KITSAULT MINE PROJECT ENVIRONMENTAL ASSESSMENT … · Figure 1: Site Location Map ... Preliminary Interramp Slope Design Curves: ... and current geologic solids provided by Avanti

KITSAULT MINE PROJECTENVIRONMENTAL ASSESSMENT

APPENDICES

VE51988 – Appendices

APPENDIX 3.0-E Feasibility Geotechnical Pit Slope Evaluation Kitsault Project British

Columbia, Canada

A V A N T I K I T S A U L T M I N E L T D

KITSAULT MOLYBDENUM PROJECT

FEASIBILITY STUDY REPORT

Project No. 165003 APPENDIX C – 1

January 2011

C.1 F E A S I B I L I T Y G E O T E C H N I C A L P I T SL O P E

E VA L U AT I O N SRK NO V 2010

Feasibility Geotechnical Pit Slope Evaluation

Kitsault Project British Columbia, Canada

Report Prepared for

Avanti Mining Inc.

Report Prepared by

November 2010

Feasibility Geotechnical Pit Slope Evaluation Kitsault Project

British Columbia, Canada

Avanti Mining Inc. 5251 DTC Parkway

Suite 405 Greenwood Village, CO

80111

SRK Consulting (U.S.), Inc. Suite 3000, 7175 West Jefferson Avenue

Denver, Colorado, USA 80235 Tel: 303.985.1333 Fax: 303.985.9947

E-mail: [email protected] Web site: www.srk.com

SRK Project Number 2CA020.004

November 2010

Author Michael Levy, P.E., P.G.

SRK Consulting Feasibility Geotechnical Pit Slope Evaluation, Kitsault Project Page i

MEL/lb 1CA020 004_Kitsault_FS_Geotechnical_Evaluation_FINAL.doc, Nov. 15, 10, 5:25 PM November 2010

Table of Contents

1 Introduction and Background ................................................................................. 1

2 Program Objectives and Work Program ................................................................ 3 2.1 Program Objectives ........................................................................................................ 3 2.2 Work Program ................................................................................................................ 3

3 Geologic Setting ...................................................................................................... 4 3.1 Local Geology ................................................................................................................ 4 3.2 Major Geologic Structures .............................................................................................. 4

4 Field Data Collection ............................................................................................... 6 4.1 Geotechnical Core Logging ............................................................................................ 6

4.1.1 Geotechnical Logging Procedures ............................................................................... 7 4.1.2 Core Drilling Method ..................................................................................................... 7

4.2 Discontinuity Orientation ................................................................................................ 8 4.3 Point Load Testing ......................................................................................................... 8 4.4 Geotechnical Observations of Existing Pit ..................................................................... 9 4.5 Packer Testing ............................................................................................................... 9

5 Laboratory Testing ................................................................................................ 11 5.1 Unconfined Compressive Strength and Elastic Properties ........................................... 11 5.2 Triaxial Compressive Strength Testing ........................................................................ 13 5.3 Direct Shear Testing..................................................................................................... 13 5.4 Direct Tensile Strength Testing .................................................................................... 15 5.5 Unit Weight Measurements .......................................................................................... 15

6 Geotechnical Model ............................................................................................... 16 6.1 Data Analysis ............................................................................................................... 16

6.1.1 Intact Rock Strength ................................................................................................... 16 6.1.2 Discontinuity Frequency ............................................................................................. 17 6.1.3 Discontinuity Shear Strength ...................................................................................... 19 6.1.4 Discontinuity Orientation ............................................................................................. 19

6.2 Rock Mass Classification ............................................................................................. 20 6.3 Geotechnical Domains ................................................................................................. 23

6.3.1 Hornfels Domain ......................................................................................................... 23 6.3.2 Intrusives Domain ....................................................................................................... 23

6.4 Rock Mass Shear Strength .......................................................................................... 26 6.5 Groundwater ................................................................................................................ 29 6.6 Design Sectors ............................................................................................................. 30

7 Interramp/Overall Slope Stability Modeling ......................................................... 35 7.1 Model Methodology ...................................................................................................... 35 7.2 Results of Interramp/Overall Stability Analysis ............................................................. 37

8 Bench Design ......................................................................................................... 40 8.1 Description of Models Used ......................................................................................... 40 8.2 Methodology ................................................................................................................. 41

8.2.1 Likelihood of Occurrence ............................................................................................ 41 8.2.2 Likelihood of Exceeding Shear Resistance ................................................................ 44 8.2.3 Likelihood of Kinematic Admissibility .......................................................................... 44

SRK Consulting Feasibility Geotechnical Pit Slope Evaluation, Kitsault Project Page ii


8.3 Results ......................................................................................................................... 47

9 Pit Slope Design Recommendations .................................................................... 48

10 Assessment of Future Geotechnical Work .......................................................... 50

11 References ............................................................................................................. 51

SRK Consulting Feasibility Geotechnical Pit Slope Evaluation, Kitsault Project Page iii


List of Tables

Table 1: Drillholes Oriented and Logged for Geotechnical Data .................................................. 7 Table 2: Summary of Discontinuity Orientation ............................................................................ 8 Table 3: Uniaxial Compressive Strength Testing ....................................................................... 12 Table 4: Triaxial Compressive Strength Testing ........................................................................ 13 Table 5: Summary of Residual Shear Strengths ........................................................................ 15 Table 6: Direct Tensile Strength Testing .................................................................................... 15 Table 7: Discontinuity Sets Delineated for Analysis ................................................................... 20 Table 8: In-situ Rock Mass Rating (IRMR) Distributions ............................................................ 21 Table 9: Secondary Hoek-Brown Parameters Stochastic Input ................................................. 26 Table 10: Results of Overall Slope Stability Modeling ............................................................... 38 Table 11: Summary of Potential Failure Forming Sets .............................................................. 42 Table 12: Summary of Discontinuity Set Spacings .................................................................... 44 Table 13: Composited Results of Backbreak Analysis .............................................................. 47 Table 14: Summary of Pit Slope Design Recommendations and Expectations ........................ 48

List of Figures

Figure 1: Site Location Map .......................................................................................................... 2 Figure 2: Location of Geotechnical Drillholes ............................................................................. 10 Figure 3: Point Load Index – UCS Correlation Factor ................................................................ 17 Figure 4: Rock Mass Parameters ............................................................................................... 18 Figure 5: Distribution of Friction Angles (Zero Cohesion) ........................................................... 19 Figure 6: Discontinuity Pole Plots ............................................................................................... 22 Figure 7: Drillhole RQD Cross-Sections ..................................................................................... 24 Figure 8: Geologic Model and Geotechnical Cross Sections ...................................................... 25 Figure 9: Rock Mass Shear Strength: Hornfels ......................................................................... 27 Figure 10: Rock Mass Shear Strength: Intrusives ..................................................................... 28 Figure 11: Summary of vibrating wire piezometer data from K09-07 (El=595.71) ...................... 31 Figure 12: Summary of vibrating wire piezometer data from K09-12 (El=548.49) ...................... 31 Figure 13: Groundwater Pressures Measured in K09-07 ........................................................... 32 Figure 14: Groundwater Pressures Measured in K09-12 ........................................................... 33 Figure 15: Pit Slope Design Sectors ........................................................................................... 34 Figure 16: Explanation of Pit Slope Terminology ........................................................................ 36 Figure 17: Preliminary Interramp Slope Design Curves: Hornfels .............................................. 39 Figure 18: Discontinuity Contour Plot for Backbreak Analysis .................................................... 43 Figure 19: Explanation of Backbreak Terminology ..................................................................... 46 Figure 20: Maximum Interramp Slope Angle Recommendations ............................................... 49

List of Appendices Appendix A: Geotechnical Core Logs Appendix B: Laboratory Testing Uniaxial Compressive Strength Testing Triaxial Compressive Strength Testing Direct Shear Testing Brazilian Disk Tension Testing Appendix C: Slope Stability Modeling

SRK Consulting Feasibility Geotechnical Pit Slope Evaluation, Kitsault Project Page iv


Limit Equilibrium Modeling Finite Element Modeling

SRK Consulting Feasibility Geotechnical Pit Slope Evaluation, Kitsault Project Page 1


1 Introduction and Background SRK Consulting (US), Inc. (SRK) was requested by Avanti Mining Inc. (Avanti) to carry out a feasibility level geotechnical evaluation for the Kitsault Project Open Pit in the British Columbia, Canada (Figure 1).

This report presents a complete description of the methods used to collect pertinent information, the information so gathered, the analytical tools employed to produce assessments of the anticipated response of the geologic environments to the development of the open pit and the recommendations based upon those assessments. The feasibility study (AMEC, 2010) ultimate pit and current geologic solids provided by Avanti were used as the basis for the evaluation.

KITSAULT PROJECT SITE

SITE LOCATION MAP

PIT SLOPE EVALUATION

DATE:

FEB. 2010FIGURE NO.:

1

SRK PROJECT NO.:1CA020.004

APPROVED:

MELREVISION NO.

AFILE NAME:

KITSAULT PROJECT



2 Program Objectives and Work Program

2.1 Program Objectives

The primary objectives of the feasibility-level geotechnical evaluation for the Kitsault project were:

To collect additional and to assimilate existing geotechnical information pertaining to the in-situ materials;

To geotechnically characterize the in-situ materials;

To undertake laboratory testing of geomechanical properties of samples of the in-situ materials;

To develop a geotechnical model to serve as the basis for geomechanical analyses;

To conduct geomechanical analyses; and,

To make recommendations pertaining to optimal slope angles and pit architecture for mine design purposes.

2.2 Work Program

The principle stages of the geotechnical evaluation work program were comprised of the following:

Recommendation of the number, location and orientation of core holes sufficient to characterize in-situ materials in the open pit area;

Geotechnical core logging and discontinuity orientation of core recovered from the drill holes;

Selection of representative drill core samples from the respective lithological units encountered in the geotechnical drill holes;

Submission of the representative samples to the University of Arizona Rock Mechanics Laboratory in Tucson, Arizona, for geomechanical testing;

Analyses and interpretation of the geotechnical data and laboratory test results to produce a comprehensive analytical model of in-situ properties;

Examination of the anticipated behavior of the geotechnical model to expected mining-induced stresses, using various analytical methods; and,

Compilation of a feasibility-level geotechnical pit slope evaluation report incorporating recommendations pertaining to optimal pit slope angles and pit architecture for mine design purposes.

As commissioned, the work reported herein was performed at a feasibility design level.



3 Geologic Setting The following description of the Kitsault geologic setting was extracted from previous work by Steininger (1981).

The Kitsault Molybdenum ore deposit is located within the Intermountain tectonic belt of the large Canadian geologic province know as the Cordillera. Rock types present within this belt range in age from Devonian to early Cenozoic, typically consisting of sedimentary, granitic, volcanic island and continental arc formations, and marine and non-marine clastics eroded mainly from uplifting of the Omineca Belt. Significant deformation has occurred in this region of the province, primarily caused by compression and extension transtensional forces.

3.1 Local Geology

The Kitsault project site is located approximately 2 km east of the Coast Plutonic Complex, consisting of a northwest trending belt of metamorphic and intrusive rocks. Hornfels is the predominant metamorphic lithology, while intrusive lithologies are typically granodiorite to quartz monzonite, with minor granite, as plutons. Intense intrusive activity within this region, including recent plateau lava flows, can be attributed to the Coast Plutonic Complex. Extensive glaciation has occurred in this area, deeply eroding valleys. Glacial remains are only present as thin alluvium veneers and swamplands covering outcrops.

The Kitsault deposit lies within the Lime Creek Intrusive Complex, hosted by the sedimentary units of Bowser Lake Group. The intrusives at the site consist of quartz diorite, granodiorite, and decreased amounts of quartz monzonite. Mineralization within the deposit is related to the last two phases of the Lime Creek Complex, i.e., the Central Stock (granodiorite) and the Northeast Porphyry (porphyritic granodiorite).

The Bowser Lake Group is primarily comprised of interbedded greywacke and argillite with bed thicknesses ranging from inches to tens of feet. The formation is primarily greywacke with all members being metamorphosed to greenschist facies. Hornfels within the Bowser Lake Formation were likely produced in reaction to intrusions along the eastern border of the Coast Plutonic Complex.

Lamprophyre dikes, occurring as numerous northeast trending swarms, are present throughout the deposit. These swarms, which are likely related to the Alice Arm Intrusives, consist of several to hundreds of dikes per mile and range in thickness from inches to 50 feet. Typically northeast trending faults, although common, appear to have had little effect on the units within the ore body.

3.2 Major Geologic Structures

Major geologic structures are those features, such as faults, dikes, shear zones, and contacts that have dimensions on the same order of magnitude as the area being characterized. These structures are treated as individual elements for design purposes, as opposed to joints, which are handled statistically.

To date, there are no known major structural features within the immediate area of the anticipated Kitsault pit. Smaller scale, high angle faulting is, however, evident in the exposed north pit wall, but it is generally oriented such that it is not expected to adversely affect pit stability.



Several smaller scale faults or shear zones have also been identified in resource and geotechnical drilling. Most of these structures are not anticipated to significantly impact pit slope stability due to their apparent lack of persistence and associated limited degree of rock degradation.

Lamprophyre dikes are exposed in existing pit walls and have been encountered in drillholes. The dikes are generally of good rock quality and are not expected to significantly impact pit slope stability.



4 Field Data Collection The field data collection program was developed with the primary objective of rock mass characterization to support development of a geotechnical model suitable for pit slope stability evaluation. Field data collection consisted of geotechnical core logging and largely subjective observations of existing pit wall conditions.

4.1 Geotechnical Core Logging

Geotechnical logging, field point load testing and discontinuity orientation of core recovered from two drill holes were conducted for this investigation. The two drill holes were designed to supplement the 2008 pre-feasibility geotechnical core logging program. In addition to the two geotechnical coreholes drilled in 2009 for this investigation, data from the six geotechnical coreholes drilled in 2008 for the previous SRK (2009) Kitsault Pre-feasibility Geotechnical Pit Slope Evaluation were also considered in the analyses.

Based on the current understanding of the deposit and mine plan, drillhole locations and orientations were selected to provide the best coverage possible of rock likely to form pit walls. The geotechnical drillhole locations were initially chosen based on preliminary and historic pit shells and, in some instances, drillhole intersections with the final pre-feasibility pit slopes were not optimal relative to the latest pit designs. It is believed, however, that this factor does not adversely impact the analyses conducted to a significant degree.

Five of the previous six geotechnical drillholes, i.e., K-08-04, K-08-09, K-08-12, K-08-14, and K-08-16, were drilled to coincide with holes planned for the Avanti 2008 resource drilling program. Based on the current understanding of the deposit, those particular five holes were selected to provide the best coverage possible of rock likely to form the Kitsault pit walls. Since no further resource drilling was planned in the area of the anticipated western pit wall, an additional hole (K-08-06) was drilled specifically to examine rock expected to comprise that wall segment.

Drillhole inclinations of approximately 60 degrees below the horizontal were selected over vertical holes since they were judged more likely to intersect geologic structures such as joints and fracture systems which, if present, will influence slope stability.

Collar locations and the drillhole azimuths of the two supplemental geotechnical holes drilled for this investigation as well as the six holes considered in the previous (SRK, 2009) investigation are summarized in Table 1 and presented on Figure 2.



Table 1: Drillholes Oriented and Logged for Geotechnical Data

Hole ID Collar Coordinates

Azimuth (deg)

Inclination (deg)

Length (m) Northing Easting Elevation

K08-04 6141730.0 473100.3 560.2 185 -58 300.5

K08-06 6141850.0 473000.0 579.0 275 -60 401.4

K08-09 6141934.7 473743.9 672.5 277 -53 433.4

K08-12 6141980.0 473300.0 594.2 002 -43 315.8

K08-14 6141850.0 473570.0 594.2 089 -43 349.6

K08-16 6141600.0 473580.0 593.6 086 -46 286.8

K09-07 6141945.9 473534.5 595.7 43 -57 400.2

K09-12 6141611.6 473249.2 550.0 180 -57 459.6

4.1.1 Geotechnical Logging Procedures

Core retrieved from the two geotechnical coreholes were logged on a 24 hour per day basis, at the rig, in the liners, or splits, prior to boxing and transporting. The geotechnical core logging program was developed to yield information pertinent to modeling of pit slope stability, such as geologic contacts, profiles of rock strength, and characterization and frequency of discontinuities. Specific parameters that were logged included:

General lithology and structures;

Total core recovery;

Rock Quality Designation (RQD);

Rock weathering and intact strength indices;

Frequency of discontinuities;

Discontinuity characteristics (type, roughness, infillings and wall condition); and,

Discontinuity orientation (when possible).

Care was taken to exclude handling or mechanically induced fracturing of the core as the inclusion of such would produce lower rock quality classifications, potentially contributing to an unnecessarily conservative slope design. Geotechnical corehole logs are presented in Appendix A.

During core logging, redundant samples of the core were collected to provide specimens for laboratory strength testing. Samples were collected at approximately 30 meter intervals, or when significant rock type or strength changes were apparent. Each sample was sealed and safely stored at the time of collection. Upon completion of the drilling, samples were shipped to SRK’s office in Denver, Colorado, for test sample selection. Select samples were then repackaged and shipped to the University of Arizona Rock Mechanics Laboratory in Tucson, Arizona, for testing.

4.1.2 Core Drilling Method

The coreholes were drilled by Driftwood Diamond Drilling, Ltd., from Smithers, British Columbia, using a skid mounted Hydracore 2000 drill rig with a 61.1mm I.D.(HQ3), 1.5m and 3.0m long triple-tube sampling barrels. The coreholes were advanced with a face discharge bit



system using a polymer mixture to facilitate core recovery. This coring method facilitated the recovery of continuous core samples as the holes advanced.

Downhole surveys were conducted by Driftwood upon completion of drilling; subsequently, the surface casing was pulled and the hole allowed to collapse. Depth to groundwater could not be determined at the time of hole advancement due to the 24 hour per day drilling schedule, with its continuous fluid injection and circulation.

4.2 Discontinuity Orientation

Orientation of discontinuities in each run was accomplished using an A.C.T. core orientation system manufactured by Reflex Instruments. The depth, alpha angle and beta angle were measured for each discontinuity on all core runs that were successfully oriented. The beta angle, i.e., the angle from the lowest part of the ellipse formed by the intersection of each discontinuity with the core, was measured from the bottom of the core in a clockwise direction when looking down hole. The alpha angle was measured as the maximum angle made by the discontinuity with respect to the core axis.

It was possible to orient a total of 1,847 discontinuities out of the total 3,360 discontinuities logged (55%) in the two supplemental geotechnical coreholes drilled for this evaluation. A summary of oriented core information by hole, including the six previous 2008 holes, is presented in Table 2.

Table 2: Summary of Discontinuity Orientation

Hole ID Drillhole Length

(m)

Core Length Oriented (m)

Total Discontinuities

Logged

Total Discontinuities

Oriented

Percentage of Discontinuitie

s Oriented

K08-04 297.8 251.2 831 722 93%

K08-06 398.4 333.0 912 740 84%

K08-09 424.3 218.8 749 383 52%

K08-12 309.7 124.2 709 351 50%

K08-14 346.1 179.4 1,181 649 55%

K08-16 281.6 141.4 543 298 55%

K09-07 400.2 362 1,661 828 50%

K09-12 459.6 412 1,699 1011 60%

4.3 Point Load Testing

Point Load Tests (PLT) were performed during core logging at a frequency of approximately one test per every 2 to 3m using a Roctest Pil-7 test machine to provide detailed and nearly continuous profiles of relative rock strength. PLTs were conducted according to International Society for Rock Mechanics (ISRM, 1985) procedures. Both axial (parallel to the long axis of the core) and diametral (perpendicular to the long axis of the core) loading tests were conducted. Axial point load testing was performed as samples suitable for testing in an axial orientation were obtained from coring or were produced by breaking especially long sticks of core in diametral tests.

A combined total of 102 point load tests were conducted on core from the two 2009 geotechnical coreholes; of those, 42 met test criteria for passing test results. Point load indices (Is(50)) were



calculated from the field PLT data using the ISRM (1985) suggested method. Calculated point load index strengths (Is(50)) ranged between 0.3 and 10.5 MPa, with an average of 5.0 MPa.

