5910 -4090 3580 13000 29700 41100 43900 August -3800 -2630 2300 7870 19200 26600 28400 September -844 -584 512 1850 6120 8470 9030 October -211 -146 128 463 1750 2420 2580 November 0 0 0 0 0 0 0 December 0 0 0 0 0 0 0 Annual -21100 -14600 12800 46300 87400 121000 129000

The surface area of the dam is 10.6 ha. This results in an average annual runoff of 13,800 m³/yr during operations and 29,600 m³/yr at closure. Tables 5.2 and 5.3 summarize the wet, dry and average year monthly-distributed runoff for operations and closure respectively.

Table 5.2 Total Annual Runoff (m³) from Alternative #1 Tailings Dam (during Operations) for Wet, Dry and Average Year

Total annual runoff from Tailings Dam Alternative #1 operations (m³)


1-in-100 dry year

1-in-10 dry year

Average year

1-in-10 wet year

1-in-100 wet year

1-in-200 wet year




Table 5.3 Total Annual Runoff (m³) from Alternative #1 Tailings (at Closure) for Wet, Dry and Average Year

Total annual runoff from Tailings Dam Alternative #1 closure (m³)


1-in-100 dry year

1-in-10 dry year

Average year

1-in-10 wet year

1-in-100 wet year

1-in-200 wet year


5.2 Alternative #2 During the pre-production phase of the mine a waste pile will be created within the inundation level of the TMF, as indicated on Figure 330-004. The pre-production waste pile will occupy an area of 12 ha but collect runoff from an area of 11.5 ha. The resulting total average annual runoff is 37,400 m³, as summarized in Table 5.4 as well as the wet and dry years and the monthly distributions.

Table 5.4 Total Annual Runoff from Pre-Production Waste Pile Alternative #2 for Wet Dry and Average Years

Total annual runoff from Pre-production Waste rock pile Alternative #2

(m³)


1-in-100 dry year

1-in-10 dry year

Average year

1-in-10 wet year

1-in-100 wet year

1-in-200 wet year




Alternative #2 has a dam surface area of 9.8 ha resulting in an average annual runoff of 12,700 m³/yr during operations and 27,300 m³/yr at closure. The wet, dry and average year monthly-distributed runoff for operations and closure is presented in Tables 5.5 and 5.6 respectively.

Table 5.5 Total Annual Runoff (m³) from Alternative #2 Tailings Dam (during Operations) for Wet, Dry and Average Year

Total annual runoff from Tailings Dam Alternative #2 operations (m³)


1-in-100 dry year

1-in-10 dry year

Average year

1-in-10 wet year

1-in-100 wet year

1-in-200 wet year


Table 5.6 Total Annual Runoff (m³) from Alternative #2 Tailings (at Closure) for Wet, Dry and Average Year

Total annual runoff from Tailings Dam Alternative #2 closure (m³)


1-in-100 dry year

1-in-10 dry year

Average year

1-in-10 wet year

1-in-100 wet year

1-in-200 wet year


6 SITE INFRASTRUCTURE



6.1 Plant Site The term plant site will refer to the mill site, fuel storage area and the small area of natural catchments that will be collected by the contact water ditches. This runoff will be collected and diverted to a sediment pond and treated. The total plant site area is 26.7 ha which result in runoff of 74,500 m³/year during an average year and 137,000 m³/year for the 1-in-100 wet year and 23,500 m³/year for the 1-in-100 dry year. The monthly breakdown for all cases can be found in Table 6.1.

Table 6.1 Annual Runoff from Plant Site for Wet, Dry and Average Years

Total annual runoff from Plant Site (m³)


1-in-100 dry year

1-in-10 dry year

Average year

1-in-10 wet year

1-in-100 wet year

1-in-200 wet year


6.2 Ore Stockpile During pre-production runoff from the ore stockpile will pass through a sediment pond and released into the environment. Once the starter dam is complete runoff from the ore stockpile will be diverted to the TMF. The area of the ore stockpile is 3.27 ha. For the ore stockpile the average annual runoff is 9130 m³/year as shown in Table 6.2 along with the monthly distribution and selected dry and wet years.

Table 6.2 Annual Runoff from Ore Stockpile for Wet, Dry and Average Years

Total annual runoff from Ore Stockpile (m³)


1-in-100 dry year

1-in-10 dry year

Average year

1-in-10 wet year

1-in-100 wet year

1-in-200 wet year




Total annual runoff from Ore Stockpile (m³)


1-in-100 dry year

1-in-10 dry year

Average year

1-in-10 wet year

1-in-100 wet year

1-in-200 wet year


6.3 Airstrip An airstrip with an area of 59.7 ha is located on a knoll in the Starichnaya River valley about 10 km north of the mine site. The location of the airstrip on the top of a knoll will allow a majority of the runoff from natural catchments surrounding the airstrip to drain away from the airstrip. The airstrip will be graded to allow runoff to flow into the tributary of the Starichnaya River north of the airstrip. The total runoff from direct precipitation on the airstrip in an average year is 167,000 m³. Table 6.3 summarizes the monthly distribution and the wet and dry years. Sediment fences or sediment ponds will be used if the TSS (Total Suspended Solids) of the runoff from the airstrip is determined to be greater than 50 mg/L.

Table 6.3 Annual Runoff from Direct Precipitation on Airstrip for Wet, Dry and Average Years

Total annual runoff from Airstrip (m³)


1-in-100 dry year

1-in-10 dry year

Average year

1-in-10 wet year

1-in-100 wet year

1-in-200 wet year


7 DESIGN CRITERIA FOR ENGINEERING STRUCTURES 7.1 Culverts All culverts on the site will be designed to convey at least a 100–year flow. This criteria was selected to minimize the need for repair or replacement.



7.2 Ditches and other water conveyance facilities The diversion ditches around the TMF will be designed to convey a 200–year flow. The chance that a 200–year flood will occur during the 7 year mine life is 3.4%. The ditches, other than diversion ditches around the TMF, will be designed to convey the 100–year flow like the culverts. Ditches will be continued until intersection with a stream or other waterbody. Unless it is deemed appropriate to allow the flow to drain naturally exit points of diversion ditches will be designed with erosion control measures. 7.3 Sedimentation Ponds Sedimentation ponds will be used to assist in treating contact water prior to the completion of the starter dam with the exception of plant site runoff which will be treated with sediment ponds for the entire life of the mine. The sediment ponds are designed for adequate operation up to a 10–year maximum 24–hour storm with average snowmelt, industry standard. A 10 micron particle size with a settling velocity of 5x10-5 m/s will be assumed for the design of sediment ponds with a lime treatment system. For the sediment ponds receiving runoff from the plant site and haul roads, a 60 micron particle size with a settling velocity of 2.2x10-2 m/s was used for design. 8 ENGINEERING STRUCTURES 8.1 Culverts It is planned to use shipping containers at most culvert crossings. Standard culvert sizes may need to be used on smaller drainages if the container culvert is deemed too large. The container culverts have interior dimensions of 2.39 m high by 2.35 m wide and 12.2 m long. For the purposes of this design container culverts are assumed to have a 1% slope. If the culvert has a slope greater than 1% the capacity of the culvert will be greater, but armouring should be placed at both the inlet and outlet. A minimal amount of armouring should be placed at both the inlet and outlet at all culvert locations to reduce erosion. For the airstrip access road culverts should be placed in the A4, A6, A7, B2, B3, B4, B5 drainages, as shown in Figure 330-001. The culverts in the A4 and A6 catchments will receive runoff from a portion of the A3 catchment and the A5 catchment. The culvert that will have the largest design flow is the one located in A4 catchment. The design flows for these catchments are summarized in Table 8.1 along with whether a shipping container culvert will convey the flow.

Table 8.1 Airstrip Access Road Culvert Summary

Culvert Location Design Flow (m³/s) Will Container

culvert pass design flow? (Y/N)

A4 6.4 Y



Culvert Location Design Flow (m³/s) Will Container

culvert pass design flow? (Y/N)

A6 3.2 Y

A7 1.1 Y

B2 2.5 Y

B3 1.1 Y

B4 1.1 Y

B5 0.92 Y

A single standard culvert design (750 mm diameter corrugated metal pipe culvert with a 1% slope) will be used for most crossings where a container is not used. A shipping container can be used in place of the 750 mm diameter corrugated metal pipe culvert. Figure 330-005 shows the locations of culverts required for the haul road to TMF alternative #1. Either a 750 mm diameter corrugated metal pipe culvert or shipping container will be adequate for these culverts. The locations of the culverts for the haul road to TMF alternative #2 are shown in Figure 330-006. A 750 mm diameter corrugated metal pipe culvert will be adequate for these culverts and therefore the shipping containers will be adequate as well. 8.2 Ditches Typically runoff water will be conveyed in roadside ditches to the appropriate locations. The exceptions are the ditches around the plant site, the TMF diversion ditches and the ditches diverting water away from the pit. The ditches in the vicinity of the pit and plant site will have a contributing area less than 1 km² Therefore the ditch dimensions presented in Table 8.2 with an assumed slope of 1% and Manning’s n of 0.35 will be able to convey the 100 year flow at most locations.

Table 8.2 General Ditch Dimensions

Base Width (m) Side-Slope Channel Depth (m) Flow (m³/s)

1 m 1.5H:1V 1. m (Includes 0.3 m freeboard)

2.3

The TMF diversion ditches are discussed in section 8.3 and will be sized larger than this typical section.



8.3 TMF Diversion ditches 8.3.1 Alternative #1 Alternative #1 has a catchment area of about 7.6 km2 upstream of the dam location. Diversions will be installed all around both sides of the tailings impoundment, at an elevation just above the projected ultimate storage level, to reduce the downslope drainage area to about 1.2 km2. Both diversion ditches will be extended downstream past the limit of the tailings dam toe to discharge to the existing drainage course. The ditches will be constructed with a design gradient along the ditch bottom of 0.5%, where practicable. These ditches will be designed to convey the 24-hour, 200-year storm event with a minimum 0.3 m of freeboard. HEC-HMS, a drainage basin runoff routing model, was used to route the rainfall event to develop peak design discharges for each tributary catchment at each ditch location. To be conservative, an SCS type II storm distribution was used to produce the largest peak for a given rainfall event and time of concentration. The input parameters for this model will be adjusted as more site stream flow, rainfall and catchment characteristic data are collected. For Alterative #1, ditches on the east side of the tailings pond (draining 1.7 km²) and on the west side (draining 4.7 km²) of the tailings pond will be constructed to collect catchment runoff and discharge it downstream of the dam toe, back into the unnamed creek. Based on the HEC-HMS model, the resulting ditch design flows for a 24-hour, 200-year storm event are:

• 3.2 m³/s for the west ditch • 1.8 m³/s for the east ditch

The resulting ditch cross-sectional configuration is trapezoidal with the dimensions presented in Table 6.3:

Table 8.3 - Alternative #1 - Diversion Ditch Dimensions




A side-discharge release structure, based on either a gated or stop-log system, will be constructed in the west diversion ditch, just upstream of the dam crest to admit summer flows into the TMF to help maintain a water balance reserve for winter mill reclaim requirements. The discharge facility will be sized to accommodate average annual spring freshet flows and will typically be operated on an as-needed basis to supplement mine operations water supply needs.



8.3.2 Alternative #2 Alternative #2 will have two diversion ditches, one on the north side, draining to the east of the tailings pond (draining 2.2 km²) and one on the west side (draining 0.33 km²) of the tailings pond. The diversion ditches will be constructed to collect catchment runoff and discharge downstream of the dam toe. Based on the HEC-HMS model, the resulting ditch design flows for a 24–hour, 200-year storm event are:

• 0.68 m³/s for the west ditch • 2.4 m³/s for the north ditch

The resulting ditch cross-sectional configuration is trapezoidal with dimensions presented in Table 6.7:

Table 8.4 – Alternative #2 - Diversion Ditch Dimensions



North 1 m 1.5H:1V 1.2 m (Includes 0.3 m freeboard)

A side-discharge release facility based on either a gated or stop-log system, will be constructed in the east diversion ditch, just upstream of the dam crest to admit summer flows into the TMF to help maintain a water balance reserve for winter mill reclaim requirements. The discharge facility will be sized to accommodate average annual spring freshet flows and will typically be operated on an as-needed basis to supplement mine operations water supply needs. 8.4 Sediment Ponds The sizes for the site sediment ponds are summarized in Table 8.5. The pit, plant site and ore stockpile sediment ponds are shown in Figure 330-002. Figure 330-005 shows the locations of the haul road sediment ponds. The design event for the pit will be retained within the pit and pumped out over a 10 day period, thus the smaller pond design. At the locations where lime mixing is used (ore stockpile and waste rock pile) a 4-day retention time is used to calculate the pond size. The plant site and haul road sediment pond are designed using a large sediment size because no lime will be added at these location. The size of the haul road and plat sediment pond may further be reduced by installing baffle and/or smaller rock weirs to settle out sediment before the water reaches the pond. The size any sediment pond can be reduced if it can be shown that the size of the particle to be settled or the settling velocity of the particle is greater than the design value.

Table 8.5 Summary of Sediment Pond Sizes

Location Width (m) Length (m)

Depth (m)

Pit 20 40 3



Location Width (m) Length (m)

Depth (m)

Plant Site 10 20 3

Ore Stockpile 10 20 3

North Haul Road

5 10 3

South Haul Road

15 25 3

Pre-production Waste Rock Pile (Alt #1)

25 50 5

Pre-production Waste Rock Pile (Alt #2)

25 45 5

8.5 Closure/Post-Closure Plan The general closure plan for the engineering structures is to fill-in the ditches and sediment ponds and re-grade to resemble existing terrain. Embankments used in the construction of sediments ponds may be used as fill material. Roadside ditches will only be filled-in and re-graded if the road is removed. At locations where culverts are removed because of the removal of a road the pre-construction terrain should be restored and where appropriate streams restored to natural pre-mine state. The TMF will be capped and graded to prevent runoff from ponding and minimize erosion caused by runoff. 9 DISCHARGE TO RECEIVING WATERS The starter dam for the TMF will not be completed until the end of 2007/beginning of 2008. Therefore, during the first two years of pit development (2006 and 2007) all contact water will be treated and discharged to the receiving body (Kaiemraveem Creek). The treatment involves the use of lime mixing and a sediment pond. A dilution pond will be utilized in situations where deemed necessary. The volume that will require treatment in an average year is 230,000 m³ in 2006 and 221,000 m³ in 2007. Multiple sediment ponds will be used for treatment. Effluent from the sewage treatment plant will be discharged to Pervyi Creek, a tributary of Kaiemraveem Creek, during these first two years, but once the starter dam is complete the effluent will be discharged into the TMF. After completion of the starter dam all contact water excluding plant site runoff, will be diverted to the TMF. The treatment volumes required for various years of operations are summarized in Table 9.1.

Table 9.1 Volume of Treatment Water (m³)

Year Alt Average Year 1-in-10 wet year

1-in-100 wet year

2007 1 244,000 351,000 438,000




1-in-100 wet year

2 235,000 334,000 416,000

2008 - 2019

Both 74,400 108,800 136,900

As discussed in Section 6.3 airstrip runoff will be discharged to the Starichnaya River system. The total discharges to Kaiemraveem Creek for selected years of operation are presented in Table 9.2.

Table 9.2 Volume of Discharge Water (m³)


1-in-100 wet year

1 301,000 408,000 495,000 2007

2 292,000 391,000 473,000

2008 - 2019

Both 74,400 108,800 136,900

Numbers presented in Tables 9.1 and 9.2 are summaries from the site water balance tables in Appendix B. 10 CONCLUSION The following conclusions can be drawn from this report:

• On a volumetric basis the treatment of the contact water during the life of the mine is manageable through the use of a lime treatment system.

• The results of the tailings water balance for either alternative show that the Tailings

Management Facility (TMF) will be able to handle all inflows and operate as a zero discharge facility.

• Shipping containers used as culverts will be able to adequately convey the design flows

at all culvert crossings around the site.



11 CLOSURE Recommendations presented herein are based on the modeling assumptions and methodologies described, and upon the baseline hydrology that forms the foundation of the site water balance. The water balance is highly integrated with the site development and mine plans, and mine plans, and must be continually updated as these plans evolve. This report has been prepared for the exclusive use of BEMA Gold Corporation for specific application to the area within this report. Any use which a third party makes of this report, or any reliance on or decisions made based on it, are the responsibility of such third parties. AMEC accepts no responsibility for damages, if any, suffered by any third party as a result of decisions made or actions based on this report. It has been prepared in accordance with generally accepted soil and foundation engineering practices. No other warranty, expressed or implied, is made. Respectfully submitted, AMEC Earth & Environmental, a division of AMEC Americas Limited

Reviewed by:

Per Greg Standen, M.Eng., EIT Water Resources Engineer

M.J. (Jim) Bester, P.Eng. Associate, Water Resources Engineer



REFERENCES

Atkin, S. 2005. Geochemical Characterization of Waste Rock and Ore Materials at the Kupol Project, Russian Far East. Prepared for Bema Gold Corporation: Vancouver, BC, Canada.

AMEC, 2005. Kupol Gold Mining Project – Tailings Dam Feasibility Report. Prepared for BEMA

gold Corporation. Vancouver, BC, Canada. Bema. 2004. Preliminary Economic Assessment: Kupol Gold Project, Far East Russia.

Prepared by Bema Gold Corporation: Vancouver, BC, Canada. Borisov, V.N., Alexeev, S.V., and Pleshevenkova, V.A. 1995. The diamond mining quarries of

East Siberia as a factor affecting surficial water quality. In: Water-Rock Interaction. Eds.: Kharaka, K.K., Chudaev, O.V., Thordsen, J.J., Armannsson, H., Breit, G.N., Evans, W.C., Keith, T.E.C., and Khakara, Y.K. Balkema: Rotterdam, The Netherlands, pp. 863-866.

Clark, I.D., Douglas, M., Raven, K., and Bottomley, D. 2000. Recharge and preservation of

Laurentide Glacial melt water in the Canadian Shield. Ground Water, 38(5), pp. 734-742.

Dalstroyizyskania. 2004. Kupol Mine: Detail Design Geotechnical Survey Report, 19004-II-2.

Prepared for BEMA Gold Corporation: Vancouver, BC, Canada. Douglas, M., Clark, I.D., Raven, K., and Bottomley, D. 2000. Groundwater mixing dynamics at

a Canadian Shield mine. Journal of Hydrology, 235, pp. 88-103. Gascoyne, M. and Kamineni, D.C. 1994. The hydrogeochemistry of fractured plutonic rocks in

the Canadian Shield. Applied Hydrogeology, 2, pp. 43-49. Sader, J.A., Leybourne, M.I., McClenaghan, M.B., Hamilton, S.B., and Robertson, K. 2003.

Field procedures and results of groundwater sampling in kimberlite from drillholes in the Kirkland Lake and Lake Timiskaming areas, northeastern Ontario. Geological Survey of Canada: Current Research, 2003-C11, pp. 1-9.



APPENDIX A : TAILINGS WATER BALANCE – Alternatives #1 and #2

APPENDIX B : SITE WATER BALANCE ANNUAL SUMMARY

Table B.11.1: Site Water Balance Summary for 2007

Annual Volume (m3)

Site Region Average 1-in-10

Wet 1-in-100

Wet Waste Rock Runoff

Pre-Production AG Waste Rock Pile (Alt. 1) 46,300 87,400 121,000 Pre-Production AG Waste Rock Pile (Alt. 2) 37,500 71,100 98,600

Plant Site Runoff 74,500 109,000 137,000 Ore Stockpile Runoff 9,130 13,300 16,800 Pit Inflows 128,000 155,300 177,800

Direct Precipitation 59,200 86,500 109,000 Groundwater Inflows 68,800 68,800 68,800

Total Treatment Volume (Alt 1) 258,000 365,000 453,000 Total Treatment Volume (Alt 2) 249,000 349,000 430,000 STP Effluent 57,100 57,100 57,100 Fresh Water Well 355,700 355,700 355,700 Potable Water Well 66,500 66,500 66,500 Total Water Discharged to Kaiemraveem (Alt. 1) 315,000 422,000 510,000 Total Water Discharged to Kaiemraveem (Alt. 2) 306,000 406,000 487,000


Annual Volume (m3) Alternative 1 Alternative 2


Wet 1-in-100

Wet Average 1-in-10

Wet 1-in-100

Wet Plant Site Runoff 74,500 109,000 137,000 74,500 109,000 137,000 TSF Inflows 1,611,330 2,450,900 3,050,300 1,324,230 1,827,700 2,240,100

Water in Process Plant Effluent 537,000 642,000 642,000 537,000 537,000 537,000 Direct Precipitation on TSF Pond 20,000 49,000 72,000 19,000 33,000 45,000 Natural Runoff Reporting to TSF 875,000 1,534,000 2,072,000 590,000 1,034,000 1,397,000 Pit Inflows Diverted to TSF 90,400 117,700 140,200 90,400 117,700 140,200

Direct Precipitation 59,200 86,500 109,000 59,200 86,500 109,000 Groundwater Inflows 31,200 31,200 31,200 31,200 31,200 31,200

Ore Stockpile Runoff 9,130 13,300 16,800 9,130 13,300 16,800 Sewage Treatment Plant 66,000 66,000 66,000 66,000 66,000 66,000 Make-up Water 0 0 0 0 0 0 Direct Precipitation on Tailings Dam 13,800 28,900 41,300 12,700 26,700 38,100

TSF Outflows 633,000 633,000 633,000 633,000 633,000 633,000

Evaporation from TSF Pond 17,000 17,000 17,000 17,000 17,000 17,000 Storage in Tailings Pore Spaces 260,000 260,000 260,000 260,000 260,000 260,000 Reclaim Water to the Process Plant 356,000 356,000 356,000 356,000 356,000 356,000

Net Change in TSF Storage 978,330 1,817,900 2,417,300 691,230 1,194,700 1,607,100 Fresh Water Well 356,000 356,000 356,000 356,000 356,000 356,000 Potable Water Well 66,500 66,500 66,500 66,500 66,500 66,500




Wet 1-in-100

Wet Average 1-in-10

Wet 1-in-100


Water in Process Plant Effluent 1,095,000 642,000 642,000 1,095,000 1,095,000 1,095,000 Direct Precipitation on TSF Pond 65,000 158,000 234,000 61,000 107,000 144,000 Natural Runoff Reporting to TSF 586,000 1,027,000 1,388,000 348,000 610,000 824,000 Pit Inflows Diverted to TSF 89,300 116,600 139,100 89,300 116,600 139,100



TSF Outflows 1,307,000 1,306,000 1,306,000 1,304,000 1,304,000 1,304,000


Net Change in TSF Storage 617,230 745,800 1,221,200 377,130 730,600 1,019,000 Fresh Water Well 356,000 356,000 356,000 356,000 356,000 356,000 Potable Water Well 66,500 66,500 66,500 66,500 66,500 66,500




Wet 1-in-100

Wet Average 1-in-10

Wet 1-in-100


Water in Process Plant Effluent 1,095,000 1,095,000 1,095,000 1,095,000 1,095,000 1,095,000 Direct Precipitation on TSF Pond 104,000 253,000 374,000 81,000 142,000 192,000 Natural Runoff Reporting to TSF 417,000 731,000 988,000 328,000 575,000 777,000 Pit Inflows Diverted to TSF 86,300 113,600 136,100 86,300 113,600 136,100



TSF Outflows 1,335,000 1,338,000 1,338,000 1,318,000 1,319,000 1,319,000


Net Change in TSF Storage 442,430 933,900 1,337,900 347,430 685,900 963,900 Fresh Water Well 356,000 356,000 356,000 356,000 356,000 356,000 Potable Water Well 66,500 66,500 66,500 66,500 66,500 66,500




Wet 1-in-100

Wet Average 1-in-10

Wet 1-in-100





TSF Outflows 1,357,000 1,358,000 1,358,000 1,328,000 1,329,000 1,329,000






Wet 1-in-100

Wet Average 1-in-10

Wet 1-in-100





TSF Outflows 1,377,000 1,358,000 1,358,000 1,338,000 1,329,000 1,329,000






Wet 1-in-100

Wet Average 1-in-10

Wet 1-in-100





TSF Outflows 1,394,000 1,358,000 1,358,000 1,347,000 1,329,000 1,329,000


Net Change in TSF Storage 398,230 1,011,700 1,484,700 323,230 701,700 997,700 Fresh Water Well 356,000 356,000 356,000 356,000 356,000 356,000 Potable Water Well 66,500 66,500 66,500 66,500 66,500 66,500




Wet 1-in-100

Wet Average 1-in-10

Wet 1-in-100



Direct Precipitation 59,200 86,500 109,000 59,200 86,500 109,000 Groundwater Inflows 100 100 100 100 100 100


TSF Outflows 1,385,000 1,385,000 1,385,000 1,331,000 1,332,000 1,332,000


Net Change in TSF Storage 146,430 559,900 896,900 239,430 541,900 791,900 Fresh Water Well 356,000 356,000 356,000 356,000 356,000 356,000 Potable Water Well 66,500 66,500 66,500 66,500 66,500 66,500

Table B.11.9: Site Water Balance Summary for 2015-2021



Wet 1-in-100

Wet Average 1-in-10

Wet 1-in-100


Water in Process Plant Effluent 730,000 730,000 730,000 730,000 730,000 730,000 Direct Precipitation on TSF Pond 334,000 812,000 1,202,000 279,000 678,000 1,004,000 Natural Runoff Reporting to TSF 54,000 95,000 128,000 38,000 67,000 90,000 Pit Inflows Diverted to TSF 59,200 86,500 109,000 59,200 86,500 109,000

Direct Precipitation 59,200 86,500 109,000 59,200 86,500 109,000 Groundwater Inflows 0 0 0 0 0 0


TSF Outflows 1,085,000 1,085,000 1,085,000 1,045,000 1,045,000 1,045,000



Appendix F AMEC Site Water Management Plan



Water Management Plan:


Submitted to:


Submitted by:


Burnaby, BC

21 June 2005

AMEC File: VM00330

Bema Gold Corporation. Water Management Plan Kupol Gold Project, Far East Russia 21 June 2005

AMEC File: VM00330 Page i

TABLE OF CONTENTS

Page

1.0 INTRODUCTION.................................................................................................................. 1

2.0 WATER MANAGEMENT OBJECTIVES, STRATEGIES, AND STANDARDS..................... 42.1 Objectives.................................................................................................................... 42.2 Strategies .................................................................................................................... 42.3 Standards .................................................................................................................... 5

3.0 WATER BALANCE .............................................................................................................. 63.1 Inflows to the Open Pit and Underground Workings ................................................... 73.2 Fresh Water Abstractions for Potable Water and the Process Plant......................... 103.3 Surface Water Inflows ............................................................................................... 103.4 Outflows .................................................................................................................... 14

4.0 WATER MANAGEMENT SYSTEM.................................................................................... 154.1 Water Supply and Distribution................................................................................... 204.2 Water Management System and Activities................................................................ 21

4.2.1 Water in the Open Pit and Underground Operations – Quantity ................... 214.2.2 Water in the Open Pit and Underground Operations – Quality ..................... 234.2.3 Surface Water – Plant and Camp Site Runoff ............................................... 254.2.4 Surface Water – Ore Stockpile ...................................................................... 264.2.5 Surface Water – Runoff from Roads and the Airstrip .................................... 274.2.6 Surface Water – AG Waste Rock Pile ........................................................... 27

4.3 Water Treatment and Discharge ............................................................................... 304.3.1 Sewage Treatment Plant ............................................................................... 304.3.2 Water Treatment Plant .................................................................................. 314.3.3 Tailings Storage Facility Pond ....................................................................... 314.3.4 Discharges to Ambient Receiving Waters ..................................................... 32

5.0 CONCLUSIONS................................................................................................................. 35

REFERENCES ........................................................................................................................... 36


AMEC File: VM00330 Page ii

LIST OF FIGURES

Figure 1.1: Regional site layout and watershed boundaries..............................................2 Figure 1.2: Plant site layout. ..............................................................................................3 Figure A.1: Conceptual water management system for 2006. .........................................39 Figure A.2: Conceptual water management system for 2007. .........................................40 Figure A.3: Conceptual water management system for 2008 through 2014....................41

LIST OF TABLES Table 4.1: Summary of applicable regulatory limits for water quality. ............................33 Table B.1: Site water balance summary for 2007. ..........................................................43 Table B.2: Site water balance summary for 2008. ..........................................................44 Table B.3: Site water balance summary for 2009. ..........................................................45 Table B.4: Site water balance summary for 2010. ..........................................................46 Table B.5: Site water balance summary for 2011. ..........................................................47 Table B.6: Site water balance summary for 2012 ...........................................................48 Table B.7: Site water balance summary for 2013. ..........................................................49 Table B.8: Site water balance summary for 2014. ..........................................................50


AMEC File: VM00330 Page 1

1.0 INTRODUCTION The Kupol Gold Project will include activities such as open pit mining, underground mining, ore stockpiling and processing, tailings and waste management, and camp operation that will generate volumes and qualities of water requiring management. The feasibility mine plan shows 7.1 million tonnes of material to be processed through the mill during a seven year mine life. The criteria provided by Bema Gold Corporation (Bema) were to develop an operations phase water management plan at a feasibility level of design. For the purposes of this plan, water management is defined as the collection, storage, treatment, discharge, and recycling of water generated at the project site, in a safe, efficient, and compliant manner. This water management plan is conceptual. Water management planning requires a multidisciplinary understanding of water-related issues (e.g., water quantity, water quality, contingency planning, environmental monitoring). A detailed plan, based on final project engineering and final water licence and land use permit conditions, will be developed for each phase of the mine: construction, operation, closure, and post-closure. Subsequent detailed plans will be used to implement appropriate water management measures. The water management plan contains three sections:

• A listing of water management objectives, strategies to implement objectives, and minimum water management standards;

• Tabulated estimate of the water balance (gains and losses of water on site) and a brief

description of each component of the water balance; and,

• An outline of the water management system. The regional site layout with watershed boundaries under tailings alternatives #1 and #2 is shown in Figure 1.1. The plant site layout for the projec is shown in Figure 1.2.



Figure 1.1: Regional site layout and watershed boundaries.



Figure 1.2: Plant site layout.