In addition to the tests routinely conducted at 2 to 3 meter intervals, at least one PLT was also performed adjacent to each UCS sample obtained for laboratory testing. The reason for the paired PLT and UCS samples was to permit estimation of a correlation factor for conversion of the field PLT tests to laboratory UCS values.

4.4 Geotechnical Observations of Existing Pit

During a site visit by SRK between September 8 and September 11, 2008, geotechnical observations of the existing pit wall conditions and performance were made and noted. The outer walls of the currently exposed pit consist primarily of a hornfels unit cut by relatively small intrusive bodies and lamprophyre dikes. The existing outer pit walls are comprised of up to approximately six 10 meter high benches separated by catch benches, resulting in interramp slope angles of approximately 43 degrees to 45 degrees over a total vertical height of 60 meters. A relatively low slope comprised of one to two benches is exposed in the interior, intrusive portion of the pit.

Based on the field observations, both the outer, hornfels slopes and the inner, intrusive slopes are in good condition, showing only minor raveling and very few observable rock displacements. The displacements observed included relatively limited plane shear and bench scale wedge failures which were noted particularly in the outer, north to northeast pit walls, and which most likely occurred during excavation when the pit was last active 26 years ago. No major fault structures were observed in the pit walls during the SRK site visit; however, some small scale, high angle faulting, as described in Section 3.2, was evident in the north pit wall.

In August, 2009, a preliminary survey of the current pit did not identify significant seeping of groundwater in the current pit walls. Observations of ‘significant’ seepage from pit walls during quarterly seepage surveys during mine reclamation studies were reported (SRK, 2004); however, no flow rates were measured. It is likely that localized inflows will vary seasonally, and be influenced by surface water flows. Current pit inflows may be recharged by surface water runoff.

4.5 Packer Testing

Hydraulic packer testing was carried out at intervals covering the full depths of the two 2009 supplemental geotechnical drill holes. This provided profiles of hydraulic conductivity necessary to evaluate hydrogeologic characteristics of the rock mass. Details of the packer testing procedures and results are presented in the (SRK, 2009) Kitsault Pre-feasibility Study Pit Hydrogeology report. Conclusions are summarized herein in Section 6.5.

473000 E

473500 E

474000 E

6141000 N

6141500 N

6142000 N

6142500 N

K08-16

K08-14

K08-09

K09-07

K08-12

K08-06

K09-12

K08-04

LEGEND

EXISTING GROUND CONTOURS (MAJOR/MINOR) 5 METER INTERVAL

LOCATION OF GEOTECHNICAL

DRILLHOLES

NOTE

1. PIT TOPOGRAPHY SHOWN IS WARDROP (2009) PRE-FEASIBILITY STUDY.

FILE NAME:

SRK JOB NO.:

DATE: APPROVED: FIGURE:

1CA020.004.Rev.A.Fig,2.Location.of.Geotech.Drillholes.2010-11-02.dwg

REVISION NO.

7175 West Jefferson Ave. Suite 3000

Denver, Colorado 80235

303-985-1333

T:\Kitsault British Columbia\!040_AutoCAD\Feasibility Pit Slopes Figures\Novemeber.2010.Updates\1CA020.004.Rev.A.Fig,2.Location.of.Geotech.Drillholes.2010-11-02.dwg

A2

ML


KITSAULT

BRITISH COLUMBIA, CANADA

1CA020.004

NOV. 2010

GEOTECHNICAL DRILLHOLE COLLAR LOCATION AND HORIZONTAL

BOREHOLE PROJECTION



5 Laboratory Testing Geomechanical testing was conducted at the University of Arizona Rock Mechanics Laboratory in Tucson, Arizona, to determine strength characteristics for the in-situ materials. The overall laboratory program consisted of direct shear, uniaxial and triaxial compressive strength, and direct tensile strength testing as well as measurements of unit weight and elastic properties. A total of 75 laboratory tests were conducted on samples selected to represent the range of the rock conditions observed in the eight geotechnical borings. After completion of the laboratory testing program, the tested samples were returned to SRK for forensic review. Raw laboratory test data is included in Appendix B.

5.1 Unconfined Compressive Strength and Elastic Properties

The uniaxial compressive strength (UCS) test involves the application of a steadily increasing axial load upon a core sample with a length-to-diameter (L/D) ratio of, ideally, between 2.0 and 2.5. The uniaxial compressive strength (in terms of stress) of the sample is the applied load that produces failure divided by the cross-sectional area of the core. For selected UCS tests, strain gauges were applied to the samples to monitor longitudinal and lateral strains which are produced in response to the axial loading. The elastic properties are derived from the strain gauge output; specifically, Young’s Modulus () is the ratio of the vertical stress to the longitudinal strain, while Poisson’s Ratio () describes the relationship between the lateral strain and the longitudinal strain.

Uniaxial compressive strength (UCS) testing was conducted on 32 samples according to ASTM Method D7012. Elastic properties (Young’s Modulus and Poisson’s Ratio) were measured for eight of the 32 UCS samples. Upon post-testing examination of the samples, it was noted that samples K08-6 at 35.8m, K08-14 at 124.1 meters and K08-12 at 192.4 meters had unusually low strengths (11.31 to 35.44 MPa) and appeared to have fractured on pre-existing discontinuities and not through the actual intact rock as should occur in a valid UCS test.

Valid tests produced UCS values ranging from 41.9 to 238.4 MPa, with a mean of 105.3 MPa; Young’s Moduli ranging from 13.7 to 69.4 GPa, with a mean value of 46.3 GPa; and, Poisson’s Ratios ranging from 0.179 to 0.302, with a mean value of 0.227. Results of the UCS and elastic properties testing from the 2008 and 2009 programs are summarized in Table 3.



Table 3: Uniaxial Compressive Strength Testing

SRK Hole ID

Sample Depth (m)

UCS (MPa)

Young’s Modulus

(GPa) Poisson’s

Ratio Unit Wt. (kN/m3) Rock Type

K08-04 39.90 70.17 14.1 0.283 25.7 Quartz Monzonite

K08-04 66.90 107.16 26.0 Quartz Monzonite




K08-06 35.77 35.44** 26.9 Quartz Monzonite

K08-06 85.06 53.75* 30.5 Diorite

K08-06 120.65 55.18* 45.4 0.179 30.9 Hornfels

K08-06 322.42 112.93 27.2 Hornfels


K08-09 157.18 89.58* 28.2 Hornfels

K08-09 252.34 135.29 27.0 Lamprophyr

K08-12 35.60 41.87 13.7 0.215 25.2 Hornfels

K08-12 157.40 238.43* 69.4 0.258 26.3 Hornfels

K08-12 311.40 71.76 26.9 Hornfels

K08-14 124.10 25.85** 26.0 Hornfels

K08-14 192.40 11.31** 30.4 Hornfels

K08-16 62.50 57.2* 25.6 Hornfels

K08-16 92.82 80.58* 26.2 Hornfels

K08-16 185.37 139.62 26.4 Hornfels



K09-07 121.22 56.78 26.1 HF Hornfels

K09-07 155.08 156.94 26.3 HF Hornfels

K09-07 182.49 49.36 26.9 Lamprophyr

K09-07 244.67 59.55 47.3 0.189 27.5 Hornfels

K09-07 338.55 126.39 28.7 Hornfels





K09-12 257.15 64.55 26.2 Quartz Monzonite * Correction factor applied to account sample L/D ratio of less than 2.0. ** UCS test results considered invalid and excluded from further analysis.



The intact Young’s Moduli determined from laboratory testing were used for empirical calculations of a rock mass deformation modulus for each domain by methods presented by Hoek and Diederichs (2006).

5.2 Triaxial Compressive Strength Testing

The triaxial compressive strength (TCS) test involves encasing a core sample in an impervious membrane and subjecting it to a selected confining pressure (3) while the sample is loaded axially (1) until failure occurs. The applied load that results in failure divided by the cross-sectional area of the core is the triaxial compressive strength given the confining pressure.

For this project, triaxial compressive strength (TCS) tests were conducted on 11 samples using ASTM Method D7012. The samples were tested at confining pressures selected to range from zero to approximately one-half of the UCS values as suggested by Hoek and Brown (1997).

TCS testing yielded compressive strengths (1) ranging between 213.8 and 294.1 MPa with a mean value of 262.1 MPa under confining pressures (3) ranging between 6.9 and 20.7 MPa, with a mean of 13.8 MPa. The results of the TCS testing are summarized in Table 4.

Table 4: Triaxial Compressive Strength Testing

Hole ID Sample

Depth (m) 3 (MPa) 1 (MPa) Unit Wt. (kN/m3)

Rock Type

K08-04 92.50 20.7 63.9 25.9 Quartz Monzonite



K08-06 293.19 10.3 321.3 26.5 Hornfels

K08-09 176.26 6.9 152.2 26.8 Hornfels

K08-12 184.65 13.8 163.0 26.2 Hornfels

K08-14 335.75 3.4 100.7 26.4 Hornfels

K08-16 154.30 20.7 223.5 26.8 Hornfels


K09-07 366.76 20.7 308.5 26.5 Hornfels


Intact rock shear strength envelopes were derived by combining tests from the respective rock types. Quartz monzonite samples yielded a peak intact friction angle of 50° and 27.7MPa cohesion. Hornfels samples yielded a combined peak intact friction angle of 50° and 20.9MPa.

5.3 Direct Shear Testing

The direct shear test involves applying a load perpendicular (normal) to a discontinuity separating two blocks of rock and continuously monitoring the shear stress necessary to displace the blocks relative to each other. To define the overall shear strength envelope, three or more normal stresses are applied to the sample and continuous displacement/shear stress data is obtained at each of the normal loads. For each normal load, the peak (maximum) and residual (steady state relative to displacement) shear stresses are recorded, thereby defining the peak and residual shear strengths given each normal stress. The relationship between an applied normal stress and the resulting shear strength defines a point on the shear strength envelope. Peak and residual shear strength envelopes can then be determined from the shear strength/normal stress points using statistical regression methods.



Direct shear testing is commonly used for estimating the expected shear strength along natural rock discontinuities such as joints, fractures and faults. Since the stress levels developed within open pits are usually much lower than the rock substance or intact strength, displacement frequently occurs along pre-existing geologic discontinuities, making the determination of discontinuity shear strength a necessity. For open pit design, direct shear testing is preferred over other methods of estimating discontinuity shear strength, such as triaxial compression testing, because direct shear testing permits a higher degree of control over the selection of the actual surface tested.

For this project, 11 core samples were selected for four point, small scale direct shear (SSDS) tests (ASTM Method D5607) to obtain discontinuity shear strength data. Natural core discontinuities preserved in the field were used for 10 of the direct shear tests. To facilitate the estimation of lower bound residual discontinuity shear strengths, a saw-cut discontinuity was created in one sample prior to testing.

The range of normal stresses applied during testing was selected to span estimated ranges of in-situ stresses that are expected to develop within the slopes and to reasonably define the characteristics of the shear strength envelopes. The selected normal loads ranged from approximately 170 to 2,070 kPa.

In order to fit a shear strength envelope to the laboratory data points, a linear or curvilinear regression analysis is typically conducted. For a linear fit, the envelope is presented according to the Mohr-Coulomb criterion, i.e., in the form of a friction angle (Φ), which corresponds to the inverse tangent of the slope of the least-squares regression line, and apparent cohesion (c), which corresponds to the shear strength intercept at zero normal stress. When conducting a linear regression with discontinuity shear strength data, the line is commonly forced through the origin simulating zero cohesion.

A curvilinear strength envelope can be presented in terms of a power curve with k and m values as described by Jeager (1971) or other nonlinear relationships such as the Hoek-Brown (Hoek, et al, 2002) criterion. For sufficiently strong rock, the curvilinear fit is considered a more realistic representation of the shear strength/normal stress relationship, particularly at relatively low normal stresses, which typify conditions in a majority of open pit mine slopes.

Although results of direct shear testing of discontinuities on some of the Kitsault samples tested demonstrated curvilinear shear strength/normal stress envelopes, most analytical stability models, including those used by SRK for backbreak analyses, utilize linear, Mohr-Coulomb parameters. Shear strengths were typified using the Mohr-Coulomb and power curve shear strength/normal stress relationships. The results are summarized in Table 5.



Table 5: Summary of Residual Shear Strengths

Hole ID Sample

Depth (m)

Linear Regression Power Regression Discontinuity Type Φ* (°) C (kPa) Φ**(°) k m

K08-04 54.63 40.4 60.7 43.7 2.0173 0.8335 Quartz Monzonite



K08-14 108.05 27.9 29.6 28.8 0.6924 0.9576 Hornfels

K08-14 249.65 37.2 0.0 37.2 0.5693 1.0558 Hornfels

K08-14 279.25 31.3 53.1 32.6 1.2843 0.8751 Hornfels

K08-16 135.90 32.4 3.4 32.5 0.6779 0.9876 Hornfels

K09-07 149.20 45.4 8.9 48.5 2.0636 0.8632 HF Hornfels

K09-07 243.29 36.0 5.5 37.2 1.2180 0.9097 Hornfels



* Best linear fit friction angle given the apparent cohesion calculated and noted ** Best linear fit friction angle assuming a zero apparent cohesion.

5.4 Direct Tensile Strength Testing

Brazilian disk tension testing according to ASTM method D3967 was conducted on 13 samples indicating intact tensile strengths ranging from 4.21 to 17.54 MPa, with a mean value of 10.48 MPa. Results of the direct tensile strength testing are summarized in Table 6.

Table 6: Direct Tensile Strength Testing

Hole ID Sample

Depth (m) Tensile

Strength (MPa) Unit Wt. (kN/m3)

Rock Type


K09-07 121.22 7.02 26.1 HF Hornfels

K09-07 155.08 17.54 26.3 HF Hornfels

K09-07 243.42 5.99 26.7 Hornfels

K09-07 338.55 14.75 28.7 Hornfels

K09-07 366.76 16.08 26.5 Hornfels








5.5 Unit Weight Measurements

Prior to actual testing of core samples, sample dimensions and weights were measured and used to calculate total unit weights for each sample. The combined data set included 54 unit weight measurements ranging from 24.7 to 30.7 kN/m3 with a mean value of 26.4 kN/m3. Unit weights are summarized along with the various strength measurements in the preceding Tables 3, 4 and 6.



6 Geotechnical Model Rock mass models were developed for Kitsault to provide a framework for interramp/overall slope stability modeling by mathematically simulating site geotechnical conditions. The term “rock mass” refers to the entire body of rock, including discontinuities. In contrast, “intact rock” or “substance strength” refers to the rock between discontinuities in a rock mass. Primary inputs to the rock mass models included intact rock strength, degree of fracturing and strength of fractures.

6.1 Data Analysis

Evaluation of the field and laboratory data collection programs indicates a high degree of variability in rock strength and geologic structure at Kitsault. This natural variation in rock strength and structure suggests that a probability-based method of analysis is most appropriate, yielding less conservative slope angles than would the selection of a unique, potentially over-conservative value as is typical in strictly deterministic analyses.

Probabilistic methods differ from deterministic methods in that each model parameter is characterized by a statistical distribution of values having a central tendency and some variation around that central tendency, rather than by a single, unique value. Further details of the probabilistic method used in this evaluation follow. Details of the data analysis methods are discussed in subsequent sections.

6.1.1 Intact Rock Strength

Intact rock strengths were assessed in the field qualitatively using ISRM (1978) methods and by conducting point load tests (PLT) as discussed in Section 4.3. Several samples of core were also selected for laboratory uniaxial compressive strength (UCS) and triaxial compressive strength (TCS) testing as described in Sections 5.1 and 5.2, respectively. UCS and Is(50) values, as well as the field estimates of intact rock strength, are plotted with depth on the geotechnical logs presented in Appendix A.

Each laboratory UCS test was paired with an adjacent field PLT Is(50) value for estimation of a correlation factor for conversion of the field PLT tests to laboratory UCS values. Overall, a relatively linear relationship was apparent between the two variables, yielding a correlation factor of 24 (UCS:Is(50)). The correlation between the laboratory UCS tests and the PLTs is demonstrated on Figure 3.



Figure 3: Point Load Index – UCS Correlation Factor

The conversion of the field PLTs to laboratory UCS values allowed nearly continuous profiles of rock strength for each corehole and provided a large population for defining UCS statistical distributions for the probabilistic analyses.

As demonstrated in the plots contained on Figure 4, both the hornfels and intrusive domains have similar ranges in UCS, however, the intrusives posses a higher mode or peak concentration (116MPa) than does the hornfels domain (61MPa).

TCS test results, as described in Section 5.2, were used for direct determination of the Hoek-Brown (Hoek, et al, 2002) material coefficient mi. As described by Hoek (1983), the Hoek-Brown constant mi is very approximately analogous to the angle of friction of the conventional Mohr-Coulomb failure criterion. Higher mi values are characteristic of brittle igneous and metamorphic rocks producing relatively steeply inclined strength envelopes and high instantaneous friction angles at lower normal stress levels. Material coefficient mi values of 28.8 and 30.2 were calculated for the hornfels and intrusive, respectively.

6.1.2 Discontinuity Frequency

The fracture (discontinuity) frequency or its inverse, fracture spacing, is a critical parameter influencing rock mass behavior. Fracture frequency is expressed as the number of fractures per unit length and fracture spacing is defined as the distance between fractures. Fracture frequency per meter was recorded during drilling for each run, thereby enabling calculation of mean fracture spacings for use in rock mass characterization and bench scale analyses, both of which are discussed in more detail in the following sections. For expedience, it was assumed that each measurement began and ended with a fracture, thereby resulting in a maximum possible spacing of about 1.5 meters, the length of the core barrel.

Intrusives Lithologic Domain Hornfels Lithologic Domain

Mean IRMR = 48 (516) Mean IRMR = 47 (904)

Mean UCS = 128 MPa (43)Mean UCS = 92 MPa (28)


Mean ff/m = 6.1 (285) Mean ff/m = 4.8 (195)

Note: Number in parenthesis represents the number of samples for the respective data set.


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As demonstrated in the plots contained on Figure 4, both the hornfels and intrusive domains display similar distributions of fracture frequency. Discontinuity spacings are discussed further in Section 8.2.3.

6.1.3 Discontinuity Shear Strength

Discontinuity shear strengths are a function of geologic history as well as rock mass weathering, alteration and/or infilling. Direct shear testing was conducted on a number of rock samples as previously discussed in Section 5.3 to provide information on the distribution of discontinuity shear strengths. Although results of direct shear testing of discontinuities on some of the samples tested demonstrated curvilinear shear strength/normal stress envelopes, most analytical stability models, including those used by SRK for backbreak analyses, utilize linear, Mohr-Coulomb parameters.

Tests results indicate similar shear strengths between the different domains and, consequently, discontinuity shear strengths were grouped together into one distribution. For the combined dataset of direct shear results, calculated friction angles (assuming zero apparent cohesion as discussed in Section 5.3) ranged from 26° to 49°, while apparent cohesion values ranged from 0 to 108kPa. The mean friction angle was 35° and the mean apparent cohesion was 25 kPa. The combined distribution of friction angles obtained from direct shear testing is shown on Figure 5.

Figure 5: Distribution of Friction Angles (Zero Cohesion)

6.1.4 Discontinuity Orientation

Geologic discontinuity influenced failure mechanisms were analyzed at both the pit wall and bench scales. The term discontinuity refers to any significant mechanical break or fracture having negligible tensile strength in the rock. Discontinuities are formed by a wide range of geological processes and can collectively include most types of joints, faults, fissures, fractures, veins, bedding planes, foliation, shear zones, dikes and contacts.

Minor discontinuities such as joints, foliation and bedding planes, represent an infinite population for practical purposes and, due to sampling limitations, are best modeled with stochastic (probabilistic) techniques. A discontinuity set denotes a grouping of discontinuities that are expected to have similar impact upon the proposed design. In open pit design, this criterion is usually modified so that all discontinuities in a similar range of orientations, i.e., dip direction and dip, are designated as a single discontinuity set.



To enable the calculation of the true dip direction and dip, the depth of intercept and the angles of the discontinuities relative to the core axis and perpendicular to the core axis, (alpha and beta angles, respectively) were measured during logging.

Accounting for the plunge and azimuth of each drillhole, discontinuity alpha and beta angles were converted to dip and dip direction using the commercially available software package, Dips developed by Rocscience, Inc. (2003). Discontinuity data from each of the geotechnical coreholes was contoured on an equal area percent plot for analysis of structural stability. In most cases, visual inspection of these plots revealed preferred discontinuity orientations. The contour plots are presented on Figure 6. A summary of discontinuity sets delineated and incorporated in the analysis of bench stability is presented in Table 7.

Table 7: Discontinuity Sets Delineated for Analysis

Set ID No. Dip Dip Direction

Mean Std. dev. Mean Std. dev.

A 496 60.7 11.0 227.0 19.4

B 473 85.0 4.4 124.9 17.1

C 281 85.3 4.7 74.8 9.4

D 130 88.4 2.7 357.9 11.5

E 1027 52.2 18.4 38.1 20.2

F 1350 58.1 12.4 311.7 26.1

G 445 23.5 7.1 277.1 45.2

H 279 40.1 14.4 117.9 16.4

6.2 Rock Mass Classification

Rock mass characterization is a largely empirical process of classification based on information obtained primarily from field data and enhanced with further data analysis and laboratory testing. The basic geotechnical parameters recorded for each core run were applied to the Laubscher (1990) In-situ Rock Mass Rating (IRMR) system, thereby creating a profile of IRMR with depth for each of the eight geotechnical holes drilled for this investigation. The Laubscher IRMR system consists of three primary parameters; intact rock strength (IRS), fracture frequency per meter (FF/m) and joint conditions (Jc). The individual parameters as well as the IRMR value out of a total of 100 for each run are displayed on the two 2009 geotechnical core logs presented in Appendix A. A large scale joint expression of slight undulation and dry conditions were assumed. It is appropriate to assign the groundwater parameter the full value when using rock mass rating systems as input to the Hoek-Brown (2002) shear strength criterion. Groundwater pressures are accounted for by using effective stress stability analyses.

The IRMR is typically adjusted to account for the expected mining environment, namely the influence of weathering, structural orientations, induced or changes to stresses and blasting to produce the Mining Rock Mass Rating (MRMR). The adjustments to the IRMR are introduced in recognition of the type of excavation proposed and the time dependant behavior of the rock mass. The potential for these adjustments were considered independently for this analysis and were not incorporated into the rock mass rating. A summary of IRMR values per domain is presented in Table 8.



Table 8: In-situ Rock Mass Rating (IRMR) Distributions

Domain Distribution Sample

No. Mean Std. Dev. Min Max

Hornfels Beta 904 47 14.3 14 81

Intrusives Weibull 516 48 12.7 15 80

471500 E

472000 E

472500 E

473000 E

473500 E

474000 E

474500 E

475000 E

6141000 N

6141500 N

6142000 N

6142500 N

K08-16

K08-14

K08-09

K09-07

K08-12

K08-06

K09-12

K08-04

LEGEND


CONTOUR PLOTS OF ORIENTED

CORE DISCONTINUITIES

K08-06

K08-12

K08-09

K08-14

K08-16

K08-04

740 POLES

MAXIMUM CONCENTRATION 5.1%

351 POLES


383 POLES


649 POLES


298 POLES


722 POLES


NOTE

1. PLOTS ARE LOWER HEMISPHERE, EQUAL AREA CONTOURED AS FISHER

CONCENTRATIONS (PERCENT OF TOTAL PER 1 PERCENT AREA)

2. PIT TOPOGRAPHY SHOWN IS WARDROP (2009) PRE-FEASIBILITY STUDY

3. GEOLOGY INTERCEPTS WITH PIT TOPOGRAPHY WERE ESTIMATED BASED

ON GEOLOGICAL MODEL PROVIDED AVANTI

X

X

X

X

X

X INDICATES MEAN DRILLHOLE

TREND AND PLUNGE

K09-07

828 POLES


K09-12

1011 POLES


X

X

X

HORNFELS UNIT

VARIOUS IGNEOUS INTRUSIONS COLLECTIVELY REFERRED TO

HEREIN AS THE INTRUSIVES UNIT.

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REVISION NO.



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NOV. 2010



6.3 Geotechnical Domains

A typical geotechnical model is composed of individual regions (domains), each of which is comprised of materials exhibiting internally similar geomechanical properties. Pertinent geotechnical parameters are assigned to each domain, based on engineering properties that are determined during field data collection and laboratory testing programs.

Based upon the IRMR as well as upon its individual components, available site geology information and laboratory test results, drill cores were divided into geotechnical intervals or domains that are expected to behave uniformly when exposed to open pit excavation-induced stresses. The materials at Kitsault were divided into two lithologic domains, i.e., intrusive and hornfels. The hornfels and intrusive domains are very similar in terms of discontinuity orientations; however, they possess distinctly different rock mass properties.

6.3.1 Hornfels Domain

The hornfels domain is generally a competent rock mass with zones of relatively intense fracturing and veining. This is evident in the bi-modal (two peaks) distribution of IRMR illustrated on Figure 4. From the geotechnical drillholes, these heavily fractured zones appeared to be concentrated around the outside of the intrusion; however, further examination using RQD data from all 2008 resource drillholes did not reveal a significant correlation between the heavily fractured zones and the distance from the intrusive core. As such, the hornfels were modeled as a single unit. Cross sections along each of the geotechnical drillhole traces showing RQD are presented in Figure 7.

6.3.2 Intrusives Domain

The intrusive domain is generally more massive and exhibits fewer signs of alteration and fracturing when compared to the hornfels domain and, consequently, possesses higher overall intact strength and IRMR values as illustrated in Figure 4.

The intrusive domain does contain intermittent zones of weaker material which typically correspond to intervals of increased fracturing, weathering and/or alteration, including minor fault zones and surface weathering. However, such intermittent weaker rock zones represent a relatively small portion of the overall intrusive rock domain and are not anticipated to adversely impact the performance of the fresh rock mass.

The geologic model was provided by Avanti as 3-dimensional wire-frames and was used to delineate the geotechnical domains. The intrusive domain and typical geotechnical cross sections are presented in Figure 8.

K08-04K08-09

K08-14

K08 06K08-06

K09-07 K09-12K09-12

K08-16

K08-12K08 12


DRILLHOLE RQD CROSS-SECTIONS


DATE:OCT. 2010

FIGURE NO.:7

SRK PROJECT NO.: 1CA020.004APPROVED:MEL

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Sector 1

n =45

Sector 4

PIT SLOPE EVALUATIONLegend:

GranodioriteGEOLOGIC MODEL AND GEOTECHNICAL

CROSS SECTIONS

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Granodiorite

Diorite

Hornfels



6.4 Rock Mass Shear Strength

The shear strength/normal stress relationship describes the ultimate shear strength available at a given point within a slope as a function of the effective normal stress acting on that point. Rock mass shear strength/normal stress relationships were developed for hornfels and intrusive domains using the Generalized Hoek-Brown criterion (Hoek, et al, 2002).

The Generalized Hoek-Brown criterion defines curvilinear shear strength envelopes that are considered effective representations of intact rock and heavily jointed rock mass behavior. Primary input parameters for the Generalized Hoek-Brown jointed rock mass criterion include the Geological Strength Index (GSI), an intact material constant (mi) and a rock mass disturbance factor (D), as defined by Hoek, et al, (2002). Probability density functions (PDF) were selected to represent stochastic (statistical) distributions of each of the primary parameters for each domain. The distributions selected were based upon the results of field and laboratory testing as well as upon SRK’s experience.

After the PDFs were selected to represent the three primary Generalized Hoek-Brown parameters (mi, GSI and D), Crystal Ball 7.3.2 (Crystal Ball), commercial software available from Oracle, was utilized to perform a large number of stochastic simulations, sampling each of the three parameter distributions during each simulation. Based on each set of primary parameters sampled, respective Hoek-Brown secondary parameters (mb, s and a) were calculated producing PDFs for each of the three, secondary parameters.

PDFs representing the UCS for each domain were also defined using a mathematical, “best-fit” technique available in Crystal Ball. The distribution types and defining parameters for the Hoek-Brown secondary parameters and for UCS selected for the analyses are summarized in Table 9.

Table 9: Secondary Hoek-Brown Parameters Stochastic Input

Domain Parameter Distribution Mean Std. Dev.

Min. Max.

Hornfels Hoek-Brown a parameter Gamma 0.5097 0.0102 0.5006 0.524

Hornfels Hoek-Brown mb parameter Lognormal 1.51 1.15 0.04 4.96

Hornfels Hoek-Brown s parameter Gamma 2.15-03 1.58E-02 0.00E+00 4.955E-02

Hornfels UCS (intact) MPa Beta 128 61 7 272

Intrusives Hoek-Brown a parameter Gamma 0.5112 0.0122 0.5001 0.5108

Intrusives Hoek-Brown mb parameter Gamma 1.5 1.3 0.00 8.21

Intrusives Hoek-Brown s parameter Lognormal 3.48E-03 4.62E-02 0.00E+00 1.425E-01

Intrusives UCS (intact) MPa Gamma 92 55 0 257

From the repeated, randomized samplings of the secondary Hoek-Brown parameters and UCS, distributions of the shear strength/normal stress relationships were calculated. Graphical representations of the range of shear strength/normal stress envelopes used by the model for each lithological domain are presented on Figures 9 and 10. In Figures 9 and 10, the 50%, 75% and 90% Upper and Lower Limits represent the ranges within which the shear strength lies, with 50%, 75% and 90% reliability, respectively.

10.0

8.0

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9.0

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h (

MP

a)

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engt

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6.5 Groundwater

Groundwater (porewater) pressure is an important component of slope stability. Porewater pressures act as buoyant forces in direct opposition to stabilizing forces, and as such, must be considered for the results of stability modeling to be realistic. A relatively free-draining slope will typically allow drawdown of the groundwater surface sufficiently deep within the slope so that porewater pressures are of minimal impact to slope stability. Since the rock mass comprising open pit benches has usually been moderately to highly disrupted by production blasting, such rock masses are usually free-draining and, in recognition, porewater pressures are seldom considered in bench scale stability analyses. However, deeper within rock masses that have been intensively weathered, altered and/or sheared, clay-filled discontinuities and/or faults are common, thereby compartmentalizing groundwater and resulting in a greatly reduced rock mass permeability. A lower permeability rock mass frequently inhibits free drainage, leading to a much steeper groundwater drawdown surface closer to the pit face. As a result, significant porewater pressures may be present on potential slip surfaces, thereby reducing effective normal stresses which, in turn, reduce resisting forces within the slope, and, consequentially, adversely impact the stability of the slope.

SRK (2009) conducted a pre-feasibility hydrogeology assessment for the proposed pit. The objective of SRK’s groundwater assessment was to characterize the hydrogeology of the proposed expansion of the Kitsault pit, to provide input to the geotechnical design for the proposed pit, to estimate pit inflow, and to assess probable inflow geochemistry.

The assessment was based on data collected from two holes drilled and instrumented in 2008, and two additional drillholes in the north-eastern and southern parts of the proposed pit. Results from the packer testing indicated that the rock mass tested has low bulk transmissivity, as is expected in the intrusive geological setting of the site.

Significant structural features, either intersecting or adjacent to the planned slopes, that could produce high inflows or residual destabilizing pressures during excavation, have not been identified in drilling to date or in Avanti’s geological model. This is supported by the 2009 drilling and testing program; however, this could not be ruled out and may be a focus of further work if additional structural interpretation does identify any major features in future studies.

The presence of upward hydraulic gradients (artesian flows) in the pit area was indicated by an elevated water table recorded in well K08-23 during sampling in July, 2009. Geotechnical drillhole K09-07 was drilled approximately 50m to the southeast of the K08-23 hole; the decision was made to install a well in K09-07 with a series of vibrating wire pressure transducers to allow pore pressures to be recorded. Artesian conditions were also encountered during drilling of geotechnical drillhole K09-12 and, consequently, a vibrating wire transducer string was also installed in that hole.

Both vibrating wire strings were equipped with data loggers to record transient changes in pressure over the inaccessible winter months. Once site access was re-established in the spring, the data was retrieved and analyzed to determine profiles of piezometric levels at each location. Results of the analysis are summarized in Figures 11 and 12 for K09-07 and K09-12 respectively.

The data generally show an increasing trend in piezometric pressure from the winter through the spring months which is believed to have occurred as a direct result of increased groundwater infiltration during the spring thaw. As such, the water levels that were calculated based on the May 2010 data readings were considered to be most representative of high ground water



conditions. When additional data from the summer and fall months is acquired from the data loggers, the data should be evaluated to verify that the May groundwater levels do in fact represent the “high water” conditions and that there isn’t a significant lag time between the spring thaw and related spikes in groundwater levels.

The calculated piezometric level for each transducer was plotted in section along the drillhole trace in order to estimate the current phreatic surface for stability modeling. The results and estimated phreatic surface for the northeast and south walls are shown on Figures 13 and 14, respectively.

Based on the relatively low in-situ hydraulic conductivities, SRK has assumed that depressurization will only occur naturally (without horizontal drains or wells) to a distance of up to approximately 20m to 60m behind the slope face for slope stability modeling. This zone of depressurization is expected to occur in response to blast induced fracturing and stress relaxation, both of which frequently result in increased apertures and lengths of joint systems near the face of large open pit slopes.

This assumption is based not only on SRK’s experience, but also on the experience of Hoek and Diederichs (2006), who suggest that the zone of heavy production blast damage and stress relieved rock can extend for 100m or more behind the crest of the slope. However, current mine plans for Kitsault include “pre-split” blasting for all final walls and temporary walls planned to stand for longer than 2 years. Consequently, the zone of disturbance and increased permeability has been reduced to a maximum distance of 20 to 60m for slope stability modeling.

6.6 Design Sectors

Slope angles within an open pit mine are influenced not only by geologic structure, rock mass strength and porewater pressures, but also by pit wall orientation and other operational considerations. The ultimate pit was divided into regions of similar structural characteristics and pit slope orientation called “design sectors”, delineating regions which are expected to exhibit similar response to pit development. Design sectors are shown on Figure 15.



Figure 11: Summary of vibrating wire piezometer data from K09-07 (El=595.71)

Figure 12: Summary of vibrating wire piezometer data from K09-12 (El=548.49)

NE SW

K09 07K09-07? ? Patsy Creek

Surface Weathering

GEOTECHNICAL PIT SLOPE REVIEW

Note: Pit geometry shown is the AMEC 2010) Feasibility ultimate pit design..

Legend:GROUNDWATER PRESSURES MEASURED IN HOLE K09-07


OCT. 2010 13MEL

KITSAULT MOLYBDENUM PROJECTSRK PROJECT NO.: 1CA020.004

FILE NAME: 196000.020REVISION NO:

A

Legend:Piezometric water levelVibrating wire transducer

N S

? ?

Patsy Creek

K09-12

Surface Weathering

GEOTECHNICAL PIT SLOPE REVIEWLegend:

Note: Pit geometry shown is the AMEC 2010) Feasibility ultimate pit design..


OCT. 2010 14MEL

SRK PROJECT NO.: 1CA020.004


A

KITSAULT MOLYBDENUM PROJECT

GROUNDWATER PRESSURES MEASURED IN HOLE K09-12

Legend:Piezometric water levelVibrating wire transducer

473000 E

473500 E

474000 E

6141000 N

6141500 N

6142000 N

6142500 N

LEGEND


PIT SLOPE DESIGN SECTORS

NOTE


FILE NAME:

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1CA020.004.Rev.A.Fig,15.Pit.Slope.Design.Sectors.2010-11-02.dwg

REVISION NO.



303-985-1333

T:\Kitsault British Columbia\!040_AutoCAD\Feasibility Pit Slopes Figures\Novemeber.2010.Updates\1CA020.004.Rev.A.Fig,15.Pit.Slope.Design.Sectors.2010-11-02.dwg

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7 Interramp/Overall Slope Stability Modeling Slope design involves analysis of the three major components of a pit slope, i.e., bench configuration, interramp angle and overall slope angle, all as defined on Figure 16. The bench configuration, which is controlled by the bench face angle, bench height, and berm width, defines the interramp angle. The overall slope angle consists of interramp sections separated by wide step-outs for haulage roads or mine infrastructure. The overall slope angles at Kitsault will be approximately equal to the corresponding, relative to design sector, interramp angles except in areas where a haul road exists on the slope or in those sectors in which both hornfels and intrusives are jointly present. In order to refine the recommendations of this study, a range of slope angles was analyzed.

SRK evaluated both global and bench scale stability for the proposed Kitsault open pit, where global failure is defined as one that occurs relatively deep through the rock mass, is pseudo-rotational, and is of sufficient scale to impact interramp and/or overall slopes. Bench scale failures typically involve only one or two bench levels and can be described as block type failures involving the translation of a block delineated by one or more structural features, such as discontinuities, within the rock mass. Techniques used by SRK for the interramp and overall slope analyses are presented in the remainder of this section. Details regarding bench scale stability analyses are presented in Section 8.

7.1 Model Methodology

The mathematical geotechnical model was input into the commercially available geotechnical modeling software packages Slide 5.039 (Rocscience, Inc., 2009) and Phase2 7.005 (Rocscience, Inc., 2009), developed by Rocscience, Inc.. Slide is a two-dimensional, limit equilibrium slope stability analysis program that analyzes slope stability by various methods of slices. Spencer’s method was selected for the limit equilibrium analyses of this evaluation due to its consideration of both force and moment equilibrium.

Phase2 is a two-dimensional, elasto-plastic finite element stress analysis code that yields a deterministic factor of safety by means of the shear strength reduction (SSR) technique. During the SSR process, the cohesion and friction angle of linear materials and the shear strength envelope of nonlinear materials are simultaneously reduced by a reduction factor until numerical convergence within the specified tolerance is no longer possible. The greatest SSR factor that allows convergence is considered the factor of safety against slope instability. The finite element method provides an alternative to limit equilibrium analysis which is based on assumptions regarding interslice forces and neglects constitutive relationships such as stress-strain behavior. In certain conditions, the finite element method more realistically models actual failure mechanisms by allowing the failure surface to implicitly emerge as strain occurs within the continuum during the shear strength reduction process.

Both limit equilibrium and finite element analyses allow for simulation of earthquake loading by application of static forces that represent seismic inertial forces resulting from potential ground accelerations caused by an earthquake. This method, known as pseudostatic analysis, simulates seismic forces in terms of a horizontal acceleration expressed as a coefficient (or percent) of gravity (g).

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Slope profiles were analyzed under static and pseudostatic conditions. Dynamic loading from potential earthquake ground acceleration was simulated using the pseudostatic technique and a Peak Ground Acceleration (PGA) expressed as a percent of gravity as previously discussed. According to the 2005 National Building Code of Canada (Institute for Research in Construction, National Research Council of Canada, 2005), the coefficient of peak horizontal acceleration that corresponds to a 10% probability of exceedance in 50 years is 0.070g for the Kitsault site.

When incorporating a PGA value as input into a slope stability model, it is common practice to reduce the PGA by a factor of 0.5 according to research conducted by the U.S. Army Corp of Engineers (Hynes-Griffen and Franklin, 1984). In summary, this reduction in horizontal acceleration is justified for earth and rock structures for the following reasons:

Realization that sustained ground acceleration is typically less than half of the PGA, which is an instantaneous acceleration; and,

Consideration that earth and rock structures effectively attenuate earthquake-induced accelerations.

Parameters describing each of the secondary parameter distributions (UCS, mb, s, and a) for each domain (Table 9) were directly input into slope stability modeling software. The Monte Carlo technique was used by the software during the analyses, randomly sampling secondary parameter and UCS distributions for each material type, yielding a normal stress/shear strength envelope for each set of parameters, for each geotechnical domain.

For each randomly generated strength envelope generated, a search of 1,331 circular surfaces (10 x 10 grid) was conducted by the software deterministically evaluating the ratio of available resisting strength to driving force (i.e. safety factor) for each valid slip surface. The critical slip surface (surface with the lowest safety factor) for each of the stochastically generated strength conditions was stored and used for calculation of the overall probability of failure. For the Rocscience software used in the limit equilibrium analyses, the overall probability of failure for a slope is defined as the percentage (of total valid samples) of valid critical surfaces yielding a safety factor of less than 1.0.

7.2 Results of Interramp/Overall Stability Analysis

The slope angles were optimized in terms of risk, i.e. probability of failure (PoF), to ensure that the design slope angles were optimal based on a quantitative evaluation of alternative designs. The PoF value incorporates the variations associated with the input parameters and the concept of risk into the design.