Bema Gold Corporation Water Management Plan Kupol Gold Project, Far East Russia 21 June 2005


2.0 WATER MANAGEMENT OBJECTIVES, STRATEGIES, AND STANDARDS 2.1 Objectives The goal of water management is to minimize the impact of the Kupol Gold Project on the local and regional aquatic ecosystems of the Kaiemraveem and Starichnaya Rivers. Based on this goal, there are two objectives of the water management plan:

1. Minimize the impacts of the project on the quantity of surface water; and,

2. Minimize the impacts of the project on the quality of surface water. 2.2 Strategies The strategies to implement the objectives of the water management plan are as follows:

1. Minimize the impacts of the project on the quantity of surface water through the following actions:

a) Reducing the intake of fresh water from local aquifers by recycling and reusing

water to the greatest extent practical; and,

b) Reducing the quantity of site contact water to the greatest extent practical.

2. Minimize the impacts of the project on the quality of surface water through the following actions:

a) Collecting, transporting, and containing inflows to the open pit and underground

workings, camp sewage, surface runoff from site infrastructure, and water coming into contact with core project facilities within the closed-circuit tailings storage facility-process plant water system to the extent practical;

b) Managing acid generating (AG) and potentially acid-generating (PAG) materials

with regard to acid rock drainage and metal leaching (ARD/ML) and other water quality concerns;

c) Managing non-acid generating (NAG) materials with regard to metal leaching and

other water quality concerns;

d) Monitoring the quantity and quality of all discharges; and,

e) Adjusting water collection, transfer, treatment, and/or disposal practices through an adaptive management program if monitoring results indicate discharge quantity and/or quality is not meeting discharge criteria.



2.3 Standards The following are the minimum standards that will be incorporated into water management planning:

1. Establish compliance with all applicable environmental legislation including, but not limited to, the following:

a) “Class 1 Fisheries Standards”. Maximum Allowable Concentrations for Rivers

and Waterways, Fisheries Standards. VNIRO: Moscow, Russia, 1999.

b) “Drinking Water and Recreational Use Standards”. Maximum Allowable Concentrations for Rivers and Waterways, Fisheries Standards. VNIRO: Moscow, Russia, 1999.

2. Cross-reference existing guidelines relevant to water management such as the following:

a) World Bank Environment, Health, and Safety Guidelines: Mining and Milling – Open Pit. World Bank: Washington, DC, USA, 1995.

b) World Bank Environment, Health, and Safety Guidelines: Mining and Milling –

Underground. World Bank: Washington, DC, USA, 1995.

c) Pollution Prevention and Abatement Handbook: General Environmental Guidelines. World Bank: Washington, DC, USA, 1998.



3.0 WATER BALANCE The primary information required for design of the water management plan is the Site Water Balance report (AMEC, 2005a). Inflows to the water management system will include the following sources:

• Annual thawback in the active layer and direct precipitation that reports to the open pit; • Annual thawback in stockpiled waste rock and ore exposed on-site;

• Annual thawback in the active layer of the underground workings;

• Surface runoff from various sources such as the plant site (including both the process

plant and camp sites), ore stockpile, fuel storage region, explosives storage site, airstrip, roads (both access and haul roads), waste rock facility, tailings storage facility, and naturally vegetated catchments in and around the project site; and,

• Fresh water intake from groundwater recharge adjacent to the Kaiemraveem River.

Surface runoff will be the largest input to the water management system. Outflows from the water management system will include the following sources:

• Contact water and treated sewage inputs to ambient receiving waters

• Contact water and treated sewage inputs to the tailings storage facility-process plant closed-circuit water balance;

• Evaporation from the tailings impoundment, open water bodies such as streams and

small ponds on-site, and from water used for dust control on haul and access roads, the airstrip, the plant site, the waste rock facility, and the tailings dam;

• Evapotranspiration from naturally vegetated surfaces residing within the site facility

catchments;

• Over-winter sublimation losses from all site regions carrying an active snowpack;

• Permanent losses to storage in the pore spaces of the tailings storage facility; Appendix B lists each component and the estimated annual quantities of water for the operational phase of the project (2006 through 2014) under average (1-in-2 wet year), 1-in-10 wet year, and 1-in-100 wet year return periods. A description of each component is listed below. Section 4.0 of this plan outlines the details of how each component will be managed.



3.1 Inflows to the Open Pit and Underground Workings Geotechnical studies at the project site indicate the thickness of the permafrost layer varies from 190 to 320 m (Dalstroyizyskania, 2004) to a depth of up to 450 m (AMEC, 2004; SRK, 2004a). Maximum permafrost depth is expected to exist on exposed elevated topographic regions, such as Kupol Mountain, with minimum permafrost depth in valley bottoms. Thus, the open pit and underground workings at Kupol are not expected to expose sub-permafrost saturated regions that could potential result in groundwater influxes to the mining operations. Groundwater quality in sub-permafrost arctic regions of Russia and North America is known to be of high salinity (Sader et al., 2003; Borisov et al., 1995; Clark et al., 2000; Douglas et al., 2000; Gascoyne and Kamineni, 1994), posing issues regarding treatment and disposal at mine sites where receiving waters generally have low dissolved solids contents. Groundwater inflows to the open pit will thus likely be restricted to a thawback of exposed permafrost in the active layer during the summer months. This process will only occur while the open pit is increasing in size and exposing fresh permafrost (2006 through 2010). Because pit development is not scheduled to start until the autumn of 2006, no significant pit inflows (from either direct on-pit precipitation or thawback inflows) are expected during this year. For the purposes of the site water balance, pit inflows are assumed to start in 2007. Once the open pit reaches its maximize size (2011 through 2014), and future operations concentrate on the underground mining activities, no net annual efflux of water from thawing of the pit wall active layer is expected to occur. Note that there is a potential for pit development to intersect a perched water table (i.e., saturated talik) between 2006 through 2010. However, the lack of hydrogeologic knowledge at the project site prevents a confident assessment as to the probability of encountering such a structure. Intersection of pit development with a saturated talik would result in unanticipated groundwater inflows not otherwise accounted for in the current water balance. As site investigations and development continue from construction through pit development, the potential to intersect a perched water table should be assessed and, if necessary, incorporated into revised site water balances. Previous modeling of thawback inflows at Kupol assumed a simplified pit geometry with wall located primarily in competent andesite having volumetric ice content of 10%. Subsequent site-specific thermal modeling indicated a water yield of about 0.272 m3/y per square meter of freshly exposed wall face competent andesite. Subsequent years of exposure were predicted to yield negligible thawback flows (AMEC, 2004). The lithology of the pit wall is not currently included in the resource model, and thus a distribution of exposed lithologies as a function of the pit life cannot be determined. For the purposes of the current water balance, thawback inflows to the open pit were assumed to arise from competent andesite only. A review of current site geological mapping indicated this assumption was reasonable based on the most recent pit shells. The previous site-specific thermal modeling also indicated the following thawback inflow yields for various exposed lithologies: clayey ash tuff, about 0.144 m3/m2/y; heavily weathered bedrock, about 0.616 m3/m2/y; and fractured rhyolite dykes, 0.126 m3/m2/y (AMEC, 2004). This previous modeling



under a prior pit design indicated that thawback inflows from competent andesite would dominate the total thawback inflow (about 90%) over each year of pit development (AMEC, 2004). Thus, the assumption of a pit wall located exclusively in competent andesite appears reasonable for the present water balance. Based on an assumed thawback inflow rate from freshly exposed competent andesite of 0.272 m3/m2/y, with negligible thaw flows from exposed walls in subsequent years following their excavation (<0.016 m3/m2/y), and the total pit wall and base areas from the current mine plan at the time of writing, the total annual thawback inflows over 2007 through 2014 are as follows: 2007, 55,300 m3; 2008, 15,700 m3; 2009, 13,300 m3; 2010, 8,800 m3; and 2011-2014, 0 m3. Note that no fresh pit wall is exposed between 2011 through 2014, and thus, thawback inflows are assumed to be negligible over this period. Further details on the derivation of these inflow volumes, and their estimated distribution on a monthly basis, are given in the Site Water Balance report (AMEC, 2005a) Thus, estimated inflows to the open pit due to thawback in exposed permafrost are expected to decrease from 55,300 m3/y in 2006 to 8,800 m3/y in 2010, and to an effective thawback inflow of 0 m3/y over the period from 2011 through 2014. Note again that no net thawback of permafrost is expected to occur between 2011 and 2014 since no fresh pit wall is being exposed during this period, and because pit walls exposed for longer than one summer season are expected to yield negligible thawback inflows. In reality, continued exposure of pit walls for longer than one summer season will result in small thawback inflow contributions. However, thawback inflows from pit wall and base surfaces exposed for longer than one season are expected to be minor in comparison to runoff from direct precipitation on the plan area of the exposed pit walls and base. Inflows from wall thawback in permafrost will also likely occur in the underground workings. Based on limited available design information, and assuming an average annual air temperature 5°C, previous site-specific thermal modeling suggested that each 100 m of galleries in competent andesite of higher volumetric ice content (10%) would yield about 300 m3/y of inflows in the year of exposure. Note that the underground workings will not be actively heated. However, average daily temperatures are above freezing from May through September (Bema, 2004), and thus, active ventilation will raise air temperatures in the underground workings above the freezing point during this period. Continuing thawback during the summer months in subsequent years following initial exposure of the wall faces is expected to yield significantly lower flow rates at <125 m3/y per 100 m of galleries (AMEC, 2004). The current primary (1°) and secondary (2°) development of underground galleries is expected to progress as follows by year: 2006, 1,090 m of 1°; 2007, 1,520 m of 1° and 2,520 m of 2°; 2008, 1,340 m of 1° and 2,140 m of 2°; 2009, 2,040 m of 1° and 2,110 m of 2°; 2010, 470 m of 1° and 3,910 m of 2°; 2011, 2,070 m of 1° and 5,350 m of 2°; 2012, 850 m of 1° and 1,220 m of 2°; 2013, 94 m of 1° and 0 m of 2°. No underground development is expected during 2014, and the late start to development in 2006 will likely prevent any significant thaw flows in the underground during this year due to cold ambient temperatures in the ventilation system. For the purposes of the preliminary underground workings water balance estimates, It is assumed that only current primary and secondary development will be ventilated. Other areas



of the underground are assumed to not be ventilated following the year of their development, and are assumed to re-freeze due to both the presence of surrounding permafrost around the workings. A thawback inflow rate of 125 m3/y per 100 m of galleries is assumed for primary and secondary development following the year of construction. For the third year after construction of new primary and secondary underground, the lack of active ventilation (especially during the summer months) is expected to result in no thaw inflows from these regions. Based on an inflow rate of 3 m3/m/y of galleries in the year of their development, 1.25 m3/m/y from galleries in the year following their development, and 0 m3/m/y from galleries in the third and future years after their development, along with the development schedule presented above, the estimated annual thawback flows in the underground workings are as follows: 2007, 13,500 m3; 2008, 15,500 m3; 2009, 16,800 m3; 2010, 18,300 m3; 2011, 27,700 m3; 2012, 15,500 m3; 2013, 2,900; and 2014, 100 m3. The 100 m3 of thawback flows in 2014 is due to reduced inputs from the 94 m of primary development taking place during 2013. Water yields from direct precipitation on the open pit were calculated as follows. Total average annual precipitation at the project site is about 380 mm. Over-winter sublimation losses total approximately 70 mm, leaving about 310 mm available for runoff. Details on the derivations of estimated annual unit runoff quantities from various surfaces are provided in the Site Water Balance (AMEC, 2005a) and Tailings Facility Feasibility Design (AMEC, 2005b) reports. With a runoff coefficient of 0.9 to account for minimal evaporative losses in the open pit, a net average annual unit runoff of 280 mm is obtained. Surface runoff in the open pit during an average year is thus expected to remain approximately constant at 59,200 m3/y over the project life, with the assumption of the pit reaching its maximum plan view areal dimension (about 212,000 m2) during 2007. Total inflows to the open pit are the sum of inflows from net thawback of exposed permafrost in both the open pit and underground workings, and surface runoff from direct precipitation on the open pit. Estimated total pit inflows during an average runoff year over the mine life as are follows: 2007, 128,000 m3; 2008, 90,400 m3; 2009, 89,300 m3; 2010, 86,300 m3; 2011, 86,900 m3; 2012, 74,700 m3; 2013, 62,100 m3; and 2014, 59,300 m3. Based on available geochemical testing and the estimated distribution of AG, PAG, and NAG rock in the pit walls (AMEC, 2005c), surface pit inflows will be of low pH, contain high concentrations of metals and dissolved solids, and be unsuitable for direct discharge without treatment based on the World Bank criteria (World Bank, 1999a and 1999b). Thus, during 2007, net runoff within the open pit and underground workings will be treated with lime to raise the pH to neutral values (pH 7-8), allowed to settle in a flow-through sedimentation pond (having a nominal hydraulic retention time [HRT] of ≥3-4 days) for metals precipitation and subsequent removal, and then released to ambient receiving waters. From late 2007 through the remainder of the project life, pit inflows will be pumped directly to the tailings storage facility for use in the process-plant closed circuit water balance. Details on the design criteria and location of the treatment/sedimentation pond for pit inflows in 2007 is given in the Site Water Balance report (AMEC, 2005a).



3.2 Fresh Water Abstractions for Potable Water and the Process Plant Fresh water will be drawn from aquifers located in permanently thawed regions (taliks) adjacent to the Kaiemraveem River and downstream of the project site. The amount of fresh water required for potable water consumption and other camp activities is estimated to be up to 160 m3/d (1.9 L/s) during the operational phase of the project (600 persons at 265 L/d consumption). During the summers of 2006 and 2007, the camp size may increase to a maximum of 900 persons. Thus, potable water requirements during the summers of 2006 and 2007 may increase to 240 m3/d (2.8 L/s). A portion of the make-up water for the process plant will also be obtained from taliks immediately adjacent to the Kaiemraveem River. The amount of make-up water required from these sources is estimated to range up to 40,000 m3/month (19 L/s) for each of tailings alternatives #1 and #2 (AMEC, 2005b). Details on the expected rates and timing for required make-up water diversions into the tailings pond are given in both the Site Water Balance (AMEC, 2005a) and Tailings Facility Feasibility Design (AMEC, 2005b) reports. Potable and process plant fresh water requirements will need to be satisfied on a continuous year-round basis. The intent is to recycle water to the greatest extent practical from the tailings impoundment in order to reduce the amount of fresh water required. Local creeks and the Kaiemraveem and Starichnaya Rivers do not flow during the over-winter period (from approximately mid-November through to the end of May) in an average climatic year. No reservoir storage of fresh water is currently planned. 3.3 Surface Water Inflows Water inflow on the surface is comprised of direct precipitation and runoff from the mine site and adjacent catchments and runoff/drainage from the waste rock and tailings storage facilities. Direct precipitation information was gathered from local, regional, and national climate records and on-site measurements. Detailed information on precipitation estimates and their derivations may be found in the Site Water Balance report (AMEC, 2005a). In an average year, approximately 74,000 m3 of direct precipitation and runoff from core plant site facilities such as the process plant, the camp site, the service complex, and the fuel storage area will be routed and temporarily stored in the sumps and sedimentation ponds located around the plant site. Following its collection and settling in sedimentation ponds, plant site runoff from all regions, other than the ore stockpile with the exception of 2007 (see below), will be allowed to drain into the headwaters of Pervyi Creek. Monitoring of runoff water quality from the plant site will be conducted over the project life. If the water quality of the plant site runoff after sedimentation is not sufficient for direct discharge based on the World Bank criteria (World Bank, 1999a and 1999b), the plant site runoff will be collected and pumped to the tailings pond for use in the process plant circuit. Runoff from the ore stockpile will be of low pH, contain high concentrations of metals and dissolved solids, and be unsuitable for direct discharge without treatment (AMEC, 2005c) based on the World Bank criteria (World Bank, 1999a and 1999b). Thus, during 2007, the small



annual quantity of runoff from the ore stockpile (about 4,400 m3/y) will be treated with lime to raise the pH to neutral values (pH 7-8), allowed to settle in a flow-through sedimentation pond (HRT≥3-4 days) for metals precipitation and subsequent removal, and then released to ambient receiving waters along with the remainder of the plant site runoff. Note that the estimated runoff quantity in the ore stockpile presented above does not include any thaw flows from ice inclusions in the mined rock. Assuming a 3,000 tpd production rate for ore during the one-year period between October 2009 (assumed start of mining activities) and October 2010 (the assumed date of completion for the starter tailings dam), about 1.1 million tonnes of ore would need to be stockpile over this time. If the ore contained 10% ice by weight (equivalent to the estimates for competent andesite), then up to 42,000 m3 of thaw flows (assuming a specific gravity of 2.65 for the rock mass) could be released from the ore stockpile during 2007. This calculation assumes the entire ore stockpile thaws during 2007, which seems unlikely given that the ore will be brought to the surface and stockpiled in a frozen state, and the a significant portion of the ore will be stockpiled from October 2009 through May 2010, during which time average air temperatures are below zero and no thawing of exposed rock is expected. However, the geochemical testing indicates the ore is expected to have elevated sulfide contents and be net acid generating. Sulfide oxidation will occur in the ore stockpile even under frozen conditions as oxygen will have ready access to the surface of the coarse rock. Thermal modeling has not been perform to determine if the potential rates of sulfide oxidation in the ore stockpile are sufficient to thaw the rock mass during summer and winter conditions. While not included in the site water balance model due to the current uncertainty of its significance, it must be noted that the potential for thaw flows from the ore stockpile in 2007 could increase the estimated average required water treatment capacity for this site region by up to an order of magnitude. Additional study with thermal modeling is recommended prior to the start of mine construction in order to examine the potential extent of thawing in the ore stockpile during 2007. This work would ensure that any additional treatment chemicals for the ore stockpile runoff are in place for 2007, and that larger runoff/thaw flow containment and treatment structures are built prior to the summer of 2007. Runoff from haul and access roads, the airstrip, and the explosives storage site will be collected via a series of ditches and culverts, routed to sedimentation ponds for settling, and allowed to directly report to the nearest natural receiving water body provided the quality is suitable for direct discharge as specified by the World Bank guidelines (World Bank, 1999a and 1999b), and that ambient fresh water regulatory limits (VNIRO, 1999) will not be exceeded downstream at the regulatory point-of-compliance. All NAG waste rock, and a significant portion of the PAG waste rock, obtained over the first four years of the project (where the open pit is being developed) will be used to construct the starter tailings dam and its subsequent raises. The two alternatives for locating the tailings storage facility are shown in Figure 1.1 and Figure 1.2, respectively. AG waste rock will be placed within the tailings storage facility, and will eventually be submerged by both saturated tailings and a free water cover in order to minimize production of acidic runoff. At closure, the free water



surface of the tailings pond will be removed. However, the AG waste rock will remain covered with saturated tailings at closure, after which time the entire tailings and waste rock storage facility is expected to freeze. The saturated tailings cover on the AG waste rock will be of a depth exceeding the expected active layer, such that thawing and potential acid generation in the buried AG waste rock will be prevented. A ditching system will be in place starting in late 2006 to divert natural runoff around the proposed tailings dam location. The storage volume behind the starter tailings dam (about 4,000,000 m3) following its completion in late 2007 will not be available for retaining contact water during this year. Thus, the 2007 site water balance must plan for no available storage behind the starter tailings dam. Of note, the diversion ditching system in-place upstream of the starter tailings dam will also function as the diversion system around the operational tailings storage facility from late 2007 through the remainder of the project life. Further details on the design and operation of the tailing storage facility are provided in the Tailings Facility Feasibility Design Report (AMEC, 2005b). During 2006 and 2007, about 877,000 tonnes and 1,585,000 tonnes, respectively, of AG waste rock will be stored upstream of the proposed starter dam locations. The total volumes of AG waste rock will cover plan areas of 290,000 m2 or 235,000 m2 at the end of 2007 under tailings alternatives #1 and #2, respectively. Because of the late start to pit development in 2006, no surface runoff from the AG waste rock pile is assumed to occur during this year. Thus, the AG waste rock will generate about 46,300 m3 or 37,500 m3 of surface runoff in 2007 during an average year under tailings alternatives #1 and #2, respectively. In addition to surface runoff arising from direct precipitation on the AG waste rock pile during 2007, thawing of entrained ice in the waste rock may contribute up to 10% by volume of the total rock mass stored at this facility. The geochemical testing indicates the AG waste rock will likely generate net acidity immediately after placement, and the exothermic nature of the sulfide oxidation process may be such to prevent the presence of unfrozen rock in the pile. The difficulty in maintaining frozen rock within net acid generating waste rock facilities in northern permafrost environments (even during the over-winter period following winter placement and burial of frozen rock) has been demonstrated elsewhere (SRK, 2004b; Rescan and SRK, 2004). Assuming all entrained ice in the AG waste rock pile melts and is released as drainage during 2007, up to 93,000 m3 of water (assuming a specific gravity of 2.65 for the rock mass) may be produced that will require collection and treatment prior to discharge. While not included in the site water balance model due to the current uncertainty of its significance, it must be noted that the potential for thaw flows from the AG waste rock stockpile in 2007 could increase the estimated average required water treatment capacity for this site region by up to a factor of 3-4. Additional study with thermal modeling is recommended prior to the start of mine construction in order to examine the potential extent of thawing in the AG waste rock stockpile during 2007. This work would ensure that any additional treatment chemicals for the stockpile runoff are in place for 2007, and that larger runoff/thaw flow containment and treatment structures are built prior to the summer of 2007. Runoff from the AG waste rock will be isolated to minimize total volumes requiring storage and treatment. Diversion ditches will be installed in early 2006 around the proposed AG waste rock



storage sites to minimize runoff. Runoff from the AG waste rock pile in 2007 will be collected and treated using a lime treatment system with a flow-through sedimentation pond (HRT≥3-4 days) as described below. Following completion of the starter tailings dam in late 2007, runoff from the AG waste rock pile will be allowed to report to the tailing pond. All AG waste rock generated after 2008 will be added to the 2006/2007 pile within the tailings storage facility. The AG waste rock placed in the tailings storage facility will be completely submerged under both tailings and a free water column cover over the operating life of the mine. Runoff and drainage from the tailings dam is comprised of runoff from direct precipitation and water used to suppress dust. No estimates of water used for dust control are currently available, but should be included in future updates to the site water balance. The tailings dam will be constructed using NAG and PAG waste rock obtained during the open pit portion of the mine life during 2006 through 2010. Details on the feasibility level tailings dam design, along with measures for reducing any risk from PAG materials used in constructing this facility, are presented in the Tailings Facility Feasibility Design Report (AMEC, 2005b). From 2011 through 2014, mining operations will be exclusively underground, and any waste rock generated over the underground portion of the mine life will be retained in unused regions of the open pit and/or used as underground backfill. Drainage can occur within the tailings dam (internal) as well as on the outside slopes (external). The amount of drainage from the tailings dam will increase as the dam size increases. The upstream liner and promoting of frozen rock fill during dam construction will reduce seepage through the dam face to near zero, thus resulting in direct precipitation (and subsequent surface runoff and infiltration) being the dominant source component of the tailings dam water balance. Because the relative amount of infiltration, and subsequent internal drainage, from the tailings dam cannot be reliably estimated without advanced modeling methods, internal and external drainage from the tailings dam is combined for the purposes of the water balance. As a conservative measure for site water balance planning purposes, the external and internal drainage from the tailings dam was estimated on the assumption of the facility reaching maximum size in late 2007 (coincident with completion of the starter tailings dam). Alternative #1 for the tailings dam has a maximum total surface area at El. 561.5 m outside the dam crest of 106,000 m2, with a plan view area of 94,800 m2. During 2006 through 2010, when the tailings dam is actively increasing in size and net wetting of the rock fill is occurring, an average year of runoff for Alternative #1 is estimated at 12,400 m3. From 2011 through 2014, where the tailings dam is complete and no further raises take place, an average year of runoff is estimated at 26,600 m3. Alternative #2 for the waste rock facility has a maximum total surface area at El. 560.5 m outside the dam crest of 97,800 m2, with a plan view area of 87,500 m2. During 2006 through 2011, when the tailings dam is actively increasing in size and net wetting of the rock is occurring, an average year of runoff for Alternative #2 is estimated at 11,400 m3. From 2011 through 2014, where the waste rock facility is closed and no further raises take place, an average year of runoff is estimated at 24,500 m3. Current geochemical and water quality modeling suggests external drainage from the downstream face of the tailings dam may be suitable for untreated release into ambient



receiving waters, as the major source of this drainage will be from water contacting NAG rock used in dam construction. However, for conservative design purposes, runoff and drainage from both tailings dam alternatives is assumed to require collection in a series of ditches, sumps, and collection ponds at the dam toe, and to be subsequently pumped into the tailings storage facility. The runoff/leachate water quality from the tailings dam will be monitored over the project life. If runoff/leachate from the tailings dam is of sufficient quality, the water will be allowed to directly discharge into ambient receiving waters. Note that these discussions do not include values for net direct precipitation on the tailings ponds and upstream face of the tailings dam, runoff from natural undiverted catchments reporting to the tailings pond, runoff intentionally diverted into the tailings pond for make-up requirements in the process plant, inputs to the tailings pond from the process plant, and losses to pore spaces in the settled tailings. For details on these portions of the tailings storage facility water balance, the reader is referred to the Tailings Facility Feasibility Design Report (AMEC, 2005b). 3.4 Outflows The majority of water inflows on site will be collected, transported, and released to ambient receiving catchments that feed into the Kaiemraveen River. Surface runoff from the airstrip, and the access road to the airstrip, will be collected, transported, and released to ambient receiving catchments that feed into the Starichnaya River. Water losses on site will include evaporation and permanent storage in the pore spaces of the tailings impoundment (over the period from 2008 through 2014 when the tailings storage facility is operational). Evaporation of water from the waste rock facility, tailings pond, and other site facilities was estimated based on climate information discussed in the Site Water Balance report (AMEC, 2005a). For further details on the tailings storage facility water balance, the reader is referred to the Tailings Facility Feasibility Design Report (AMEC, 2005b). Note that the tailings storage facility is designed as a zero-discharge impoundment for the operational life of the mine.



4.0 WATER MANAGEMENT SYSTEM The water management system is the infrastructure and practices that are designed to physically, chemically, and biologically manage water. The design of the system is founded on the water management objectives, strategies, and standards discussed above in Section 2.0, and also based on sound environmental and engineering considerations. Contingencies for unexpected events and emergencies will be built into the system, and an adaptive management approach will be taken in order to best respond to, and address, expected changes in the mine plan and processing operations over the project life. The water management system is as follows:

• Ditches and culverts will be sized to convey the 1-in-200 wet year freshet (defined as melting of the average year spring snow-water equivalent within a 24 hour period plus a 1-in-200 year 24 hour rainfall event occurring at the same time, i.e. “a 1-in-200 wet year rain-on-snow event”). Sedimentation ponds will be sized to store the 1-in-200 wet year freshet volume from the relevant catchment area.

• Fresh water will be drawn from taliks located immediately adjacent to the Kaiemraveem

River and piped to a storage location at the plant site. Fresh water required for the process plant and other site non-potable requirements (e.g., wash-water, equipment cleaning) will not be treated prior to use. Fresh water used for drinking needs will be treated with chlorine and then distributed to process plant and camp facilities as potable water.

• During the period from 2008 through 2014, thawback inflows and surface runoff within

the open pit and underground workings will be pumped to the tailings storage facility for use as reclaim water in the process plant, and for permanent storage in the tailings pore spaces. Under current estimates for inflow rates and volumes to the open pit and underground workings, no inflows to these regions will be released to ambient receiving waters between 2008 and 2014.

• During 2007, pit inflows will be treated with a lime-addition system in a flow-through

sedimentation pond for pH adjustment. Treated pit inflows will be allowed to settle for removal of particulate phase contaminants (both suspended sediments and particulate/colloidal phase metals), and released to ambient receiving waters in an unnamed right tributary of the Kaiemraveem River near the open pit (see the Site Water Balance report (AMEC, 2005a) for locations of the treatment/sedimentation ponds and discharge locations). The quality and quantity of this treated water will be monitored to ensure it complies with applicable direct discharge limits under the World Bank criteria (World Bank, 1999a and 1999b), and that sufficient dilutional capacity exists so ambient receiving water quality limits at the regulatory point-of-compliance (VNIRO, 1999) in the Kaiemraveem River will not be exceeded.

• Over the project life, surface runoff from the plant and camp sites (with the exception of

runoff from the ore stockpile), the AG waste rock facility, the outside crest of the tailings



dam, and other core site facilities will be collected in sumps, ponds, and ditches, transferred to sedimentation ponds for removal of suspended solids and particulate/colloidal phase contaminants, and allowed to report to ambient receiving water bodies (tributaries of the Kaiemraveem River). The quantity and quality of runoff from these regions will be monitored to ensure it complies with applicable direct discharge limits under the World Bank criteria (World Bank, 1999a and 1999b), and that sufficient dilutional capacity exists so ambient receiving water quality limits at the regulatory point-of-compliance (VNIRO, 1999) in the Kaiemraveem River will not be exceeded.

Treated runoff from the plant site will be allowed to report to the headwaters of Pervyi Creek (a right tributary of the Kaiemraveem River upstream of the project site). Treated runoff from the AG waste rock facility and outside crest of the tailings dam will be allowed to report to either Tretyi Creek (for tailings alternative #1) or Vtoroi Creek (for tailings alternative #2).

• During the period from 2008 through 2014, surface runoff from the ore stockpile will be

pumped to the tailings storage facility for use as reclaim water in the process plant, and for permanent storage in the tailings pore spaces. Under current estimates for runoff from the ore stockpile, no water yields from this region will be released to ambient receiving waters between 2008 and 2014.

• During late 2006 and all of 2007, the open pit will generate ore for future processing, but

the process plant will not become operational until either late in 2007 or early in 2008. Thus, all ore produced during 2006 and 2007 will need to be stockpiled at the plant site for subsequent processing that will begin in 2008. Surface runoff from the ore stockpile is expected to be acidic with high concentrations of metals and major ions such as sulfate, iron, and calcium.

The ore stockpile runoff will be treated with a lime-addition system in a flow-through sedimentation pond for pH adjustment, allowed to settle for removal of particulate phase contaminants (both suspended sediments and particulate/colloidal phase metals), and pumped to the runoff collection system for the plant site. Thus, treated runoff from the ore stockpile will be mixed with runoff from the plant and camp sites (thus being subjected to an additional settling period in the plant and camp site sedimentation ponds for maximum removal of particulate colloidal contaminants resulting from the lime treatment process), and allowed to report to the ambient headwater receiving environment of Pervyi Creek. The quality and quantity of this treated water will be monitored to ensure it complies with applicable direct discharge limits under the World Bank criteria (World Bank, 1999a and 1999b), and that sufficient dilutional capacity exists so ambient receiving water quality limits at the regulatory point-of-compliance (VNIRO, 1999) in the Kaiemraveem River will not be exceeded.