Based on accepted engineering experience, interramp/overall slope designs that yield probabilities of failure of up to 30% for slopes with low failure consequences and approximately 10% for high failure consequences are appropriate for most open pit mines. Slopes of high failure consequence are generally those slopes that are critical to mine operations, such as those on which major haul roads are established, those providing ingress or egress points to the pit, or those underlying infrastructure such as processing facilities or structures.

Given the wide range of interramp slope heights in the hornfels lithologic domain, interramp slopes were initially modeled for a variety of slope heights and interramp angles in order to develop a series of preliminary interramp slope design curves, as illustrated on Figure 17. The curves assumed a 50m depressurized zone behind the slope face based on rock mass disturbance that typically results from large scale production blasting (see Section 6.5). However, SRK



understands that mine plans now include pre-shear blasting for final walls and interim walls planned to stand for longer than two years. The expected reduction in rock mass disturbance resulting from pre-split blasting will reduce drainage somewhat but will also increase shear strengths. As such, the preliminary design curves were considered suitable for preliminary guidance of interramp slope angles.

Initial interramp slope angles (IRA) were selected for each sector based on the preliminary IRA design curves and bench stability considerations, as discussed in Section 8. Overall slope models were then constructed for the critical design sectors to confirm the stability of overall slopes. The overall slopes were modeled varying search limits to confirm stability of the high interramp slopes as well. Critical surfaces were also evaluated both at the toe of the slope and at the interface between the hornfels and intrusive domains.

Results of the analyses including the mean factor of safety (FoS) and probability of failure (PoF) are summarized in Table 10. A horizontal seismic coefficient of 0.035g was assumed for the pseudostatic analyses.

Table 10: Results of Overall Slope Stability Modeling

Sector IRA (deg) Overall Slope

Angle (°)

Overall Height

(m)

Static Pseudostatic

Mean FoS PoF Mean FoS PoF

1 48 45 525 1.6 4% 1.5 8%

2 48 46 410 N/A 1 N/A 1 N/A 1 N/A 1

3 52 50 370 1.5 5% 1.4 10%

4 50 45 420 1.7 3% 1.5 7%

5 54 46 425 N/A 2 N/A 2 N/A 2 N/A 2

6 54 45 335 1.7 2% 1.6 4%

7 54 50 290 N/A 3 N/A 3 N/A 3 N/A 3

8 52 48 250 1.7 3% 1.6 4%

9 56 56 195 2.1 1% 1.9 2%

10 54 56 210 N/A 4 N/A 4 N/A 4 N/A 4 1 The Sector 1 overall slope model was also used as the basis for Sector 2 design. 2 The Sector 4 overall slope model was also used as the basis for Sector 5 design. 3 The Sector 6 overall slope model was also used as the basis for Sector 7 design. 4 The Sector 9 overall slope model was used as the basis of Sector 10 design.

Results of the individual interramp and overall slope analyses are presented graphically in Appendix C, where the highlighted surface is the Critical Deterministic Surface, which is defined as the slip surface with the lowest safety factor when all the input parameters are equal to their mean values. The remaining surfaces shown on the diagrams in Appendix C correspond to the Global Minimum Surfaces, which correspond to the critical surfaces for each of the random samples of material parameters.

20% POF

25% POF

15% POF

10% POF

5% POF

Note:Probabilities of failure assume a 0.035 horizontal pseudostatic coefficient and a 50m depressurized zone behind slope face.

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8 Bench Design The consequences of an overall or high interramp slope failure on a final pushback commonly produce significant impact on mine economics, in that a substantial quantity of ore is frequently rendered uneconomic by the additional, unanticipated cost of removing the resulting failed wall material. The evaluation of the anticipated stability of final design slopes is therefore necessary and must be incorporated into final design recommendations. Of similar importance and impact on the project economics, though not nearly as dramatic as large scale slope failures, are the design and excavation of the benches and lower height interramp slopes, i.e., those slopes comprised of two to three benches. The overall slope designs cannot be achieved if benches cannot be safely and effectively established.

Although it was determined that the expected performance of the overall and higher interramp slopes comprising the proposed Kitsault open pit would best be predicted and subsequently examined using rock mass failure models, the anticipated behavior of the bench and lower interramp slopes was judged to be most realistically assessed using analytical models that incorporate structurally controlled failure mechanisms. This decision was made in recognition of the presence of the project’s pervasive and ubiquitous rock structure, i.e., joints and other non-fault discontinuities which will likely facilitate structurally controlled failures, and of the site materials’ relatively high rock mass strengths. The latter factor essentially precludes the development of rock mass failure in benches and in lower height interramp slopes; consequently, as described below, the evaluation of structurally controlled failure potential of benches and lower height interramp scale slopes played the predominant role in the formulation of bench design recommendations.

In recognition of the documented tolerance of lower-height slopes to earthquake-induced ground accelerations on the order of those reasonably expected at the site, only static analyses were conducted. It is anticipated that large scale production blasting typical to large open pits will induce additional fracturing of bench rock such that any groundwater will sufficiently drain from benches and; consequently, groundwater forces were not considered in the bench scale analyses.

8.1 Description of Models Used

Rock structure controlled primary failure mechanisms are generally simplified into one of three categories, i.e.:

Plane shear failure, defined as translation (sliding) of a failure mass on a single geologic structure oriented essentially parallel with the slope being modeled;

Simple wedge failure, defined as translation of a tetrahedral-shaped failure mass (bounded by the slope face, the essentially flat upper surface and two geologic structures, each oriented obliquely to the slope face) on either of the two lower bounding geologic structures or on both, i.e., down the line of their intersection; and

Step path failure, defined as a translation similar to the plane shear mode but due to insufficient length of the plane shear set to foster a through-going failure surface, the failure path must “step” up to another discontinuity of the same set via a sub-perpendicular discontinuity joint set failing in tension. This process of sliding and stepping occurs until a discontinuity daylights at the top of the bench.

In all instances, the failure mechanism is only viable if:

A potential failure mass exists, that is, if a geologic structure or structure set is present in a plane shear orientation relative to the slope being evaluated, or if two geologic structures or



structure sets are present and oriented relative to each other and to the slope being studied in such a manner that a wedge tetrahedron or step path is formed; and

Either the plane shear or step path surface(s), or the wedge line of intersection (between the two geologic structures) is “daylighted”, i.e., if the potential plane shear or step path surface(s) or the wedge intersection line intersects the slope above ground (bench) level, and the plunge of either, in the plane perpendicular to the slope face, is flatter than the slope angle.

Geologic structure controlled failure is only possible if discontinuities occur spatially so that a potential failure mass exists, and if the mass is unconstrained at the slope face. Once it is determined that a viable potential failure mass does exist, the likelihood of geologic structure controlled failure can be assessed by determining if the maximum shear resistance, which can be developed along the potential failure surface or surfaces, is greater than the driving forces acting to destabilize the mass.

The maximum available shear resistance is a function of both the structure continuity and its shear strength. For failure to occur along a discontinuous geologic structure, the non-fractured rock forming the intact portion(s) of the potential failure surface must fail. Intact rock strength for all but the weakest of rock tends to be higher than the stresses developed in bench and lower height interramp slopes; consequently, the strength of intact rock is rarely exceeded. Seldom do failures of such scale occur along surfaces that are not comprised of through-going, continuous geologic structures. The models used for bench and lower height interramp slopes within the proposed Kitsault open pit assumed that failures could only develop along geologic structures which were continuous, or in the case of step path failure, consisting of a combined continuous path through the slopes. The available shear resistance, then, becomes a function solely of the discontinuity shear strength along the surface, given that a continuous structure(s) are present.

With the assumption of the necessity of continuous geologic structure(s) relative to the slope under consideration, the likelihood of the failure of bench or lower height interramp slope simplifies, becoming a function of:

The likelihood that potential failure surface forming geologic structure(s) occur in the required orientation relative to the slope and, in the case of wedge and step path failures, to each other;

The likelihood that the shear stresses along the potential failure surface(s) exceed the maximum available shear resistance;

The likelihood of geologic structure(s) continuity or, in the case of step path failure, combined path continuity on the appropriate scale;

The likelihood of the potential failure surface(s) being daylighted; and

The number of such structure(s).

8.2 Methodology

The methodology used to determine the five components of the bench face and lower interramp slope likelihoods of failure is discussed below. All analyses conducted in the bench and lower interramp height slope stability modeling was conducted using Oracle’s Crystal Ball software.

8.2.1 Likelihood of Occurrence

Lower hemisphere, equal area polar plots depicting the orientations (dip directions and dips) of geologic structures measured during the 2008 and 2009 geotechnical core drilling programs were



developed for each of the rock types and drillholes for analysis. Upon visual inspection, taking into consideration the bias against encountering structures with orientations sub-parallel to the bearing of each corehole, it was determined that the variations in structure orientation concentrations between holes were insufficient to warrant a separation of the structures either laterally, within the geotechnical domains, or with depth. As a result of that determination, the structure orientations from all eight coreholes were combined to produce a project-wide geologic structure base, as depicted in Figure 18. Statistics for each discontinuity set are presented in Table 7.

The combined data set of discontinuities was divided into categories which, given sufficient persistence, had the potential to create structurally controlled failures. Plane shear and wedge type failures were evaluated for pit sectors assuming an average orientation of the pit walls in each sector. The potential wedge and plane shear forming discontinuity sets are summarized for each sector in Table 11.

Table 11: Summary of Potential Failure Forming Sets

Sector Mean DDR Sub-sector Potential

Plane Shear Set(s)

Potential Wedge

Sets

1 225 a & b A -

2 270 - - F/A

3 305 - F -

4 005 - - E/F ,F/C

5 050 - E E/B

6 090 - - E/H

7 135 a & b H A/H

8 180 - - A/H, A/B

9-1b 225 a & b A -

10-7b 145 a & b H A/H

Once the potential wedge and plane shear forming structures were delineated for each design sector, the correlation between dip and dip direction was calculated for each structure to determine if the distribution of dips was dependent upon dip direction. Invariably, the correlation was sufficiently low, thereby justifying the sampling of wedge set dips and dip directions as independent variables. Recognizing that the ranges of dip directions for the potential plane shear forming structures were limited by the definition of the set, it was similarly assumed that plane shear set dips were independent of dip direction.

The initial step in the analyses consisted of the simulation of a large number of potential plane shear surfaces and wedge geometries by Monte Carlo sampling of the dip distribution of the plane shear orientations and of the dip and dip direction distributions of the wedge sets members. In the case of the wedge geometries, the lines of their intersections were also calculated and, according to another common convention, if the line of intersection was within +/- 20º of being parallel with the mean face of the design sector, the geometry was considered to constitute that of a non-viable wedge. Recognizing one more common convention, all potential wedges whose dihedral angle (the angle formed between the “right” and the “left” wedge members) was less than 20º, thereby indicating a large surface area to mass ratio, were also considered to constitute non-viable wedges.

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8.2.2 Likelihood of Exceeding Shear Resistance

In accordance with common practice for these types of analyses, it was assumed that the shear strength along discontinuities consisted only of a frictional component, represented by the friction angle, and that there was no cohesion or apparent cohesion component. For the Kitsault project, the direct shear test results were used to approximate the distribution of friction angles of discontinuities as described in Sections 5.3 and 6.1.3.

The likelihoods of the shear resistance being exceeded were calculated by a Monte Carlo sampling of the friction angle distributions described above for each of the plane shear and wedge geometries sampled, per the methodology described in Section 8.2.1. In the case of the plane shear geometries, if the realization of the friction angle sampling exceeded that of the dip, the geometry was considered to be safe from sliding. Similarly, if the realization of the friction angle sampling exceeded the plunge of the line of intersection of a potential wedge geometry, that geometry was considered stable. The number of potential plane shear surfaces for which the dip of the surface exceeded the sampled friction angle and the number of potential wedges for which the plunge of the line of intersection exceeded its sampled friction angle were divided by the total number of surfaces of each type, to provide the likelihood that the shear resistance would be exceeded. This was done for each design sector judged to have potential for the respective failure modes. The distribution selected to represent the shear strength is presented on Figure 5.

8.2.3 Likelihood of Kinematic Admissibility

Fracture spacing (the perpendicular distance between two adjacent geologic structures of similar orientation) distributions were derived for each set from the oriented core data and fracture frequency observations made as part of the core logging. The mean fracture spacing calculated for each discontinuity set is summarized in Table 12.

Table 12: Summary of Discontinuity Set Spacings

Pit Area Sections of Core

Mean Spacings per Discontinuity Set (m)

A B C D E F G H

For W, SW, and S

Quadrants

Entire Hole Length

8.34 3.87 8.64 13.17 0.4 0.65 2.58 3.33

For N, NE, E, and SE Quadrants

“Fracture Zones”

0.44 1.11 4.08 3.77 0.21 0.28 0.54 0.69

Outside “Fracture Zones”

0.44 2.13 3.5 9.07 0.81 0.49 0.87 1.66

Entire Hole Length

0.44 1.545 3.43 5.65 0.38 0.38 0.7 1.06

Using the mean “fracture zone” spacings, exponential distributions were selected to represent discontinuity spacings for the plane shear structures and for both the “right” and “left” wedge structures for each set. The exponential distribution, which has only one parameter, i.e., the mean, is a commonly accepted method with which fracture spacings are characterized. For the Kitsault project, fracture lengths were also characterized with exponential distributions, an equally common practice.

Using Monte Carlo, random sampling from the exponential distributions of fracture spacing, large numbers of bench faces (typically 1000 or more), each with a unique pattern of either plane shear or “right” and “left” wedge structure intersections with the bench face were simulated. For the



plane shear analyses, at each point of intersection between the bench face and a plane shear structure, a dip was sampled from the best fit distribution to the actual data, a structure length was sampled from an exponential distribution, and a uniformly distributed (between 0 and 1) offset ratio was sampled. By multiplying the sampled length by the offset ratio, and considering the sampled dip, it could be determined whether a simulated fracture which intersected the bench face was sufficiently long to intersect the bench along the crest of the bench, thereby forming a viable failure geometry. This process was conducted at each intersection point. For those structures sufficiently long to form viable failure geometries, the maximum backbreak, as graphically defined in Figure 19, was determined for each simulation of a bench face. The distribution of the maximum backbreak from each simulation was then calculated and plotted for each design sector.

Similarly, virtual bench face simulations were performed for potential wedge failure geometries; the only significant difference being that lengths were sampled for both the “right” and “left” structures, and both had to be sufficiently long to extend from the face up to the crest of the bench to constitute a viable failure geometry. A distribution of backbreak from wedge development was also calculated and plotted for each design sector analyzed.

As a final step, for design sectors susceptible to both plane shear and wedge type failures, the plane shear and wedge backbreak distributions were composited to produce distributions of effective bench face angles representing both failure modes simultaneously.

Effective bench face angle distributions were produced for each design sector assuming mean structure lengths of 10m and 25m and double (20m high) benches. The two mean lengths (10m and 25m) were used to evaluate bench performance under likely conditions (10m mean length) and under the anticipated least advantageous conditions (25m mean length). These values were selected on the basis of available information on site geology and reconnaissance level observations of accessible outcrops exposed at the site. The development of better, more representative, estimates of the actual mean lengths of the geologic structure within the Kitsault pit were precluded by the relative lack of existing bench face exposures.

The double bench option was examined in recognition that, unless the mean lengths of the geologic structure appreciably exceed the height of the benches, bench stacking can produce remaining bench widths in excess of those that would be achieved with single benching. This results from the fact that, unless the geologic structures are long when compared with the bench height, it becomes less likely that a structure will be sufficiently long to daylight near the toe of a slope and extend to daylight at the crest of the slope, thereby delineating a viable failure. The fact that most open pit benches degrade primarily in their upper reaches and, usually, not from the toe of the bench to the top further demonstrates this principle.


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8.3 Results

Table 13 presents the results of the backbreak analyses, as the 50%, 95% and 98% probabilities that the noted effective bench face angles will be exceeded. For example, given a mean geologic structure length of 10m, the average effective bench face angle of double (20m high) benches in Sector 2 should be on the order of 70º, while along 95% of the benches, the effective bench face angles should be greater than approximately 52º, and, along 98% of the benches, the effective bench face angles should be greater than approximately 48º.

Table 13: Composited Results of Backbreak Analysis

Sector 10m Mean Discontinuity

Length 25m Mean Discontinuity

Length Anticipated Mode(s) of Failure

50% 95% 98% 50% 95% 98%

1 & 9 79 59 54 70 52 49 Plane Shear

2 70 52 48 51 42 41 Wedge

3 83 62 57 71 51 47 Plane Shear

4 68 52 49 51 39 36 Wedge

5 73 55 50 62 46 43 Wedge & Plane Shear

6 79 58 52 65 42 39 Wedge

7 & 10 73 56 50 60 43 38 Wedge & Plane Shear

8 70 56 52 54 43 40 Wedge

Common practice is to design the benches based on the 50% probability angles and to ensure that the overall/high interramp slopes are no steeper than the 95% or 98% angles, on the basis of the fact that, should the overall/high interramp angles exceed those, benches will too frequently degrade all the way back to the toe of the overlying bench, thereby leaving no catch bench and, in the worst case, jeopardizing the stability of the overlying bench.

As demonstrated in Table 13, effective bench face angles are heavily impacted by the mean discontinuity lengths. Results also indicate that beyond a mean length of about 25m, bench stacking may no longer be an effective means of increasing achievable bench face angles.



9 Pit Slope Design Recommendations For certain geologic environments, the combination of the average anticipated bench face angle and the preferred interramp angle, based on global (interramp/overall) stability considerations, alone, do not provide a sufficiently wide average catch bench width to effectively control rockfall and/or overbank slough accumulation. In such instances, recommended interramp angles are flattened sufficiently to provide adequately wide average catch benches, as was the case for Kitsault sectors 4 through 8 and 10. This was primarily determined by the analytic indications that a bench will be totally lost and the overlying bench will be undercut approximately 2% of the time. Pit slope design recommendations for each sector are summarized in Table 14 and shown on Figure 20.

Table 14: Summary of Pit Slope Design Recommendations and Expectations

Sector Max. Slope Height (m)

Interramp Slope

Angle (°)

Average Bench Face

Angle (°)

Bench Height (m)

Average Berm

Width (m)

1 520 48 79 20 14.1

2 425 48 70 20 10.7

3 370 52 83 20 13.1

4 430 50 68 20 8.7

5 425 54 73 20 8.4

6 345 54 73 20 8.4

7 370 54 73 20 8.4

8 350 52 70 20 8.4

9 195 56 79 20 9.6

10 210 54 73 20 8.4

Recommendations for interramp and overall slope angles are based on the assumption that the rock up to approximately 20m to 60m behind slope faces will be depressurized. Based on the current understanding of the rock mass conditions and mine plans, SRK feels that this range of depressurization should be achievable naturally, without having to resort to the implementation of active dewatering measures, i.e. horizontal drains and pumping wells. However, piezometers are recommended to be installed by the early stages of development to verify and monitor pit wall depressurization on a global scale as mining progresses.

If active measures are required to achieve sufficient depressurization, it will most likely be in the northeaster portion of the pit (Sectors 1 and 2) where, although not encountered by the recent hydrogeology programs, the potential for excess pore pressures exists due to steep topography northeast of the pit. Active depressurization measures would likely consist of a series of horizontal drains and/or vertical pumping wells behind pit wall crests.

473000 E

473500 E

474000 E

6141000 N

6141500 N

6142000 N

6142500 N

MAXIMUM INTERRAMP SLOPE

ANGLE RECOMMENDATIONS

NOTE


FILE NAME:

SRK JOB NO.:


1CA020.004.Rev.A.Fig,20.Pit.Slope.Design.Recommendations.2010-11-02.dwg

REVISION NO.



303-985-1333

T:\Kitsault British Columbia\!040_AutoCAD\Feasibility Pit Slopes Figures\Novemeber.2010.Updates\1CA020.004.Rev.A.Fig,20.Pit.Slope.Design.Recommendations.2010-11-02.dwg

A20

ML


KITSAULT


1CA020.004

NOV. 2010

48°

52°

*50°

54°

52°

56°

54°

LEGEND


HORNFELS UNIT

VARIOUS IGNEOUS INTRUSIONS COLLECTIVELY REFERRED TO

HEREIN AS THE INTRUSIVES UNIT.

* - INDICATES OPPORTUNITY TO INCREASE INTER-RAMP SLOPE ANGLES WITH BENCH MAPPING DURING PRODUCTION



10 Assessment of Future Geotechnical Work SRK recommends the following be completed during the detailed design and development stages:

Review of detailed design plans for the Patsy Creek diversion across the south pit wall. Additional stability analyses may be necessary at that time to confirm stability of final design.