• Surface runoff from haul and access roads and the airstrip over the project life will be

collected in a series of ditches and released to ambient receiving waters. If necessary, especially on steep slopes where erosion may be problematic, small ponds that function as both equalization basins (to attenuate peak flows released to ambient receiving water



bodies) and sediment traps will be installed within the road and airstrip runoff collection systems.

The quantity and quality of runoff from these regions will be monitored to ensure it complies with applicable direct discharge limits under the World Bank criteria (World Bank, 1999a and 1999b), and that sufficient dilutional capacity exists so ambient receiving water quality limits at the regulatory points-of-compliance (VNIRO, 1999) in the Kaiemraveem and Starichnaya Rivers will not be exceeded. Best management practices will also be followed to help minimize erosion of roads and the airstrip, and thus reduce sediment loadings to ambient receiving water bodies and/or associated mitigative treatment costs.

• During 2006 and 2007, the open pit will generate AG waste rock, but the starter tailings

dam will not be completed until late 2007. Surface runoff from the AG waste rock is expected to be acidic with elevated concentrations of metals and major ions such as sulfate, iron, and calcium. The autumn start date of pit development in 2006 will result in no anticipated runoff from the AG waste rock pile during this year. During late 2006 and all of 2007, AG waste rock from the open pit will be deposited in the valley of the tailings storage facility (for either option) at a point upstream of the starter tailings dam construction. For tailings alternative #1, the AG waste rock pile (and tailings dam) will be located in the catchment of Tretyi Creek. For tailings alternative #2, the AG waste rock pile (and tailings dam) will be located in the catchment of Vtoroi Creek.

Diversion ditching around the AG waste rock pile will limit net runoff to the exposed plan area of this location to about 37,000 m3 or 46,000 m3 in 2007 under tailings alternatives #1 and #2, respectively. The AG waste rock pile runoff will be treated with a lime-addition system in a flow-through sedimentation pond for pH adjustment, allowed to settle for removal of particulate phase contaminants (both suspended sediments and particulate/colloidal phase metals), and released to ambient receiving waters in Tretyi Creek (for tailings alternative #1) or Vtoroi Creek (for tailings alternative #2). The quality and quantity of this treated water will be monitored to ensure it complies with applicable direct discharge limits under the World Bank criteria (World Bank, 1999a and 1999b), and that sufficient dilutional capacity exists so ambient receiving water quality limits at the regulatory point-of-compliance (VNIRO, 1999) in the Kaiemraveem River will not be exceeded.

• Prior to completion of the starter tailings dam in late 2007, natural runoff outside the AG waste rock storage areas in either tailings option catchment will be diverted away from the waste rock pile via a series of ditches, and allowed to report to ambient receiving waters in Tretyi Creek (for tailings alternative #1) or Vtoroi Creek (for tailings alternative #2).

• Following completion of the starter tailings dam in late 2007, runoff within a reduced

portion of the respective contributing catchments upstream of either tailings dam option will be allowed to report to the tailings pond. The remainder of the natural surface runoff in the catchments upstream of the tailings dam will be diverted around the tailings storage facility via a series of ditches, and allowed to report to ambient receiving waters



in Tretyi Creek (for tailings alternative #1) or Vtoroi Creek (for tailings alternative #2). Water stored in the tailings storage facility will be recycled and reused to the greatest extent possible for make-up water to the process plant, dust suppression on roads, the airstrip, the downstream face of the tailings dam, and other site facilities, and for fire suppression.

• Sewage will be treated in a sewage treatment plant (STP). During 2006 and 2007, and

prior to construction of the starter tailings dam, effluent from the STP will be discharged year-round into the headwater region of the Pervyi Creek immediately adjacent to the plant site. Discharge during the over-winter period will result in about 32,000 m3 of ice accumulation near the end-of-pipe in the Pervyi Creek headwaters. The over-winter treated sewage ice accumulation will melt during the subsequent freshet period in May/June, during which time STP discharges over ensuing the open-water season from May through October into the Pervyi Creek headwaters will be occurring. The quantity and quality of STP discharges will be monitored to ensure it complies with applicable direct discharge limits under the World Bank criteria (World Bank, 1999a and 1999b), and that sufficient dilutional capacity exists so ambient receiving water quality limits at the regulatory point-of-compliance (VNIRO, 1999) in the Kaiemraveem River will not be exceeded. Note that the World Bank criteria for direct discharge are end-of-pipe guidelines.

Following completion of the starter tailings dam in late 2007, the STP effluent will be pumped to the tailings storage facility for use in the process plant closed-circuit water balance. The diversion of STP effluent into the tailings storage facility from late 2007 through the remainder of the project life is intended to minimize any potential risks of drinking water contamination in the potable water wells located adjacent to the Kaiemraveem River approximately 15 km downstream (along watercourse paths in the Pervyi Creek and Kaiemraveem River) of the proposed STP discharge. During 2006 and 2007, it is expected the STP will achieve sufficient treatment efficiency for pathogen removal in its tertiary level UV unit to negate any potential risks for downstream potable water abstractions. In an average year, the dilution factor for STP effluent at the proposed potable water abstraction wells in the Kaiemraveem downstream of the project site is a factor of about 300. This degree of dilution, combined with the tertiary level STP treatment process, the long travel time of STP effluent in the Kaiemraveem River prior to any downstream potable water abstractions (thereby allowing natural attenuation to further lower any residual pathogen and/or contaminant levels), and the fact that STP effluent will not be released to the Kaiemraveem River during the low-flow winter period (i.e., the effluent will remain frozen in the headwaters of the Pervyi Creek), will likely result in a minimal risk of pathogen/contaminant presence near the potable water intakes in the Kaiemraveem River. Further riverbank filtration processes (the potable water wells are located in the consolidated sand and gravel aquifer talik zone adjacent to the Kaiemraveem River), and subsequent chlorination prior to distribution in the camp and process plant sites, will also act to minimize the potential risk to the quality of the site drinking water source.



Sludge produced by the STP will be sent to a lined landfill facility during 2006 and 2007. Following start-up of the tailings storage facility in late 2007, sewage sludge will be sent to the AG waste rock pile within the tailings storage facility.

The components and practices of water management system vary with each phase of the project. An outline of the system during each phase of the project will be described in detail following the description of the supply and distribution of water. The supply and distribution will be discussed separately since they remain consistent regardless of the phase of the project.



4.1 Water Supply and Distribution Fresh water for potable consumption will be drawn from aquifers located in taliks immediately adjacent to the Kaiemraveem River and piped to a storage location where it will be treated with chlorine and then distributed to camp facilities as potable water. The estimated required volume of potable water ranges from a maximum of 230 m3/d during the summers of 2006 and 2007 (254 L/p/d at a 900 man camp) to approximately 150 m3/d during the remainder of the operational phase (254 L/p/d at a nominal average 600 man camp). These are conservative estimates provided mainly to allow sizing of the potable water distribution system and the sewage treatment plant. Increases in potable water use efficiency over the operational period should reduce the volume of fresh water required. Fresh water obtained via wells adjacent to the Kaiemraveem River will also be used for make-up requirements in the process plant-tailings storage facility closed circuit water balance, as well as for fire suppression. Process plant fresh water requirements from the Kaiemraveem River are estimated at about 82,000 m3 during an average year for tailings alternatives #1 and #2 (Orocon, 2005). The fresh water intake and pumphouse will be located on the western shore of the Kaiemraveem River approximately 6.5 km downstream of the project site. The intake will consist of vertical filtration wells fitted with vertical turbine pumps that supply water on demand. Fresh water for potable purposes will be pumped via an overland heat-traced pipeline to the fresh water storage tank in the mill from where it will be pumped to a potable water tank via a treatment plant. The pipeline will consist of insulated and heat-traced, high-density polyethylene pipe, on surface adjacent to the service road. Potable water will be sampled daily to ensure it meets health regulations for total and residual chlorine and microbiological parameters. Treated water will be piped to the accommodations and service complexes. Fresh water obtained from the Kaiemraveem River for make-up purposes in the process plant will be pumped directly to the mill fresh water tank without treatment. A second set of fresh water pumps will allow up to 2,730 m3/d to be pumped to the mill (in the same line) in case the reclaim pump system fails, or when natural diversions from the catchment upstream of the tailings storage facility are unable to meet water demands. Permanent storage of water in the tailings pond void spaces will comprise a significant loss pathway for water pumped to the tailings storage facility. Evaporation from the open water surface in the tailings storage facility will also result in water losses from this facility during its operational life. Water for plant site fire suppression will be pumped from the process plant storage tank through a pressurized system to adjacent areas of the process plant and to the accommodations, service complexes, power and water treatment plants, and the utilidors. The volume of water and duration of flow in the pressurized system will be standardized to relevant fire codes. An adequate volume of water will be maintained in the tank to ensure availability of water in the event of a fire.



4.2 Water Management System and Activities 4.2.1 Water in the Open Pit and Underground Operations – Quantity The objectives for managing pit inflows during the operations phase include the following:

• Ensure minimal disruption of mining activities due to water reporting to the open pits; and,

• From late 2007 through the remainder of the project life, to maximize timely collection

and transport of combined pit inflows to the tailings storage facility for use as reclaim water in the process plant, thereby minimizing the need to obtain make-up water for the process plant from other surface or ground water sources.

During the operational phase of the project, there will be the following three main sources of water into the open pits: 1. Net runoff via direct precipitation; 2. Thawback inflows from the exposed pit walls; and, 3. Thawback inflows from the exposed walls of the underground operations As discussed above, thawback inflows to the open pit and underground workings have been estimated from thermal modeling and on-site investigations of ice content in representative wall rocks (AMEC, 2004). Net runoff via direct precipitation on the open pits under varying return periods has been estimated using baseline hydrologic data (AMEC, 2005a) and an assumed runoff coefficient of 0.90. For an average year, net unit runoff within the open pits is estimated at 280 mm. The total average annual volume of net runoff in the open pits can then be estimated by multiplying the plan area of the open pits by the 280 mm of net unit annual runoff. Details regarding the source of climate data and derivations in annual unit runoff values for various land use types are given in the Site Water Balance (AMEC, 2005a) and Tailings Facility Feasibility Design (AMEC, 2005b) reports. Under the base case, combined inflows to the open pit and underground workings will not be segregated prior to collection and removal. A system of ditches and sumps will be constructed and continuously maintained and upgraded throughout the operation phase of the project in order to ensure optimized collection of pit inflows such that excess water within the pits does not interfere with mining operations. It is understood that no pit inflows will occur during the winter months, and thus, heat-traced piping will not be installed to transfer pit inflows either directly to the lime treatment system adjacent to the open pit during 2007, or to the tailings storage facility from late 2007 through the remainder of the project. Sufficient storage will be provided in the open pit to store the 1-in-200 wet year freshet volume (defined above) plus anticipated thawback inflows during the year of occurrence. A pumping system capable of removing the 1-in-200 wet year freshet volume (including thawback inflows over this time) within 15 days of its occurrence is recommended as a risk management



measure. This type of system will ensure minimum risk to a disruption of mining activities from excessive water collection in the open pit and underground workings during the project’s operational phase. Perimeter berms will be constructed around the circumference of the open pit in order to reduce surface runoff inputs from outside the pit rim. The perimeter berms will not be impermeable and will allow a small amount of runoff from outside the pit rim to report to the open pit during runoff events where the local topography favours such patterns. The quantity of external runoff reporting to the open pit during these periods has not been estimated, but is not anticipated to have a significant influence on the water balance within the open pit and underground workings. The quantity and quality of water reporting the open pit and the underground workings will be monitored over the operational phase of the project. The quantity of inflows to the open pit and underground workings will be estimated using flow records from in-pit pumps. The water quality of these water sources will be estimated by regular sampling of the composite water collecting in sumps at the base of the open pit over the mine’s operational phase. Selected sampling of water quality in the underground workings may be undertaken. In the event that inflows to the open pit are of a significantly higher quantity than anticipated, contingency measures will include the expansion of sumps at the base of the open pit and/or installation of higher pumping capacity. The volume of water entering the open pits will change over time as the open pit grows in areal and volumetric extent between 2006 and 2010, and as more underground workings are brought on-line over the project life. Water entering the open pit will be routed by ditches to a series of sumps. Temporary sumps will be developed in working areas that will allow initial settlement of coarse suspended solids from the water. From the temporary sumps, water will be directed through a combination of ditches and pipelines to main sumps equipped with multiple storage areas and pumps to direct water to the pit inflow lime treatment and sedimentation pond system during 2007, or directly to the tailings storage facility (via an overland pumping and pipeline system) from late 2007 through the remainder of the project life. Pipelines from the pumps will be routed along access roads into the open pit. One pipeline from the open pit will route water collected in the main sump to either the pit inflow lime treatment and sedimentation pond system during 2007, or directly to the tailings storage facility (via an overland pumping and pipeline system) from late 2007 through the remainder of the project life. In the event of electrical or mechanical failures, pumps may be non-functional and flooding of sumps could result. To address this, a limited amount of storage capacity will be provided in the open pit to prevent flooding of sumps and working areas.



4.2.2 Water in the Open Pit and Underground Operations – Quality Composite inflows to the open pit and underground workings are expected to be acidic (pH 3-4), contain high dissolved and total metal loadings and nutrient concentrations (including elevated nitrate and ammonia concentrations from blasting residues), and have high TDS loadings. Thawback inflows from the exposed pit walls, thawback inflows in the underground workings that are subsequently pumped to surface sumps in the open pits, along with surface runoff from on-pit precipitation will mix to form a composite pit inflow water. This water will also contact residual explosives and fine blast rock in the open pit, leading to additional loadings of TDS, major anions/cations, ammonia/nitrate, and metals. Coarse material suspended in the pit inflows will settle in the sumps. During 2007 through 2010, when the open pit is still growing, coarse suspended material in pit inflows will be removed with the waste rock and disposed of either in the tailings dam (for NAG and PAG waste rock) or in the AG waste rock pile. From 2011 through the remainder of the project, coarse suspended material in the pit inflows will settle in the pit sumps. Removal of this coarse material will be periodically required to ensure adequate capacity in the pit inflow collection and storage system. The coarse material collected in the pit water management system will be either stockpiled in unused portions of the pit, used as backfill for mined-out portions of the underground operation, or disposed of in the tailings storage facility. Fine sediment remaining in the decanted pit inflows will be either sent to the lime treatment flow-through sedimentation pond for further treatment during 2007, or will be sent directly without treatment to the tailings storage facility from late 2007 through the remainder of the project life. During 2007, lime neutralization (with potential flocculation if needed), and extended hydraulic retention times in the lime treatment pond adjacent to the open pit, will ensure residual fine sediments from pit inflows are removed prior to direct discharge into ambient receiving waters. Further information on the expected quality of the pit inflows is available in the Surface Water Quality Modeling Report (AMEC, 2005c). Based on the geochemical testing results for the AG, PAG, and NAG rock to be exposed in the pit walls (Atkin 2005), as well as experience with acidic drainage at similar northern mine sites, treatment of the pit inflows is expected to require a calcium oxide (CaO; or pebble lime) dose of 2.5 kg/m3. The relative distribution of AG, PAG, and NAG in the final pit wall exposure is estimated as follows: AG (42%), PAG (11%), and NAG (47%) (Wardrop, 2005). The laboratory based humidity cell testing (Atkin, 2005) suggests there is insufficient alkalinity produced from the NAG waste rock to neutralize the acidity produced from the AG waste rock when approximately equal quantities of drainage from these material types are combined. Thus, approximately equal quantities of exposed AG and NAG waste rock are expected to yield an overall acidic runoff with a pH in the range of 3-4 and high dissolved metal and major ion concentrations. The water quality of pit inflows (including underground workings that are subsequently mixed with open pit inflows in surface sumps) will not likely be suitable for direct discharge to ambient receiving waters without prior treatment. Thus, during 2007, all pit inflows will be treated with lime neutralization and sedimentation prior to direct discharge into receiving waters. From late



2007 through the remainder of the project life, the pit inflows will not be treated, and will instead be sent directly to the tailings storage facility for retential in the process plant-tailings pond closed-circuit water balance. Geochemical testing of the waste rock (Atkin, 2004), as well as experience with acidic drainage at similar northern mine sites, suggests pebble lime treatment of the pit inflows will require a calcium oxide (CaO; or pebble lime) dose of 2.5 kg/m3. Lime treatment with flocculant addition should be able to reduce all metals of concern in the pit inflows to less than the World Bank direct discharge limits (World Bank, 1999a and 1999b). A sedimentation pond with a minimum hydraulic retention time of 3 to 4 days would be required for efficient metal removal. This minimum HRT requirement is based on similar treatment experience at other mine sites in northern environments. Pit inflows are predicted to contain an appreciable amount of dissolved iron. Subsequent neutralization of the inflows, and oxidation of the dissolved iron, will result in precipitation as ferric hydroxide. Ferric hydroxide precipitation will assist to co-precipitate arsenic to concentrations (likely <0.3 mg/L) below the World Bank direct discharge limits (World Bank, 1999a and 1999b). Note that, in general, iron would otherwise be added to the treatment system for arsenic removal if high iron concentrations were not already present in the pit inflows. pH adjustment with ferric co-precipitation will help remove other metals of concern, such as copper and zinc. Dissolved aluminum is also present at elevated concentrations in the pit inflows. Dissolved aluminum will likely precipitate as the corresponding hydroxide at neutral pH, with assistance from ferric hydroxide residues. Aluminum concentrations in the pit inflows (likely <1 mg/L)are expected to be suitable for direct discharge to ambient receiving waters based on the World Bank criteria (World Bank, 1999a and 1999b). High phosphorus concentrations are also possible in the pit inflows. As with arsenic and aluminum, the elevated iron concentrations in the feed water will likely precipitate phosphorus to low effluent concentrations. The lime treatment system will not likely be effective in treating for elevated ammonia and nitrate concentrations from blasting residues that are expected to be present in the pit inflows. Proper management practices in controlling the use and dispersion of blasting materials and residues will likely maintain low nitrate and ammonia levels in pit inflows that do not require further treatment. Oil and grease from used explosives, accidental spills, or vehicle leaks will also collect in the sumps. Oil and grease will be skimmed and/or sorbed from the sumps, brought to the surface, and either placed in the lined landfill during 2006 and 2007 or in the tailings facility from 2008 through 2013, or incinerated on-site. The method of disposal of hydrocarbon contaminated materials will be determined at a later date.



4.2.3 Surface Water – Plant and Camp Site Runoff The objective for managing net surface runoff from the plant site during the operations phase is to ensure minimal disruption of mining activities due to surface runoff. During the operational phase of the project, precipitation on the plant site will yield net runoff requiring collection, transfer, and treatment. Net runoff via direct precipitation on the plant site under varying return periods has been estimated using baseline hydrologic data (AMEC, 2005a) and an assumed runoff coefficient of 0.90. For an average year, net unit runoff from the plant site is estimated at 280 mm. The total average annual volume of net runoff from the plant site can thus be estimated by multiplying the plan area of the plant site (about 265,000 m2) by the 280 mm of net unit annual runoff. In an average year, approximately 74,000 m3 of runoff from the plant site is expected. A system of ditches and collection ponds will be constructed and continuously maintained and upgraded throughout the operation phase of the project in order to ensure optimized collection of plant and camp site runoff such that excess water does not interfere with mining operations (AMEC, 2005a). The conveyance capacity of ditches in the plant and camp site will be sized to accommodate the estimated flows from a 1-in-200 year 24 hour precipitation event. Sufficient storage capacity will be installed in sedimentation ponds around the plant and camp site to ensure storage of a 1-in-200 wet year freshet (AMEC, 2005a). This system will ensure minimum risk to a disruption of mining activities from excessive water collection and movement within the plant site area during the project’s operational phase. Further details on site drainage collection, transport, and disposal are given in the Site Water Balance report (AMEC, 2005a). The quantity and quality of surface runoff from the plant site will be monitored over the operation phase of the project. The quantity of surface runoff will be estimated using flow records from pumps in the collection ponds. The water quality of surface runoff from the plant site will be estimated by regular sampling of the composite water collecting in sedimentation ponds over the mine’s operational phase. While there are no defined streams in the core facilities area, there are natural drainage areas that will be re-contoured in order to construct site facilities. Rockfill ditches and grading will direct the runoff from the re-contoured areas towards collection ponds. The collection ponds will either be periodically pumped out, or allowed to drain naturally if possible, into the nearby ambient receiving water catchment of Pervyi Creek. As noted elsewhere in the current report, sedimentation ponds will collect sediment generated from runoff in outlying areas such as access and haul roads, the airstrip, and the explosives storage area. The sedimentation ponds will be located at points of runoff concentration where overflow water will be allowed to flow to adjacent watercourses. Solids in the sedimentation ponds will be periodically removed, if necessary, to ensure continuing adequate storage capacity. Runoff from these regions is not anticipated to yield contaminated sediments that would not be suitable for disposal on nearby natural surfaces or adjacent to roadways. Any major spills on roads and the airstrip will be contained, collected, and remediated as per local and national guidelines and standard mining best management practices.



4.2.4 Surface Water – Ore Stockpile During 2006 and 2007, the open pit will generate ore for future processing, but the process plant will not become operational until early in 2008. All ore produced during 2006 and 2007 will need to be stockpiled at the plant site for subsequent processing that will begin in 2008. Development of the open pit is not expected to begin until autumn of 2006, and thus, runoff from the ore stockpile is not anticipated during this year. Surface runoff from the ore stockpile (about 16,000 m2 in plan view, for an average annual surface runoff water yield of about 4,400 m3) is expected to be acidic with high concentrations of metals and major ions. In addition to surface runoff from direct precipitation on the ore stockpile, the geochemical nature of the ore (Atkin, 2005) will be such to promote the start of sulfide oxidation immediately after removal in the pit and exposure to atmospheric conditions. The rapid onset of exothermic sulfide oxidation in the ore will may be such to thaw the stockpile on a continuous basis over the project life. Based on a nominal process plant throughput of up to 3,000 tonnes per day, approximately 42,000 m3/y of thaw flows are expected from the ore stockpile. Therefore, the total volume of ore stockpile drainage requiring treatment and discharge may range up to 46,000 m3 in an average year. However, as noted above, due to the uncertainty regarding the timing and extent of any thaw flows from the ore stockpile, this water source has not been included in the site water balance. The ore stockpile runoff will be treated with a lime-addition system in a flow-through type sedimentation pond for pH adjustment, allowed to settle for removal of particulate phase contaminants (both suspended sediments and particulate/colloidal phase metals), and pumped to the runoff collection system for the plant site. Treated runoff from the ore stockpile will be mixed with runoff from the plant site (thus being subjected to an additional settling period in the plant site sedimentation ponds for maximum removal of particulate colloidal contaminants resulting from the lime treatment process), and allowed to report to the ambient headwater receiving environment of Pervyi Creek (a right tributary of the Kaiemraveem River upstream of the project site). The quality and quantity of this treated water will be monitored to ensure it complies with applicable direct discharge limits under the World Bank criteria (World Bank, 1999a and 1999b), and that sufficient dilutional capacity exists so ambient receiving water quality limits at the regulatory point-of-compliance (VNIRO, 1999) in the Kaiemraveem River will not be exceeded. Geochemical testing of the raw ore (Atkin, 2004), as well as experience with acidic drainage at similar northern mine sites, suggests pebble lime treatment of the ore stockpile runoff will require a calcium oxide (CaO; or pebble lime) dose of 2.5 kg/m3. As noted above with regard to treatment of pit inflows, lime treatment with flocculant addition should be able to reduce all metals of concern in the ore stockpile runoff to less than the applicable direct discharge limits.



4.2.5 Surface Water – Runoff from Roads and the Airstrip An airstrip has an exposed plan area of 600,000 m2 and is located on a knoll in the Starichnaya River valley about 10 km north of the plant site. The location of the airstrip on the top of a knoll will allow a majority of the runoff from natural catchments surrounding the airstrip to drain away from the airstrip. The airstrip will be graded to allow runoff to flow into the tributary of the Starichnaya River north of the airstrip. The total runoff from direct precipitation on the airstrip in an average year is 167,000 m³. Sediment fences and/or sediment ponds will be used if suspended solids concentrations in runoff from the airstrip are greater than 50 mg/L. Ditches will be installed surrounding haul and access roads. Where necessary, the ditching system will convey stormwater runoff through culverts designed to convey the 1-in-200 year 24 hour precipitation event. Ditches will be continued until intersection with a stream or other water body. If deemed appropriate, ditches could be terminated and the flow allowed to drain naturally. Exit points of diversion ditches will be designed with erosion control measures where required. Details on the design specifications, locations, and flow capacities for culverts installed on haul and access roads are provided in the Site Water Balance report (AMEC, 2005a). Runoff collected from haul and access roads will be allowed to report to ambient receiving waters provided it meets direct discharge criteria. No significant chemical water quality issues are expected from road material because NAG rock will be used for construction. Where suspended solids concentrations in roadside runoff are found to exceed 50 mg/L at the point of discharge into ambient receiving waters, sedimentation ponds will be installed to reduce suspended solids levels below the direct discharge limit. 4.2.6 Surface Water – AG Waste Rock Pile The AG waste rock pile is the surface containment region for 8.5 million tonnes of waste rock (AMEC, 2005d). The AG waste rock pile will be located within the final tailings impoundment . During late 2006 and all of 2007, the open pit will generate AG waste rock, but the starter tailings dam will not be completed until late 2007. Based on available geochemical testing (Atkin, 2005), surface runoff from the AG waste rock is expected to be acidic with elevated concentrations of metals and major ions such as sulfate, iron, and calcium. During 2006 and 2007, AG waste rock from the open pit will be deposited in the valley of the tailings storage facility (for either option) at a point upstream of the starter tailings dam construction. For tailings alternative #1, the AG waste rock pile (and tailings dam) will be located in the valley of Tretyi Creek. For tailings alternative #2, the AG waste rock pile (and tailings dam) will be located in the catchment of Vtoroi Creek. Diversion ditching around the AG waste rock pile will limit net runoff to the exposed plan area of this location. Development of the open pit will not occur until autumn 2006, and thus, runoff from the AG waste rock pile during this year is assumed to be negligible. During 2006 and 2007, about 877,000 tonnes and 1,585,000 tonnes, respectively, of AG waste rock will be stored upstream of the proposed starter dam locations. The total volumes of AG waste rock will cover plan areas of 290,000 m2 or 235,000 m2 at the end of 2007 under tailings alternatives #1 and



#2, respectively. Because of the late start to pit development in 2006, no surface runoff from the AG waste rock pile is assumed to occur during this year. Thus, the AG waste rock will generate about 46,300 m3 or 37,500 m3 of surface runoff in 2007 during an average year under tailings alternatives #1 and #2, respectively. In addition to surface runoff arising from direct precipitation on the AG waste rock pile during 2007, thawing of entrained ice in the waste rock may contribute up to 10% by volume of the total rock mass stored at this facility. The geochemical testing indicates the AG waste rock will likely generate net acidity immediately after placement, and the exothermic nature of the sulfide oxidation process may be such to prevent the presence of unfrozen rock in the pile. The difficulty in maintaining frozen rock within net acid generating waste rock facilities in northern permafrost environments (even during the over-winter period following winter placement and burial of frozen rock) has been demonstrated elsewhere (SRK, 2004b; Rescan and SRK, 2004). Assuming all entrained ice in the AG waste rock pile melts and is released as drainage during 2007, up to 93,000 m3 of water (assuming a specific gravity of 2.65 for the rock mass) may be produced that will require collection and treatment prior to discharge. While not included in the site water balance model due to the current uncertainty of its significance, it must be noted that the potential for thaw flows from the AG waste rock stockpile in 2007 could increase the estimated average required water treatment capacity for this site region by up to a factor of 3-4. Additional study with thermal modeling is recommended prior to the start of mine construction in order to examine the potential extent of thawing in the AG waste rock stockpile during 2007. This work would ensure that any additional treatment chemicals for the stockpile runoff are in place for 2007, and that larger runoff/thaw flow containment and treatment structures are built prior to the summer of 2007. Runoff from the AG waste rock will be isolated to minimize total volumes requiring storage and treatment. Diversion ditches will be installed in early 2006 around the proposed AG waste rock storage sites to minimize runoff. Runoff from the AG waste rock pile in 2007 will be collected and treated using a lime treatment system with a flow-through sedimentation pond (HRT≥3-4 days) as described below. Following completion of the starter tailings dam in late 2007, runoff from the AG waste rock pile will be allowed to report to the tailing pond. All AG waste rock generated after 2008 will be added to the 2006/2007 pile within the tailings storage facility. The AG waste rock placed in the tailings storage facility will be completely submerged under both tailings and a free water column cover over the operating life of the mine. The AG waste rock pile runoff will be treated with a lime-addition system in a flow-through sedimentation pond for pH adjustment, allowed to settle for removal of particulate phase contaminants (both suspended sediments and particulate/colloidal phase metals), and released to ambient receiving waters in Tretyi Creek (for tailings alternative #1) or Vtoroi Creek (for tailings alternative #2). The quality and quantity of this treated water will be monitored to ensure it complies with applicable direct discharge limits under the World Bank criteria (World Bank, 1999a and 1999b), and that sufficient dilutional capacity exists so ambient receiving water quality limits at the regulatory point-of-compliance (VNIRO, 1999) in the Kaiemraveem River will not be exceeded.



Geochemical testing of the AG waste rock (Atkin, 2005), as well as experience with acidic drainage at similar northern mine sites, suggests pebble lime treatment of the AG waste rock runoff will require a calcium oxide (CaO; or pebble lime) dose of 2.5 kg/m3. As noted above with regard to treatment of pit inflows and the ore stockpile runoff, lime treatment with flocculant addition should be able to reduce all metals of concern in the ore stockpile runoff to less than the applicable direct discharge limits.



4.3 Water Treatment and Discharge 4.3.1 Sewage Treatment Plant A sewage treatment plant (STP) is currently installed at the project site. The system configuration is as follows: settling chamber → pre-aeration chamber → treatment chamber → pump chamber → filter unit → UV disinfection. The tertiary level nature of the system will provide a high level of physical, chemical, and biological treatment efficiency. By 2006, upgrades to the STP will result in a capacity of 150 m3/d (based on a 600 man camp at 265 L/p/d) at an organic loading rate of 800 mg/L BOD5 (128 kg BOD5/d maximum loading). During two summer month periods during 2006 and 2007, an additional 300 men may increase the camp’s total population to 900 men for a short period of time (e.g., June through October). In such a case, the STP capacity would be increased to 230 m3/d. The peak hourly flow capacity of the STP will be 2.5 times the daily flow rate, or from about 400 m3/d (600 man camp) to 600 m3/d (900 man camp). The combined BOD5 loading capacity of the STP is a composite of 250 mg/L BOD5 from domestic residential grade wastewater (40% nominal average of total daily flows) and 1,200 mg/L BOD5 of high-strength food preparation wastewater (60% nominal average of total daily flows). The average influent design capacities are 250 mg/L for suspended solids, 50 mg/L for fats, oils, and grease, and 50 mg/L for total Kjeldahl nitrogen. Effluent characteristics from the STP will be as follows: BOD5≤30 mg/L and TSS≤30 mg/L. No specified effluent criteria for ammonia, nitrate, total nitrogen, fecal coliforms, phosphorus, and dissolved oxygen are available. It is assumed the STP effluent will meet the World Bank direct discharge water quality standards to allow summer discharge direct to receiving water bodies during 2006 and 2007 without dilution, including the following parameters of note with regard to sewage treatment: pH between 6 to 9; BOD5≤50 mg/L; TSS≤50 mg/L; oil and grease≤20 mg/L; and temperature<5°C above ambient temperature of receiving waters (<3°C increase where receiving water temperature is >28°C). Elevated metal concentrations are not anticipated in the domestic wastewater and high-strength food preparation wastewater sources to the STP. From late 2007 through the remainder of the project life, the STP effluent will be pumped to the tailings pond and permanently retained in the closed-circuit process plant water balance. Thus, direct discharge criteria will not apply to the STP effluent over this period of the project, thereby allowing for potential cost savings by minimizing or eliminating the use of secondary and tertiary processes. STP discharge will be directed into the catchment of the Pervyi Creek near its headwaters for both the summer and winter periods of 2006 until late 2007. Over-winter discharge will remain frozen near the point of discharge. The STP discharge piping will not be heat-traced. At a peak over-winter camp size of 600 men, and assuming a 212 day ice-covered season (where nearby creeks do not flow) from November through May, approximately 32,000 m3 of frozen treated sewage will accumulate near the STP discharge point. This frozen effluent will thaw over the spring/early summer freshet and achieve maximum dilution during this time. Summer STP



discharge when rivers are flowing will be at the same location as with the over-winter discharge regime. From late 2007 through the remainder of the project life, the STP effluent will be pumped directly to the tailings pond and permanently retained in the closed-circuit process plant water balance. 4.3.2 Water Treatment Plant No permanent water treatment plant will be installed at the project site. Runoff from site facilities will be monitored to ensure it meets World Bank direct discharge (World Bank, 1999a and 1999b) and ambient receiving water quality limits (VNIRO, 1999). Where surface runoff contains elevated suspended solids concentrations above the 50 mg/L direct discharge limit, sedimentation ponds will be installed to reduce suspended solids levels below 50 mg/L, after which time the water will be allowed to report to ambient receiving waters. For the ore stockpile runoff, the AG waste rock pile runoff (in 2007 only), and the pit inflows (in 2007 only), lime treatment and subsequent sedimentation (with flocculant addition if required) will be used to neutralize the acidic runoff expected from these site regions. Lime-based neutralization of the ore stockpile runoff, AG waste rock pile runoff, and pit inflows is also expected to lower metal concentrations in these feed water to levels below the World Bank direct discharge limits (World Bank, 1999a and 1999b). All acidic drainage on site will be managed according to best management practices to minimize, collect, and treat such waters so as to comply with regulatory guidelines. Water flows directed to the tailings storage facility during the operational phase will be retained in the closed tailings pond and process plant circuit. Process plant effluent will be subjected to a cyanide recovery and subsequent hypochlorite destruction process prior to release into the tailings pond. Treatment of process plant effluent for cyanide removal is expected to lower total cyanide levels in the tailings pond to <5 with a pH of 7-8. Water in the tailings pond will either be lost to voids or evaporation, or will be used as reclaim water for the process plant. Water remaining in the tailings pond (about 2-3 Mm3) at mine closure will be discharged to ambient receiving waters (e.g., Kaiemraveem River) provided it meets both World Bank direct discharge (World Bank, 1999a and 1999b) and ambient fresh water receiving criteria (VNIRO, 1999) at the respective points-of-compliance. Treatment of the tailings water at closure would be undertaken if the water quality was not suitable for discharge into the Kaiemraveem River. 4.3.3 Tailings Storage Facility Pond The tailings impoundment is designed as a closed circuit process with no discharge, so the impoundment must have enough capacity to store average annual net water inflows, a 1-in-200 wet year return period, and a Probable Maximum Flood (PMF) event with adequate freeboard during the mine life. On an annual basis, the level of risk exposure for the tailings storage facility will be 0.5%. To allow for water to be added to the tailings pond during spring runoff, the west diversion ditch of tailings alternative #1 or the east diversion ditch of tailings alternative #2 will be constructed



with a side-discharge release structure immediately upstream of the tailing dam crest’s abutment (AMEC, 2005b). Releases through this facility will be conducted during the spring/summer months (May to September) on an annual or as needed basis. Water levels will be continuously monitored to determine the volume of water released into the tailings impoundment. Water management guidelines for the tailings storage facility will be developed during the mine’s operation. The planned releases into the pond will provide buffer storage for winter month water withdrawals while maintaining the minimum 4 m depth over the tailings level. Additional make-up water will be pumped from process water wells as needed. Seepage rates are assumed to be zero, as the dam embankment is lined with a bituminous geomembrane and the foundation soils are shown to freeze and to maintain its frozen state during summer months, based on thermal modeling. A monthly water/mass balance model was developed to simulate the hydrologic behaviour of the tailings impoundment. Annual summary water balance spreadsheets for tailings alternatives #1 and #2 are presented as part of the site water balance summary in Appendix B. The base case scenario of the tailings storage facility water balance model employs average annual precipitation, runoff and evaporation rates for the duration of the mine life. The starter tailings dam is assumed to be completed by October 2007. Milling start-up is scheduled for June 2008. Inflow sources to the tailings storage facility include water in the process plant effluent, direct precipitation on the tailings pond, natural runoff reporting to the tailings pond, pit inflows, effluent from the STP, and direct precipitation on the tailings dam. Note that if direct precipitation on the downstream face of the tailings dam yields net runoff with water quality suitable for direct discharge to ambient receiving waters according to the World Bank criteria (World Bank, 1999a and 1999b), this drainage will not be returned via pumping into the tailings pond. Outflows consist of evaporation, losses to void spaces in the settled tailings, and reclaim water for the process plant. Further details on the tailings storage facility water balance, water diversion system, and design criteria are provided in the Tailings Facility Feasibility Design Report (AMEC, 2005). 4.3.4 Discharges to Ambient Receiving Waters Site discharges outside the tailings pond-process plant closed circuit water balance will be allowed to report to the Kaiemraveem and Starichnaya Rivers via natural watercourses. The water quality of all discharges to receiving waters will be monitored to ensure compliance with direct discharge and ambient fresh water regulatory guidelines. Applicable regulatory limits for water quality at the project site are given in Table 4.1. The regulatory point-of-compliance for direct discharges is at the end-of-pipe. The regulatory point-of-compliance for ambience fresh water receiving guidelines in the Kaiemraveem River is assumed to be 500 m downstream of the last site discharge into the Kaiemraveem River, which is anticipated to be at a point 500 m downstream of the confluence between Vtoroi Creek and the Kaiemraveem River. The regulatory point-of-compliance for ambience fresh water receiving guidelines in the Starichnaya River is assumed to be 500 m downstream of the last site discharge into the Starichnaya River, which is anticipated to be at a point 500 m downstream of the airstrip.



Table 4.1: Summary of applicable regulatory limits for water quality.

World Bank Direct Discharge

Limitsa Russian Class 1

Fisheries Standardsb Bulk Parameters

pH 6-9 6.5-8.5 Turbidity n/ac 0.25 NTU Total Suspended Solids 50 mg/L n/a

Major Ions Calcium n/a 180 mg/L Chloride n/a 300 mg/L Fluoride n/a 0.05 mg/L Magnesium n/a 40 mg/L Potassium n/a 50 mg/L Sodium n/a 120 mg/L Sulfate n/a 100 mg/L

Nutrients Ammonia n/a 0.5 mg/L Nitrate n/a 40 mg/L Nitrite n/a 0.08 mg/L Phosphate n/a 0.2 mg/L

Cyanide Thiocyanate n/a 0.19 mg/L Free Cyanide 0.1 mg/L n/a (0.022 mg/Ld) WAD Cyanide 0.5 mg/L n/a (0.022 mg/Ld) Total Cyanide 1 mg/L 0.05 mg/L (0.022 mg/Ld)

Metals Aluminum n/a 0.04 mg/L Antimonye n/a 0.005 mg/Le Arsenic 1 mg/L 0.05 mg/L Cadmium 0.1 mg/L 0.005 mg/L Chromium 1 mg/L (0.05 mg/Lf) 0.07 mg/L Cobalt n/a 0.01 mg/L Copper 0.3 mg/L 0.001 mg/L Iron 2 mg/L 0.1 mg/L Lead 0.6 mg/L 0.006 mg/L Manganese n/a 0.01 mg/L Mercury 0.002 mg/L 0.00001 mg/L Molybdenum n/a 0.001 mg/L Nickel 0.5 mg/L 0.01 mg/L Selenium n/a 0.002 mg/L Silver n/a 0.01 mg/L Strontium n/a 0.4 mg/L Thallium n/a 0.0001 mg/L Zinc 1 mg/L 0.01 mg/L



a World Bank Environment, Health, and Safety Guidelines for direct discharges from open-pit milling and mining operations (World Bank, 1999a) and underground mines (World Bank, 1999b). Guidelines apply to effluent discharged to receiving waters from tailings impoundments, mine drainage, sedimentation basins, sewage systems, and storm water drainage. Where background concentrations exceed these levels, the discharge may contain concentrations up to natural background levels. Concentrations up to 110% of natural background can be accepted if no significant adverse impact can be demonstrated. b Russian Class 1 fisheries limits are shown for all parameters except where noted (VNIRO, 1999). c “n/a” indicates no applicable guideline for a particular parameter under the given regulatory class. d No Russian Class 1 fisheries limits apply for free and weak acid dissociable (WAD) cyanide (VNIRO, 1999). World Bank Environment, Health, and Safety Guidelines for ambient receiving waters at the edge of the mixing zone (World Bank, 1999a, 1999b) are shown in parentheses. e There is no Russian Class 1 fisheries limit for antimony (VNIRO, 1999). The drinking water/recreational use limit is shown (VNIRO, 1999). f Value in parentheses is the World Bank Environment, Health, and Safety Guidelines limit for hexavalent chromium in direct discharges. The 1 mg/L limit is for total chromium (World Bank, 1999a, 1999b).



5.0 CONCLUSIONS Recommendations presented herein are based on an evaluation of the findings of the site investigation noted. If conditions other than those reported are noted during subsequent phases of the project, AMEC should be notified and be given the opportunity to review and revise the current recommendations, if necessary. Recommendations presented herein may not be valid if an adequate level of review or inspection is not provided during construction. This report has been prepared for the exclusive use of BEMA Gold, Inc. for specific application to the area within this report. Any use which a third party makes of this report, or any reliance on or decisions made based on it, are the responsibility of such third parties. AMEC accepts no responsibility for damages, if any, suffered by any third party as a result of decisions made or actions based on this report. It has been prepared in accordance with generally accepted engineering practices. No other warranty, expressed or implied, is made. Respectfully submitted, AMEC Earth & Environmental, a division of AMEC Americas Limited

Per Sierra Rayne, Ph.D., P.Ag., P.Chem., A.Sc.T. Environmental Scientist Per Peter Lighthall, M.Eng., P.Eng. Vice-President, Mining Project Manager

Reviewed by: Per Jim Bester, P.Eng. Senior Water Resources Engineer



REFERENCES

AMEC. 2004. Kupol Gold Mining Project – Proposed Open Pit Preliminary Assessment of Net Water Inflows. Prepared for Bema Gold Corporation: Vancouver, BC, Canada. AMEC. 2005a. Site Water Balance: Kupol Gold Project. Prepared for Bema Gold Corporation: Vancouver, BC, Canada. AMEC. 2005b. Tailings Facility Feasibility Design Report: Kupol Gold Project. Prepared for Bema Gold Corporation: Vancouver, BC, Canada. AMEC. 2005c. Surface Water Quality Modeling Report: Kupol Gold Project. Prepared for Bema Gold Corporation: Vancouver, BC, Canada. AMEC. 2005d. Waste Rock Management Report: Kupol Gold Project. Prepared for Bema Gold Corporation: Vancouver, BC, Canada. Atkin, S. 2005. Geochemical Characterization of Waste Rock and Ore Materials at the Kupol Project, Russian Far East. Prepared for Bema Gold Corporation: Vancouver, BC, Canada. Bema. 2004. Preliminary Economic Assessment: Kupol Gold Project, Far East Russia. Prepared by Bema Gold Corporation: Vancouver, BC, Canada. Borisov, V.N., Alexeev, S.V., and Pleshevenkova, V.A. 1995. The diamond mining quarries of East Siberia as a factor affecting surficial water quality. In: Water-Rock Interaction. Eds.: Kharaka, K.K., Chudaev, O.V., Thordsen, J.J., Armannsson, H., Breit, G.N., Evans, W.C., Keith, T.E.C., and Khakara, Y.K. Balkema: Rotterdam, The Netherlands, pp. 863-866. Clark, I.D., Douglas, M., Raven, K., and Bottomley, D. 2000. Recharge and preservation of Laurentide Glacial melt water in the Canadian Shield. Ground Water, 38(5), pp. 734-742. Dalstroyizyskania. 2004. Kupol Mine: Detail Design Geotechnical Survey Report, 19004-II-2. Prepared for Bema Gold Corporation: Vancouver, BC, Canada. Douglas, M., Clark, I.D., Raven, K., and Bottomley, D. 2000. Groundwater mixing dynamics at a Canadian Shield mine. Journal of Hydrology, 235, pp. 88-103. Gascoyne, M. and Kamineni, D.C. 1994. The hydrogeochemistry of fractured plutonic rocks in the Canadian Shield. Applied Hydrogeology, 2, pp. 43-49. Orocon. 2005. Mill Site Water Balance Spreadsheet (May 20, 2005). Prepared for Bema Gold Corporation: Vancouver, BC, Canada. Rescan and SRK. 2004. 2004 Waste Rock Storage Area Seepage Survey Report. Prepared for BHP Billiton Diamonds, Inc.: Yellowknife, NWT, Canada.



Sader, J.A., Leybourne, M.I., McClenaghan, M.B., Hamilton, S.B., and Robertson, K. 2003. Field procedures and results of groundwater sampling in kimberlite from drillholes in the Kirkland Lake and Lake Timiskaming areas, northeastern Ontario. Geological Survey of Canada: Current Research, 2003-C11, pp. 1-9. SRK. 2004a. General Geotechnical Core Logging Manual – Kupol Site, 2004 Drilling Season. Prepared for Bema Gold Corporation: Vancouver, BC, Canada. SRK. 2004b. 2003 Waste Rock Storage Area Seepage and Waste Rock Survey Report. Prepared for BHP Billiton Diamonds, Inc.: Yellowknife, NWT, Canada. VNIRO. 1999. Maximum Allowable Concentrations for Rivers and Waterways, Fisheries Standards. VNIRO: Moscow, Russia, 1999. Wardrop. 2005. E-mail from Gordon Zurowski (Wardrop) to Sierra Rayne (AMEC) dated May 4, 2005. Subject heading “Final ARD Surface Percentages.xls”. Prepared for Bema Gold Corporation: Vancouver, BC, Canada. World Bank. 1998. Pollution Prevention and Abatement Handbook: Toward Cleaner Production. The World Bank Group: Washington, DC, USA. World Bank. 1999a. World Bank Environment, Health, and Safety Guidelines: Open Pit Mines. The World Bank Group: Washington, DC, USA. World Bank. 1999b. World Bank Environment, Health, and Safety Guidelines: Underground Mines. The World Bank Group: Washington, DC, USA.



APPENDIX A : CONCEPTUAL WATER MANAGEMENT SYSTEM



Figure A.1: Conceptual water management system for 2006.

STP Effluent Pervyi Creek

Plant Site Runoff Sedimentation Pond

Camp Site Runoff Sedimentation Pond

Potable Water Inputs FromKaiemraveem River Kaiemraveem River



Figure A.2: Conceptual water management system for 2007.

STP Effluent Pervyi Creek

Plant Site Runoff

Ore Stockpile Runoff

Unnamed Right Tributary ofKaiemraveem River

(Between Pervyi and Tretyi Creeks)

Lime Treatment System /Sedimentation Pond

Sedimentation Pond


Pit Inflows Lime Treatment System /Sedimentation Pond

Vtoroi Creek (Option 1) orTretii Creek (Option 2)

AG Waste RockStockpile Runoff

Lime Treatment System /Sedimentation Pond

Potable Water Inputs FromKaiemraveem River

Kaiemraveem River



Figure A.3: Conceptual water management system for 2008 through 2014

STP Effluent

Pervyi Creek

Plant Site Runoff

Ore Stockpile Runoff Lime Treatment System /Sedimentation Pond

Sedimentation Pond


Pit Inflows

Tailings StorageFacility

AG Waste RockStockpile Runoff

Potable Water Inputs FromKaiemraveem River

Kaiemraveem River

Runoff from Upstream Faceof Tailings Dam

ProcessPlant

Fresh Water Inputs FromKaiemraveem River

Natural Surface Diversions FromUpstream Catchments

Direct Precipitation onTailings Pond

Runoff/Seepage fromDownstream Face of Tailings Dam

Vtoroi Creek (Option 1) orTretii Creek (Option 2)

Dependent onWater Quality



APPENDIX B : SITE WATER BALANCE ANNUAL SUMMARY



Table B.1: Site water balance summary for 2007.

Annual Volume (m3)


Wet 1-in-100

Wet Waste Rock Runoff

Initial AG Waste Rock Pile (Alt. 1) 46,300 87,400 121,000Initial AG Waste Rock Pile (Alt. 2) 37,500 71,100 98,600

Plant Site Runoff 74,500 109,000 137,000Ore Stockpile Runoff 9,130 13,300 16,800Pit Inflows 128,000 155,300 177,800

Direct Precipitation 59,200 86,500 109,000Groundwater Inflows 68,800 68,800 68,800

Total Treatment Volume (Alt 1) 258,000 365,000 453,000Total Treatment Volume (Alt 2) 249,000 349,000 430,000STP Effluent 57,100 57,100 57,100Fresh Water Well 355,700 355,700 355,700Potable Water Well 66,500 66,500 66,500Total Water Discharged to Kaiemraveem (Alt. 1) 315,000 422,000 510,000Total Water Discharged to Kaiemraveem (Alt. 2) 306,000 406,000 487,000






Wet 1-in-100

Wet Average 1-in-10

Wet 1-in-100


Water in Process Plant Effluent 537,000 642,000 642,000 537,000 537,000 537,000Direct Precipitation on TSF Pond 20,000 49,000 72,000 19,000 33,000 45,000Natural Runoff Reporting to TSF 875,000 1,534,000 2,072,000 590,000 1,034,000 1,397,000Pit Inflows Diverted to TSF 90,400 117,700 140,200 90,400 117,700 140,200

Direct Precipitation 59,200 86,500 109,000 59,200 86,500 109,000Groundwater Inflows 31,200 31,200 31,200 31,200 31,200 31,200

Ore Stockpile Runoff 9,130 13,300 16,800 9,130 13,300 16,800Sewage Treatment Plant 66,000 66,000 66,000 66,000 66,000 66,000Make-up Water 0 0 0 0 0 0Direct Precipitation on Tailings Dam 13,800 28,900 41,300 12,700 26,700 38,100

TSF Outflows 633,000 633,000 633,000 633,000 633,000 633,000

Evaporation from TSF Pond 17,000 17,000 17,000 17,000 17,000 17,000Storage in Tailings Pore Spaces 260,000 260,000 260,000 260,000 260,000 260,000Reclaim Water to the Process Plant 356,000 356,000 356,000 356,000 356,000 356,000

Net Change in TSF Storage 978,330 1,817,900 2,417,300 691,230 1,194,700 1,607,100 Fresh Water Well 356,000 356,000 356,000 356,000 356,000 356,000Potable Water Well 66,500 66,500 66,500 66,500 66,500 66,500






Wet 1-in-100

Wet Average 1-in-10

Wet 1-in-100


Water in Process Plant Effluent 1,095,000 642,000 642,000 1,095,000 1,095,000 1,095,000Direct Precipitation on TSF Pond 65,000 158,000 234,000 61,000 107,000 144,000Natural Runoff Reporting to TSF 586,000 1,027,000 1,388,000 348,000 610,000 824,000Pit Inflows Diverted to TSF 89,300 116,600 139,100 89,300 116,600 139,100



TSF Outflows 1,307,000 1,306,000 1,306,000 1,304,000 1,304,000 1,304,000


Net Change in TSF Storage 617,230 745,800 1,221,200 377,130 730,600 1,019,000 Fresh Water Well 356,000 356,000 356,000 356,000 356,000 356,000Potable Water Well 66,500 66,500 66,500 66,500 66,500 66,500






Wet 1-in-100

Wet Average 1-in-10

Wet 1-in-100


Water in Process Plant Effluent 1,095,000 1,095,000 1,095,000 1,095,000 1,095,000 1,095,000Direct Precipitation on TSF Pond 104,000 253,000 374,000 81,000 142,000 192,000Natural Runoff Reporting to TSF 417,000 731,000 988,000 328,000 575,000 777,000Pit Inflows Diverted to TSF 86,300 113,600 136,100 86,300 113,600 136,100



TSF Outflows 1,335,000 1,338,000 1,338,000 1,318,000 1,319,000 1,319,000


Net Change in TSF Storage 442,430 933,900 1,337,900 347,430 685,900 963,900 Fresh Water Well 356,000 356,000 356,000 356,000 356,000 356,000Potable Water Well 66,500 66,500 66,500 66,500 66,500 66,500






Wet 1-in-100

Wet Average 1-in-10

Wet 1-in-100





TSF Outflows 1,357,000 1,358,000 1,358,000 1,328,000 1,329,000 1,329,000





Table B.6: Site water balance summary for 2012



Wet 1-in-100

Wet Average 1-in-10

Wet 1-in-100





TSF Outflows 1,377,000 1,358,000 1,358,000 1,338,000 1,329,000 1,329,000








Wet 1-in-100

Wet Average 1-in-10

Wet 1-in-100





TSF Outflows 1,394,000 1,358,000 1,358,000 1,347,000 1,329,000 1,329,000


Net Change in TSF Storage 398,230 1,011,700 1,484,700 323,230 701,700 997,700 Fresh Water Well 356,000 356,000 356,000 356,000 356,000 356,000Potable Water Well 66,500 66,500 66,500 66,500 66,500 66,500






Wet 1-in-100

Wet Average 1-in-10

Wet 1-in-100



Direct Precipitation 59,200 86,500 109,000 59,200 86,500 109,000Groundwater Inflows 100 100 100 100 100 100


TSF Outflows 1,385,000 1,385,000 1,385,000 1,331,000 1,332,000 1,332,000


Net Change in TSF Storage 146,430 559,900 896,900 239,430 541,900 791,900 Fresh Water Well 356,000 356,000 356,000 356,000 356,000 356,000Potable Water Well 66,500 66,500 66,500 66,500 66,500 66,500

Appendix G AMEC Surface Water Quality Modeling Report



Surface Water Quality Modeling Report


Submitted to:


Submitted by:


Burnaby, BC

21 June 2005

AMEC File: VM00330

Bema Gold Corporation. Surface Water Quality Modeling Report Kupol Gold Project, Far East Russia 21 June 2005

AMEC File: VM00330 Page i

TABLE OF CONTENTS

Page

EXECUTIVE SUMMARY .............................................................................................................IV

1.0 INTRODUCTION.................................................................................................................. 1

2.0 METHODS ........................................................................................................................... 4

3.0 RESULTS AND DISCUSSION........................................................................................... 303.1 Direct Discharge of Runoff to Receiving Waters ....................................................... 303.2 Ambient Receiving Water Quality of Site Discharges ............................................... 32

3.2.1 Tailings Alternative #1 ................................................................................... 323.2.1.1 General ........................................................................................... 323.2.1.2 Calendar Year 2006........................................................................ 333.2.1.3 Calendar Year 2007........................................................................ 343.2.1.4 Calendar Years 2008-2014............................................................. 37

3.2.2 Tailings Alternative #2 ................................................................................... 393.2.2.1 General ........................................................................................... 393.2.2.2 Calendar Year 2006........................................................................ 393.2.2.3 Calendar Year 2007........................................................................ 413.2.2.4 Calendar Years 2008-2014............................................................. 44

3.2.3 Summary Comparison With Receiving Water Criteria................................... 46

4.0 POTENTIAL CLOSURE/POST-CLOSURE WATER QUALITY ISSUES........................... 47

5.0 CONCLUSIONS................................................................................................................. 48

6.0 CLOSURE.......................................................................................................................... 50

REFERENCES ........................................................................................................................... 51


AMEC File: VM00330 Page ii

LIST OF FIGURES Figure 1.1: Regional site layout and watershed boundaries..............................................2 Figure 1.2: Plant site layout. ..............................................................................................3

LIST OF TABLES Table 1.1: Project reports pertaining to water management.............................................1 Table 2.1: Summary of estimated background water quality in the Kaiemraveem River

and its contributing catchments near the project site. Applicable ambient receiving water quality limits are shown for comparison.................................................................................5

Table 2.2: Late-stage equilibrium pH values and elemental production rates from the individual NAG rock humidity cells. All values are in units of mg/kg/wk, except for pH, which is presented in traditional logarithmic units..................................................................9

Table 2.3: Summary of estimated runoff water quality from exposed unsaturated and saturated NAG rock. Applicable direct discharge and Russian Class 1 fisheries standards and the corresponding exceedence factors are shown for comparison. .............................12

Table 2.4: Late-stage equilibrium pH values and elemental production rates from the individual AG rock humidity cells. All values are in units of mg/kg/wk, except for pH, which is presented in traditional logarithmic units..........................................................................14

Table 2.5: Summary of estimated runoff water quality from exposed unsaturated and saturated AG rock. Applicable direct discharge and Russian Class 1 fisheries standards and the corresponding exceedence factors are shown for comparison. .............................16

Table 2.6: Estimated solid and dissolved phase concentrations, with major speciation contributions, of water quality parameters from exposed unsaturated NAG rock. Applicable regulatory limits and the corresponding exceedence factors for dissolved phase concentrations are shown for comparison. ..........................................................................19

Table 2.7: Estimated solid and dissolved phase concentrations, with major speciation contributions, of water quality parameters from exposed saturated NAG rock. Applicable regulatory limits and the corresponding exceedence factors for dissolved phase concentrations are shown for comparison. ..........................................................................20

Table 2.8: Estimated solid and dissolved phase concentrations, with major speciation contributions, of water quality parameters from exposed unsaturated AG rock. Applicable regulatory limits and the corresponding exceedence factors for dissolved phase concentrations are shown for comparison. ..........................................................................21

Table 2.9: Estimated solid and dissolved phase concentrations, with major speciation contributions, of water quality parameters from exposed saturated AG rock. Applicable regulatory limits and the corresponding exceedence factors for dissolved phase concentrations are shown for comparison. ..........................................................................22

Table 2.10: Estimated solid and dissolved phase concentrations, with major speciation contributions, of water quality parameters from exposed unsaturated AG rock following pebble lime (CaO) neutralization to pH 7. Applicable regulatory limits and the corresponding exceedence factors for dissolved phase concentrations are shown for comparison….......................................................................................................................24

Table 3.1: Predicted average annual concentrations in the Kaiemraveem River during


AMEC File: VM00330 Page iii

2006 at the ambient point-of-compliance for tailings alternative #1.....................................33 Table 3.2: Predicted average annual concentrations in the Kaiemraveem River during


2008 through 2014 at the ambient point-of-compliance for tailings alternative #1. .............37 Table 3.4: Predicted average annual concentrations in the Kaiemraveem River during



2008 through 2014 at the ambient point-of-compliance for tailings alternative #2. .............44 Table 3.7: Potential Russian Class 1 fisheries standard exceedences (with exceedence

factors in parentheses) in the Kaiemraveem River over the project life...............................46


AMEC File: VM00330 Page iv

EXECUTIVE SUMMARY

The Kupol Gold Project will include activities such as open pit mining, underground mining, ore stockpiling and processing, tailings and waste management, and camp operation that will generate volumes and qualities of water requiring management. The feasibility mine plan shows 7.1 million tonnes of material to be processed through the mill during a seven year mine life. The criteria provided by Bema Gold Corporation (Bema) were to develop a water management plan, site water balance, and conduct surface water quality modeling efforts consistent with a feasibility level of analysis and two alternative tailings storage scenarios. This report presents the results of preliminary mass balance and geochemical surface water quality modeling activities for discharges reporting to ambient receiving waters near the proposed Kupol Gold Project. Concentrations of all relevant water quality parameters were compared against the World Bank direct discharge criteria for open pit (World Bank, 1999a) and underground (World Bank, 1999b) mines, and against the Russian Class 1 fisheries standards (and drinking water/recreational use standards where no fisheries standard was available) (VNIRO, 1999). The following three general project phases were considered: (1) the initial year of site construction in 2006; (2) the first full year of pit development and prior to construction of the starter tailings dam in 2007; and (3) the operations phase of the project from 2008 through 2014 after the closed-circuit process plant-tailings pond water balance and process mill start-up has occurred. Two tailings alternatives were also considered separately in the modeling approach. The following four distinct source profiles were used in the coupled mass balance-geochemical ambient surface water quality model: (1) background concentrations in the Kaiemraveem River; (2) background concentrations in contributing catchments of the Kaiemraveem River and from natural surface runoff; (3) concentrations in runoff from NAG waste/quarried rock; and (4) concentrations in runoff from treated and untreated AG waste rock. Background concentrations were obtained from the baseline water quality and hydrology program conducted by Bema (2005). Concentrations in NAG and AG runoff were estimated based on knowledge of baseline and mine site hydrology, and the comprehensive 2003-2004 geochemical testing program (Atkin, 2005). Modeling of secondary geochemical reactions, and the potential alterations to estimated runoff water quality from NAG and AG rock surfaces due to particulate/colloidal phase retention in pore spaces and settling ponds, were investigated using an equilibrium based approach. Dissolved phase concentrations from unsaturated and saturated NAG rock, and unsaturated AG rock, were then input to a series of linked mass balance calculations coupled to the site water balance model. Using this approach, and assuming complete mixing of all water sources prior to the ambient regulatory point-of-compliance on the Kaiemraveem River at a location 500 m downstream of the last potential input from project site runoff, the resulting concentrations of the various water quality parameters of interest were calculated. During 2006, runoff will occur from the following site facilities: plant and camp site, haul and access roads, the airstrip, and the explosives storage facility. No direct discharge exceedences into either the Kaiemraveem or Starichnaya Rivers, or their contributing catchments, under


AMEC File: VM00330 Page v

either of the tailings alternatives are expected from site runoff in 2006 under the World Bank criteria (World Bank, 1999a and 1999b). In addition, no Russian Class 1 fisheries standard exceedences are expected from site runoff into either the Kaiemraveem or Starichnaya Rivers under either of the tailings alternatives at the ambient regulatory points-of-compliance located 500 m downstream of the last project discharge into either of these rivers during 2006 under the Russian Class 1 fisheries standards (VNIRO, 1999). During 2007, runoff will occur from the following site facilities: plant and camp site, haul and access roads, the airstrip, the explosives storage facility, the AG waste rock stockpile, the ore stockpile, and pit inflows. Runoff from the AG waste rock stockpile, the ore stockpile, and pit inflows will be treated with lime neutralization prior to release into receiving waters. No direct discharge exceedences into either the Kaiemraveem or Starichnaya Rivers, or their contributing catchments, under either of the tailings alternatives are expected from site runoff in 2007 under the World Bank criteria (World Bank, 1999a and 1999b). In addition, with the exception of zinc, no Russian Class 1 fisheries standard exceedences are expected from site runoff into either the Kaiemraveem or Starichnaya Rivers under either of the tailings alternatives at the ambient regulatory points-of-compliance located 500 m downstream of the last project discharge into either of these rivers under the Russian Class 1 fisheries standards (VNIRO, 1999). During 2007, zinc concentrations at the ambient point-of-compliance in the Kaiemraveem River are expected to exceed the Russian Class 1 fisheries standards (VNIRO, 1999) by about 10% under both tailings alternatives. The minimal exceedence factor for this parameter, as well as its limited spatial and temporal duration (≤4-5 months over a 14 year project life), combined with generally conservative regulatory levels for zinc in freshwaters suggests that the short-term elevated zinc level in the Kaiemraveem River will not pose a significant threat to aquatic life. Over the period from 2008 through 2014, runoff will occur from the following site facilities: plant and camp site, haul and access roads, the airstrip, the explosives storage facility, and the downstream crest of the tailings dam. No direct discharge exceedences into either the Kaiemraveem or Starichnaya Rivers, or their contributing catchments, under either of the tailings alternatives are expected from site runoff from 2008 through 2014 under the World Bank criteria (World Bank, 1999a and 1999b). In addition, no Russian Class 1 fisheries standard exceedences are expected from site runoff into either the Kaiemraveem or Starichnaya Rivers under either of the tailings alternatives at the ambient regulatory points-of-compliance located 500 m downstream of the last project discharge into either of these rivers from 2008 through 2014 under the Russian Class 1 fisheries standards (VNIRO, 1999). The current water and waste management planning for the project, as indicated by the results of the current modeling exercise, suggest there will be no significant water quality impacts on the Kaiemraveem and Starichnaya Rivers, including their contributing catchments, from development of the Kupol Gold Project between 2006 and 2014.