Installation of a system of differential piezometers by early stages of development to verify and monitor pit wall depressurization during mine operation. There is potential to steepen interramp and overall slope angles in Sectors 1 and 2 if actual groundwater drawdown is greater than expected; and,

Institution and performance of a geologic structure bench mapping program to verify and optimize structural models used for the analyses, with special emphasis on the determination of structure persistence.

An ongoing geotechnical data collection program should be undertaken during pit development, to include geotechnical mapping and documentation of rock mass (particularly structure) and ground water conditions. The geotechnical data collection should concentrate on providing important data such as discontinuity persistence, spacing and variations in orientation that will allow further refinement of the bench design. The data collected should be used to confirm parameters used in the geotechnical models contained herein and to further refine the analyses providing more accurate estimates of anticipated slope behavior.

Analyses and recommendations presented herein are based on ultimate pit designs as described in this report, and, as such, any significant changes to mine plans or pit architecture should be reviewed by SRK to verify that recommendations will remain valid for the new mine plans.



11 References AMEC Earth and Environmental, Kitsault Project Feasibility Study, In progress.

Hoek E., Strength of jointed rock masses. Twenty-third Rankine Lecture, Geotechnique, 1983:23 (3):187-223.

Hoek E., Brown E.T., Practical Estimates of Rock Mass Strength. International Journal of Rock Mechanics and Mining Sciences & Geomechanics Abstracts, 1997:34 (8):1165-1186.

Hoek E., Carranza-Torres CT, Corkum B., Hoek-Brown Failure Criterion – 2002 Edition. In: Proceedings of the Fifth International North American Rock Mechanics Symposium, Toronto, Canada, Vol. 1, 2002. p. 267-273.

Hoek E., Diederichs M.S. Empirical Estimation of Rock Mass Modulus, International Journal of Rock Mechanics and Mining Sciences, 2006:43(2):203-215.

Hynes-Griffen, M.E. and Franklin, A.G.; Rationalizing the Seismic Coefficient Method; United States Army Corps of Engineers, Waterways Experiment Station, CWIS Work Unit 31145, 1984.

Institute for Research in Construction, National Research Council of Canada, 2005. The 2005 National Building Code of Canada.

International Society for Rock Mechanics Commission on Standardization of Laboratory and Field Tests, Suggested Methods for the Quantitative Descriptions of Discontinuities in Rock Masses, International Journal of Rock Mechanics, Mining Sciences and Geomechanics Abstracts, Vol. 15, 1978. P. 319-368.

International Society for Rock Mechanics Commission on Testing, Suggested Method for Determining Point Load Strength, International Journal of Rock Mechanics and Mining Sciences, Vol. 22, No. 2, 1985. p. 52-60.

Jaeger, J.C. Friction of rocks and stability of rock slopes. Geotechnique, 1971: v.21, n.2, 97-103.

Laubscher D.H., A geomechanics classification system for the rating of rock mass in mine design. Journal of South African Mining and Metallurgy, Vol. 90, No. 10, October 1990. pp 257-273.

Maptek Pty. Ltd., Lakewood, Colorado, 2008. Vulcan 7.5.0.798

Rocscience, Inc., Toronto, Ontario, 2003. Dips 5.106, visualization and analysis software for orientation based geologic data.

Rocscience, Inc., Toronto, Ontario, 2009. Slide 5.039, 2-dimensional limit equilibrium slope stability analysis software.

Rocscience, Inc., Toronto, Ontario, 2009. Phase2 7.005, 2-dimensional, elasto-plastic finite element stress analysis software.

SRK Consulting Inc., 2004. Annual Reclamation Report.



SRK Consulting Inc., NI 43-101 Preliminary Economic Assessment, Kitsault Molybdenum Property, Effective November 3rd, 2008.

SRK Consulting Inc., Pre-feasibility Geotechnical Pit Slope Design, Kitsault Molybdenum Project, June, 2009.

SRK Consulting Inc., Kitsault Pre-feasibility Study Pit Hydrogeology, December, 2009.

Steininger, Roger C., Geology of the Kitsault Molybdenum Deposit, March, 1981.

Wardrop, NI 43-101 Pre-feasibility Study – Avanti Mining Inc. Kitsault Molybdenum Property, British Columbia, Canada, December, 2009.

Appendices

Appendix A: Geotechnical Core Logs

25

50

75

100

125

150

25

50

75

100

125

150

DEP

TH -

ft

2525

DEP

TH -

m

OB

GDQM

DESCRIPTION

G1

SYM

BO

L

50 100

150

200

(MPa)

50 100

150

200

(MPa)

IRS (est.)UCS

PLT (CF=23)Alp

ha

Maj

orSt

ruct

ures

7.52.521.95.51.69.8

3.3

6.6

3.3

2.0

23.0

23.0

9.8

13.8

31.5

19.0

5.9

2.6

11.8

4.6

3.9

4.6

7.9

2.6

1.3

3.3

3.3

3.9

15.1

4.6

5.9

21.0

3.9

FF/m

151212101012

11

17

15

15

10

10

10

10

10

10

9

10

13

12

15

16

15

15

21

21

13

13

7

22

10

17

9

Join

t Con

ditio

nR

atin

g (4

0)

10 20 3010 20 30

FF/m

10 20 3010 20 30

FF/m

10 20 3010 20 30FF/m CJ+J

525539385645

49

50

45

49

25

25

33

26

19

30

46

48

29

40

50

52

41

55

68

59

39

46

26

52

38

28

43

RM

R

20 40 60 8020 40 60 8020 40 60 8020 40 60 80

646537296647

54

56

36

35

12

10

19

3

2

18

49

46

10

31

56

54

32

65

88

72

26

48

15

52

33

11

36

RMS (MPa)

25 50 75

(%)25 50 75

(%)

TCRRQD

6141945.85N

2009-08-302009-08-19

KITSAULT (1CA020.004)

Kitsault Feasibility Geotechnical Assessment

473534.52E

TO

OF

GEOTECHNICAL CORE LOG

1 9Kitsault, BC

STRATIGRAPHY

IRS: Intact Rock Strength (subjective)UCS: Uniaxial Compressive Strength (MPa)Pt LOAD: Point Load Test (MPa)FF/m: Fracture Frequency per m

TCR: Total Core RecoveryRQD: Rock Quality DesignationRMR: Rock Mass RatingRMS: Rock Mass Strength

GougeSheared

JointedBroken

DEFINITIONS LEGEND OF MAJOR STRUCTURES

C:\G

eote

c77L

OG

_ST

YLE

_FIL

E_L

90_M

L.st

y P

LOT

TE

D: 2

009-

12-0

1 15

:13h

rs

-60.00 40.00

PROJECT:

LOCATION:

SITE & PROJECT No:

BORING DATE:

AZIMUTH:

DATUM:

DIP:

BOREHOLE:

PAGE:

CLIENT HOLE ID:

COORDINATES:

DRILL: Diamond Drill

DRILL TYPE:

K09-07-GT

0 - 20 21

- 40

41 - 6

0

81 - 1

00

61 - 8

0

LEGEND OF RMR (90)

PLAN No:Nad83

90

175

200

225

250

275

300

325

175

200

225

250

275

300

325

DEP

TH -

ft

7575

DEP

TH -

m

HFF

GDQM

HFF

DESCRIPTIONSYM

BO

L

50 100

150

200

(MPa)

50 100

150

200

(MPa)

IRS (est.)UCS

PLT (CF=23)Alp

ha

Maj

orSt

ruct

ures

38.328.42.5

5.3

2.6

15.7

11.8

5.3

4.6

2.0

4.6

3.3

5.3

6.64.9

3.9

6.6

2.0

36.1

7.9

2.6

2.0

3.9

5.3

3.3

2.6

2.6

2.6

2.6

3.3

1.3

1.3

4.6

3.3

2.6

FF/m

109

16

10

13

9

11

13

9

10

12

12

10

1711

12

12

10

7

10

10

8

9

8

10

9

14

11

15

9

10

6

13

13

12

Join

t Con

ditio

nR

atin

g (4

0)

10 20 3010 20 30

FF/m

10 20 3010 20 30

FF/m

10 20 3010 20 30FF/m CJ+J

212455

42

51

31

25

46

33

40

38

42

42

484443

37

42

16

30

48

47

36

28

46

45

56

49

53

43

45

43

47

54

52

RM

R

20 40 60 8020 40 60 8020 40 60 8020 40 60 80

91255

33

58

19

9

45

9

16

21

33

33

474534

28

29

6

15

51

41

21

10

45

39

66

50

59

38

28

27

46

63

53

RMS (MPa)

25 50 75

(%)25 50 75

(%)

TCRRQD

6141945.85N

2009-08-302009-08-19



473534.52E

TO

OF


2 9Kitsault, BC

STRATIGRAPHY



GougeSheared

JointedBroken


C:\G

eote

c77L

OG

_ST

YLE

_FIL

E_L

90_M

L.st

y P

LOT

TE

D: 2

009-

12-0

1 15

:13h

rs

-60.00 40.00

PROJECT:

LOCATION:

SITE & PROJECT No:

BORING DATE:

AZIMUTH:

DATUM:

DIP:

BOREHOLE:

PAGE:

CLIENT HOLE ID:

COORDINATES:


DRILL TYPE:

K09-07-GT

0 - 20 21

- 40

41 - 6

0

81 - 1

00

61 - 8

0

LEGEND OF RMR (90)

PLAN No:Nad83

90

350

375

400

425

450

475

350

375

400

425

450

475

DEP

TH -

ft

125125

DEP

TH -

m

HFF

DESCRIPTIONSYM

BO

L

50 100

150

200

(MPa)

50 100

150

200

(MPa)

IRS (est.)UCS

PLT (CF=23)Alp

ha

Maj

orSt

ruct

ures

2.0

2.6

1.3

3.3

13.8

3.9

2.6

2.0

3.9

0.7

1.3

2.6

3.3

2.6

3.3

8.5

12.5

2.0

3.3

9.2

2.0

3.3

4.6

0.7

22.3

5.9

19.7

3.3

7.9

3.9

2.6

2.0

3.9

FF/m

17

9

9

17

12

19

13

10

10

7

9

13

9

10

12

11

10

11

10

10

10

10

10

14

10

9

10

8

10

14

13

23

13

Join

t Con

ditio

nR

atin

g (4

0)

10 20 3010 20 30

FF/m

10 20 3010 20 30

FF/m

10 20 3010 20 30FF/m CJ+J

60

43

46

43

25

58

53

54

52

53

52

51

41

46

50

34

26

58

53

29

46

44

42

64

21

42

20

29

28

40

41

58

36

RM

R

20 40 60 8020 40 60 8020 40 60 8020 40 60 80

70

36

32

23

7

114

81

90

76

48

51

52

36

40

70

24

13

75

77

16

39

40

37

71

7

37

2

6

10

12

9

38

14

RMS (MPa)

25 50 75

(%)25 50 75

(%)

TCRRQD

6141945.85N

2009-08-302009-08-19



473534.52E

TO

OF


3 9Kitsault, BC

STRATIGRAPHY



GougeSheared

JointedBroken


C:\G

eote

c77L

OG

_ST

YLE

_FIL

E_L

90_M

L.st

y P

LOT

TE

D: 2

009-

12-0

1 15

:13h

rs

-60.00 40.00

PROJECT:

LOCATION:

SITE & PROJECT No:

BORING DATE:

AZIMUTH:

DATUM:

DIP:

BOREHOLE:

PAGE:

CLIENT HOLE ID:

COORDINATES:


DRILL TYPE:

K09-07-GT

0 - 20 21

- 40

41 - 6

0

81 - 1

00

61 - 8

0

LEGEND OF RMR (90)

PLAN No:Nad83

90

500

525

550

575

600

625

650

500

525

550

575

600

625

650

DEP

TH -

ft

175175

DEP

TH -

m

LAMP

HFF

LAMPHF

DESCRIPTIONSYM

BO

L

50 100

150

200

(MPa)

50 100

150

200

(MPa)

IRS (est.)UCS

PLT (CF=23)Alp

ha

Maj

orSt

ruct

ures

2.0

0.7

1.3

3.9

3.9

2.0

3.9

2.0

2.0

6.6

2.0

3.3

0.7

1.3

0.7

0.7

2.0

2.6

2.0

0.7

0.7

0.7

2.0

2.0

0.7

9.2

3.9

2.6

3.3

3.9

3.3

2.6

4.6

FF/m

12

8

13

9

10

8

9

9

10

10

10

7

13

13

9

11

9

9

9

22

22

22

22

22

22

9

9

7

9

9

11

12

13

Join

t Con

ditio

nR

atin

g (4

0)

10 20 3010 20 30

FF/m

10 20 3010 20 30

FF/m

10 20 3010 20 30FF/m CJ+J

44

48

56

42

36

36

31

41

42

31

40

33

51

47

49

51

42

37

48

74

74

74

67

64

72

27

31

35

34

32

35

46

45

RM

R

20 40 60 8020 40 60 8020 40 60 8020 40 60 80

31

21

58

37

10

10

5

23

16

12

15

13

14

15

22

20

18

18

40

98

98

98

86

80

84

8

7

8

11

15

17

39

42

RMS (MPa)

25 50 75

(%)25 50 75

(%)

TCRRQD

6141945.85N

2009-08-302009-08-19



473534.52E

TO

OF


4 9Kitsault, BC

STRATIGRAPHY



GougeSheared

JointedBroken


C:\G

eote

c77L

OG

_ST

YLE

_FIL

E_L

90_M

L.st

y P

LOT

TE

D: 2

009-

12-0

1 15

:13h

rs

-60.00 40.00

PROJECT:

LOCATION:

SITE & PROJECT No:

BORING DATE:

AZIMUTH:

DATUM:

DIP:

BOREHOLE:

PAGE:

CLIENT HOLE ID:

COORDINATES:


DRILL TYPE:

K09-07-GT

0 - 20 21

- 40

41 - 6

0

81 - 1

00

61 - 8

0

LEGEND OF RMR (90)

PLAN No:Nad83

90

675

700

725

750

775

800

675

700

725

750

775

800

DEP

TH -

ft

225225

DEP

TH -

m

LAMP

HF

DESCRIPTIONSYM

BO

L

50 100

150

200

(MPa)

50 100

150

200

(MPa)

IRS (est.)UCS

PLT (CF=23)Alp

ha

Maj

orSt

ruct

ures

3.9

4.6

7.2

4.6

3.9

3.9

4.6

5.3

3.9

0.94.42.6

2.0

5.9

4.6

2.0

1.3

2.0

3.9

4.6

1.3

2.0

0.7

0.7

0.7

2.6

3.3

2.6

2.6

1.3

4.6

1.3

2.0

FF/m

10

11

10

8

9

8

10

11

11

221111

11

14

20

16

12

20

19

18

19

19

19

19

19

16

18

20

10

15

12

21

16

Join

t Con

ditio

nR

atin

g (4

0)

10 20 3010 20 30

FF/m

10 20 3010 20 30

FF/m

10 20 3010 20 30FF/m CJ+J

41

43

36

40

38

41

40

44

44

684349

54

46

58

60

59

64

58

56

66

66

73

73

73

58

56

62

54

62

52

70

63

RM

R

20 40 60 8020 40 60 8020 40 60 8020 40 60 80

37

39

29

33

29

37

29

42

40

753953

61

49

90

120

117

132

114

108

138

138

159

159

159

114

68

126

102

126

74

150

129

RMS (MPa)

25 50 75

(%)25 50 75

(%)

TCRRQD

6141945.85N

2009-08-302009-08-19



473534.52E

TO

OF


5 9Kitsault, BC

STRATIGRAPHY



GougeSheared

JointedBroken


C:\G

eote

c77L

OG

_ST

YLE

_FIL

E_L

90_M

L.st

y P

LOT

TE

D: 2

009-

12-0

1 15

:13h

rs

-60.00 40.00

PROJECT:

LOCATION:

SITE & PROJECT No:

BORING DATE:

AZIMUTH:

DATUM:

DIP:

BOREHOLE:

PAGE:

CLIENT HOLE ID:

COORDINATES:


DRILL TYPE:

K09-07-GT

0 - 20 21

- 40

41 - 6

0

81 - 1

00

61 - 8

0

LEGEND OF RMR (90)

PLAN No:Nad83

90

850

875

900

925

950

975

850

875

900

925

950

975

DEP

TH -

ft

275275

DEP

TH -

m

DESCRIPTIONSYM

BO

L

50 100

150

200

(MPa)

50 100

150

200

(MPa)

IRS (est.)UCS

PLT (CF=23)Alp

ha

Maj

orSt

ruct

ures

1.3

1.3

2.6

1.3

2.6

2.0

3.3

2.0

1.3

2.6

2.0

0.7

0.71.62.23.3

13.1

11.8

1.3

1.3

5.3

1.3

6.6

3.9

3.3

7.2

11.2

0.7

2.0

0.7

5.3

0.7

2.6

0.7

FF/m

19

12

21

28

11

17

12

20

12

5

10

10

13221617

11

12

15

19

12

21

13

17

17

20

19

22

23

22

22

24

19

22

Join

t Con

ditio

nR

atin

g (4

0)

10 20 3010 20 30

FF/m

10 20 3010 20 30

FF/m

10 20 3010 20 30FF/m CJ+J

66

61

61

75

53

59

52

62

61

49

57

64

67

705957

31

28

60

68

49

68

42

54

60

54

48

76

67

76

59

78

63

76

RM

R

20 40 60 8020 40 60 8020 40 60 8020 40 60 80

138

123

123

165

76

117

96

126

95

67

93

132

141

150117111

18

9

74

144

87

144

41

102

120

102

84

168

141

168

117

174

129

168

RMS (MPa)

25 50 75

(%)25 50 75

(%)

TCRRQD

6141945.85N

2009-08-302009-08-19



473534.52E

TO

OF


6 9Kitsault, BC

STRATIGRAPHY



GougeSheared

JointedBroken


C:\G

eote

c77L

OG

_ST

YLE

_FIL

E_L

90_M

L.st

y P

LOT

TE

D: 2

009-

12-0

1 15

:13h

rs

-60.00 40.00

PROJECT:

LOCATION:

SITE & PROJECT No:

BORING DATE:

AZIMUTH:

DATUM:

DIP:

BOREHOLE:

PAGE:

CLIENT HOLE ID:

COORDINATES:


DRILL TYPE:

K09-07-GT

0 - 20 21

- 40

41 - 6

0

81 - 1

00

61 - 8

0

LEGEND OF RMR (90)

PLAN No:Nad83

90

1000

1025

1050

1075

1100

1125

1000

1025

1050

1075

1100

1125

DEP

TH -

ft

325325

DEP

TH -

m

DESCRIPTIONSYM

BO

L

50 100

150

200

(MPa)

50 100

150

200

(MPa)

IRS (est.)UCS

PLT (CF=23)Alp

ha

Maj

orSt

ruct

ures

1.3

0.7

2.6

2.0

2.0

2.6

2.0

2.0

3.9

2.0

1.3

0.7

2.6

1.3

1.1

1.6

2.6

0.7

2.6

2.6

3.9

0.7

3.9

6.6

0.7

0.7

5.9

0.7

0.7

3.9

1.3

2.6

6.6

FF/m

15

30

11

13

10

10

10

16

13

17

11

11

11

10

16

10

10

22

18

13

20

19

19

19

22

19

18

16

22

19

19

20

19

Join

t Con

ditio

nR

atin

g (4

0)

10 20 3010 20 30

FF/m

10 20 3010 20 30

FF/m

10 20 3010 20 30FF/m CJ+J

62

84

53

57

54

52

52

60

50

61

60

65

53

57

62

58

52

76

62

55

59

73

58

54

76

67

54

70

76

61

68

62

54

RM

R

20 40 60 8020 40 60 8020 40 60 8020 40 60 80

97

134

83

70

82

78

96

120

90

95

120

135

99

70

77

114

96

168

126

105

117

159

114

67

168

69

102

150

168

123

144

126

102

RMS (MPa)

25 50 75

(%)25 50 75

(%)

TCRRQD

6141945.85N

2009-08-302009-08-19



473534.52E

TO

OF


7 9Kitsault, BC

STRATIGRAPHY



GougeSheared

JointedBroken


C:\G

eote

c77L

OG

_ST

YLE

_FIL

E_L

90_M

L.st

y P

LOT

TE

D: 2

009-

12-0

1 15

:13h

rs

-60.00 40.00

PROJECT:

LOCATION:

SITE & PROJECT No:

BORING DATE:

AZIMUTH:

DATUM:

DIP:

BOREHOLE:

PAGE:

CLIENT HOLE ID:

COORDINATES:


DRILL TYPE:

K09-07-GT

0 - 20 21

- 40

41 - 6

0

81 - 1

00

61 - 8

0

LEGEND OF RMR (90)

PLAN No:Nad83

90

1175

1200

1225

1250

1275

1300

1175

1200

1225

1250

1275

1300

DEP

TH -

ft

375375

DEP

TH -

m

DESCRIPTIONSYM

BO

L

50 100

150

200

(MPa)

50 100

150

200

(MPa)

IRS (est.)UCS

PLT (CF=23)Alp

ha

Maj

orSt

ruct

ures

0.7

0.7

0.7

0.7

3.9

2.0

2.6

1.3

4.6

2.0

0.7

2.0

1.3

3.9

3.9

1.3

1.3

2.6

6.6

3.9

1.3

3.3

0.7

9.2

5.9

3.3

2.6

1.3

1.3

2.0

3.3

0.7

2.6

FF/m

22

19

24

10

10

10

17

17

10

12

22

10

10

11

17

15

13

12

17

19

21

19

21

14

19

18

18

19

20

17

24

21

17

Join

t Con

ditio

nR

atin

g (4

0)

10 20 3010 20 30

FF/m

10 20 3010 20 30

FF/m

10 20 3010 20 30FF/m CJ+J

76

73

76

62

43

52

59

62

46

54

74

55

55

48

57

62

58

52

53

56

66

60

75

43

51

59

56

68

67

62

55

73

55

RM

R

20 40 60 8020 40 60 8020 40 60 8020 40 60 80

168

159

102

77

40

60

72

77

49

63

98

65

65

52

68

77

70

60

61

66

84

74

132

42

58

72

66

96

86

77

35

96

65

RMS (MPa)

25 50 75

(%)25 50 75

(%)

TCRRQD

6141945.85N

2009-08-302009-08-19



473534.52E

TO

OF


8 9Kitsault, BC

STRATIGRAPHY



GougeSheared

JointedBroken


C:\G

eote

c77L

OG

_ST

YLE

_FIL

E_L

90_M

L.st

y P

LOT

TE

D: 2

009-

12-0

1 15

:13h

rs

-60.00 40.00

PROJECT:

LOCATION:

SITE & PROJECT No:

BORING DATE:

AZIMUTH:

DATUM:

DIP:

BOREHOLE:

PAGE:

CLIENT HOLE ID:

COORDINATES:


DRILL TYPE:

K09-07-GT

0 - 20 21

- 40

41 - 6

0

81 - 1

00

61 - 8

0

LEGEND OF RMR (90)

PLAN No:Nad83

90

1325

1350

1375

1400

1425

1450

1475

1325

1350

1375

1400

1425

1450

1475

DEP

TH -

ft

425425

DEP

TH -

m

DESCRIPTIONSYM

BO

L

50 100

150

200

(MPa)

50 100

150

200

(MPa)

IRS (est.)UCS

PLT (CF=23)Alp

ha

Maj

orSt

ruct

ures

FF/m

Join

t Con

ditio

nR

atin

g (4

0)

10 20 3010 20 30

FF/m

10 20 3010 20 30

FF/m

10 20 3010 20 30FF/m CJ+J

RM

R

20 40 60 8020 40 60 8020 40 60 8020 40 60 80

RMS (MPa)

25 50 75

(%)25 50 75

(%)

TCRRQD

6141945.85N

2009-08-302009-08-19



473534.52E

TO

OF


9 9Kitsault, BC

STRATIGRAPHY



GougeSheared

JointedBroken


C:\G

eote

c77L

OG

_ST

YLE

_FIL

E_L

90_M

L.st

y P

LOT

TE

D: 2

009-

12-0

1 15

:13h

rs

-60.00 40.00

PROJECT:

LOCATION:

SITE & PROJECT No:

BORING DATE:

AZIMUTH:

DATUM:

DIP:

BOREHOLE:

PAGE:

CLIENT HOLE ID:

COORDINATES:


DRILL TYPE:

K09-07-GT

0 - 20 21

- 40

41 - 6

0

81 - 1

00

61 - 8

0

LEGEND OF RMR (90)

PLAN No:Nad83

90

25

50

75

100

125

150

25

50

75

100

125

150

DEP

TH -

ft

2525

DEP

TH -

m

OB

GDQM

LAMP

DESCRIPTION

G1

SYM

BO

L

50 100

150

200

(MPa)

50 100

150

200

(MPa)

IRS (est.)UCS

PLT (CF=23)Alp

ha

Maj

orSt

ruct

ures

11.2

3.3

3.9

5.3

1.3

10.5

0.7

1.3

8.5

1.3

0.7

2.6

40.0

3.9

3.3

3.3

22.3

5.3

3.3

4.6

3.9

3.9

2.6

5.3

2.6

10.5

4.6

4.6

FF/m

24

21

26

25

21

24

30

22

17

18

22

19

13

10

20

11

7

19

24

22

21

24

23

22

21

21

19

24

Join

t Con

ditio

nR

atin

g (4

0)

10 20 3010 20 30

FF/m

10 20 3010 20 30

FF/m

10 20 3010 20 30FF/m CJ+J

51

62

66

58

66

52

82

67

48

63

74

51

25

45

56

39

22

44

55

46

48

51

63

60

63

41

49

52

RM

R

20 40 60 8020 40 60 8020 40 60 8020 40 60 80

58

77

84

70

84

60

112

86

52

79

98

45

14

43

66

24

8

27

35

28

30

32

79

74

79

27

41

39

RMS (MPa)

25 50 75

(%)25 50 75

(%)

TCRRQD

6141611.59N

2009-08-092009-08-31



473249.24E

TO

OF


1 10Kitsault, BC

STRATIGRAPHY



GougeSheared

JointedBroken


C:\G

eote

c77L

OG

_ST

YLE

_FIL

E_L

90_M

L.st

y P

LOT

TE

D: 2

009-

12-0

2 09

:01h

rs

-60.00 180.00

PROJECT:

LOCATION:

SITE & PROJECT No:

BORING DATE:

AZIMUTH:

DATUM:

DIP:

BOREHOLE:

PAGE:

CLIENT HOLE ID:

COORDINATES:


DRILL TYPE:

K09-12-GT

0 - 20 21

- 40

41 - 6

0

81 - 1

00

61 - 8

0

LEGEND OF RMR (90)

PLAN No:Nad83

90

175

200

225

250

275

300

325

175

200

225

250

275

300

325

DEP

TH -

ft

7575

DEP

TH -

m

GDQM

LAMP

GDQM

DESCRIPTIONSYM

BO

L

50 100

150

200

(MPa)

50 100

150

200

(MPa)

IRS (est.)UCS

PLT (CF=23)Alp

ha

Maj

orSt

ruct

ures

3.3

7.2

3.9

3.9

8.5

3.3

7.2

12.5

0.7

2.6

3.3

1.3

1.3

2.0

1.3

0.7

0.7

3.3

0.7

1.3

3.9

3.3

4.6

2.6

2.0

2.6

3.9

4.6

3.3

2.6

1.3

2.6

2.6

FF/m

24

17

17

22

17

19

14

19

22

11

1

5

20

18

21

26

21

20

22

21

19

21

21

20

6

21

21

15

7

9

13

11

24

Join

t Con

ditio

nR

atin

g (4

0)

10 20 3010 20 30

FF/m

10 20 3010 20 30

FF/m

10 20 3010 20 30FF/m CJ+J

65

37

44

49

36

52

38

33

64

43

33

52

65

60

66

78

73

58

74

66

56

62

55

58

35

57

58

49

39

35

46

37

58

RM

R

20 40 60 8020 40 60 8020 40 60 8020 40 60 80

82

22

27

31

17

41

29

9

42

27

23

60

82

74

84

105

96

70

98

84

66

77

65

70

12

58

70

47

30

10

13

11

54

RMS (MPa)

25 50 75

(%)25 50 75

(%)

TCRRQD

6141611.59N

2009-08-092009-08-31



473249.24E

TO

OF


2 10Kitsault, BC

STRATIGRAPHY



GougeSheared

JointedBroken


C:\G

eote

c77L

OG

_ST

YLE

_FIL

E_L

90_M

L.st

y P

LOT

TE

D: 2

009-

12-0

2 09

:01h

rs

-60.00 180.00

PROJECT:

LOCATION:

SITE & PROJECT No:

BORING DATE:

AZIMUTH:

DATUM:

DIP:

BOREHOLE:

PAGE:

CLIENT HOLE ID:

COORDINATES:


DRILL TYPE:

K09-12-GT

0 - 20 21

- 40

41 - 6

0

81 - 1

00

61 - 8

0

LEGEND OF RMR (90)

PLAN No:Nad83

90

350

375

400

425

450

475

350

375

400

425

450

475

DEP

TH -

ft

125125

DEP

TH -

m

LAMPGDQM

LAMP

GDQM

DESCRIPTIONSYM

BO

L

50 100

150

200

(MPa)

50 100

150

200

(MPa)

IRS (est.)UCS

PLT (CF=23)Alp

ha

Maj

orSt

ruct

ures

2.6

3.3

3.3

3.3

2.6

23.0

28.2

22.3

7.2

2.0

1.3

0.7

2.6

3.3

1.3

2.0

5.3

0.7

1.3

3.3

10.5

4.6

3.3

3.3

4.6

3.9

2.0

3.3

3.3

2.0

4.6

1.3

2.0

FF/m

8

10

15

13

10

8

8

8

8

19

17

22

7

12

22

22

15

21

17

8

8

10

13

18

15

12

26

10

8

16

10

21

9

Join

t Con

ditio

nR

atin

g (4

0)

10 20 3010 20 30

FF/m

10 20 3010 20 30

FF/m

10 20 3010 20 30FF/m CJ+J

35

39

45

43

40

18

15

18

24

64

64

74

47

50

67

67

48

73

62

41

28

38

54

50

45

41

58

42

40

52

40

60

41

RM

R

20 40 60 8020 40 60 8020 40 60 8020 40 60 80

13

21

32

30

24

0

0

0

6

80

80

98

51

56

86

86

52

96

77

30

16

29

63

40

36

28

36

37

31

46

33

46

25

RMS (MPa)

25 50 75

(%)25 50 75

(%)

TCRRQD

6141611.59N

2009-08-092009-08-31



473249.24E

TO

OF


3 10Kitsault, BC

STRATIGRAPHY



GougeSheared

JointedBroken


C:\G

eote

c77L

OG

_ST

YLE

_FIL

E_L

90_M

L.st

y P

LOT

TE

D: 2

009-

12-0

2 09

:01h

rs

-60.00 180.00

PROJECT:

LOCATION:

SITE & PROJECT No:

BORING DATE:

AZIMUTH:

DATUM:

DIP:

BOREHOLE:

PAGE:

CLIENT HOLE ID:

COORDINATES:


DRILL TYPE:

K09-12-GT

0 - 20 21

- 40

41 - 6

0

81 - 1

00

61 - 8

0

LEGEND OF RMR (90)

PLAN No:Nad83

90

500

525

550

575

600

625

650

500

525

550

575

600

625

650

DEP

TH -

ft

175175

DEP

TH -

m

LAMPGDQM

DESCRIPTIONSYM

BO

L

50 100

150

200

(MPa)

50 100

150

200

(MPa)

IRS (est.)UCS

PLT (CF=23)Alp

ha

Maj

orSt

ruct

ures

2.0

3.3

3.3

3.3

3.3

2.6

1.3

2.0

1.3

2.0

3.3

0.7

1.3

1.3

0.7

0.7

2.6

0.7

0.7

0.7

6.6

0.7

0.7

0.7

1.3

0.7

0.7

2.0

0.7

5.3

2.6

3.3

6.6

FF/m

9

7

8

9

5

15

18

14

10

16

20

22

21

12

26

22

21

18

19

24

19

22

22

21

13

22

21

12

21

18

16

21

21

Join

t Con

ditio

nR

atin

g (4

0)

10 20 3010 20 30

FF/m

10 20 3010 20 30

FF/m

10 20 3010 20 30FF/m CJ+J

36

33

34

45

42

53

63

56

49

54

59

72

62

51

76

72

61

70

71

76

46

74

74

73

60

74

73

57

73

53

56

59

54

RM

R

20 40 60 8020 40 60 8020 40 60 8020 40 60 80

6

17

16

43

38

54

79

66

41

61

63

82

65

46

96

90

75

91

93

102

40

98

98

96

74

98

96

68

96

61

66

72

63

RMS (MPa)

25 50 75

(%)25 50 75

(%)

TCRRQD

6141611.59N

2009-08-092009-08-31



473249.24E

TO

OF


4 10Kitsault, BC

STRATIGRAPHY



GougeSheared

JointedBroken


C:\G

eote

c77L

OG

_ST

YLE

_FIL

E_L

90_M

L.st

y P

LOT

TE

D: 2

009-

12-0

2 09

:01h

rs

-60.00 180.00

PROJECT:

LOCATION:

SITE & PROJECT No:

BORING DATE:

AZIMUTH:

DATUM:

DIP:

BOREHOLE:

PAGE:

CLIENT HOLE ID:

COORDINATES:


DRILL TYPE:

K09-12-GT

0 - 20 21

- 40

41 - 6

0

81 - 1

00

61 - 8

0

LEGEND OF RMR (90)

PLAN No:Nad83

90

675

700

725

750

775

800

675

700

725

750

775

800

DEP

TH -

ft

225225

DEP

TH -

m

LAMP

GDQM

LAMPGDQM

DESCRIPTIONSYM

BO

L

50 100

150

200

(MPa)

50 100

150

200

(MPa)

IRS (est.)UCS

PLT (CF=23)Alp

ha

Maj

orSt

ruct

ures

3.9

2.6

1.0

3.3

1.3

0.7

2.6

1.3

3.0

2.0

2.6

3.0

3.0

3.9

2.6

0.7

1.6

FF/m

17

21

7

9

13

17

9

9

8

20

16

17

10

17

14

25

21

Join

t Con

ditio

nR

atin

g (4

0)

10 20 3010 20 30

FF/m

10 20 3010 20 30

FF/m

10 20 3010 20 30FF/m CJ+J

54

63

53

35

55

69

49

52

41

60

54

54

47

48

52

74

67

RM

R

20 40 60 8020 40 60 8020 40 60 8020 40 60 80

63

79

52

18

65

89

54

52

36

74

63

63

51

43

60

98

86

RMS (MPa)

25 50 75

(%)25 50 75

(%)

TCRRQD

6141611.59N

2009-08-092009-08-31



473249.24E

TO

OF


5 10Kitsault, BC

STRATIGRAPHY



GougeSheared

JointedBroken


C:\G

eote

c77L

OG

_ST

YLE

_FIL

E_L

90_M

L.st

y P

LOT

TE

D: 2

009-

12-0

2 09

:01h

rs

-60.00 180.00

PROJECT:

LOCATION:

SITE & PROJECT No:

BORING DATE:

AZIMUTH:

DATUM:

DIP:

BOREHOLE:

PAGE:

CLIENT HOLE ID:

COORDINATES:


DRILL TYPE:

K09-12-GT

0 - 20 21

- 40

41 - 6

0

81 - 1

00

61 - 8

0

LEGEND OF RMR (90)

PLAN No:Nad83

90

850

875

900

925

950

975

850

875

900

925

950

975

DEP

TH -

ft

275275

DEP

TH -

m

DESCRIPTIONSYM

BO

L

50 100

150

200

(MPa)

50 100

150

200

(MPa)

IRS (est.)UCS

PLT (CF=23)Alp

ha

Maj

orSt

ruct

ures

2.3

3.3

5.6

3.0

4.6

3.9

1.3

0.3

5.3

1.3

2.0

1.6

3.3

3.9

3.6

2.0

1.3

FF/m

22

19

19

14

21

21

22

19

20

21

12

19

9

8

16

17

12

Join

t Con

ditio

nR

atin

g (4

0)

10 20 3010 20 30

FF/m

10 20 3010 20 30

FF/m

10 20 3010 20 30FF/m CJ+J

54

50

53

51

55

56

67

76

53

66

52

62

41

41

51

62

57

RM

R

20 40 60 8020 40 60 8020 40 60 8020 40 60 80

44

29

61

58

65

66

86

102

61

84

60

77

36

36

58

77

68

RMS (MPa)

25 50 75

(%)25 50 75

(%)

TCRRQD

6141611.59N

2009-08-092009-08-31



473249.24E

TO

OF


6 10Kitsault, BC

STRATIGRAPHY



GougeSheared

JointedBroken


C:\G

eote

c77L

OG

_ST

YLE

_FIL

E_L

90_M

L.st

y P

LOT

TE

D: 2

009-

12-0

2 09

:01h

rs

-60.00 180.00

PROJECT:

LOCATION:

SITE & PROJECT No:

BORING DATE:

AZIMUTH:

DATUM:

DIP:

BOREHOLE:

PAGE:

CLIENT HOLE ID:

COORDINATES:


DRILL TYPE:

K09-12-GT

0 - 20 21

- 40

41 - 6

0

81 - 1

00

61 - 8

0

LEGEND OF RMR (90)

PLAN No:Nad83

90

1000

1025

1050

1075

1100

1125

1000

1025

1050

1075

1100

1125

DEP

TH -

ft

325325

DEP

TH -

m

DESCRIPTIONSYM

BO

L

50 100

150

200

(MPa)

50 100

150

200

(MPa)

IRS (est.)UCS

PLT (CF=23)Alp

ha

Maj

orSt

ruct

ures

1.3

0.7

1.3

1.0

0.3

0.3

1.3

3.0

1.5

0.3

1.0

14.4

5.6

6.2

9.2

4.6

FF/m

18

21

13

21

21

22

15

17

15

23

19

9

9

15

9

16

Join

t Con

ditio

nR

atin

g (4

0)

10 20 3010 20 30

FF/m

10 20 3010 20 30

FF/m

10 20 3010 20 30FF/m CJ+J

60

73

58

68

78

79

62

56

62

81

69

21

25

46

28

44

RM

R

20 40 60 8020 40 60 8020 40 60 8020 40 60 80

74

96

70

88

105

107

77

66

77

110

89

7

2

49

10

34

RMS (MPa)

25 50 75

(%)25 50 75

(%)

TCRRQD

6141611.59N

2009-08-092009-08-31



473249.24E

TO

OF


7 10Kitsault, BC

STRATIGRAPHY



GougeSheared

JointedBroken


C:\G

eote

c77L

OG

_ST

YLE

_FIL

E_L

90_M

L.st

y P

LOT

TE

D: 2

009-

12-0

2 09

:01h

rs

-60.00 180.00

PROJECT:

LOCATION:

SITE & PROJECT No:

BORING DATE:

AZIMUTH:

DATUM:

DIP:

BOREHOLE:

PAGE:

CLIENT HOLE ID:

COORDINATES:


DRILL TYPE:

K09-12-GT

0 - 20 21

- 40

41 - 6

0

81 - 1

00

61 - 8

0

LEGEND OF RMR (90)

PLAN No:Nad83

90

1175

1200

1225

1250

1275

1300

1175

1200

1225

1250

1275

1300

DEP

TH -

ft

375375

DEP

TH -

m

DESCRIPTIONSYM

BO

L

50 100

150

200

(MPa)

50 100

150

200

(MPa)

IRS (est.)UCS

PLT (CF=23)Alp

ha

Maj

orSt

ruct

ures

6.9

2.0

3.3

3.0

2.6

5.3

8.9

3.3

6.2

4.3

3.0

6.2

3.0

4.3

3.0

3.0

3.3

FF/m

20

9

19

19

17

14

11

14

14

14

12

10

7

11

13

13

14

Join

t Con

ditio

nR

atin

g (4

0)

10 20 3010 20 30

FF/m

10 20 3010 20 30

FF/m

10 20 3010 20 30FF/m CJ+J

51

49

57

56

57

47

28

44

35

42

51

33

36

41

46

48

52

RM

R

20 40 60 8020 40 60 8020 40 60 8020 40 60 80

58

50

68

66

68

51

13

30

20

28

58

20

24

36

42

45

60

RMS (MPa)