1.0 INTRODUCTION The Kupol Gold Project will include activities such as open pit mining, underground mining, ore stockpiling and processing, tailings and waste management, and camp operation that will generate volumes and qualities of water requiring management. The feasibility mine plan shows 7.1 million tonnes of material to be processed through the mill during a seven year mine life. The criteria provided by Bema Gold Corporation (Bema) were to develop a water management plan, site water balance, and conduct surface water quality modeling efforts consistent with a 12 million tonne mine plan (or approximately 14 years of mill operation) and two alternative tailings storage scenarios as shown in Figures 1.1 and 1.2. Key site locations, hydrologic features, catchment names and areas, and regulatory points-of-compliance are shown in these figures and are referred to throughout the current report. This report presents the results of preliminary mass balance and geochemical surface water quality modeling activities for discharges reporting to ambient receiving waters near the proposed Kupol Gold Project. This study is one of several reports pertaining to water management for the project, as listed in Table 1.1. The Surface Water Quality Modeling Report presents the results from preliminary modeling activities under the current Water Management Plan (AMEC, 2005a) and Site Water Balance (AMEC, 2005b) during the operational phase of the project.

Table 1.1: Project reports pertaining to water management.

Report Issue Date Comments

Tailings Facility Design June 2005 Designs for location, construction, and operation of the tailings facility

Waste Rock Management June 2005 Management plans for waste rock over the life of the project.

Site Water Balance Report June 2005 A description of the site water balance model and the water balance projections, on a monthly basis, for the duration of the project.

Water Management Plan June 2005 A general overview of the overall site water management strategy and objectives, with a summary of key water balance issues.

Ambient Surface Water Quality Modeling – This Study June 2005

Water quality projections provided for discharge to Kaiemraveem Creek, and for runoff from various terrain types (e.g., waste rock, plant site, natural landscape). The water quality modeling is integrated within the water balance model for the site.

Closed Circuit Tailings Pond Water Quality Modeling June 2005

Water quality projections for the closed-circuit operations phase water balance between the process plant and tailings pond.



Figure 1.1: Regional site layout and watershed boundaries.



Figure 1.2: Plant site layout.



2.0 METHODS The following four distinct source profiles were used in the coupled mass balance-geochemical ambient surface water quality model: Background concentrations in the Kaiemraveem River: Water quality samples were collected upstream of the project site at stations K101 and K102 on the Kaiemraveem River during 2004. Sampling dates at each station were as follows: K101 – June 21, July 17, September 13, and October 1; K102 – July 17, September 13, and October 1 (no sample on June 21 was taken at site K102). For the purposes of modeling, the average background concentration in the Kaiemraveem River was taken as the arithmetic mean of these samples for each parameter (Table 2.1). Details on the sampling locations and dataset can be obtained elsewhere (Bema, 2005). Sites K104 and K105 on the Kaiemraveem River were not used to assess background water quality profiles. These sites contain elevated sulfate (58-276 mg/L, versus 4.9-22 mg/L upstream at K101 and K102) and calcium (17-72 mg/L, versus 3-7 mg/L upstream at K101 and K102) concentrations suggesting current impacts at the project site (i.e., pre-2006 impacts) on downstream water quality in the Kaiemraveem River. Further sampling is recommended to better characterize baseline water quality in the Kaiemraveem River (and Starichnaya River). Background concentrations in contributing catchments of the Kaiemraveem River and from natural surface runoff: Water quality samples were taken near the project site in contributing catchments of the Kaiemraveem River at stations K202, K302, K401, K402, and K601 during 2004. As with samples K104 and K105 on the Kaiemraveem River, samples from sites K202, K302, K401, and K402 appear impacted by current activities at the project site (elevated calcium, magnesium, and sulfate concentrations, and depressed pH values), and are not representative of unimpacted background runoff contributing to the Kaiemraveem River. Samples from K601 were collected on July 17, September 13, and October 1 (no sample on June 21 was taken at site K601). For the purposes of modeling, the average background concentration in the contributing catchments of the Kaiemraveem River and from natural surface runoff on, and around, the project site was taken as the arithmetic mean of these samples at site K601 for each parameter (Table 2.1). Details on the sampling locations and dataset can be obtained elsewhere (Bema, 2005). Note that background concentrations in the Kaiemraveem River and its contributing catchments and natural surface runoff were assumed equivalent to the corresponding background levels in the Starichnaya River system. No background water quality samples were available for comparison in the Starichnaya River system.



Table 2.1: Summary of estimated background water quality in the Kaiemraveem River and its contributing catchments near the project site. Applicable ambient receiving water

quality limits are shown for comparison.

KaiemraveemRiver

Background Natural Surface Runoff

Russian Class 1 Fisheries Standarda

Bulk Parameters pH 7.29 7.39 6.5-8.5 Turbidity (NTU) 4.9b 4.8b 0.25 Total Dissolved Solids (mg/L) 74 66 n/ac Petroleum Products (mg/L) <0.05 <0.05 0.05 Detergents (mg/L) <0.01 <0.01 0.5 Phenols (mg/L) <0.0005 <0.0005 0.001 Major Ions Sodium (mg/L) 3.8 3.9 120 Potassium (mg/L) 0.65 0.23 50 Calcium (mg/L) 4.0 7.8 180 Magnesium (mg/L) 1.4 3.0 40 Bicarbonate (mg CaCO3/L) 21 27 n/ac Carbonate (mg CaCO3/L) <1 <1 n/ac Chloride (mg/L) <1 <1 300 Sulphate (mg/L) 10 19 100 Nutrients BOD (mg/L) 0.52 0.50 n/ac Ammonia (mg/L) 0.51b 0.23 0.5 Nitrate (mg/L) 0.18 0.23 40 Nitrite (mg/L) 0.040 0.037 0.08 Phosphate (mg/L) <0.05 <0.05 0.2 Metals Aluminum (mg/L) n/ad n/ad 0.04 Antimony (mg/L)c <0.01e <0.01e 0.005f Arsenic (mg/L) <0.05 <0.05 0.05 Cadmium (mg/L) <0.01e <0.01e 0.005 Chromium (mg/L) <0.02 <0.02 0.07 Cobalt (mg/L) <0.01 <0.01 0.01 Copper (mg/L) 0.0029b 0.0012b 0.001 Iron (mg/L) 0.59b 0.052 0.1 Lead (mg/L) <0.002 <0.002 0.006 Manganese (mg/L) 0.019b <0.01 0.01 Mercury (mg/L) <0.0015e <0.0015e 0.00001 Molybdenum (mg/L) n/ad n/ad 0.001



KaiemraveemRiver

Background Natural Surface Runoff

Russian Class 1 Fisheries Standarda

Nickel (mg/L) <0.01 <0.01 0.01 Selenium (mg/L) <0.02e <0.02e 0.002 Silver (mg/L) n/ad n/ad 0.01 Strontium (mg/L) <0.05 <0.05 0.4 Thallium (mg/L) n/ad n/ad 0.0001 Zinc (mg/L) 0.0080 0.011b 0.01

a Russian Class 1 fisheries limits are shown for all parameters except where noted (VNIRO. 1999). b Background concentrations in the Kaiemraveem River and/or its contributing catchments exceed Russian Class 1 fisheries standard guidelines. Assessment of potential project impacts on ambient water quality must be considered in this context. c No applicable guideline is available under the given regulatory class. d Concentrations in the Kaiemraveem River and/or its contributing catchments are not available for this parameter. e Method detection limits (MDLs) for background water quality samples in the Kaiemraveem River and/or its contributing catchments exceed Russian Class 1 fisheries standard guidelines. Assessment of potential project impacts on ambient water quality must be considered in this context. Values at one-half the MDL were used for modeling for these parameters. f There is no Russian Class 1 fisheries limit for antimony (VNIRO, 1999). The drinking water/recreational use limit is shown (VNIRO, 1999).



Note that background turbidity, biochemical oxygen demand (BOD), and concentrations of copper, iron, and manganese currently exceed Russian Class 1 fisheries standard guidelines in either, or both, the Kaiemraveem River and its contributing catchments. Assessment of potential project impacts on ambient water quality must be considered in this context. In some cases, the project may reduce a water quality exceedence for a parameter that exceeds Russian Class 1 fisheries standard guidelines, but the resulting concentration may still remain above the regulatory limit. In such an event, the project would be deemed as not being out of regulatory compliance, and can be thought of as creating a net improvement in regional water quality for the parameter of interest. For antimony, cadmium, mercury, and selenium, the MDL in background water quality samples exceeds the corresponding Russian Class 1 fisheries standard guidelines. A reliable assessment of whether background water quality currently meets the Russian Class 1 fisheries standard guidelines for these parameters cannot be made. For the purposes of modeling, a value of one-half the MDL was used for background and natural surface runoff concentrations. Thus, estimates of resulting water quality in the Kaiemraveem River during the operations phase of the project may be conservative and represent upper bounds. If the modeled concentrations for these parameters (assuming background concentrations at one-half the MDL) are of concern, further water quality analyses for these source types may be required using methods with lower detection limits. For aluminum, molybdenum, and silver, no analyses were conducted in background water sources. Without background levels for these analytes, modeling of ambient receiving water concentrations cannot take place. However, concentrations of these parameters in site runoff (see below) can be compared to direct discharge standards. Further water quality analyses may be required for these analytes should they pose a potential concern in receiving waters. Concentrations in runoff from NAG waste/quarried rock: Construction of haul and access roads, plant and camp site basing, the airstrip, and the surface of the tailings dam will utilize NAG rock obtained either as waste from the open pit development, or quarried from nearby sources. Note that there is a lack of geochemical data for barren borrow material, and there is a likely difference between carbonate and metals concentrations in the NAG waste rock and in barren surface rocks (e.g., barren basalt and/or andesite). As such, the current approach for quarried materials is an approximation, and selected ABA testing of quarried materials should be undertaken during operations to ensure this rock is suitable for construction without harmful water quality effects. A comprehensive geochemical testing program was undertaken as part of the project feasibility level assessment (Atkin, 2005). A suite of representative NAG waste rock samples was subjected to 20 weeks of humidity cell testing. Over the final five weeks of humidity cell testing, normalized concentrations of water quality parameters (termed production rates and expressed in units of mg of analyte per kg of rock sample per week of testing; mg/kg/wk) were stable and termed to be at equilibrium conditions. Only the following five NAG humidity cells which maintained pH values in the leachate above 6.5 (the minimum Russian Class 1 fisheries standard regulatory limit for pH) were used to calculate representative NAG production rates:



HC6, HC8, HC12, HC19, and HC22. Individual and arithmetically averaged production rates in the NAG humidity cells are given in Table 2.2.



Table 2.2: Late-stage equilibrium pH values and elemental production rates from the individual NAG rock humidity cells. All values are in units of mg/kg/wk, except for pH,

which is presented in traditional logarithmic units.

HC6 HC8 HC12 HC19 HC22 Average pH 7.11 8.73 7.18 7.25 7.81 7.44 Chloride 0.0031 0.013 0.012 0.012 0.013 0.011 Fluoride 0.0016 0.0015 0.0015 0.0015 0.0015 0.0015 Sulfate 1.56 1.07 3.08 5.14 4.54 3.08 Nitrite 0.000031 0.000031 0.00016 0.00013 0.000045 0.000078 Nitrate 0.0015 0.0025 0.00030 0.0020 0.0032 0.0019 Aluminum 0.0012 0.0018 0.0013 0.0011 0.0056 0.0022 Antimony 0.00038 0.00020 0.00028 0.000058 0.0014 0.00047 Arsenic 0.00014 0.00031 0.00030 0.000029 0.022 0.0045 Cadmium 0.0000031 0.0000031 0.0000030 0.0000029 0.0000030 0.0000030 Calcium 0.21 0.069 0.28 0.31 0.25 0.23 Chromium 0.000016 0.000016 0.000015 0.000015 0.000015 0.000015 Cobalt 0.000016 0.000016 0.000015 0.000015 0.000015 0.000015 Copper 0.000016 0.000016 0.000015 0.000015 0.000023 0.000017 Iron 0.00078 0.00077 0.00075 0.00073 0.00076 0.00076 Lead 0.000016 0.000016 0.000015 0.000015 0.000015 0.000015 Magnesium 0.046 0.0080 0.019 0.015 0.049 0.027 Manganese 0.000078 0.00011 0.00010 0.00012 0.00026 0.00013 Mercury 0.00000031 0.00000031 0.00000030 0.00000029 0.00000030 0.00000030Molybdenum 0.00048 0.000025 0.0000075 0.000093 0.00099 0.00032 Nickel 0.000016 0.000016 0.000015 0.000015 0.000015 0.000015 Phosphate 0.019 0.0023 0.0022 0.0022 0.0023 0.0056 Potassium 0.042 0.017 0.054 0.026 0.098 0.047 Selenium 0.000016 0.000016 0.000015 0.000015 0.000015 0.000015 Silver 0.0000039 0.0000039 0.0000037 0.0000037 0.0000038 0.0000038 Sodium 0.075 0.69 0.033 0.038 0.23 0.21 Strontium 0.0033 0.0017 0.0045 0.0026 0.0080 0.0040 Thallium 0.000013 0.0000015 0.0000022 0.0000015 0.0000015 0.0000038 Zinc 0.000078 0.000077 0.000075 0.000073 0.000076 0.000076 Alkalinity 3.4 6.6 3.1 2.5 4.7 4.0 Acidity (to pH 4.5) 0 0 0 0 0 0 Acidity (to pH 8.3) 0.8 0 1.0 1.4 1.5 1.3



Average elemental production rates for individual parameters from the NAG waste rock humidity cells were converted to expected runoff water quality from haul and access roads, the explosives storage area, the plant and camp site, the airstrip, and the downstream crest of the tailings dam (and other exposed NAG rock on-site) by the following approach. During the year of rock placement, annual average unit runoff from disturbed surfaces was estimated at 280 mm (AMEC, 2005b) and was calculated by subtracting over-winter sublimation losses (70 mm; Zhang et al., 2003; Zhang et al., 2004) from the average annual total precipitation (380 mm), and multiplying by a runoff coefficient of 0.9. In addition, net wetting of exposed rock by 1% is expected to occur, based on assumptions from similar northern sites (Diavik, 2000). With an average rock density of 2,650 kg/m3, a porosity of 25%, and an active layer/infiltration depth of 1 m, each 1 m2 of exposed rock surface will consume 20 mm of precipitation in the first full year after placement. Furthermore, the disturbance during placement will allow additional water storage in regions of ponding and in uncompacted pore spaces. Subsequent evaporation losses under this scenario was assumed to equal the estimated annual site evapotranspiration of 120 mm (Berezovskaya et al., 2004; Kubota et al., 2004; Yamazaki et al., 2004). This approach yields an expected average annual unit runoff of 130 mm from unsaturated disturbed surfaces during the year of construction. For all subsequent years, the disturbed region is assumed to be both saturated and undergoing compaction/consolidation, thereby reducing the ability of the active layer to either consume water as part of net-wetting of pore spaces, or to hold water for a period sufficient to allow significant evaporative losses. In such a case, the saturated rock yields an average annual unit runoff of 280 mm. Average elemental production rates (Pavg) were then converted to corresponding runoff concentrations (C) via the following formula,

2

3

3

2

3

3avg

mmR

L 1000m 1

mm10.75

mkg2,6500.2 wk52

wkkgmgP

C(mg/L)××××××

×=

where 52 weeks is a multiplier to convert weekly average production rates to annual average production rates, 0.2 is a multiplier to account for the approximately 20% fraction of rock surface with weathering products that is flushed by runoff/infiltration each year (Morin, 1997), 2,650 kg/m3 is the assumed density of rock, 0.75 accounts for the 75% of each subsurface m3 occupied by rock (i.e., 25% porosity), 1 m3/m2 is the volume of rock in the active/infiltration layer normalized per unit area, R is the average annual unit runoff for unsaturated (0.130 m or 130 mm) and saturated (0.280 mm or 280 m) rock, and 1 m3/1000 L is a conversion factor to ensure the units of C are in mg/L. Using this approach, the estimated average runoff water quality from unsaturated and saturated NAG rock is given in Table 2.3. Direct discharge limits referenced here and throughout the



report are from the World Bank guidelines for underground (World Bank, 1999a) and open pit (World Bank, 1999b) mines.



Table 2.3: Summary of estimated runoff water quality from exposed unsaturated and saturated NAG rock. Applicable direct discharge and Russian Class 1 fisheries standards and the corresponding exceedence factors are shown for comparison.

NAG Rock (Unsaturated) NAG Rock (Saturated)

Runoff Water Quality Exceedence Factor (Direct Discharge)

Exceedence Factor (Ambient Receiving Water) Runoff Water Quality

Exceedence Factor (Direct Discharge)

Exceedence Factor (Ambient Receiving Water)

Direct Discharge Limit

Ambient Receiving Water Limit

pH 7-8 <<1 <<1 7-8 <<1 <<1 6-9 6.5-8.5 Alkalinity (mg/L) 640 n/a n/a 300 n/a n/a n/a n/a Sulfate (mg/L) 490b n/a 4.9 230b n/a 2.3 n/a 100 Chloride (mg/L) 1.7 n/a 0.0056 0.79 n/a 0.0026 n/a 300 Fluoride (mg/L) <0.24b n/a <4.8 <0.11b n/a <2.2 n/a 0.05 Nitrite (mg/L) 0.012 n/a 0.16 0.0058 n/a 0.072 n/a 0.08 Nitrate (mg/L) 0.30 n/a 0.0076 0.14 n/a 0.0035 n/a 40 Aluminum (mg/L) 0.35b n/a 8.7 0.16b n/a 4.0 n/a 0.04 Antimony (mg/L) 0.075b n/a 15 0.035b n/a 7.0 n/a 0.005 Arsenic (mg/L) 0.71b 0.71 14 0.33b 0.33 6.6 1 0.05 Cadmium (mg/L) <0.00048 <0.0048 <0.096 <0.00022 <0.0022 <0.045 0.1 0.005 Calcium (mg/L) 36 n/a 0.20 17 n/a 0.093 n/a 180 Chromium (mg/L) <0.0024 <0.0024 <0.034 <0.0011 <0.0011 <0.016 1 0.07 Cobalt (mg/L) <0.0024 n/a <0.24 <0.0011 n/a <0.11 n/a 0.01 Copper (mg/L) <0.0027b <0.0088 <2.6 <0.0012b <0.0041 <1.2 0.3 0.001 Iron (mg/L) <0.12b <0.060 <1.2 <0.056 <0.028 <0.56 2 0.1 Lead (mg/L) <0.0024 <0.0040 <0.40 <0.0011 <0.0019 <0.19 0.6 0.006 Magnesium (mg/L) 4.4 n/a 0.11 2.0 n/a 0.051 n/a 40 Manganese (mg/L) 0.021b n/a 2.1 0.0099b n/a 0.99 n/a 0.01 Mercury (mg/L) <0.000048b <0.024 <4.8 <0.000022b <0.011 <2.2 0.002 0.00001 Molybdenum (mg/L) 0.051b n/a 51 0.024b n/a 24 n/a 0.001 Nickel (mg/L) 0.0024 0.0048 0.24 0.0011 0.0022 0.11 0.5 0.01 Phosphate (mg/L) <0.88b n/a <4.4 <0.41b n/a <2.1 n/a 0.2 Potassium (mg/L) 7.5 n/a 0.15 3.5 n/a 0.070 n/a 50 Selenium (mg/L) <0.0024b n/a <1.2 <0.0011 n/a <0.56 n/a 0.002 Silver (mg/L) <0.00060 n/a <0.060 <0.00028 n/a <0.028 n/a 0.01 Sodium (mg/L) 34 n/a 0.28 16 n/a 0.13 n/a 120 Strontium (mg/L) 0.64b n/a 1.6 0.29 n/a 0.74 n/a 0.4 Thallium (mg/L) <0.00061b n/a <6.1 <0.00028b n/a <2.8 n/a 0.0001 Zinc (mg/L) <0.012b <0.012 <1.2 <0.0056 <0.0056 <0.56 1 0.01 a Meets or exceeds direct discharge limit and requires dilution and/or treatment prior to discharge. b Meets or exceeds Russian Class 1 fisheries limit and requires dilution and/or treatment prior to edge of mixing zone.



The runoff from either unsaturated or saturated NAG rock is not expected to exceed direct discharge limits for any parameter. Thus, prior treatment of NAG rock runoff is not required before discharge to receiving waters provided sufficient dilution is available prior to the edge of the mixing zone, or reaching the ambient point-of-compliance, such that Russian Class 1 fisheries standard limits are met. Concentrations in runoff from AG waste rock: During 2006 and 2007, AG waste rock will be stockpiled in the catchment upstream of the tailings dam prior to starter dam completion in late 2007. Runoff from this stockpile will be minimized, collected, and treated as discussed in the Water Management Plan (AMEC, 2005a). Note that due to the late start of pit construction in 2006 (about October), there will be negligible runoff from the AG waste rock stockpile during this year, and a treatment system will not be required for this source until spring 2007, and will only operate during 2007. AG rock will also be exposed in the pit walls during 2006 and 2007 and will yield runoff from direct precipitation on the open pit and thawback inflows from the pit walls as discussed in the Site Water Balance (AMEC, 2005b). Pit inflows will be minimized, collected, as treated as discussed in the Water Management Plan (AMEC, 2005a). The late start of pit development in 2007 will result in negligible pit inflows requiring management during 2006. Thus, the pit inflow treatment system need only operate for 2007. From late 2007 through the duration of the project’s operational phase (about 2014), all pit inflows will be sent to the tailings pond, and any AG waste rock stockpile runoff will be contained behind the tailings dam. Thus, a water treatment system will not be required from late 2007 through 2014. A comprehensive geochemical testing program was undertaken as part of the project feasibility level assessment (Atkin, 2005). A suite of representative AG waste rock samples was subjected to 20 weeks of humidity cell testing. Over the final five weeks of humidity cell testing, normalized concentrations of water quality parameters were stable and termed to be at equilibrium conditions. Only the following seven AG humidity cells which maintained pH values in the leachate below 5 were used to calculate representative AG production rates: HC4, HC5, HC7, HC11, HC13, HC14, and HC15. Individual and arithmetically averaged production rates in the AG humidity cells are given in Table 2.4.



Table 2.4: Late-stage equilibrium pH values and elemental production rates from the individual AG rock humidity cells. All values are in units of mg/kg/wk, except for pH, which is presented in traditional logarithmic units.

HC4 HC5 HC7 HC11 HC13 HC14 HC15 Average pH 2.66 3.45 3.41 3.95 2.30 2.16 2.87 2.69 Chloride 0.0031 0.0031 0.028 0.0085 0.0073 0.0079 0.0031 0.0087 Fluoride 0.019 0.0015 0.015 0.0014 0.0077 0.010 0.0112 0.0095 Sulfate 280 68 210 37 420 590 610 320 Nitrite 0.00020 0.000031 0.000028 0.000028 0.0011 0.0011 0.00051 0.00042 Nitrate 0.00031 0.00031 0.0019 0.00028 0.00076 0.00072 0.0010 0.00076 Aluminum 0.71 0.017 0.055 0.0046 0.53 0.64 0.85 0.40 Antimony 0.00036 0.00011 0.000014 0.000014 0.000015 0.000014 0.000015 0.000077 Arsenic 0.0045 0.023 0.000014 0.000014 0.0054 0.019 0.00020 0.0074 Cadmium 0.000019 0.0000031 0.00010 0.0000028 0.0000031 0.0000043 0.0000031 0.000019 Calcium 0.54 1.0 3.3 0.59 2.6 2.6 11 3.1 Chromium 0.00069 0.000015 0.000014 0.000014 0.00034 0.00036 0.00022 0.00023 Cobalt 0.0045 0.0046 0.0044 0.00015 0.0027 0.0030 0.00094 0.0029 Copper 0.0034 0.0010 0.00030 0.00047 0.0047 0.0052 0.0052 0.0029 Iron 5.5 0.28 0.079 0.00069 5.5 7.6 0.38 2.8 Lead 0.000016 0.000015 0.000014 0.000014 0.000015 0.000014 0.000015 0.000015 Magnesium 0.10 0.11 1.5 0.10 0.018 0.029 0.17 0.28 Manganese 0.11 0.14 0.66 0.0078 0.0024 0.0023 0.030 0.14 Mercury 0.00000031 0.00000031 0.00000028 0.00000028 0.00000030 0.00000029 0.00000031 0.00000030 Molybdenum 0.00016 0.0000077 0.0000069 0.0000069 0.0020 0.000081 0.0000077 0.00033 Nickel 0.0015 0.0016 0.0022 0.000069 0.0014 0.0014 0.00051 0.0013 Phosphate 0.0023 0.0023 0.0021 0.0021 0.28 0.39 0.0023 0.098 Potassium 0.013 0.091 0.012 0.014 0.032 0.013 0.0077 0.026 Selenium 0.000039 0.000015 0.000021 0.000014 0.00028 0.000080 0.00014 0.000083 Silver 0.0000039 0.0000039 0.0000034 0.0000035 0.0000038 0.0000062 0.0000039 0.0000041 Sodium 0.0080 0.0076 0.085 0.018 0.0041 0.0068 0.082 0.030 Strontium 0.013 0.017 0.0011 0.000069 0.030 0.036 0.020 0.017 Thallium 0.000067 0.000029 0.0000066 0.0000014 0.000032 0.0000036 0.0000015 0.000020 Zinc 0.030 0.024 0.0061 0.00069 0.00062 0.0056 0.0097 0.011 Alkalinity <MDLa <MDL <MDL <MDL <MDL <MDL <MDL <MDL Acidity (to pH 4.5) 78 4.2 5.1 0.2 210 330 67 99 Acidity (to pH 8.3) 300 21 20 4.9 350 570 150 1,400 a Concentrations were below method detection limits over the last five weeks of testing



The same approach taken to convert elemental production rates for the NAG humidity cells (described above) into corresponding runoff water quality was also applied to the AG humidity cell dataset. Estimated concentrations of various water quality parameters were calculated for both saturated and unsaturated AG rock. Unsaturated AG rock will be represented by the AG waste rock stockpile behind the tailings dam while pit development is occurring (2006-2010) and fresh waste rock is being place in this location, as well as freshly exposed AG pit walls during continuing pit development from 2006 through 2010. Once placement of AG waste rock behind the tailings dam is completed in late 2010, any exposed AG waste rock in the tailings pond will yield saturated AG rock water quality. Based on the geochemical testing results for the AG, PAG, and NAG rock to be exposed in the pit walls (Atkin, 2005), as well as experience with acidic drainage at similar northern mine sites, treatment of the pit inflows is expected to require a calcium oxide (CaO; or pebble lime) dose of 2.5 kg/m3. The relative distribution of AG, PAG, and NAG in the final pit wall exposure is estimated as follows: AG (42%), PAG (11%), and NAG (47%) (Wardrop, 2005). The laboratory based humidity cell testing (Atkin, 2005) suggests there is insufficient alkalinity produced from the NAG waste rock to neutralize the acidity produced from the AG waste rock when approximately equal quantities of drainage from these material types are combined. Thus, approximately equal quantities of exposed AG and NAG waste rock are expected to yield an overall acidic runoff with a pH in the range of 3-4 and high dissolved metal and major ion concentrations. For the purposes of the site water quality model, equal quantities of exposed unsaturated NAG and AG rock were assumed in the open pit during its development in 2007. Using the same approach taken for the NAG rock cells, the estimated average runoff water quality from unsaturated and saturated AG rock is given in Table 2.5.