25 50 75

(%)25 50 75

(%)

TCRRQD

6141611.59N

2009-08-092009-08-31



473249.24E

TO

OF


8 10Kitsault, BC

STRATIGRAPHY



GougeSheared

JointedBroken


C:\G

eote

c77L

OG

_ST

YLE

_FIL

E_L

90_M

L.st

y P

LOT

TE

D: 2

009-

12-0

2 09

:01h

rs

-60.00 180.00

PROJECT:

LOCATION:

SITE & PROJECT No:

BORING DATE:

AZIMUTH:

DATUM:

DIP:

BOREHOLE:

PAGE:

CLIENT HOLE ID:

COORDINATES:


DRILL TYPE:

K09-12-GT

0 - 20 21

- 40

41 - 6

0

81 - 1

00

61 - 8

0

LEGEND OF RMR (90)

PLAN No:Nad83

90

1325

1350

1375

1400

1425

1450

1475

1325

1350

1375

1400

1425

1450

1475

DEP

TH -

ft

425425

DEP

TH -

m

LAMP

GDQM

DESCRIPTIONSYM

BO

L

50 100

150

200

(MPa)

50 100

150

200

(MPa)

IRS (est.)UCS

PLT (CF=23)Alp

ha

Maj

orSt

ruct

ures

4.3

3.9

3.3

3.0

3.6

3.6

7.5

4.3

3.0

6.6

4.9

3.3

3.6

3.0

8.2

2.0

FF/m

12

17

9

12

13

15

11

14

13

12

16

14

16

10

12

11

Join

t Con

ditio

nR

atin

g (4

0)

10 20 3010 20 30

FF/m

10 20 3010 20 30

FF/m

10 20 3010 20 30FF/m CJ+J

46

48

43

47

46

44

29

48

52

37

45

50

47

43

37

49

RM

R

20 40 60 8020 40 60 8020 40 60 8020 40 60 80

45

42

39

44

44

37

14

52

60

26

42

56

44

38

30

47

RMS (MPa)

25 50 75

(%)25 50 75

(%)

TCRRQD

6141611.59N

2009-08-092009-08-31



473249.24E

TO

OF


9 10Kitsault, BC

STRATIGRAPHY



GougeSheared

JointedBroken


C:\G

eote

c77L

OG

_ST

YLE

_FIL

E_L

90_M

L.st

y P

LOT

TE

D: 2

009-

12-0

2 09

:01h

rs

-60.00 180.00

PROJECT:

LOCATION:

SITE & PROJECT No:

BORING DATE:

AZIMUTH:

DATUM:

DIP:

BOREHOLE:

PAGE:

CLIENT HOLE ID:

COORDINATES:


DRILL TYPE:

K09-12-GT

0 - 20 21

- 40

41 - 6

0

81 - 1

00

61 - 8

0

LEGEND OF RMR (90)

PLAN No:Nad83

90

1500

1525

1550

1575

1600

1625

1500

1525

1550

1575

1600

1625

DEP

TH -

ft

475475

DEP

TH -

m

DESCRIPTIONSYM

BO

L

50 100

150

200

(MPa)

50 100

150

200

(MPa)

IRS (est.)UCS

PLT (CF=23)Alp

ha

Maj

orSt

ruct

ures

3.6

4.3

2.6

FF/m

12

11

9

Join

t Con

ditio

nR

atin

g (4

0)

10 20 3010 20 30

FF/m

10 20 3010 20 30

FF/m

10 20 3010 20 30FF/m CJ+J

49

41

47

RM

R

20 40 60 8020 40 60 8020 40 60 8020 40 60 80

54

36

51

RMS (MPa)

25 50 75

(%)25 50 75

(%)

TCRRQD

6141611.59N

2009-08-092009-08-31



473249.24E

TO

OF


10 10Kitsault, BC

STRATIGRAPHY



GougeSheared

JointedBroken


C:\G

eote

c77L

OG

_ST

YLE

_FIL

E_L

90_M

L.st

y P

LOT

TE

D: 2

009-

12-0

2 09

:02h

rs

-60.00 180.00

PROJECT:

LOCATION:

SITE & PROJECT No:

BORING DATE:

AZIMUTH:

DATUM:

DIP:

BOREHOLE:

PAGE:

CLIENT HOLE ID:

COORDINATES:


DRILL TYPE:

K09-12-GT

0 - 20 21

- 40

41 - 6

0

81 - 1

00

61 - 8

0

LEGEND OF RMR (90)

PLAN No:Nad83

90

Appendix B: Laboratory Testing

Uniaxial Compressive Strength Testing

Project # ClientDate LocationTechnician Sample #

Sample # Rock TypeDensity : 166.6 (pcf)

Fail Stress psi 2,668.0 (kg/m3)

MpaFail Stress 16,203 (psi)

Sample # : Modulus psi 111.7 Mpa

Rock Type: Poisson's

Hole # :Depth : Disp. Rate : 0.0003 (in/sec)Alterations: Load Rate : (lbs/sec)Diameter : 2.400 (in) Gage Reading : 73,300 (lbs)Height : 5.169 (in) Mode of Failure FractureWeight : 1022.40 (gm) Test Duration : (sec)Area : 4.524 (in2) 2:1 Correction : 1Volume : 23.384 (in3)

Fracture XX

Intact

Both

Dia. 1 2.400 Ht. 1 5.181Dia. 2 2.400 Ht. 2 5.169 Fail Load 73300 lbsDia. 3 2.400 Ht. 3 5.161Dia. 4 2.400 Ht. 4 5.166Dia. 5 2.400 Weight (gm) 1022.40Dia. 6 2.400 Sample # 07GT-01U

30.08-30.27

07GT-01U

Worksheet

Test Data:

Sample Data :

16,203

K09-07GT-01

111.74

07GT-01U

1CA020.00411/19/2009D.Streeter

07GT-01U

University of ArizonaGEOMECHANICAL LABORATORY

TUCSON, ARIZONA USA

Uniaxial Compression Test Results

Mode of Failure :

SRKKITSAULT

Pre-Failure Sketch Post-Failure Sketch







Hole # :Depth : Disp. Rate : 0.0003 (in/sec)Alterations: Load Rate : (lbs/sec)Diameter : 2.406 (in) Gage Reading : 45,700 (lbs)Height : 4.701 (in) Mode of Failure FractureWeight : 926.26 (gm) Test Duration : (sec)Area : 4.547 (in2) 2:1 Correction : 0.997157Volume : 21.375 (in3)

Fracture XX

Intact

Both


58.98-59.16

07GT-02U

Worksheet

Test Data:

Sample Data :

10,021

K09-07GT-02

69.11

07GT-02U

1CA020.00411/19/2009D.Streeter

07GT-02U


TUCSON, ARIZONA USA


Mode of Failure :

SRKKITSAULT









Fracture XX

Intact

Both



TUCSON, ARIZONA USA


Mode of Failure :

SRKKITSAULT07GT-04U

1CA020.00411/19/2009D.Streeter

07GT-04U

Test Data:

Sample Data :

8,232

K09-07-GT

56.78

121.13-121.31

07GT-04U

Worksheet








Hole # :Depth : Disp. Rate : 0.0003 (in/sec)Alterations: Load Rate : (lbs/sec)Diameter : 2.407 (in) Gage Reading : 104,000 (lbs)Height : 4.645 (in) Mode of Failure FractureWeight : 927.94 (gm) Test Duration : (sec)Area : 4.550 (in2) 2:1 Correction : 0.995662Volume : 21.136 (in3)

Fracture XX

Intact

Both



TUCSON, ARIZONA USA


Mode of Failure :

SRKKITSAULT07GT-05U

1CA020.00411/19/2009D.Streeter

07GT-05U

Test Data:

Sample Data :

22,757

K09-07-GT

156.94

154.99-155.18

07GT-05U

Worksheet









Fracture XX

Intact

Both



TUCSON, ARIZONA USA


Mode of Failure :

SRKKITSAULT07GT-06U

1CA020.00411/19/2009D.Streeter

07GT-06U

Test Data:

Sample Data :

7,157

K09-07GT-06

49.36

182.39-182.58

07GT-06U

Worksheet









Fracture XX

Intact

Both

Dia. 1 2.393 Ht. 1 5.384Dia. 2 2.394 Ht. 2 5.387 Fail Load 38800 lbsDia. 3 2.389 Ht. 3 5.400Dia. 4 2.391 Ht. 4 5.393Dia. 5 2.393 Weight (gm) 1113.77Dia. 6 2.392 Sample # 07GT-07E

244.57-244.76

07GT-07E

Worksheet

Test Data:

Sample Data :

8,635

K09-07-GT

59.55

6.86E+06

0.189

07GT-07E

1CA020.00411/19/2009D.Streeter

07GT-07E


TUCSON, ARIZONA USA


Mode of Failure :

SRKKITSAULT








Hole # :Depth : Disp. Rate : 0.0003 (in/sec)Alterations: Load Rate : (lbs/sec)Diameter : 2.400 (in) Gage Reading : 82,900 (lbs)Height : 5.144 (in) Mode of Failure BothWeight : 1114.72 (gm) Test Duration : (sec)Area : 4.524 (in2) 2:1 Correction : 1Volume : 23.268 (in3)

Fracture

Intact

Both XX



TUCSON, ARIZONA USA


Mode of Failure :

SRKKITSAULT07GT-09U

1CA020.00411/19/2009D.Streeter

07GT-09U

Test Data:

Sample Data :

18,326

K09-07-GT

126.39

338.42-338.67

07GT-09U

Worksheet








Hole # :Depth : Disp. Rate : 0.0003 (in/sec)Alterations: Load Rate : (lbs/sec)Diameter : 2.398 (in) Gage Reading : 74,100 (lbs)Height : 5.246 (in) Mode of Failure IntactWeight : 1020.92 (gm) Test Duration : (sec)Area : 4.519 (in2) 2:1 Correction : 1Volume : 23.704 (in3)

Fracture

Intact XX

Both



TUCSON, ARIZONA USA


Mode of Failure :

SRKKITSAULT12GT-01U

1CA020.00411/19/2009D.Streeter

12GT-01U

Test Data:

Sample Data :

16,399

K09-12-GT

113.10

34.66-34.90

12GT-01U

Worksheet









Fracture

Intact XX

Both



TUCSON, ARIZONA USA


Mode of Failure :

SRKKITSAULT12GT-02U

1CA020.00411/19/2009D.Streeter

12GT-02U

Test Data:

Sample Data :

27,484

K09-12-GT

189.55

90.10-90.30

12GT-02U

Worksheet









Fracture

Intact XX

Both

Dia. 1 2.402 Ht. 1 5.014Dia. 2 2.404 Ht. 2 5.021 Fail Load 120000 lbsDia. 3 2.405 Ht. 3 5.025Dia. 4 2.405 Ht. 4 5.017Dia. 5 2.404 Weight (gm) 1001.41Dia. 6 2.402 Sample # 12GT-03E

124.36-124.54

12GT-03E

Worksheet

Test Data:

Sample Data :

26,449

K09-12-GT

182.41

9.41E+06

0.258

12GT-03E

1CA020.00411/19/2009D.Streeter

12GT-03E


TUCSON, ARIZONA USA


Mode of Failure :

SRKKITSAULT









Fracture

Intact

Both XX


183.80-184.01

12GT-05U

Worksheet

Test Data:

Sample Data :

13,867

K09-12-GT

95.64

12GT-05U

1CA020.00411/19/2009D.Streeter

12GT-05U


TUCSON, ARIZONA USA


Mode of Failure :

SRKKITSAULT









Fracture

Intact

Both XX



TUCSON, ARIZONA USA


Mode of Failure :

SRKKITSAULT12GT-07U

1CA020.00411/19/2009D.Streeter

12GT-07U

Test Data:

Sample Data :

13,709

K09-12-GT

94.55

257.04-257.25

12GT-07U

Worksheet


Triaxial Compressive Strength Testing


Rock TypeDensity : 165.9 (pcf)

Sigma 3 Sigma 1 2,657.2 (kg/m3)(psi) (psi)

1,500 39,955 Peak

KITSAULT07GT-03T

1CA020.00411/25/2009D.Streeter Failure Data:

U.S. Standard

Residuals

Test Data:Sample Data :


TUCSON, ARIZONA USA

Triaxial Compression Test Results

Sample # 07GT-03T

SRK

07GT-03TSample # : 500 5,541 Disp. Rate : 0.0003 (in/sec)Rock Type 1,000 9,139 Load Rate : (lbs/sec)

Hole # : 1,500 12,395 Gage Reading : 181,000 (lbs)Depth : #VALUE! Mode of Failure IntactAlterations Test Duration : (sec)Diameter : 2.402 (in)Height : 5.007 (in)Weight : 987.61 (gm) Sigma 3 Sigma 1

Area : 4.530 (in2) (MPa) (MPa)Volume : 22.680 (in3) 10.34 275.5 Peak

K09-07-GT

Residuals

Metric Standard

87.94-88.15

Residuals

07GT-03T

3.45 38.26.90 63.0

10.34 85.5#VALUE! #VALUE!

Fracture

Intact XX

Both

Dia. 1 2.400 Ht. 1 5.005 Sigma 3 Fail Load

Dia. 2 2.399 Ht. 2 5.004 (psi) gage (lbs)Dia. 3 2.401 Ht. 3 5.010 1,500 181,000Dia. 4 2.407 Ht. 4 5.008 500 25,100Dia. 5 2.404 Weight (gm) 987.61 1,000 41,400Dia. 6 2.399 Sample # 07GT-03T 1,500 56,150

Mode of Failure :

Worksheet

Residuals

Pre-Failure Sketch Post-Failure

+




3,000 44,746 Peak

KITSAULT07GT-10T


U.S. Standard

Residuals



TUCSON, ARIZONA USA


Sample # 07GT-10T

SRK




K09-07-GT

Residuals

Metric Standard

366.68-366.85

Residuals

07GT-10T

3.45 38.06.90 53.2


Fracture

Intact XX

Both



Mode of Failure :

Worksheet

Residuals


+

Figure 1: Linear Failure Envelope

0

2000

4000

6000

8000

10000

12000

0 200 400 600 800 1000 1200 1400 1600 1800 2000

Fa

ilu

re S

tre

ss

(p

si)

Confining Stress (psi)

Test Points Linear Regression -1 STD +1 STD

+95% Conf. -95% Conf. Linear (-1 STD) Linear (+1 STD)

Linear Regression: Sig1 = A + B*Sig3A = 3175.0 (psi)B = 4.6060Phi = 40.03Cohesion = 739.69

Project: Sample: 07GT-10TRocktype:

CALL & NICHOLAS, INC.




2,500 45,838 Peak

KITSAULT12GT-08T


U.S. Standard

Residuals



TUCSON, ARIZONA USA


Sample # 12GT-08T

SRK




K09-12-GT

Residuals

Metric Standard

291.70-291.88

Residuals

12GT-08T

3.45 41.16.90 71.4


Fracture

Intact XX

Both



Mode of Failure :

Worksheet

Residuals


+

Figure 1: Linear Failure Envelope

0

2000

4000

6000

8000

10000

12000

14000

0 200 400 600 800 1000 1200 1400 1600 1800 2000

Fa

ilu

re S

tre

ss

(p

si)

Confining Stress (psi)


+95% Conf. -95% Conf. Linear (-1 STD) Linear (+1 STD)

Linear Regression: Sig1 = A + B*Sig3A = 2482.7 (psi)B = 7.4160Phi = 49.67Cohesion = 455.83

Project: Sample: 12GT-08TRocktype:

CALL & NICHOLAS, INC.

Date: 11/25/2009

Project: Number of Points: 3

Sample: 7GT-03T Mean A: 2171.00 S.D. A: 213.27

Rock Type: Mean B: 6.8540 S.D. B: 0.1975

Data File: Mean Phi: 48.19 Mean Cohesion: 414.63

R: 1.000 R 2 : 0.999

Sum X: 3000.00 Sum Y: 27075.00

Confining Failure Sum X 2 : 3500000.0 Sum Y 2 : 267860027.0

Test Stress Stress X Mean: 1000.0 Y Mean: 9025.0

Number (psi) (psi) X Var: 250000.0 Y Var: 11754076.0

1 500.00 5541.00 XY Var: 1713500.0 Std Err of Est.: 19494.0

2 1000.00 9139.00

3 1500.00 12395.00

Number of Points: 20

Mean A: 1990.49 S.D. A: 10.25

Mean B: 6.9336 S.D. B: 0.0123

Mean Phi: 48.41 Mean Cohesion: 377.96

R: 1.000 R 2 : 1.000

X Mean: 712.5 Y Mean: 6930.7

X Var: 196875.0 Y Var: 9465368.1

XY S.D.: 1365060.0 Std Err of Est.: 565.4


Mean A: 2351.51 S.D. A: 10.25

Mean B: 6.7744 S.D. B: 0.0123


R: 1.000 R 2 : 1.000

X Mean: 712.5 Y Mean: 7178.2

X Var: 196875.0 Y Var: 9035518.6


K: 56.9330

M: 0.7358

K: 31.0764

M: 0.8102

C: 765.5517

Laboratory Test Values

Linear Regression: Sig 1 = A + B*Sig 3

Linear - Minus 1 S.D. Tau

Power w/Intercept Regression: Y = C + KX M

Power Regression: Sig 1 = K*Sig 3M

Linear - Plus 1 S.D. Tau

Date: 11/25/2009

Project: Number of Points: 3

Sample: ALL Mean A: 35244.00 S.D. A: 4797.05

Rock Type: Mean B: 3.5496 S.D. B: 1.9862

Data File: Mean Phi: 34.08 Mean Cohesion: 9353.35

R: 0.873 R 2 : 0.762

Sum X: 7000.00 Sum Y: 130579.00

Confining Failure Sum X 2 : 17500000.0 Sum Y 2 : 5702926785.0

Test Stress Stress X Mean: 2333.3 Y Mean: 43526.3

Number (psi) (psi) X Var: 583333.3 Y Var: 9650852.3

1 1500.00 39995.00 XY Var: 2070583.3 Std Err of Est.: 4602337.8

2 2500.00 45838.00

3 3000.00 44746.00


Mean A: 30886.59 S.D. A: 162.98

Mean B: 4.8357 S.D. B: 0.0978


R: 0.996 R 2 : 0.993

X Mean: 1425.0 Y Mean: 37777.5

X Var: 787500.0 Y Var: 18550477.2



Mean A: 39601.41 S.D. A: 162.98

Mean B: 2.2634 S.D. B: 0.0978


R: 0.984 R 2 : 0.968

X Mean: 1425.0 Y Mean: 42826.8

X Var: 787500.0 Y Var: 4169972.4


K: 10788.0397

M: 0.1805

K: 10655.1447

M: 0.1820

C: 70.5624

Linear Regression: Sig 1 = A + B*Sig 3

Linear - Minus 1 S.D. Tau

Power w/Intercept Regression: Y = C + KX M

Power Regression: Sig 1 = K*Sig 3M

Linear - Plus 1 S.D. Tau

Laboratory Test Values

Direct Shear Testing

Date Project # 1CA020.004Technician Client SRK

Location NormalSample # Stress Friction Angle 39.19 degRock Type (psi) Cohesion 0.57 psiDrill Hole 25 124 56.1 12GT-02S ADepth 50 247 112.2 12GT-02S B

75 371 168.3 12GT-02S C K 0.8486 (for X in psi)Shear Plane 100 495 224.4 12GT-02S D M 0.9928Surface PrepShapeTest Speed 0.025 in/min K 0.8160 (for X in psi)Area 4.947 in 2 M 0.9998Diameter 2.510 in C 0.5579 psiRadius 1.255 inTilt Correction 0.668 deg

Sample:

Notes :DENSITY= 163.2 PCF

Load Filename(.dat) (Lbs) / (kgs)

Modified Power: Y = KX M + CInsitu

Digitized/Circular

11/13/09D.STREETER

Linear: Y = BX + C

Plot of Raw Trace Data

KITSAULT12GT-02S

Sample Data Trace Information

Power: Y = KX MK09-12-GT

139.96-140.21Test Data

Joint

University of ArizonaGeomechanical Laboratory

Tucson, Arizona USA

NormalResults

Area & Load Data for SSDS

University of Arizona

0

100

200

300

400

500

600

0 0.05 0.1 0.15 0.2 0.25 0.3

Horizontal Displacement (in)

Shea

r Fo

rce

(lbs)

A.LST (25 psi) B.LST (50 psi) C.LST (75 psi) D.LST (100 psi)

Project: SRK Sample: 12GT-02SRocktype:


0.167 in.105.0 lbf

0.135 in.190.0 lbf

0.132 in.305.0 lbf

0.129 in.395.0 lbf

0.0

10.0

20.0

30.0

40.0

50.0

60.0

70.0

80.0

90.0

100.0

0.0 20.0 40.0 60.0 80.0 100.0 120.0

Normal Stress (psi)

Shea

r St

reng

th (p

si)


Linear Regression: Y = C + BX C = 0.5668 (psi) B = 0.8152 Phi = 39.19



0.0

10.0

20.0

30.0

40.0

50.0

60.0

70.0

80.0

90.0

100.0

0.0 20.0 40.0 60.0 80.0 100.0 120.0

Normal Stress (psi)

Shea

r St

reng

th (p

si)

Test Points Power Regression -1 STD +1 STD

Power Regression: Y = KX M

K = 0.8486 M = 0.9928



0.0

10.0

20.0

30.0

40.0

50.0

60.0

70.0

80.0

90.0

100.0

0.0 20.0 40.0 60.0 80.0 100.0 120.0

Normal Stress (psi)

Shea

r St

reng

th (p

si)

Test Points ModPower Regression -1 STD +1 STD

ModPower Regression: Y = C + KX M

K = 0.8160 M = 0.9998 C = 0.5579 (psi)






Sample: 07GT-05S




Digitized/Circular

11/24/09D.STREETER

Linear: Y = BX + C


KITSAULT07GT-05S



243.29-243.54Test Data

Joint


Tucson, Arizona USA

NormalResults



0

100

200

300

400

500

600

700

800

900

1000

0 0.05 0.1 0.15 0.2 0.25 0.3


Shea

r Fo

rce

(lbs)




0.140 in.81.0 lbf

0.143 in.155.0 lbf

0.157 in.370.0 lbf

0.154 in.