Table 2.5: Summary of estimated runoff water quality from exposed unsaturated and saturated AG rock. Applicable direct discharge and Russian Class 1 fisheries standards and the corresponding exceedence factors are shown for comparison.

AG Rock (Unsaturated) AG Rock (Saturated)

Runoff Water Quality Exceedence Factor (Direct Discharge)

Exceedence Factor (Ambient Receiving Water) Runoff Water Quality

Exceedence Factor(Direct Discharge)


Direct DischargeLimit

Ambient Receiving Water Limit

pH 3-4a,b 100-1,000 320-2,800 3-4a,b 100-1,000 320-2,800 6-9 6.5-8.5 Alkalinity (mg/L) n/a n/a n/a n/a n/a n/a n/a n/a Sulfate (mg/L) 50,000b n/a 500 23,000b n/a 230 n/a 100 Chloride (mg/L) 1.4 n/a 0.0046 0.64 n/a 0.0021 n/a 300 Fluoride (mg/L) 1.5b n/a 30 0.70b n/a 14 n/a 0.05 Nitrite (mg/L) 0.067 n/a 0.83 0.031 n/a 0.39 n/a 0.08 Nitrate (mg/L) 0.12 n/a 0.0030 0.056 n/a 0.0014 n/a 40 Aluminum (mg/L) 64b n/a 1,600 30b n/a 740 n/a 0.04 Antimony (mg/L) 0.012b n/a 2.5 0.0057b n/a 1.1 n/a 0.005 Arsenic (mg/L) 1.2a,b 1.2 23 0.54b 0.54 11 1 0.05 Cadmium (mg/L) 0.0031 0.031 0.62 0.0014 0.014 0.29 0.1 0.005 Calcium (mg/L) 500b n/a 2.8 230b n/a 1.3 n/a 180 Chromium (mg/L) 0.037 0.037 0.53 0.017 0.017 0.25 1 0.07 Cobalt (mg/L) 0.46b n/a 46 0.21b n/a 22 n/a 0.01 Copper (mg/L) 0.46a,b 1.5 460 0.21b 0.72 210 0.3 0.001 Iron (mg/L) 440a,b 220 4,400 203a,b 100 2,000 2 0.1 Lead (mg/L) <0.0024 <0.0039 <0.39 <0.0011 <0.0018 <0.18 0.6 0.006 Magnesium (mg/L) 45b n/a 1.12 21 n/a 0.52 n/a 40 Manganese (mg/L) 22b n/a 2,200 10b n/a 1,000 n/a 0.01 Mercury (mg/L) <0.000047b <0.024 <4.7 <0.000022b <0.011 <2.2 0.002 0.00001 Molybdenum (mg/L) 0.052b n/a 52 0.024b n/a 24 n/a 0.001 Nickel (mg/L) 0.20b 0.40 20 0.093b 0.19 9.3 0.5 0.01 Phosphate (mg/L) 16b n/a 78 7.2b n/a 36 n/a 0.2 Potassium (mg/L) 4.1 n/a 0.083 1.9 n/a 0.038 n/a 50 Selenium (mg/L) 0.013b n/a 6.6 0.0061b n/a 3.1 n/a 0.002 Silver (mg/L) <0.00065 n/a <0.065 <0.00030 n/a <0.030 n/a 0.01 Sodium (mg/L) 4.8 n/a 0.040 2.2 n/a 0.019 n/a 120 Strontium (mg/L) 2.7b n/a 6.6 1.2b n/a 3.1 n/a 0.4 Thallium (mg/L) 0.0032b n/a 32 0.0015b n/a 15 n/a 0.0001 Zinc (mg/L) 1.7a,b 1.7 170 0.81b 0.81 81 1 0.01

a Meets or exceeds direct discharge limit and requires dilution and/or treatment prior to discharge. b Meets or exceeds Russian Class 1 fisheries limit and requires dilution and/or treatment prior to edge of mixing zone.



The estimated concentrations of water quality parameters in NAG and AG runoff from unsaturated and saturated rock given above are conservative mass balance estimates that assume no removal or retention of particulate or colloidal species prior to runoff leaving the region of interest. Thus, these concentrations represent upper-bound estimates not likely to be observed under field conditions. Secondary geochemical reactions will begin to occur immediately after the weathering and leaching process releases compounds into the runoff and/or infiltration. Modeling of secondary geochemical reactions, and the potential alterations to estimated runoff water quality from NAG and AG rock surfaces due to particulate/colloidal phase retention in pore spaces and settling ponds, were investigated using an equilibrium based approach and the CHEAQS Pro v.2005.1 software program (Verweij, 2005). CHEAQS is a program that can calculate chemical equilibria in aquatic systems. It includes a database with many complexes, redox equilibria, gas solution equilibria, saturation solids, adsorption of cations and ligands. A number of peer-reviewed publications in the open scientific literature have used CHEAQS for geochemical modeling. Regulatory agencies in North America are cognizant as to the over-conservatism brought into modeling the potential environmental impacts of a large resource project when only mass-balance approaches are used for estimating runoff water quality. Geochemical modeling approaches (using software such as, e.g., CHEAQS, PHREEQC, and MINEQL) are encouraged so that a more realistic assessment of resulting water quality can be made. Note that any forward equilibrium thermodynamic model assumes all reactions are instantaneous and go to completion. Both assumptions are difficult to achieve under field conditions for surface processes. Thus, assumptions regarding relative precipitation of particulate and colloidal species may represent upper bounds, and an increased contribution from the dissolved fraction may be present under field conditions due to kinetic constraints. The water quality profiles for exposed unsaturated and saturated NAG and AG rock presented in Table 2.3 and Table 2.5 were input to the CHEAQS program, and the system allowed to reach geochemical equilibrium according to the following constraints. The pH of NAG and AG waste rock was assumed to be 7 and 3, respectively, as per the equilibrium pH values from the 20 weeks of humidity cell testing (Atkin, 2005). Organic complexation of metals (predominantly divalent cations such as copper and zinc) via dissolved organic carbon was ignored. Redox equilibria was modeled assuming a pE value of 8.45 (500 mV), which is representative of oxygenated surface water environments (Schulz, 2004). Saturation solids were enabled during calculation, thereby facilitating an understanding as to the relative saturation indices of various parameters. Partial pressures of atmospheric gases such as CO2, NH3, H2S, and SO2 were ignored. Adsorption phasing and subsequent particulate/colloidal scavenging was ignored in the modeling approach, and the convergence criterion was 0.01%. The results of the geochemical modeling on unsaturated and saturated exposed NAG and AG rock are given in Table 2.6, Table 2.7, Table 2.8, and Table 2.9. The major contributing species (and their percentage contribution towards total particulate and dissolved source concentrations in parentheses) are shown by element for each runoff type, along with comparisons of dissolved



concentrations to direct discharge and ambient receiving water limits by way of exceedence factors.



Table 2.6: Estimated solid and dissolved phase concentrations, with major speciation contributions, of water quality parameters from exposed unsaturated NAG rock. Applicable regulatory limits and the corresponding exceedence factors for dissolved phase concentrations are shown for comparison.

Solid Species Dissolved Species

Parameter

Total Concentration (mg/L)

Concentration (mg/L)

Major Speciation


Major Speciation



Sodium 34 0 n/a 34 Na+ (95%), NaSO4- (4%) n/a 0.28

Magnesium 4.4 0 n/a 4.4 Mg2+ (57%), MgSO4(aq) (34%) n/a 0.11 Aluminum 0.35 0.34 Al(OH)3 (98%) 0.0085 Al(OH)4

- (2%) n/a 0.21 Potassium 7.5 0 n/a 7.5 K+ (95%), KSO4

- (5%) n/a 0.15 Calcium 36 28 CaCO3 (72%), Ca2(HPO4)(OH)2 (6%) 7.8 Ca2+ (11%), CaSO4(aq) (8%) n/a 0.043 Chromium <0.0024 0 n/a <0.0024 Cr4(OH)6

6+ (46%), Cr(OH)2+ (40%) <0.0024 <0.034

Manganese 0.021 0.021 MnO2 (100%) 0 n/a 0 0 Iron <0.12 <0.12 Fe2O3 (100%) 0 n/a 0 0 Cobalt <0.0024 0 n/a <0.0024 CoHCO3

+ (33%), Co2+ (30%) n/a <0.24 Nickel 0.0024 0 n/a 0.0024 NiHCO3

+ (36%), NiCO3 (27%) 0.0048 0.24 Copper <0.0027 0 n/a <0.0027a CuCO3 (69%), Cu(CO3)2

2- (30%) <0.009 <2.7 Zinc <0.012 0 n/a <0.012a ZnCO3 (44%), Zn2+ (22%) <0.012 <1.2 Strontium 0.64 0 n/a 0.64a Sr2+ (53%), SrSO4 (34%) n/a 1.6 Cadmium <0.00048 0 n/a <0.00048 Cd2+ (36%), CdCO3 (28%) <0.0048 <0.096 Mercury <0.000048 0 n/a <0.000048a Hg(OH)2 (72%), Hg(CO3)(OH)- (24%) <0.024 <4.8 Lead <0.0024 0 n/a <0.0024 Pb(CO3) (66%), Pb(CO3)2

2- (30%) <0.004 <0.40 Silver <0.00064 0 n/a <0.00064 Ag+ (72%), Ag(CO3)- (13%) n/a <0.064 Fluoride <0.24 0 n/a <0.24a F- (99.5%), MgF+ (0.4%) n/a <4.8 Arsenic 0.17 0 n/a 0.17a HAsO4

2- (91%), H2AsO4- (8%) 0.17 3.4

Selenium <0.0024 0 n/a <0.0024a SeO42- (99.7%), CaSeO4 (0.2%) n/a <1.2

Molybdenum 0.051 0 n/a 0.051a MoO42- (100%) n/a 51

Phosphate <0.88 <0.84 Ca2(HPO4)(OH)2 (95%) <0.043 HPO42- (4%) n/a <0.22

Sulfate 490 0 n/a 490a SO42- (100%) n/a 4.9

Alkalinity 640 0 n/a 640 n/a n/a n/a a Meets or exceeds Russian Class 1 fisheries limit and requires dilution and/or treatment prior to regulatory point-of-compliance.



Table 2.7: Estimated solid and dissolved phase concentrations, with major speciation contributions, of water quality parameters from exposed saturated NAG rock. Applicable regulatory limits and the corresponding exceedence factors for dissolved phase concentrations are shown for comparison.


Parameter



Major Speciation


Major Speciation



Sodium 16 0 n/a 16 Na+ (97%), NaSO4- (2%) n/a 0.13

Magnesium 2 0 n/a 2 Mg2+ (66%), MgSO4(aq) (27%) n/a 0.050 Aluminum 0.16 0.15 Al(OH)3 (95%) 0.0080 Al(OH)4

- (5%) n/a 0.20 Potassium 3.5 0 n/a 3.5 K+ (97%), KSO4

- (3%) n/a 0.070 Calcium 17 6.2 CaCO3 (31%), Ca2(HPO4)(OH)2 (6%) 11 Ca2+ (38%), CaSO4(aq) (19%) n/a 0.061 Chromium <0.0011 0 n/a <0.0011 CrO4

2- (97%), HCrO4- (3%) <0.0011 <0.016

Manganese 0.0099 0.0099 MnO2 (100%) 0 n/a 0 0 Iron <0.056 <0.056 Fe2O3 (100%) 0 n/a 0 0 Cobalt <0.0011 0 n/a <0.0011 Co2+ (41%), CoHCO3

+ (26%) n/a <0.11 Nickel 0.0011 0 n/a 0.0011 Ni2+ (31%), NiHCO3

+ (31%) 0.0022 0.11 Copper <0.0012 0 n/a <0.0012a CuCO3 (83%), Cu(CO3)2

2- (15%) <0.0040 <1.2 Zinc <0.0056 0 n/a <0.0056 ZnCO3 (38%), Zn2+ (32%) <0.0056 <0.56 Strontium 0.29 0 n/a 0.29 Sr2+ (63%), SrSO4 (28%) n/a 0.73 Cadmium <0.00022 0 n/a <0.00022 Cd2+ (48%), CdCO3 (22%) <0.0022 <0.044 Mercury <0.000022 0 n/a <0.000022a Hg(OH)2 (84%), Hg(CO3)(OH)- (13%) <0.011 <2.2 Lead <0.0011 0 n/a <0.0011 Pb(CO3) (80%), Pb(CO3)2

2- (15%) <0.0018 <0.18 Silver <0.00024 0 n/a <0.00024 Ag+ (83%), Ag(CO3)- (8%) n/a <0.024 Fluoride <0.11 0 n/a <0.11a F- (99.6%), MgF+ (0.2%) n/a <2.2 Arsenic 0.33 0 n/a 0.33a HAsO4

2- (88%), H2AsO4- (9%) 0.33 6.6

Selenium <0.0011 0 n/a <0.0011 SeO42- (99.5%), CaSeO4 (0.4%) n/a <0.56

Molybdenum 0.024 0 n/a 0.024a MoO42- (100%) n/a 24

Phosphate <0.41 <0.40 Ca2(HPO4)(OH)2 (98%) <0.0090 HPO42- (1.8%) n/a <0.045

Sulfate 230 0 n/a 230a SO42- (100%) n/a 2.3

Alkalinity 300 0 n/a 300 n/a n/a n/a a Meets or exceeds Russian Class 1 fisheries limit and requires dilution and/or treatment prior to regulatory point-of-compliance.



Table 2.8: Estimated solid and dissolved phase concentrations, with major speciation contributions, of water quality parameters from exposed unsaturated AG rock. Applicable regulatory limits and the corresponding exceedence factors for dissolved phase concentrations are shown for comparison.


Parameter



Major Speciation


Major Speciation



Sodium 4.8 0 n/a 4.8 NaSO4- (95%), Na+ (4%) n/a 0.040

Magnesium 45 0 n/a 45b MgSO4(aq) (99.93%), Mg2+ (0.07%) n/a 1.1 Aluminum 64 0 n/a 64b AlSO4

+ (99%), AlF2+ (1%) n/a 1,600

Potassium 4.1 0 n/a 4.1 KSO4- (96%), K+ (4%) n/a 0.082

Calcium 500 270 CaSO4·2H2O (55%) 230b CaSO4(aq) (45%), Ca2+ (0.02%), n/a 1.3 Chromium 0.037 0 n/a 0.037 CrSO4

+ (97%), CrSO4OH (3%) 0.037 0.53 Manganese 22 0 n/a 22b MnSO4 (100%) n/a 2200 Iron 440 190 Fe2O3 (42%) 250a,b Fe(II)SO4 (54%), Fe(III)SO4

+ (4%) 125 2,500 Cobalt 0.46 0 n/a 0.46b CoSO4 (100%) n/a 46 Nickel 0.20 0 n/a 0.20b Ni(SO4)2

2- (91%), NiSO4 (9%) 0.40 20 Copper 0.46 0 n/a 0.46a,b CuSO4 (100%) 1.5 460 Zinc 1.7 0 n/a 1.7a,b Zn(SO4)2

2- (94%), ZnSO4 (4%) 1.7 170 Strontium 2.7 0 n/a 2.7b SrSO4 (100%) n/a 6.8 Cadmium 0.0031 0 n/a 0.0031 Cd(SO4)2

2- (86%), Cd(SO4)34- (8%) 0.031 0.62

Mercury <0.000047 0 n/a <0.000047b HgCl2 (94%), Hg(SO4)22- (5%) <0.024 <4.7

Lead <0.0024 0 n/a <0.0024 Pb(SO4)22- (99%), PbSO4 (1%) <0.0040 <0.40

Silver <0.00065 0 n/a <0.00065 AgSO4- (90%), Ag(SO4)2

3- (9%) n/a <0.065 Fluoride 1.5 0 n/a 1.5b AlF2

+ (63%), MgF+ (0.4%) n/a 30 Arsenic 1.2 0 n/a 1.2a,b H2AsO4

- (84%), H3AsO4 (16%) 1.2 24 Selenium 0.013 0 n/a 0.013b HSeO3

- (68%), H2SeO3 (32%) n/a 6.6 Molybdenum 0.052 0 n/a 0.052b H2MoO4 (91%), HMoO4

- (8%) n/a 52 Phosphate 16 0 n/a 16 H2(PO4)- (86%), H3(PO4) (14%) n/a 80 Sulfate 50,100 0 n/a 50,100a SO4

2- (91%), HSO4- (9%) n/a 501

Alkalinity 0 0 n/a 0 n/a n/a n/a a Meets or exceeds direct discharge limit and requires dilution and/or treatment prior to discharge. b Meets or exceeds Russian Class 1 fisheries limit and requires dilution and/or treatment prior to regulatory point-of-compliance.



Table 2.9: Estimated solid and dissolved phase concentrations, with major speciation contributions, of water quality parameters from exposed saturated AG rock. Applicable regulatory limits and the corresponding exceedence factors for dissolved phase concentrations are shown for comparison.


Parameter



Major Speciation


Major Speciation




Magnesium 21 0 n/a 21 MgSO4(aq) (97%), Mg2+ (3%) n/a 0.53 Aluminum 30 0 n/a 30b AlSO4

+ (98%), AlF2+ (1%) n/a 750 Potassium 1.9 0 n/a 1.9 KSO4

- (72%), K+ (28%) n/a 0.038 Calcium 230 0 n/a 230b CaSO4(aq) (98%), Ca2+ (2%), n/a 1.3 Chromium 0.017 0 n/a 0.017 CrSO4

+ (98%), CrSO4OH (2%) 0.017 0.24 Manganese 10 0 n/a 10b MnSO4 (97%), Mn2+ (3%) n/a 1,000 Iron 200 180 Fe2O3 (87%) 27a,b Fe(II)SO4 (13%) 14 270 Cobalt 0.21 0 n/a 0.21b CoSO4 (98%), Co2+ (2%) n/a 21 Nickel 0.093 0 n/a 0.093b Ni(SO4)2

2- (84%), NiSO4 (16%) 0.19 9.3 Copper 0.21 0 n/a 0.21b CuSO4 (98%), Cu2+ (2%) 0.70 210 Zinc 0.81 0 n/a 0.81b Zn(SO4)2

2- (74%), Zn(SO4)34- (20%) 0.81 81

Strontium 1.2 0 n/a 1.2b SrSO4 (98%), Sr2+ (2%) n/a 3.0 Cadmium 0.0014 0 n/a 0.0014 Cd(SO4)2

2- (47%), Cd(SO4)34- (47%) 0.014 0.28

Mercury <0.000022 0 n/a <0.000022b HgCl2 (94%), Hg(SO4)22- (3%) <0.011 <2.2

Lead <0.0011 0 n/a <0.0011 Pb(SO4)22- (98%), PbSO4 (2%) <0.0018 <0.18

Silver <0.00030 0 n/a <0.00030 AgSO4- (73%), Ag(SO4)2

3- (17%), Ag+ (10%) n/a <0.030 Fluoride 0.70 0 n/a 0.70b AlF2

+ (45%), AlF2+ (38%) n/a 14 Arsenic 0.54 0 n/a 0.54b H2AsO4

- (89%), H3AsO4 (11%) 0.54 11 Selenium 0.0061 0 n/a 0.0061b HSeO3

- (75%), H2SeO3 (25%) n/a 3.1 Molybdenum 0.024 0 n/a 0.024b H2MoO4 (87%), HMoO4

- (12%) n/a 24 Phosphate 7.2 0 n/a 7.2b H2(PO4)- (90%), H3(PO4) (10%) n/a 36 Sulfate 23,000 0 n/a 23,000 SO4

2- (94%), HSO4- (5%) n/a 230

Alkalinity 0 0 n/a 0 n/a n/a n/a a Meets or exceeds direct discharge limit and requires dilution and/or treatment prior to discharge. b Meets or exceeds Russian Class 1 fisheries limit and requires dilution and/or treatment prior to regulatory point-of-compliance.



Note that runoff from the AG waste rock stockpile and the exposed AG portions of the pit walls may need to be treated in 2007. No AG rock runoff is expected to require treatment during 2006, or between 2008 through 2014. If necessary, treatment of AG rock runoff in 2007 will be via pebble lime (CaO) application to neutralize the pH to near-neutral values (pH 7-8), followed by mixing and settling in a sedimentation pond with a nominal hydraulic retention time (HRT) of 3-4 days to maximize removal of particulate and colloidal contaminants. Details as to sizing of the AG rock runoff treatment system for the pit inflows and AG waste rock stockpile during 2007 are given in the Water Management Plan (AMEC, 2005a). Locations of the sedimentation ponds for treatment, and ditches for conveying water on-site and to receiving waters, are given in the Site Water Balance (AMEC, 2005b). Assuming that potential treatment of the unsaturated source AG rock runoff (i.e., Table 2.5) that is to occur in 2007 will result in an effluent stream of pH 7, the results of the geochemical modeling under this neutralization scenario are given in Table 2.10. Addition of calcium oxide (pebble lime) at an average rate of 2.5 g/L for neutralization of mineral and proton acidity will increase expected total concentrations of calcium ions in the resulting treated AG runoff by approximately 17,800 mg/L, to a total concentration 18,400 mg/L. The increased concentration of calcium ions in the treated AG runoff has been included in the geochemical modeling approach for this effluent stream.



Table 2.10: Estimated solid and dissolved phase concentrations, with major speciation contributions, of water quality parameters from exposed unsaturated AG rock following pebble lime (CaO) neutralization to pH 7. Applicable regulatory limits and the corresponding exceedence factors for dissolved phase concentrations are shown for comparison.


Parameter



Major Speciation

Concentration(mg/L)

Major Speciation




Magnesium 45 0 n/a 45b MgSO4(aq) (99.7%), Mg2+ (0.3%) n/a 1.1 Aluminum 64 64 Al(OH)3 (86%), Al3(PO4)2(OH)3(H2O)5 (13%) 0.21b AlPO4 (0.32%) n/a 5.3 Potassium 4.1 0 n/a 4.1 KSO4

- (91.2%), K+ (8.8%) n/a 0.082 Calcium 18,400 18,200 Ca(SO4)(H2O)2 (98.8%) 230b CaSO4(aq) (1.2%) n/a 1.3 Chromium 0.037 0 n/a 0.037 CrSO4OH (46%), CrO4

2- (38%), HCrO4- (15%) 0.037 0.53

Manganese 22 20 MnO2 (93%) 1.6b MnSO4 (7.2%) n/a 160 Iron 440 440 Fe2O3 (100%) 0 n/a 0 0 Cobalt 0.46 0 n/a 0.46b CoSO4 (99.7%), Co2+ (0.3%) n/a 46 Nickel 0.20 0 n/a 0.20b Ni(SO4)2

2- (90%), NiSO4 (10%) 0.40 20 Copper 0.46 0 n/a 0.46a,b CuSO4 (99.6%), Cu2+ (0.22%), CuHPO4 (0.13%) 1.5 460 Zinc 1.7 0 n/a 1.7a,b Zn(SO4)2

2- (89%), Zn(SO4)34- (6%), Zn(SO4) (4%) 1.7 170

Strontium 2.7 0 n/a 2.7b SrSO4 (99.7%), Sr2+ (0.3%) n/a 6.8 Cadmium 0.0031 0 n/a 0.0031 Cd(SO4)2

2- (74%), Cd(SO4)34- (20%), Cd(SO4) (6%) 0.031 0.62

Mercury <0.000047 0 n/a <0.000047b Hg(OH)2 (90%), HgCl(OH) (10%) <0.012 <2.4 Lead <0.0024 0 n/a <0.0024 Pb(SO4)2

2- (98.6%), PbSO4 (1.4%) <0.0020 <0.20 Silver <0.00065 0 n/a <0.00065 AgSO4

- (84%), Ag(SO4)23- (13%), Ag+ (3%) n/a <0.033

Fluoride 1.5 1.4 Ca10(PO4)6F2 (96.4%) 0.054b F- (3.6%) n/a 1.1 Arsenic 1.2 0 n/a 1.2a,b H2AsO4

- (55%), HAsO42- (44%) 1.2 24

Selenium 0.013 0 n/a 0.013b SeO42- (80%), HSeO3

- (19%) n/a 6.6 Molybdenum 0.052 0 n/a 0.052b MoO4

2- (99.8%), HMoO4- (0.2%) n/a 52

Phosphate 16 14 Ca10(PO4)6F2 (44%), Al3(PO4)2(OH)3(H2O)5 (40%) 2.5b H2(PO4)- (9.2%), HPO42- (4.7%) n/a 13

Sulfate 50,000 15,000 Ca(SO4)(H2O)2 (29%) 35,000b SO42- (70.4%) n/a 350

Chloride 1.4 0 n/a 1.4 Cl- (100%) n/a 0.0047 a Meets or exceeds direct discharge limit and requires dilution and/or further treatment prior to discharge. b Meets or exceeds Russian Class 1 fisheries limit and requires dilution and/or further treatment prior to regulatory point-of-compliance.



Dissolved phase concentrations from unsaturated and saturated NAG rock, and unsaturated AG rock, were then input to a series of linked mass balance calculations coupled to the site water balance model. Details regarding the site water balance model are provided in the Site Water Balance report (AMEC, 2005b). The conceptual site water system on an annual basis over the project life, indicating linkages between water sources and components, is provided in both the Site Water Balance report (AMEC, 2005a) and the Water Management Plan (AMEC, 2005a). The Site Water Balance report (AMEC, 2005b) also provides estimated flow volumes from, and to, each component of the water management system. The reader is referred to these documents for details on the site water management system. The present report will not discuss the detailed quantitative hydrological aspects of the site water system, or the corresponding derivations. Using this approach, and assuming complete mixing of all water sources prior to the ambient regulatory point-of-compliance on the Kaiemraveem River at a location 500 m downstream of the last potential input from project site runoff, the resulting concentrations of the various water quality parameters of interest were calculated. Further geochemical modeling on the resulting mixed whole-water profile was not conducted, as the mixing of site runoff from the various sources with the background waters of the Kaiemraveem River and its contributing catchments generally resulted in a reduction of concentrations. Thus, saturation compounds would not likely be reached in the Kaiemraveem River and such advanced modeling methods are not warranted. Concentrations of the following parameters were modeled in the Kaiemraveem River at the ambient regulatory point-of-compliance by year: 2006 – sulphate, phosphate, antimony, arsenic, copper, manganese, mercury, and selenium; 2007 – sulphate, calcium, magnesium, phosphate, antimony, arsenic, cobalt, copper, iron, manganese, mercury, nickel, selenium, strontium, and zinc; and 2008-2014 – sulphate, phosphate, antimony, arsenic, copper, manganese, and mercury. The selection criteria for modeling was that an analyte must have an estimated concentration in one of the site source waters (unsaturated or saturated NAG rock runoff, or unsaturated AG rock runoff during 2007) that was expected to either exceed direct discharge limits or would require dilution in receiving waters in order to meet Russian Class 1 fisheries standard limits (i.e., exceeds Russian Class 1 fisheries standard at source). Note that while aluminum in unsaturated AG rock runoff is expected to reach 0.21 mg/L, or 5.3 fold higher than the Russian Class 1 fisheries standard limit of 0.04 mg/l, the concentration of this element cannot be reliably modeled in the Kaiemraveem River due to the lack of analyses for aluminum in background water quality. However, if values of one-half the Russian Class 1 fisheries standard limit for aluminum are used as background and natural surface runoff concentrations (i.e., 0.02 mg/L as source concentration from these regions) in the modeling approach, aluminum concentrations only increase from 0.020 mg/L upstream of the project site to values of 0.0203 and 0.0205 mg/L (both concentrations remain below the Russian Class 1 fisheries standard limit), respectively, for tailings alternatives #1 and #2 at the corresponding ambient regulatory points-of-compliance downstream of the project. These changes to approximated background concentrations of aluminum are 1.5% and 2.5%, respectively, and would not significantly affect baseline aluminum concentrations in this aquatic



system. Thus, regardless of the absence of background aluminum concentrations in the Kaiemraveem River and its contributing catchments, aluminum will not be a water quality parameter of concern over the project life. The small increase in aluminum concentrations (<2.5%) in the Kaiemraveem River downstream of the project will not likely be measurable by conventional analytical methods, and will not pose a risk to water quality. Similarly, note that while fluoride in unsaturated AG rock runoff is expected to reach 0.054 mg/L, or 1.1 fold higher than the Russian Class 1 fisheries standard limit of 0.05 mg/l, the concentration of this element cannot be reliably modeled in the Kaiemraveem River due to the lack of analyses for fluoride in background water quality. However, if values of one-half the Russian Class 1 fisheries standard limit for fluoride are used as background and natural surface runoff concentrations (i.e., 0.025 mg/L as source concentration from these regions) in the modeling approach, fluoride concentrations only increase from 0.025 mg/L upstream of the project site to values of 0.0252 and 0.0253 mg/L (both concentrations remain below the Russian Class 1 fisheries standard limit), respectively, for tailings alternatives #1 and #2 at the corresponding ambient regulatory points-of-compliance downstream of the project. These changes to approximated background concentrations of fluoride are 0.8% and 1.2%, respectively, and would not significantly affect baseline fluoride concentrations in this aquatic system. Thus, regardless of the absence of background fluoride concentrations in the Kaiemraveem River and its contributing catchments, fluoride will not be a water quality parameter of concern over the project life. The small increase in fluoride concentrations (≤1%) in the Kaiemraveem River downstream of the project will not likely be measurable by conventional analytical methods, and will not pose a risk to water quality. Finally, while molybdenum in unsaturated AG rock runoff is expected to reach 0.052 mg/L (52 fold higher than the Russian Class 1 fisheries standard limit of 0.001 mg/l), and in unsaturated NAG rock runoff molybdenum concentrations are expected to reach 0.051 mg/L (51 times the ambient limit), the concentration of this element cannot be reliably modeled in the Kaiemraveem River due to the lack of analyses for molybdenum in background water quality. However, if values of one-half the Russian Class 1 fisheries standard limit for molybdenum are used as background and natural surface runoff concentrations (i.e., 0.0005 mg/L as source concentration from these regions) in the modeling approach, molybdenum concentrations only increase from 0.0005 mg/L upstream of the project site to values of 0.00071 and 0.00082 mg/L (both concentrations remain below the Russian Class 1 fisheries standard limit), respectively, for tailings alternatives #1 and #2 at the corresponding ambient regulatory points-of-compliance downstream of the project. These changes to approximated background concentrations of molybdenum are 42% and 64%, respectively, and based on the assumptions given above, would not result in exceedences of the Russian Class 1 fisheries standards downstream of the project site. The modest increase in molybdenum concentrations (~50%) in the Kaiemraveem River will likely not pose a risk to water quality. The remaining following water quality parameters that were not modeled during any year of the project are not expected to exceed direct discharge or ambient receiving water quality limits at the point of generation, and thus, do not require further consideration: turbidity and total



suspended solids (sedimentation ponds on-site are designed to meet direct discharge and Russian Class 1 fisheries standard; see Site Water Balance (AMEC, 2005b) for design criteria and locations); chloride, potassium, sodium, ammonia, nitrate, nitrite, cyanide, thiocyanate, free cyanide, WAD cyanide, total cyanide, cadmium, chromium, lead, silver, and thallium. The modeling approach has the following limitations which may be addressed in a future series of impact assessment quality models. The model simplifies the hydrology of the Kaiemraveem Creek and its contributing catchments. Complete mixing in the Kaiemraveem River is assumed prior to the Russian Class 1 fisheries standard point-of-compliance, thereby ignoring advection, dispersion, and diffusion effects – along with thermal and/or density based vertical stratification – that may be important in determining the fine structure of spatial and temporal contaminant patterns within local aquatic systems. One- or two-dimension hydrodynamic modeling in the Kaiemraveem River (e.g., using software such as QUAL2E) would be required to investigate, and potentially account for, incomplete mixing processes in receiving waters. A sewage treatment plant (STP) is currently installed at the project site, and will be operated continuously over the project life. The system configuration is as follows: settling chamber → pre-aeration chamber → treatment chamber → pump chamber → filter unit → UV disinfection. The tertiary level nature of the system will provide a high level of physical, chemical, and biological treatment efficiency. By 2006, upgrades to the STP will result in a capacity of 160 m3/d (based on a 600 man camp at 265 L/p/d) at an organic loading rate of 800 mg/L BOD5 (128 kg BOD5/d maximum loading). During two summer month periods during 2006 and 2007, an additional 300 men may increase the camp’s total population to 900 men for a short period of time (e.g., June through October). In such a case, the STP capacity would be increased to 240 m3/d. The peak hourly flow capacity of the STP will be 2.5 times the daily flow rate, or from 400 m3/d (600 man camp) to 600 m3/d (900 man camp). The combined BOD5 loading capacity of the STP is a composite of 250 mg/L BOD5 from domestic residential grade wastewater (40% nominal average of total daily flows) and 1,200 mg/L BOD5 of high-strength food preparation wastewater (60% nominal average of total daily flows). The average influent design capacities are 250 mg/L for suspended solids, 50 mg/L for fats, oils, and grease, and 50 mg/L for total Kjeldahl nitrogen. Effluent characteristics from the STP will be as follows: BOD5≤30 mg/L and TSS≤30 mg/L. No specified effluent criteria for ammonia, nitrate, total nitrogen, fecal coliforms, phosphorus, and dissolved oxygen are available. It is assumed the STP effluent will meet the World Bank direct discharge water quality standards to allow summer discharge direct to receiving water bodies during 2006 and 2007 without dilution, including the following parameters of note with regard to sewage treatment: pH between 6 to 9; BOD5≤50 mg/L; TSS≤50 mg/L; oil and grease≤20 mg/L; and temperature<5°C above ambient temperature of receiving waters (<3°C increase where receiving water temperature is >28°C). Elevated metal concentrations are not anticipated in the domestic wastewater and high-strength food preparation wastewater sources to the STP. From late 2007 through the remainder of the project life, the STP effluent will be pumped to the tailings pond and permanently retained in the



closed-circuit process plant water balance. Thus, direct discharge criteria will not apply to the STP effluent over this period of the project, thereby allowing for potential cost savings by minimizing or eliminating the use of secondary and tertiary processes. STP discharge will be directed into the catchment of the Pervyi Creek near its headwaters for both the summer and winter periods of 2006 through late 2007. Over-winter discharge will remain frozen near the point of discharge. The STP discharge piping will not be heat-traced. At a peak over-winter camp size of 600 men, and assuming a 212 day ice-covered season (where nearby creeks do not flow) from November through May, approximately 32,000 m3 of frozen treated sewage will accumulate near the STP discharge point. This frozen effluent will thaw over the spring/early summer freshet and achieve maximum dilution during this time. Summer STP discharge when rivers are flowing will be at the same location as with the over-winter discharge regime. From late 2007 through the remainder of the project life, the STP effluent will be pumped directly to the tailings pond and permanently retained in the closed-circuit process plant water balance. The lack of effluent specification for key nutrient parameters such as phosphate, ammonia, nitrate, and nitrite, for which Russian Class 1 fisheries standard of 0.2 mg/L, 0.5 mg/L, 40 mg/L, and 0.08 mg/L, respectively, exist, prevents inclusion of STP effluent into the current modeling approach. Once this data is available, a revised model should be constructed to ensure that the STP effluent will not adversely affect nutrient concentrations in the Kaiemraveem river at the regulatory point-of-compliance. The total annual volume of STP effluent discharge that reaches the Kaiemraveem River is expected to be <87,000 m3 in 2006/2007. With an annual throughflow of 23,000,000 m3 during an average year in the Kaiemraveem River at the ambient point-of-compliance, the STP will be diluted by, on average, a factor of >260 prior to reaching the point-of-compliance. Thus, adequate dilution, and the extended residence time in the river during its about 10 km journey (along the natural watercourse in Pervyi Creek and Kaiemraveem River) from the STP discharge point to the ambient point-of-compliance, will likely reduce the concentrations of physical, chemical, and biological parameters in the STP effluent to levels below concern. Note that potable drinking water is also intended to be obtained from riverbank filtration wells immediately adjacent to the Kaiemraveem River near the point-of-compliance for tailings alternative #2. However, for both tailings alternatives, the potable drinking water wells are located on the Kaiemraveem River upstream of where runoff/seepage from the tailings dam would enter the Kaiemraveem River. Thus, no impacts on drinking water quality are expected due to potential seepage from the lined tailings impoundment. In addition, the dilution and travel time in the Kaiemraveem River will, in combination with the filtering effects of riverbank wells, act to minimize any potential risk to drinking water quality from upstream STP discharges in 2007. The mass balance model in the Kaiemraveem River (using geochemical modeling based inputs) also ignores biogeochemical reactions that may degrade compounds and enhance or decrease the solubility and/or bioavailability of water quality parameters within the river itself.



Chemical/biological reactivity with water quality parameters cannot readily be input into such a simplified receiving water mass balance model, and more advanced modeling techniques are required (e.g., use of software such as QUAL2E as mentioned above). This lack of accounting for biogeochemical reactivity will result in potentially overestimated concentrations of redox active elements such as iron and manganese, which will not likely pose a significant resulting water quality problem in the oxygenated conditions of surface water bodies such the Kaiemraveem River. Concentrations of nutrients such as nitrogen, phosphorus, and carbon (and the various species that make up total concentrations of these parameters) cannot be readily modeled without first constructing a full limnological model of the system that includes biological cycling of nutrients. The minimal nutrient concentrations in STP effluent released to the Kaiemraveem River in 2007 will not likely lead to trophic changes in this system. For conservative (unreactive) parameters such as chloride, fluoride, and sodium (and total dissolved solids (TDS)) – where adsorption, precipitation, and biological uptake/removal mechanisms are usually negligible – the mass balance modeling technique will likely accurately capture the lack of inherent reactivity for such species. As such, the highest likely source of error in modeling conservative elements via the methods used in the current study lies in not accounting for incomplete mixing and storage within water bodies. In addition, the model is not calibrated, or able to be calibrated. This is a flaw in any modeling approach, as the reality of various simplifying assumptions used to construct the model cannot be tested using available baseline training sets. For a model with the ability to be calibrated (i.e., run to a baseline steady-state and compared to observed baseline steady-state conditions), more detailed site hydrology and water quality information would be required. Inclusion of this data in the modeling approach would likely present significant time and cost barriers. The length of the modeling period is 14 years and is consistent with site activities (e.g., open pit mining from 2006 through 2010, underground mining from 2008 through 2014, construction of the starter tailings dam in late 2007, etc.) as set out in the Water Management Plan (AMEC, 2005a) and Site Water Balance (AMEC, 2005b). The reader is referred to these documents for further details on the mine plan and project hydrology, water linkages, flow rates, and volumes.



3.0 RESULTS AND DISCUSSION 3.1 Direct Discharge of Runoff to Receiving Waters The mass balance and geochemical modeling results indicate there are no expected World Bank direct discharge limit (World Bank, 1999a and 1999b) exceedences for runoff from either unsaturated or saturated NAG rock exposed on-site. Thus, runoff from haul and access roads, the plant and camp site, the outside crest of the tailings dam, the airstrip, and the explosives storage facility will be suitable for direct discharge to ambient receiving waters. A similar modeling approach was undertaken on the unsaturated and saturated AG rock runoff, as well as on the unsaturated AG rock runoff following pebble lime (CaO) neutralization to pH 7. The results for unsaturated AG rock runoff prior to treatment suggest direct discharge exceedences (with predicted exceedence factors in parentheses) for the following analytes: pH, iron (125), copper (1.5), zinc (1.7), and arsenic (1.2). In the saturated AG rock runoff, pH and iron (with an exceedence factor of 14) are expected to exceed the World Bank direct discharge limits (World Bank, 1999a and 1999b). The difference in the number and extent of direct discharge exceedences between the unsaturated and saturated AG rock runoff illustrates the strong dependence of water quality on quantity of runoff and infiltration from AG rock surfaces, and also indicates the low exceedence factors for copper, zinc, and arsenic in unsaturated AG runoff. Following lime neutralization of the unsaturated AG rock runoff, as set out in the Water Management Plan (AMEC, 2005a), the geochemical modeling approach predicts that copper, zinc, and arsenic will have residual dissolved concentrations above the World Bank direct discharge limits (World Bank, 1999a and 1999b). Predicted exceedence factors for these three parameters using the geochemical modeling technique are as follows: copper (1.5), zinc (1.7), and arsenic (1.2). However, while the geochemical modeling method accounts for equilibrium-based speciation and saturation solids in an oxygenated environment, the technique does not account for secondary adsorption scavenging by saturation solids. This adsorption scavenging by precipitates such as hematite (Fe(OH)3), doyleite (Al(OH)3), gypsum (Ca(SO4)(H2O)2), and pyrolusite (MnO2) will also likely reduce residual concentrations of arsenic, copper, and zinc to below the World Bank direct discharge limits (World Bank, 1999a and 1999b). If field experience on mine start-up indicates lime neutralization and subsequent secondary mineral precipitation does not sufficiently lower concentrations of these analytes below World Bank direct discharge limits (World Bank, 1999a and 1999b), additional make-up water from surrounding regions generating natural runoff could be used to dilute the resultant treated AG runoff stream to a level that meets the World Bank direct discharge limits. The locations of the treatment/sedimentation ponds for the AG waste rock stockpile runoff and the pit inflows in 2007 are near the points of runoff generation (see Site Water Balance (AMEC, 2005b) for details and location), thereby facilitating the collection, transfer, and mixing of dilution water with treated AG rock runoff. Note that only in runoff from the AG waste rock stockpile will likely generate water quality having the potential to exceed World Bank direct discharge limits after treatment, thereby requiring



dilution prior to reaching receiving waters. The pit wall runoff will be an approximately equal volumetric mixed composite of unsaturated AG and NAG rock runoff (as the pit wall modeling indicates approximately equal portions of these geochemical types). Thus, no parameters are expected to exceed World Bank direct discharge limits in treated pit inflows. For comparison with the treated AG waste rock stockpile runoff, concentrations (with exceedence factors in parentheses) of copper, zinc, and arsenic in the treated pit inflows are expected to reach 0.23 mg/L (0.77), 0.85 mg/L (0.85), and 0.96 mg/L (0.96). Similarly, the ore stockpile runoff will be a composite mix (approximately 80:20 on a volumetric basis) of ore and unsaturated AG runoff, respectively, due to a predicted dilution of 19% in the open pit (Bema, 2004). The single humidity cell test on an ore sample suggests that water quality in runoff from the ore will be significantly better that is expected from the AG rock runoff. For example, comparative production rates of copper, zinc, and arsenic from the AG waste runoff and ore are as follows: copper (AG, 0.0029 mg/kg/wk; ore, <0.000063 mg/kg/wk); zinc (AG, 0.011 mg/kg/wk; ore, <0.000063 mg/kg/wk); and arsenic (AG, 0.0074 mg/kg/wk; ore, 0.00044 mg/kg/wk). With production rates from 46-175 fold higher by element in the AG rock humidity cells versus the ore, and with the minimal direct discharge exceedences expected in treated AG rock runoff for these parameters (1.2-1.7), an 80:20 volumetric mix of ore:AG rock runoff (as is expected from the ore stockpile) would readily meet all direct discharge criteria following lime neutralization and settling. Thus, there are no direct discharge concerns under the World Bank limits for the ore stockpile runoff following treatment.



3.2 Ambient Receiving Water Quality of Site Discharges 3.2.1 Tailings Alternative #1 3.2.1.1 General Under tailings alternative #1, the tailings storage facility will be located in the headwaters of Tretii Creek. The regulatory Russian Class 1 fisheries standard point-of-compliance under this site plan is 500 m downstream of the confluence of Tretii Creek and Kaiemraveem River (Figure 1.1). At this location, the total upstream natural surface catchment area of Kaiemraveem River is about 121 km2. Assuming an average annual unit runoff of 190 mm for natural surfaces (AMEC, 2005b), the average annual volume of water flowing in Kaiemraveem River at the Russian Class 1 fisheries standard point-of-compliance is 23.0 Mm3. In an average year, the peak flow rate in the Kaiemraveem River near the project site is estimated at 80 m3/s (6.9 Mm3/d), with a 95th percentile minimum flow rate of 0.13 m3/s (0.011 Mm3/d) and 95th percentile maximum flow rate of 90 m3/s (7.8 Mm3/d) (Bema, 2005). A factor of 690 exists between the 95th percentile maximum and minimum estimated annual flow rates. Due to the lack of a hydrologic monitoring station on the Kaiemraveem River or its tributaries, the annual hydrographs cannot be incorporated into the current surface water quality modeling approach. Instead, the estimated water quality in the Kaiemraveem River at the ambient point-of-compliance will be examined under annual average flow conditions, and assuming that relative runoff generation patterns are uniform over the course of the year from all land use types. The airstrip runoff will be discharged in the Starichnaya River following sedimentation in a settling pond. In an average year, the airstrip will generate about 170,000 m3 of runoff (AMEC, 2005b). It is assumed the airstrip access road from the plant site will contribute an approximately equal quantity of annual runoff into the Starichnaya River. The contributing catchment area of the Starichnaya River upstream of the airstrip is 185 km2 (Bema, 2005). Assuming an average annual unit runoff of 190 mm for natural surfaces (AMEC, 2005b), the average annual volume of water flowing in Starichnaya River at the Russian Class 1 fisheries standard point-of-compliance located 500 m downstream of the airstrip is 35.3 Mm3. Thus, the airstrip and access road runoff will be diluted by a factor of about 100 prior to the Starichnaya River ambient regulatory point-of-compliance. No ambient receiving water exceedence factors for saturated or unsaturated NAG rock (assumed to be geochemically equivalent to the airstrip construction material) are greater than 51 (for molybdenum, all other exceedence factors are <4), and thus, the airstrip and access road runoff will be sufficiently diluted in the Starichnaya River such that no exceedences of ambient water quality criteria will occur. Direct discharge limits discussed above do not depend on flow rates in receiving waters, and are thus insensitive to the hydrologic assumptions beyond end-of-pipe considerations. As noted above, no direct discharge water quality issues are expected from unsaturated and saturated exposed NAG or AG rock over the project life.



3.2.1.2 Calendar Year 2006 During 2006, runoff will occur from the following site facilities: plant and camp site, haul and access roads, the airstrip, and the explosives storage facility. All runoff from each of these facilities is assumed to be from unsaturated NAG rock or its geochemical and hydrological equivalent. Concentrations of the eight water quality parameters whose levels in unsaturated NAG rock runoff are above the Russian Class 1 fisheries standards, and which must be diluted prior to the regulatory point-of-compliance in the Kaiemraveem River, are given in Table 3.1. All other water quality parameters not listed in Table 3.1 are not expected to have concentrations exceeding Russian Class 1 fisheries standards at the point of generation, and thus do not require extensive modeling efforts or discussion.

Table 3.1: Predicted average annual concentrations in the Kaiemraveem River during 2006 at the ambient point-of-compliance for tailings alternative #1.

Parameter Background Concentration in Kaiemraveem River (mg/L)

Predicted Concentration in Kaiemraveem River (mg/L)

Ambient Freshwater Regulatory Limit (mg/L)

Exceedence Factor

Sulphate 10 16 100 0.16 Phosphate <0.05a <0.05 (<0.025)b 0.2 <0.13c Antimony <0.01a <0.01 (<0.0051)b 0.005 <1.0 (<1.02)c Arsenic <0.05a <0.05 (<0.026)b 0.05 <0.52c Copper 0.0029 0.0020d 0.001 1.95d Manganese 0.019 0.011d 0.01 1.12d Mercury <0.0015a <0.0015 (<0.00075)b 0.00001 <1.0 (<75)c Selenium <0.02a <0.02 (<0.010)b 0.002 <1.0 (<4.98)c

a Average background concentration is below MDL as shown. Background concentration in Kaiemraveem River was assumed to equal one-half the MDL for modeling purposes. b Modeled concentration is presented in parentheses. Modeled concentration is below MDL for background concentration in Kaiemraveem River because the one-half MDL value for background concentration was used in the modeling efforts. c Background concentration in Kaiemraveem River is below the MDL value. Thus, modeled concentrations and their corresponding exceedence factors represent upper estimates. Where predicted concentrations are below MDL values for background levels, exceedence factors (calculated based and presented in parentheses) should be assumed as <1. d Background concentration for this parameter exceeds Russian Class 1 fisheries standard and project does not lead to increase in concentration of this analyte. Formal regulatory exceedence does not occur. Based on the modeling results, concentrations of sulfate, phosphate, and arsenic are not expected to exceed Russian Class 1 fisheries standards at the regulatory point-of-compliance in the Kaiemraveem River. Copper and manganese are predicted to have concentrations greater than the Russian Class 1 fisheries standards at the point-of-compliance, with corresponding exceedence factors of 1.9



and 1.1, respectively. However, the predicted copper and manganese concentrations in the Kaiemraveem River during 2006 are below background concentrations for these parameters. Thus, natural concentrations of copper and manganese in the Kaiemraveem River already exceed Russian Class 1 fisheries standards prior to project development. Formal regulatory exceedences do not occur where the concentration of an analyte is above a regulatory limit under background/baseline conditions, and where project development lowers the analyte’s concentrations in receiving waters relative to background conditions. For antimony, mercury, and selenium, MDLs for background concentrations in the Kaiemraveem River, its contributing catchments, and natural surface runoff were above the corresponding Russian Class 1 fisheries standard. Reliable predictions of resultant concentrations in the Kaiemraveem River at the regulatory point-of-compliance thus cannot be made. Where MDLs for background conditions exceed regulatory limits, the impacts of a project on concentrations of these parameters cannot be assessed. For antimony, mercury, and selenium, concentrations at the point-of-compliance in the Kaiemraveem River should be considered to be at, or below, the corresponding regulatory limits. Based on the results and reasoning presented above, there are no expected regulatory exceedences of the Russian Class 1 fisheries standards at the point-of-compliance in the Kaiemraveem River during 2006 under tailings alternative #1. 3.2.1.3 Calendar Year 2007 During 2007, runoff will occur from the following site facilities: plant and camp site, haul and access roads, the airstrip, the explosives storage facility, the AG waste rock stockpile, the ore stockpile, and pit inflows. Runoff from all site facilities except the AG waste rock stockpile, the ore stockpile, and pit inflows is assumed to be from saturated NAG rock or its geochemical and hydrological equivalent. Runoff from the AG waste rock stockpile and ore stockpile was assumed to be from unsaturated AG rock, while pit inflows were assumed to be from an equal volumetric mixture of unsaturated AG and NAG rock or their geochemical and hydrological equivalents. Limited kinetic testing has been initiated on representative samples of the ore, with only a single humidity cell test conducted for a 20 week duration (Atkin, 2005). In addition, the Preliminary Economic Assessment (Bema, 2004) suggests an average dilution in the final pit model of 19%. Given the structural geology of the site (Bema, 2004; Rhys, 2004), dilutional waste rock near the ore vein is expected to be dominantly AG material. Since the limited humidity cell testing on the ore suggests that AG waste rock will have significantly poorer water quality (both with regard to depressed pH values and elevated metals concentrations), for conservative ambient water quality modeling purposes, unsaturated AG rock runoff was used as a surrogate for a composite 80:20 split of ore:AG rock runoff from the ore stockpile during 2007. Note that despite this conservative geochemical and water quality approach for the ore stockpile, the low annual runoff volume from this source in 2007 (about 9,000 m3) compared to pit inflows (about 100,000 m3, of which 50% is from AG rock) and AG waste rock stockpile runoff (about 45,000 m3) over the same period means that the conservative modeling approach



for the ore stockpile does not significantly affect predicted water quality in the Kaiemraveem River, particularly at the ambient point-of-compliance. Any predicted ambient regulatory exceedences are not avoided if a more refined (and less conservative) approach towards estimating the ore stockpile runoff is taken. Concentrations of the 15 water quality parameters whose levels in saturated NAG rock runoff and/or unsaturated AG rock runoff are above Russian Class 1 fisheries standards, and which must be diluted prior to the regulatory point-of-compliance in the Kaiemraveem River, are given in Table 3.2. All other water quality parameters not listed in Table 3.2 are not expected to have concentrations exceeding Russian Class 1 fisheries standards at the point of generation, and thus do not require extensive modeling efforts or discussion.





Exceedence Factor

Sulphate 10 48 100 0.48 Calcium 4.0 6.1 180 0.034 Magnesium 1.4 2.3 40 0.058 Phosphate <0.05a <0.05 (<0.027)b 0.2 <0.14c Antimony <0.01a <0.01 (<0.0052)b 0.005 <1.0 (<1.05)c Arsenic <0.05a <0.05 (<0.028)b 0.05 <0.56c Cobalt <0.01a <0.01 (<0.0054)b 0.01 <0.54c Copper 0.0029 0.0024d 0.001 2.37d Iron 0.59 0.29d 0.1 2.91d Manganese 0.019 0.013d 0.01 1.26d Mercury <0.0015a <0.0015 (<0.00075)b 0.00001 <1.0 (<75)c Nickel <0.01a <0.01 (<0.0052)b 0.01 <0.52c Selenium <0.02a <0.02 (<0.010)b 0.002 <1.0 (<4.98)c Strontium <0.05a <0.05 (<0.029)b 0.4 <0.07c Zinc 0.008 0.011 0.01 1.12

a Average background concentration is below MDL as shown. Background concentration in Kaiemraveem River was assumed to equal one-half the MDL for modeling purposes. b Modeled concentration is presented in parentheses. Modeled concentration is below MDL for background concentration in Kaiemraveem River because the one-half MDL value for background concentration was used in the modeling efforts. c Background concentration in Kaiemraveem River is below the MDL value. Thus, modeled concentrations and their corresponding exceedence factors represent upper estimates. Where predicted concentrations are below MDL values for background levels, exceedence factors (calculated based and presented in parentheses) should be assumed as <1. d Background concentration for this parameter exceeds Russian Class 1 fisheries standards and project does not lead to increase in concentration of this analyte. Formal regulatory exceedence does not occur.



Based on the modeling results, concentrations of sulfate, calcium, magnesium, phosphate, arsenic, cobalt, nickel, and strontium are not expected to exceed Russian Class 1 fisheries standard at the regulatory point-of-compliance in the Kaiemraveem River. Copper, iron, manganese, and zinc are predicted to have concentrations greater than the Russian Class 1 fisheries standards at the point-of-compliance, with corresponding exceedence factors of 2.4, 2.9, 1.3, and 1.1 respectively. However, the predicted copper, iron, and manganese concentrations in the Kaiemraveem River during 2007 are below background concentrations for these parameters. Thus, natural concentrations of copper, iron, and manganese in the Kaiemraveem already exceed Russian Class 1 fisheries standards prior to project development. Formal regulatory exceedences do not occur where the concentration of an analyte is above a regulatory limit under background/baseline conditions, and where project development lowers the analyte’s concentrations in receiving waters relative to background conditions. The concentration of zinc at the ambient point-of-compliance is expected to exceed the Russian Class 1 fisheries limit by about 12%. The short duration of this exceedence (i.e., ≤4-5 months over a 14 year project life), the low exceedence factor (1.1), and the generally conservative regulatory levels for zinc in freshwaters suggests that the short-term elevated zinc level in the Kaiemraveem River will not pose a significant threat to aquatic life. In addition, concentrations of zinc downstream of the project site will decline below the Russian Class 1 fisheries limit due to influxes of natural surface runoff. Thus, the elevated zinc levels in 2007 are limited in time and extent. As discussed above, for the purposes of modeling, zinc concentrations in treated AG rock runoff were not adjusted downwards to account for potential adsorptive scavenging by particulate and colloidal phases in the treatment settling ponds. Thus, zinc concentrations from treated AG rock runoff represent conservative upper-bound estimates. It is likely that adsorptive scavenging of zinc in the AG rock treatment system will remove sufficient quantities of this parameter such that Russian Class 1 fisheries standards will be met at the regulatory point-of-compliance. In the event that regulatory agencies do not accept a reliance on adsorptive scavenging as a means of meeting Russian Class 1 fisheries standards, consideration should be given towards developing a site-specific Russian Class 1 fisheries standard limit for zinc. The minimal predicted exceedence factor for zinc (1.1), and the limited duration and extent of elevated zinc levels, would strongly favour the granting of a short-term site-specific water quality guideline for this parameter. For antimony, mercury, and selenium, MDLs for background concentrations in the Kaiemraveem River, its contributing catchments, and natural surface runoff were above the corresponding Russian Class 1 fisheries standards. Reliable predictions of resultant concentrations in the Kaiemraveem River at the regulatory point-of-compliance thus cannot be made. Where MDLs for background conditions exceed Russian Class 1 fisheries standards, the impacts of a project on concentrations of these parameters cannot be assessed. For antimony, mercury, and selenium, concentrations at the point-of-compliance in the Kaiemraveem River should be considered to be at, or below, the corresponding Russian Class 1 fisheries standards.



Based on the results and reasoning presented above, there is only a single expected regulatory exceedence of the Russian Class 1 fisheries standards at the point-of-compliance in the Kaiemraveem River during 2007 under tailings alternative #1. Zinc concentrations may exceed the Russian Class 1 fisheries standard by a factor of 1.1. 3.2.1.4 Calendar Years 2008-2014 Over the period from 2008 through 2014, runoff will occur from the following site facilities: plant and camp site, haul and access roads, the airstrip, the explosives storage facility, and the downstream crest of the tailings dam. All runoff from each of these facilities is assumed to be from saturated NAG rock or its geochemical and hydrological equivalent. Concentrations of the seven water quality parameters whose levels in saturated NAG rock runoff are above Russian Class 1 fisheries standards, and which must be diluted prior to the regulatory point-of-compliance in the Kaiemraveem River, are given in Table 3.3. All other water quality parameters not listed in Table 3.3 are not expected to have concentrations exceeding Russian Class 1 fisheries standards at the point of generation, and thus do not require extensive modeling efforts or discussion.

Table 3.3: Predicted average annual concentrations in the Kaiemraveem River during 2008 through 2014 at the ambient point-of-compliance for tailings alternative #1.




Exceedence Factor

Sulphate 10 16 100 0.16 Phosphate <0.05a <0.05 (<0.025)b 0.2 <0.13c Antimony <0.01a <0.01 (0.0051)b 0.005 <1.0 (<1.03)c Arsenic <0.05a <0.05 (0.026)b 0.05 <0.53c Copper 0.0029 0.0019b 0.001 1.95d Manganese 0.019 0.011d 0.01 1.1d Mercury <0.0015a <0.0015 (<0.00075)b 0.00001 <1.0 (<75)c

a Average background concentration is below MDL as shown. Background concentration in Kaiemraveem River was assumed to equal one-half the MDL for modeling purposes. b Modeled concentration is presented in parentheses. Modeled concentration is below MDL for background concentration in Kaiemraveem River because the one-half MDL value for background concentration was used in the modeling efforts. c Background concentration in Kaiemraveem River is below the MDL value. Thus, modeled concentrations and their corresponding exceedence factors represent upper estimates. Where predicted concentrations are below MDL values for background levels, exceedence factors (calculated based and presented in parentheses) should be assumed as <1. d Background concentration for this parameter exceeds Russian Class 1 fisheries standards and project does not lead to increase in concentration of this analyte. Formal regulatory exceedence does not occur.



Based on the modeling results, concentrations of sulfate, phosphate, and arsenic are not expected to exceed Russian Class 1 fisheries standards at the regulatory point-of-compliance in the Kaiemraveem River. Copper and manganese are predicted to have concentrations greater than the Russian Class 1 fisheries standards at the point-of-compliance, with corresponding exceedence factors of 1.9 and 1.1, respectively. However, the predicted copper and manganese concentrations in the Kaiemraveem River over the period from 2008 through 2014 are below background concentrations for these parameters. Thus, natural concentrations of copper and manganese in the Kaiemraveem River already exceed Russian Class 1 fisheries standards prior to project development. Formal regulatory exceedences do not occur where the concentration of an analyte is above a regulatory limit under background/baseline conditions, and where project development lowers the analyte’s concentrations in receiving waters relative to background conditions. For antimony and mercury, MDLs for background concentrations in the Kaiemraveem River, its contributing catchments, and natural surface runoff were above the corresponding Russian Class 1 fisheries standards. Reliable predictions of resultant concentrations in the Kaiemraveem River at the regulatory point-of-compliance thus cannot be made. Where MDLs for background conditions exceed Russian Class 1 fisheries standards, the impacts of a project on concentrations of these parameters cannot be assessed. For antimony and mercury, concentrations at the point-of-compliance in the Kaiemraveem River should be considered to be at, or below, the corresponding Russian Class 1 fisheries standards. Based on the results and reasoning presented above, there are no expected regulatory exceedences of the Russian Class 1 fisheries standards at the point-of-compliance in the Kaiemraveem River over the period from 2008 through 2014 under tailings alternative #1.



3.2.2 Tailings Alternative #2 3.2.2.1 General Under tailings alternative #2, the tailings storage facility will be located in the headwaters of Vtoroi Creek. The regulatory Russian Class 1 fisheries standard point-of-compliance under this site plan is 500 m downstream of the confluence of Vtoroi Creek and Kaiemraveem River (Figure 1.1). At this location, the total upstream natural surface catchment area of Kaiemraveem River is about 79 km2. Assuming an average annual unit runoff of 190 mm for natural surfaces (AMEC, 2005b), the average annual volume of water flowing in Kaiemraveem River at the Russian Class 1 fisheries standard point-of-compliance is 15.0 Mm3. In an average year, the peak flow rate in the Kaiemraveem near the project site is estimated at 80 m3/s (6.9 Mm3/d), with a 95th percentile minimum flow rate of 0.13 m3/s (0.011 Mm3/d) and 95th percentile maximum flow rate of 90 m3/s (7.8 Mm3/d) (Bema, 2005). A factor of 690 exists between the 95th percentile maximum and minimum estimated annual flow rates. Due to the lack of a hydrologic monitoring station on the Kaiemraveem River or its tributaries, the annual hydrographs cannot be incorporated into the current surface water quality modeling approach. Instead, the estimated water quality in the Kaiemraveem River at the ambient point-of-compliance will be examined under annual average flow conditions. The airstrip runoff will be discharged in the Starichnaya River following sedimentation in a settling pond. In an average year, the airstrip will generate about 170,000 m3 of runoff (AMEC, 2005b). It is assumed the airstrip access road from the plant site will contribute an approximately equal quantity of annual runoff into the Starichnaya River. The contributing catchment area of the Starichnaya River upstream of the airstrip is 185 km2 (Bema, 2005). Assuming an average annual unit runoff of 190 mm for natural surfaces (AMEC, 2005b), the average annual volume of water flowing in Starichnaya River at the Russian Class 1 fisheries standard point-of-compliance located 500 m downstream of the airstrip is 35.3 Mm3. Thus, the airstrip and access road runoff will be diluted by a factor of about 100 prior to the Starichnaya River ambient regulatory point-of-compliance. No ambient receiving water exceedence factors for saturated or unsaturated NAG rock (assumed to be geochemically equivalent to the airstrip construction material) are greater than 51 (for molybdenum, all other exceedence factors are <4), and thus, the airstrip and access road runoff will be sufficiently diluted in the Starichnaya River such that no exceedences of ambient water quality criteria will occur. Direct discharge limits discussed above do not depend on flow rates in receiving waters, and are thus insensitive to the hydrologic assumptions beyond end-of-pipe considerations. As noted above, no direct discharge water quality issues are expected from unsaturated and saturated exposed NAG or AG rock over the project life. 3.2.2.2 Calendar Year 2006 During 2006, runoff will occur from the following site facilities: plant and camp site, haul and access roads, the airstrip, and the explosives storage facility. All runoff from each of these facilities is assumed to be from unsaturated NAG rock or its geochemical and hydrological equivalent. Concentrations of the eight water quality parameters whose levels in unsaturated



NAG rock runoff are above Russian Class 1 fisheries standards, and which must be diluted prior to the regulatory point-of-compliance in the Kaiemraveem River, are given in Table 3.4. All other water quality parameters not listed in Table 3.4 are not expected to have concentrations exceeding Russian Class 1 fisheries standards at the point of generation, and thus do not require extensive modeling efforts or discussion.





Exceedence Factor

Sulphate 10 14 100 0.14 Phosphate <0.05a <0.05 (<0.025)b 0.2 <0.13c Antimony <0.01a <0.01 (<0.0051)b 0.005 <1.0 (<1.03)c Arsenic <0.05a <0.05 (<0.026)b 0.05 <0.53c Copper 0.0029 0.0024d 0.001 2.34d Manganese 0.019 0.015d 0.01 1.45d Mercury <0.0015a <0.0015 (<0.00074)b 0.00001 <1.0 (<74)c Selenium <0.02a <0.02 (<0.010)b 0.002 <1.0 (<4.96)c

a Average background concentration is below MDL as shown. Background concentration in Kaiemraveem River was assumed to equal one-half the MDL for modeling purposes. b Modeled concentration is presented in parentheses. Modeled concentration is below MDL for background concentration in Kaiemraveem River because the one-half MDL value for background concentration was used in the modeling efforts. c Background concentration in Kaiemraveem River is below the MDL value. Thus, modeled concentrations and their corresponding exceedence factors represent upper estimates. Where predicted concentrations are below MDL values for background levels, exceedence factors (calculated based and presented in parentheses) should be assumed as <1. d Background concentration for this parameter exceeds Russian Class 1 fisheries standards and project does not lead to increase in concentration of this analyte. Formal regulatory exceedence does not occur. Based on the modeling results, concentrations of sulfate, phosphate, and arsenic are not expected to exceed Russian Class 1 fisheries standards at the regulatory point-of-compliance in the Kaiemraveem River. Copper and manganese are predicted to have concentrations greater than the Russian Class 1 fisheries standards at the point-of-compliance, with corresponding exceedence factors of 2.3 and 1.5, respectively. However, the predicted copper and manganese concentrations in the Kaiemraveem River during 2006 are below background concentrations for these parameters. Thus, natural concentrations of copper and manganese in the Kaiemraveem River already exceed Russian Class 1 fisheries standards prior to project development. Formal regulatory exceedences do not occur where the concentration of an analyte is above a regulatory limit under background/baseline conditions, and where project development lowers the analyte’s concentrations in receiving waters relative to background conditions.



For antimony, mercury, and selenium, MDLs for background concentrations in the Kaiemraveem River, its contributing catchments, and natural surface runoff were above the corresponding Russian Class 1 fisheries standards. Reliable predictions of resultant concentrations in the Kaiemraveem River at the regulatory point-of-compliance thus cannot be made. Where MDLs for background conditions exceed regulatory limits, the impacts of a project on concentrations of these parameters cannot be assessed. For antimony, mercury, and selenium, concentrations at the point-of-compliance in the Kaiemraveem River should be considered to be at, or below, the corresponding Russian Class 1 fisheries standards. Based on the results and reasoning presented above, there are no expected regulatory exceedences of the Russian Class 1 fisheries standards at the point-of-compliance in the Kaiemraveem River during 2006 under tailings alternative #2. 3.2.2.3 Calendar Year 2007 During 2007, runoff will occur from the following site facilities: plant and camp site, haul and access roads, the airstrip, the explosives storage facility, the AG waste rock stockpile, the ore stockpile, and pit inflows. Runoff from all site facilities except the AG waste rock stockpile, the ore stockpile, and pit inflows is assumed to be from saturated NAG rock or its geochemical and hydrological equivalent. Runoff from the AG waste rock stockpile and ore stockpile was assumed to be from unsaturated AG rock, while pit inflows were assumed to be from an equal volumetric mixture of unsaturated AG and NAG rock or their geochemical and hydrological equivalents. Limited kinetic testing has been initiated on representative samples of the ore, with only a single humidity cell test conducted for a 20 week duration (Atkin, 2005). In addition, the Preliminary Economic Assessment (Bema, 2004) suggests an average dilution in the final pit model of 19%. Given the structural geology of the site (Bema, 2004; Rhys, 2004), dilutional waste rock near the ore vein is expected to be dominantly AG material. Since the limited humidity cell testing on the ore suggests that AG waste rock will have significantly poorer water quality (both with regard to depressed pH values and elevated metals concentrations), for conservative ambient water quality modeling purposes, unsaturated AG rock runoff was used as a surrogate for a composite 80:20 split of ore:AG rock runoff from the ore stockpile during 2007. Note that despite this conservative geochemical and water quality approach for the ore stockpile, the low annual runoff volume from this source in 2007 (about 9,000 m3) compared to pit inflows (about 100,000 m3, of which 50% is from AG rock) and AG waste rock stockpile runoff (about 45,000 m3) over the same period means that the conservative modeling approach for the ore stockpile does not significantly affect predicted water quality in the Kaiemraveem River, particularly at the ambient point-of-compliance. Any predicted Russian Class 1 fisheries standard exceedences are not avoided if a more refined (and less conservative) approach towards estimating the ore stockpile runoff is taken. Concentrations of the 15 water quality parameters whose levels in saturated NAG rock runoff and/or unsaturated AG rock runoff are above Russian Class 1 fisheries standards, and which



must be diluted prior to the regulatory point-of-compliance in the Kaiemraveem River, are given in Table 3.5. All other water quality parameters not listed in Table 3.5 are not expected to have concentrations exceeding Russian Class 1 fisheries standards at the point of generation, and thus do not require extensive modeling efforts or discussion.





Exceedence Factor

Sulphate 10 63 100 0.63 Calcium 4.0 5.3 180 0.029 Magnesium 1.4 1.9 40 0.048 Phosphate <0.05a <0.05 (<0.032)b 0.2 <0.16c Antimony <0.01a <0.01 (<0.0053)b 0.005 <1.0 (<1.05)c Arsenic <0.05a <0.05 (<0.031)b 0.05 <0.61c Cobalt <0.01a <0.01 (<0.0062)b 0.01 <0.62c Copper 0.0029 0.0030 0.001 2.98 Iron 0.59 0.42d 0.1 4.17d Manganese 0.019 0.017d 0.01 1.67d Mercury <0.0015a <0.0015 (<0.00074)b 0.00001 <1.0 (<74)c Nickel <0.01a <0.01 (<0.0055)b 0.01 <0.55c Selenium <0.02a <0.02 (<0.010)b 0.002 <1.0 (<4.98)c Strontium <0.05a <0.05 (<0.032)b 0.4 <0.08c Zinc 0.008 0.011 0.01 1.12

a Average background concentration is below MDL as shown. Background concentration in Kaiemraveem River was assumed to equal one-half the MDL for modeling purposes. b Modeled concentration is presented in parentheses. Modeled concentration is below MDL for background concentration in Kaiemraveem River because the one-half MDL value for background concentration was used in the modeling efforts. c Background concentration in Kaiemraveem River is below the MDL value. Thus, modeled concentrations and their corresponding exceedence factors represent upper estimates. Where predicted concentrations are below MDL values for background levels, exceedence factors (calculated based and presented in parentheses) should be assumed as <1. d Background concentration for this parameter exceeds Russian Class 1 fisheries standards and project does not lead to increase in concentration of this analyte. Formal regulatory exceedence does not occur. Based on the modeling results, concentrations of sulfate, calcium, magnesium, phosphate, arsenic, cobalt, nickel, and strontium are not expected to exceed Russian Class 1 fisheries standards at the regulatory point-of-compliance in the Kaiemraveem River. Copper, iron, manganese, and zinc are predicted to have concentrations greater than the Russian Class 1 fisheries standards at the point-of-compliance, with corresponding exceedence



factors of 3.0, 4.2, and 1.1 respectively. However, the predicted iron and manganese concentrations in the Kaiemraveem River during 2007 are at or below background concentrations for these parameters. Thus, natural concentrations of iron and manganese in the Kaiemraveem already exceed Russian Class 1 fisheries standards prior to project development. Formal regulatory exceedences do not occur where the concentration of an analyte is above a regulatory limit under background/baseline conditions, and where project development lowers the analyte’s concentrations in receiving waters relative to background conditions. The concentrations of copper and zinc at the ambient point-of-compliance are expected to exceed the corresponding Russian Class 1 fisheries standards by about 200% and 12%, respectively. The short duration of these exceedences (i.e., ≤4-5 months over a 14 year project life), the low to moderate exceedence factors (1.1 for zinc, 3.0 for copper), and the generally conservative regulatory levels for copper and zinc in freshwaters suggests that the short-term elevated levels of these elements in the Kaiemraveem River may not pose a significant threat to aquatic life. In addition, concentrations of copper and zinc downstream of the project site will decline below the Russian Class 1 fisheries standards due to influxes of natural surface runoff. Thus, the elevated copper and zinc levels in 2007 are limited in time and extent. Furthermore, background concentrations of copper already exceed the Russian Class 1 fisheries standard by a factor of 2.9 (or 190%), and thus, an increase in copper concentration to a level 200% above the 0.001 mg/L limit will not likely be measurable, and not likely meet the criteria as having a significant environmental impact. The 3.4% predicted increase in copper concentrations, where background levels of this element were already 190% above the Russian Class 1 fisheries standard, is not considered a regulatory exceedence in the context of the current report. Scenarios where background concentrations already exceed regulatory limits prior to the project start, and the project impacts only act to increase concentrations by a few percent, are generally not viewed as being out of regulatory compliance. As discussed above, for the purposes of modeling, copper and zinc concentrations in treated AG rock runoff were not adjusted downwards to account for potential adsorptive scavenging by particulate and colloidal phases in the treatment settling ponds. Thus, concentrations of these elements in treated AG rock runoff represent conservative upper-bound estimates. It is likely that adsorptive scavenging of copper and zinc in the AG rock treatment system will remove sufficient quantities of these parameters such that Russian Class 1 fisheries standards will be met at the regulatory point-of-compliance. In the event that regulatory agencies do not accept a reliance on adsorptive scavenging as a means of meeting Russian Class 1 fisheries standards, consideration should be given towards developing site-specific Russian Class 1 fisheries standards for copper and zinc. The minimal predicted exceedence factor for zinc (1.1), with background copper levels already 190% higher than the corresponding limit (and a minimal expected increase in background copper concentrations during 2007), and the limited duration and extent of elevated copper and zinc levels, would strongly favour the granting of a short-term site-specific water quality guidelines for these parameters.



For antimony, mercury, and selenium, MDLs for background concentrations in the Kaiemraveem River, its contributing catchments, and natural surface runoff were above the corresponding Russian Class 1 fisheries standards. Reliable predictions of resultant concentrations in the Kaiemraveem River at the regulatory point-of-compliance thus cannot be made. Where MDLs for background conditions exceed regulatory limits, the impacts of a project on concentrations of these parameters cannot be assessed. For antimony, mercury, and selenium, concentrations at the point-of-compliance in the Kaiemraveem River should be considered to be at, or below, the corresponding Russian Class 1 fisheries standards. Based on the results and reasoning presented above, there is only one expected regulatory exceedence of the Russian Class 1 fisheries standards at the point-of-compliance in the Kaiemraveem River during 2007 under tailings alternative #2. Zinc concentrations may exceed Russian Class 1 fisheries standard by a factor of 1.1. 3.2.2.4 Calendar Years 2008-2014 Over the period from 2008 through 2014, runoff will occur from the following site facilities: plant and camp site, haul and access roads, the airstrip, the explosives storage facility, and the downstream crest of the tailings dam. All runoff from each of these facilities is assumed to be from saturated NAG rock or its geochemical and hydrological equivalent. Concentrations of the seven water quality parameters whose levels in saturated NAG rock runoff are above Russian Class 1 fisheries standards, and which must be diluted prior to the regulatory point-of-compliance in the Kaiemraveem River, are given in Table 3.6. All other water quality parameters not listed in Table 3.6 are not expected to have concentrations exceeding Russian Class 1 fisheries standards at the point of generation, and thus do not require extensive modeling efforts or discussion.

Table 3.6: Predicted average annual concentrations in the Kaiemraveem River during 2008 through 2014 at the ambient point-of-compliance for tailings alternative #2.




Exceedence Factor

Sulphate 10 8.8 100 0.09 Phosphate <0.05a <0.05 (<0.015)b 0.2 <0.08c Antimony <0.01a <0.0032 (<0.01)b 0.005 <0.64c Arsenic <0.05a <0.017 (<0.05)b 0.05 <0.33c Copper 0.0029 0.0014d 0.001 1.43d Manganese 0.019 0.0089 0.01 0.90 Mercury <0.0015a <0.0015 (<0.00045)b 0.00001 <45c

a Average background concentration is below MDL as shown. Background concentration in Kaiemraveem River was assumed to equal one-half the MDL for modeling purposes. b Modeled concentration is presented in parentheses. Modeled concentration is below MDL for background concentration in Kaiemraveem River because the one-half MDL value for background concentration was used in the modeling efforts.



c Background concentration in Kaiemraveem River is below the MDL value. Thus, modeled concentrations and their corresponding exceedence factors represent upper estimates. Where predicted concentrations are below MDL values for background levels, exceedence factors (calculated based and presented in parentheses) should be assumed as <1. d Background concentration for this parameter exceeds Russian Class 1 fisheries standards and project does not lead to increase in concentration of this analyte. Formal regulatory exceedence does not occur. Based on the modeling results, concentrations of sulfate, phosphate, antimony, arsenic, and manganese are not expected to exceed Russian Class 1 fisheries standards at the regulatory point-of-compliance in the Kaiemraveem River. Copper is predicted to have a concentration greater than the Russian Class 1 fisheries standard limit at the point-of-compliance, with a corresponding exceedence factor of 1.4. However, the predicted copper concentrations in the Kaiemraveem River over the period from 2008 through 2014 are below background concentrations for this parameter. Thus, natural concentrations of copper in the Kaiemraveem River already exceed the Russian Class 1 fisheries standard prior to project development. Formal regulatory exceedences do not occur where the concentration of an analyte is above a regulatory limit under background/baseline conditions, and where project development lowers the analyte’s concentrations in receiving waters relative to background conditions. For mercury, MDLs for background concentrations in the Kaiemraveem River, its contributing catchments, and natural surface runoff were above the corresponding Russian Class 1 fisheries standard. Reliable predictions of resultant concentrations in the Kaiemraveem River at the regulatory point-of-compliance thus cannot be made. Where MDLs for background conditions exceed regulatory limits, the impacts of a project on concentrations of these parameters cannot be assessed. For mercury, concentrations at the point-of-compliance in the Kaiemraveem River should be considered to be at, or below, the corresponding Russian Class 1 fisheries standard. Based on the results and reasoning presented above, there are no expected regulatory exceedences of the Russian Class 1 fisheries standards at the point-of-compliance in the Kaiemraveem River over the period from 2008 through 2014 under tailings alternative #2.



3.2.3 Summary Comparison With Receiving Water Criteria Based on the results of the ambient freshwater quality modeling exercise presented above, a summary of the expected Russian Class 1 fisheries standard exceedences over the project life under tailings alternatives #1 and #2 is presented in Table 3.7.

Table 3.7: Potential Russian Class 1 fisheries standard exceedences (with exceedence factors in parentheses) in the Kaiemraveem River over the project life.

Year Alternative #1 Alternative #2 2006 No formal

regulatory exceedences

No formal regulatory exceedences

2007 Zinc (1.1) Zinc (1.1) 2008-2014 No formal

regulatory exceedences

No formal regulatory exceedences

Under both tailings alternatives, Russian Class 1 fisheries standard exceedences are only expected during 2007, when treated AG rock runoff must be discharged into nearby catchments that feed into the Kaiemraveem River. As is noted in the discussions above, conservative upper-bound estimates for copper and zinc in the treated AG rock runoff. The potential effects of adsorptive scavenging on these elements were not considered. However, regulatory agencies may not accept reliance on adsorptive scavenging as a means of further reducing the estimated concentrations of copper and zinc in treated AG runoff to levels sufficiently low that downstream Russian Class 1 fisheries standard exceedences in the Kaiemraveem River do not occur.



4.0 POTENTIAL CLOSURE/POST-CLOSURE WATER QUALITY ISSUES Discharges from exposed NAG material used for plant site construction, roads, and the airstrip are not likely to present any water quality issues. It is assumed these regions can be reclaimed with minimal effort and no long-term monitoring requirements at the end of mine operations. At the end of the mine life, remaining free water in the tailings storage facility will be discharged into ambient receiving waters (likely the Kaiemraveem River). If concentrations of any parameters are above World Bank direct discharge limits, or if insufficient dilution is available in receiving waters in order to meet the Russian Class 1 fisheries standards at the downstream regulatory point-of-compliance, water treatment activities may be initiated. In particular, concentrations of residual cyanide (and daughter products) may not meet World Bank direct discharge criteria or the Russian Class 1 standards downstream in the receiving waters. Following treatment and discharge of free water in the tailings storage facility, this facility will be capped with inert rock to a depth greater than the active layer, and allowed to freeze. No water quality issues are expected following capping of the tailings facility provided suitable NAG rock is used as capping material. The current mine plan calls for an open pit geometry with a total storage volume of about 450,000 m3 below the spill point at 582 masl. Runoff water quality in the open pit at closure is expected to approximate that observed during operations. Thus, closure/post-closure runoff in the open pit will likely be acidic with elevated concentrations of metals and major ions such as sulfate, iron, and calcium. This runoff will collect at the base of the pit and will likely require treatment prior to discharge into ambient receiving waters. The average annual runoff in the open pit of 60,000 m3 can readily be contained within the available storage at the pit base below its spill point. Up to seven years of average runoff (not assuming net losses/gains to or from the underground workings) can potentially be contained within the open pit prior to its spilling into receiving waters. With the assumption that the exposed sulfides in the pit walls will not completely oxidize prior to closure, annual or multi-annual (at less than 6-7 year intervals) batch treatment of pit runoff at closure/post-closure will likely be required until the water quality in the pit runoff meets World Bank direct discharge criteria. Similar treatment methods as discussed in the Water Management Plan for the AG waste rock stockpile, ore stockpile, and open pit runoff in 2007 are recommended for the closure/post-closure open pit runoff. Note that due to the batch nature of closure/post-closure open pit water treatment, flow-through treatment facilities are not likely to be recommended. Thus, in addition to the required chemical dosing and mixing equipment, a pumping and pipeline system may be required to convey treated waters out of the open pit and into ambient receiving waters. Such treatment is expected to produce an effluent that meets the World Bank direct discharge criteria. Further geochemical modeling using the available (and/or extended) humidity cell testing results is recommended in order to better understand the nature, distribution, and temporal properties of closure/post-closure open pit runoff. The results will better guide closure planning and costing efforts, and ideally indicate the potential required length of post-closure site presence and water treatment activities.



5.0 CONCLUSIONS The potential surface water quality impacts of the Kupol Gold Project on the Kaiemraveem and Starichnaya Rivers was investigated using a coupled mass balance and geochemical modeling approach. The water quality modeling methods were based on knowledge regarding background water chemistry and hydrology and an understanding of the planned project activities and conceptual water and waste management plans that may impact nearby receiving waters. Concentrations of all relevant water quality parameters were compared against the World Bank direct discharge criteria for open pit (World Bank, 1999a) and underground (World Bank, 1999b) mines, and against the Russian Class 1 fisheries standards (and drinking water/recreational use standards where no fisheries standard was available) (VNIRO, 1999). The following three general project phases were considered: (1) the initial year of site construction in 2006; (2) the first full year of pit development and prior to construction of the starter tailings dam in 2007; and (3) the operations phase of the project from 2008 through 2014 after the closed-circuit process plant-tailings pond water balance and process mill start-up has occurred. Two tailings alternatives were also considered separately in the modeling approach. During 2006, runoff will occur from the following site facilities: plant and camp site, haul and access roads, the airstrip, and the explosives storage facility. No World Bank direct discharge exceedences into either the Kaiemraveem or Starichnaya Rivers, or their contributing catchments, under either of the tailings alternatives are expected from site runoff in 2006. In addition, no Russian Class 1 fisheries standard exceedences are expected from site runoff into either the Kaiemraveem or Starichnaya Rivers under either of the tailings alternatives at the ambient regulatory points-of-compliance located 500 m downstream of the last project discharge into either of these rivers during 2006. During 2007, runoff will occur from the following site facilities: plant and camp site, haul and access roads, the airstrip, the explosives storage facility, the AG waste rock stockpile, the ore stockpile, and pit inflows. Runoff from the AG waste rock stockpile, the ore stockpile, and pit inflows will be treated with lime neutralization prior to release into receiving waters. No direct World Bank discharge exceedences into either the Kaiemraveem or Starichnaya Rivers, or their contributing catchments, under either of the tailings alternatives are expected from site runoff in 2007. In addition, with the exception of zinc, no Russian Class 1 fisheries standard exceedences are expected from site runoff into either the Kaiemraveem or Starichnaya Rivers under either of the tailings alternatives at the ambient regulatory points-of-compliance located 500 m downstream of the last project discharge into either of these rivers. During 2007, zinc concentrations at the ambient point-of-compliance in the Kaiemraveem River are expected to exceed the Russian Class 1 fisheries standards by about 10% under both tailings alternatives. The minimal exceedence factor for this parameter, as well as its limited spatial and temporal duration (≤4-5 months over a 14 year project life), combined with generally conservative regulatory levels for zinc in freshwaters suggests that the short-term elevated zinc level in the Kaiemraveem River will not pose a significant threat to aquatic life.



Over the period from 2008 through 2014, runoff will occur from the following site facilities: plant and camp site, haul and access roads, the airstrip, the explosives storage facility, and the downstream crest of the tailings dam. No World Bank direct discharge exceedences into either the Kaiemraveem or Starichnaya Rivers, or their contributing catchments, under either of the tailings alternatives are expected from site runoff from 2008 through 2014. In addition, no Russian Class 1 fisheries standard exceedences are expected from site runoff into either the Kaiemraveem or Starichnaya Rivers under either of the tailings alternatives at the ambient regulatory points-of-compliance located 500 m downstream of the last project discharge into either of these rivers from 2008 through 2014.



6.0 CLOSURE Recommendations presented herein are based on the modeling assumptions and methodologies described, and upon the baseline hydrology that forms the foundation of the site water balance and, hence, the site water quality mass balance and geochemical models. The water balance and water quality model is highly integrated with the site development and mine plans, and must be continually updated as these plans evolve. This report has been prepared for the exclusive use of Bema Gold, Inc. for specific application to the area within this report. Any use which a third party makes of this report, or any reliance on or decisions made based on it, are the responsibility of such third parties. AMEC accepts no responsibility for damages, if any, suffered by any third party as a result of decisions made or actions based on this report. It has been prepared in accordance with generally accepted engineering and environmental practices. No other warranty, expressed or implied, is made. Respectfully submitted, AMEC Earth & Environmental, a division of AMEC Americas Limited

Per Sierra Rayne, Ph.D., P.Ag., P.Chem., A.Sc.T. Environmental Scientist Per Peter Lighthall, M.Eng., P.Eng. Vice-President, Mining Project Manager

Reviewed by: Per Larry Connell, P.Eng. Senior Mining Environmental Consultant



REFERENCES

AMEC. 2005a. Water Management Plan: Kupol Project – Chukotka Autonomous Okrug, Far East Russia. Prepared for Bema Gold, Inc.: Vancouver, BC, Canada. AMEC. 2005b. Site Water Balance: Kupol Project – Chukotka Autonomous Okrug, Far East Russia. Prepared for Bema Gold, Inc.: Vancouver, BC, Canada. Atkin, S. 2005. Geochemical Characterization of Waste Rock and Ore Materials at the Kupol Project, Russian Far East. Prepared for Bema Gold Corporation: Vancouver, BC, Canada. Bema. 2004. Preliminary Economic Assessment: Kupol Gold Project, Far East Russia. Prepared by Bema Gold Corporation: Vancouver, BC, Canada. Bema. 2005. Hydrological Baseline Studies, VNII-1: . Prepared by Bema Gold, Inc.: Vancouver, BC, Canada. Berezovskaya, S., Yang, D., and Kane, D.L. “Compatibility analysis of precipitation and runoff trends over the large Siberian watersheds.” Geophysical Research Letters, 31, L21502, 1-4. VNIRO. 1999. Maximum Allowable Concentrations for Rivers and Waterways, Fisheries Standards. VNIRO: Moscow, Russia, 1999. Diavik. 2000. Site Water Balance: Version 2, December 2000. Diavik Diamonds Project: Yellowknife, NWT, Canada. Kubota, J., Suzuki, K., Yamazaki, Y., Ohata, T., and Vuglinsky, V. 2004. “Water and energy budget in the southern mountainous region of eastern Siberia.” The 6th International Study Conference on GEWEX in Asia and GAME, Kyoto International Community House, Kyoto, Japan. Morin, K.A., and Hutt, N.M. 1997. Environmental Geochemistry of Minesite Drainage: Practical Theory and Case Studies. Minesite Drainage Assessment Group (MDAG): Vancouver, BC, Canada. Rhys, D.A. 2004. Structural Study of the Kupol Deposit, Chukotka Autonomous Okrug, Eastern Russia. Prepared for Bema Gold Corporation: Vancouver, BC, Canada. Schulz, K.L. 2004. Redox Reactions (SUNY College of Environmental Science and Forestry). Available on-line at http://www.esf.edu/efb/schulz/Limnology/redox.html. Site accessed: June 10, 2005. Verweij, W. 2005. CHEAQS Pro v.2005.1. Available on-line at http://home.tiscali.nl/cheaqs/author.html. Site accessed: March 21, 2005.



Wardrop. 2005. E-mail from Gordon Zurowski (Wardrop) to Sierra Rayne (AMEC) dated May 4, 2005. Subject heading “Final ARD Surface Percentages.xls”. World Bank. 1999a. World Bank Environment, Health, and Safety Guidelines: Open Pit Mines. The World Bank Group: Washington, DC, USA. World Bank. 1999b. World Bank Environment, Health, and Safety Guidelines: Underground Mines. The World Bank Group: Washington, DC, USA. Yamazaki, T., Yabuki, H., Ishii, Y., Ohta, T., and Ohata, T. “Water and energy exchanges at forests and a grassland in eastern Siberia evaluated using a one-dimensional land surface model.” Journal of Hydrometeorology, 5(3), pp. 504-515. Zhang, Y., Ohata, T., Ersi, K., and Tandong, Y. 2003. “Observation and estimation of evaporation from the ground surface of the cryosphere in eastern Asia.” Hydrological Processes, 17, pp. 1135-1147. Zhang, Y., Suzuki, K., Kadota, T., and Ohata, T. 2003. “Sublimation from snow surface in southern mountain taiga of eastern Siberia.” Journal of Geophysical Research, 109, D21103, pp. 1-10.

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