0.0

20.0

40.0

60.0

80.0

100.0

120.0

140.0

160.0

180.0

200.0

0.0 50.0 100.0 150.0 200.0 250.0

Normal Stress (psi)

Shea

r St

reng

th (p

si)





0.0

20.0

40.0

60.0

80.0

100.0

120.0

140.0

160.0

180.0

200.0

0.0 50.0 100.0 150.0 200.0 250.0

Normal Stress (psi)

Shea

r St

reng

th (p

si)



K = 1.2180 M = 0.9097



0.0

20.0

40.0

60.0

80.0

100.0

120.0

140.0

160.0

180.0

200.0

0.0 50.0 100.0 150.0 200.0 250.0

Normal Stress (psi)

Shea

r St

reng

th (p

si)



K = 1.1707 M = 0.9174 C = 0.0263 (psi)






Sample: 07GT-03S




Digitized/Circular

11/24/09D.STREETER

Linear: Y = BX + C


KITSAULT07GT-03S



149.2-149.48Test Data

Joint


Tucson, Arizona USA

NormalResults



0

100

200

300

400

500

600

0 0.05 0.1 0.15 0.2 0.25 0.3


Shea

r Fo

rce

(lbs)




0.130 in.139.0 lbf

0.129 in.253.0 lbf

0.141 in.350.0 lbf

0.148 in.458.0 lbf

0.0

20.0

40.0

60.0

80.0

100.0

120.0

0.0 20.0 40.0 60.0 80.0 100.0 120.0

Normal Stress (psi)

Shea

r St

reng

th (p

si)





0.0

20.0

40.0

60.0

80.0

100.0

120.0

0.0 20.0 40.0 60.0 80.0 100.0 120.0

Normal Stress (psi)

Shea

r St

reng

th (p

si)



K = 2.0636 M = 0.8632



0.0

20.0

40.0

60.0

80.0

100.0

120.0

0.0 20.0 40.0 60.0 80.0 100.0 120.0

Normal Stress (psi)

Shea

r St

reng

th (p

si)



K = 1.4839 M = 0.9262 C = 4.5067 (psi)






Sample: 12GT-05S




Digitized/Circular

11/24/09D.STREETER

Linear: Y = BX + C


KITSAULT12GT-05S



284.52-284.70Test Data

Joint


Tucson, Arizona USA

NormalResults



0

100

200

300

400

500

600

700

800

0 0.05 0.1 0.15 0.2 0.25 0.3


Shea

r Fo

rce

(lbs)




0.158 in.92.0 lbf

0.163 in.168.0 lbf

0.177 in.344.0 lbf

0.197 in.720.0 lbf

0.0

50.0

100.0

150.0

200.0

250.0

300.0

0.0 50.0 100.0 150.0 200.0 250.0 300.0

Normal Stress (psi)

Shea

r St

reng

th (p

si)





0.0

50.0

100.0

150.0

200.0

250.0

300.0

0.0 50.0 100.0 150.0 200.0 250.0 300.0

Normal Stress (psi)

Shea

r St

reng

th (p

si)



K = 0.8937 M = 0.9582



0.0

50.0

100.0

150.0

200.0

250.0

300.0

0.0 50.0 100.0 150.0 200.0 250.0 300.0

Normal Stress (psi)

Shea

r St

reng

th (p

si)



K = 0.6884 M = 0.9998 C = 5.0340 (psi)



Brazilian Disk Tension Testing


Sample # Rock Type

T psi psi

Mpa

T psi 1,017 (psi)

Sample # : T= Indirect tensile strength 7.0 Mpa

Rock Type: Hole # :Depth : DENSI Disp. Rate : Alterations: Load Rate : 82 (lbs/sec)Diameter : 2.408 (in) Gage Reading : 4,590 (lbs)Length: 1.194 (in) Density: 166.3 pcf

Pre-existing Weakness PlanePost Failure Fracture

Dia. 1 2.406 Ht. 1 1.185Dia. 2 2.412 Ht. 2 1.195 Fail Load 4590 lbs ForceDia. 3 2.406 Ht. 3 1.201 Sample # 07GT-04B

Front view Back View


TUCSON, ARIZONA USA

Brazilian Disk Test Results

1CA020.00411/13/2009D.Streeter

SRKKITSAULT07GT-04B

07GT-04B

Test Data:

Sample Data :

1,017

K09-07-GT

7.02

121.13-12131

07GT-04B

Worksheet


+ +

-

- -


Sample # Rock Type

T psi psi

Mpa

T psi 2,544 (psi)


Rock Type: Hole # :Depth : Disp. Rate : Alterations: Load Rate : 76 (lbs/sec)Diameter : 2.412 (in) Gage Reading : 10,600 (lbs)Length: 1.100 (in) Density: 167.3 pcf





TUCSON, ARIZONA USA


1CA020.00411/13/2009D.Streeter

SRKKITSAULT07GT-05B

07GT-05B

Test Data:

Sample Data :

2,544

K09-07-GT

17.54

154.99-155.18

07GT-05B

Worksheet


+ +

-

- -


Sample # Rock Type

T psi psi

Mpa

T psi 2,139 (psi)







TUCSON, ARIZONA USA


1CA020.00411/13/2009D.Streeter

SRKKITSAULT07GT-09B

07GT-09B

Test Data:

Sample Data :

2,139

K09-07-GT

14.75

338.42-338.67

07GT-09B

Worksheet


+ +

-

- -


Sample # Rock Type

T psi psi

Mpa

T psi 1,622 (psi)







TUCSON, ARIZONA USA


1CA020.00411/13/2009D.Streeter

SRKKITSAULT12GT-02B

12GT-02B

Test Data:

Sample Data :

1,622

K09-12-GT

11.18

90.10-90.30

12GT-02B

Worksheet


+ +

-

- -


Sample # Rock Type

T psi psi

Mpa

T psi 1,199 (psi)



QUARTZ





TUCSON, ARIZONA USA


1CA020.00411/13/2009D.Streeter

SRKKITSAULT12GT-05B

12GT-05B

Test Data:

Sample Data :

1,199

K09-12-GT

8.27

183.80-184.01

12GT-05B

Worksheet


+ +

-

- -


Sample # Rock Type

T psi psi

Mpa

T psi 1,736 (psi)







TUCSON, ARIZONA USA


1CA020.00411/13/2009D.Streeter

SRKKITSAULT12GT-07B

12GT-07B

Test Data:

Sample Data :

1,736

K09-12-GT

11.97

257.04-257.25

12GT-07B

Worksheet


+ +

-

- -


Sample # Rock Type

T psi psi

Mpa

T psi 1,158 (psi)







TUCSON, ARIZONA USA


1CA020.00411/25/2009D.Streeter

SRKKITSAULT07GT-01B

07GT-01B

Test Data:

Sample Data :

1,158

K09-07-GT

7.99

30.08-30.27

07GT-01B

Worksheet


+ +

-

- -


Sample # Rock Type

T psi psi

Mpa

T psi 869 (psi)


Rock Type: Hole # :Depth : Disp. Rate : Alterations: Load Rate : 97 (lbs/sec)Diameter : 2.379 (in) Gage Reading : 4,620 (lbs)Length: 1.423 (in) Density: 169.8pcf


Dia. 1 2.379 Ht. 1 1.415Dia. 2 2.378 Ht. 2 1.419 Fail Load 4620 lbs ForceDia. 3 2.381 Ht. 3 1.435 Sample # 07GT-05sB

243.29-243.54

07GT-05sB

Worksheet

Test Data:

Sample Data :

869

K09-07-GT

5.99

SRKKITSAULT07GT-05sB

1CA020.00411/25/2009D.Streeter

07GT-05sB



TUCSON, ARIZONA USA



+ +

-

- -


Sample # Rock Type

T psi psi

Mpa

T psi 2,331 (psi)







TUCSON, ARIZONA USA


1CA020.00411/25/2009D.Streeter

SRKKITSAULT07GT-10B

07GT-10B

Test Data:

Sample Data :

2,331

K09-07-GT

16.08

366.68-366.85

07GT-10B

Worksheet


+ +

-

- -


Sample # Rock Type

T psi psi

Mpa

T psi 1,193 (psi)


Rock Type: Hole # :Depth : Disp. Rate : Alterations: Load Rate : 92 (lbs/sec)Diameter : 2.400 (in) Gage Reading : 6,030 (lbs)Length: 1.342 (in) Density: 164.1pcf





TUCSON, ARIZONA USA


1CA020.00411/25/2009D.Streeter

SRKKITSAULT12GT-01B

12GT-01B

Test Data:

Sample Data :

1,193

K09-12-GT

8.23

34.66-34.90

12GT-01B

Worksheet


+ +

-

- -


Sample # Rock Type

T psi psi

Mpa

T psi 611 (psi)





139.96-140.21

12GT-02sB

Worksheet

Test Data:

Sample Data :

611

K09-12-GT

4.21


1CA020.00411/25/2009D.Streeter

12GT-02sB



TUCSON, ARIZONA USA



+ +

-

- -


Sample # Rock Type

T psi psi

Mpa

T psi 1,599 (psi)







TUCSON, ARIZONA USA


1CA020.00411/25/2009D.Streeter


12GT-05sB

Test Data:

Sample Data :

1,599

K09-12-GT

11.03

284.52-284.70

12GT-05sB

Worksheet


+ +

-

- -


Sample # Rock Type

T psi psi

Mpa

T psi 1,732 (psi)







TUCSON, ARIZONA USA


1CA020.00411/25/2009D.Streeter

SRKKITSAULT12GT-08B

12GT-08B

Test Data:

Sample Data :

1,732

K09-12-GT

11.94

291.70-291.88

12GT-08B

Worksheet


+ +

-

- -

Appendix C: Slope Stability Modeling

Limit Equilibrium Modeling

Safety Factor00 Safety Factor1.900

2.033

2.167

2.300

2.433

2.567

2.700

2 833

160

1400

FS (deterministic) = 1.985FS (mean) = 1.624PF = 1.835%RI (normal) = 2.099




Overall Slope ResultsFS (mean) = 1.579PF = 3.261%

2.833

2.967

3.100

3.233

3.367

3.500+

1200

1000

RI (lognormal) = 2.579RI (lognormal) = 2.579

W

RI (lognormal) = 2.579RI (lognormal) = 2.579RI (normal) = 1.663RI (lognormal) = 1.988

800

Surface Weathering

W

W

600

400

200

0

GEOTECHNICAL PIT SLOPE DESIGN

0

-800 -600 -400 -200 0 200 400 600 800 1000 1200 1400 1600 1800 2000

Legend:SECTOR 1 (NORTHEAST)

STATIC ANALYSISAPPROVED: FIGURE:DATE:

OCT. 2010 A1MEL



A

Legend:HornfelsIntrusives

Safety FactorSafety Factor1.800

1.933

2.067

2.200

2.333

2.467

2.600

2 733

1600

1400

0.035

Overall Slope ResultsFS (mean) = 1.463PF = 6.828%

2.733

2.867

3.000

3.133

3.267

3.400+

1200

00

FS (deterministic) = 1.833FS (mean) = 1.486PF = 4.535%RI (normal) = 1.673RI (lognormal) = 1.948




RI (normal) = 1.273RI (lognormal) = 1.43110

080

0

Surface Weathering

W

600

400

W

200

GEOTECHNICAL PIT SLOPE DESIGNLegend:

0

-800 -600 -400 -200 0 200 400 600 800 1000 1200 1400 1600 1800 20

SECTOR 1 (NORTHEAST)PSEUDOSTATIC ANALYSIS


OCT. 2010 A2MEL



A


Safety Factor1.900

2.008

2.117

2.225

2.3331000

FS (deterministic) = 1.978FS (mean) = 1.575PF = 1.628%




2.442

2.550

2.658

2.767

2.875

2.983

3.092

3.200+

800

RI (normal) = 1.917RI (lognormal) = 2.313RI (normal) = 1.917RI (lognormal) = 2.313

W

RI (normal) = 1.917RI (lognormal) = 2.313RI (normal) = 1.917RI (lognormal) = 2.313Overall Slope Results

FS (mean) = 1.516PF = 5.145%RI (normal) = 1.537RI (lognormal) = 1.792

3.200+

600

Surface Weathering

W

400

200

0


-600 -400 -200 0 200 400 600 800 1000 1200 1400 1600

SECTOR 3 (SOUTHEAST)STATIC ANALYSIS


OCT. 2010 A3MEL



A


Safety Factor1.800

1.925

2.050

2.175

2.300

2.425

1000

0.035





Overall Slope Results

2.550

2.675

2.800

2.925

3.050

3.175

3.300+

800

W

Overall Slope ResultsFS (mean) = 1.417PF = 9.339%RI (normal) = 1.280RI (lognormal) = 1.422

600

Surface Weathering

W

400

00200


-400 -200 0 200 400 600 800 1000 1200 1400 1600

SECTOR 3 (SOUTHEAST)PSEUDOSTATIC ANALYSIS


OCT. 2010 A4MEL



A


Safety Factor2.000

2.158

2.317

2.475

2.633

2.792

1100

1000

FS (deterministic) = 2.015FS (mean) = 1.688PF 2 311%



FS (deterministic) = 2.015FS (mean) = 1.688PF 2 311%O ll Sl l

2.950

3.108

3.267

3.425

3.583

3.742

3.900+

900

800

PF = 2.311%RI (normal) = 1.683RI (lognormal) = 2.074


W




060

0

Surface Weathering

500

400

W

300

200

0


100

-600 -500 -400 -300 -200 -100 0 100 200 300 400 500 600 700 800 900 1000 1100 1200 1300 1400

SECTOR 4 (SOUTH)STATIC ANALYSIS


OCT. 2010 A5MEL



A


Safety Factor1.900

2.050

2.200

2.350

2.500

2.650

1100

1000

0.035

FS (deterministic) = 1.903FS (mean) = 1.560PF = 6.211%RI ( l) 1 518



FS (deterministic) = 1.903FS (mean) = 1.560PF = 6.211%RI ( l) 1 518O ll Sl R lt

2.800

2.950

3.100

3.250

3.400

3.550

3.700+

900

800


W



700

600

Surface Weathering

W

500

400

W

300

200

00


10

-600 -500 -400 -300 -200 -100 0 100 200 300 400 500 600 700 800 900 1000 1100 1200

SECTOR 1 (SOUTH)PSEUDOSTATIC ANALYSIS


OCT. 2010 A6MEL



A


Safety Factor00 Safety Factor2.100

2.308

2.517

2.725

2.933

3.142

3.350

1110

000

FS (d t i i ti ) 2 180FS (d t i i ti ) 2 180FS (d t i i ti ) 2 180FS (d t i i ti ) 2 180Overall Slope ResultsFS (mean) = 1 699

3.558

3.767

3.975

4.183

4.392

4.600+

900

800





FS (mean) 1.699PF = 2.477%RI (normal) = 1.610RI (lognormal) = 1.981

700

600

Surface Weathering

W500

400

W

300

200


100

-200 -100 0 100 200 300 400 500 600 700 800 900 1000 1100 1200 1300 1400 1500

SECTOR 6 (WEST)STATIC ANALYSIS


OCT. 2010 A7MEL



A


Safety Factor2.000

2.158

2.317

2.475

2.633

2.792

2.950

1000

900

0.035

FS (deterministic) = 2.059FS (deterministic) = 2.059FS (deterministic) = 2.059FS (deterministic) = 2.059Overall Slope ResultsFS (mean) = 1.676

3.108

3.267

3.425

3.583

3.742

3.900+

800

700





FS (mean) 1.676PF = 3.125%RI (normal) = 1.209RI (lognormal) = 1.427

600

0

W

5040

0

Surface Weathering

W

300

200

100

-200 -100 0 100 200 300 400 500 600 700 800 900 1000 1100 1200 1300


SECTOR 6 (WEST)PSEUDOSTATIC ANALYSIS


OCT. 2010 A8MEL



A


Safety Factory2.300

2.433

2.567

2.700

2.833

2.967

3.100

3 233

1000

900





Overall Slope ResultsFS (mean) = 1.703

2 990%

3.233

3.367

3.500

3.633

3.767

3.900+

800

700


W



600

500

W

400

0

Surface Weathering

300

200

100

-200 -100 0 100 200 300 400 500 600 700 800 900 1000 1100 1200 1300


SECTOR 8 (NORTH)STATIC ANALYSIS


NOV. 2010 A9MEL



A


0.035

Safety Factor2.200

2.317

2.433

2.550

2.667

2.783

2.900

1000

900

FS (deterministic) = 2.222FS (mean) = 1.681PF = 1 000%




Overall Slope ResultsFS (mean) = 1 602

3.017

3.133

3.250

3.367

3.483

3.600+

800

700



W




600

00

Surface Weathering

W

540

030

020

0


100

-300 -200 -100 0 100 200 300 400 500 600 700 800 900 1000 1100

SECTOR 8 (NORTH)PSEUDOSTATIC ANALYSIS


NOV. 2010 A10MEL



A


Finite Element Modeling

Critical SRF: 1.64

0.035

MaximumShear Strain

0.00e+0008.50e-0021.70e-0012.55e-0013.40e-001

1000

900

4.25e-0015.10e-0015.95e-0016.80e-0017.65e-0018.50e-0019.35e-0011.02e+0001.11e+0001.19e+0001.28e+000

980

070

0

1

1.28e+0001.36e+0001.45e+0001.53e+0001.62e+0001.70e+000

600

500

Surface Weathering 1

400

300

200

100


-100 0 100 200 300 400 500 600 700 800 900 1000 1100 1200 1300 1400 1500 1600 1700 1800

SECTOR 1 (NORTHEAST)PSEUDOSTATIC ANALYSIS


NOV. 2010 A11MEL



A

Critical SRF: 1.84MaximumShear Strain

0.00e+0002.00e-0014 00 001

700

1 0.035

4.00e-0016.00e-0018.00e-0011.00e+0001.20e+0001.40e+0001.60e+0001.80e+0002.00e+0002.20e+0002.40e+0002 60 000

600

500

2.60e+0002.80e+0003.00e+0003.20e+0003.40e+0003.60e+0003.80e+0004.00e+000

400

0

Surface Weathering

1300

200

100

0


-200 -100 0 100 200 300 400 500 600 700 800 900 1000 1100 1200 1300

SECTOR 3 (SOUTHEAST)PSEUDOSTATIC ANALYSIS


NOV. 2010 A12MEL



A


0.00e+0002.00e-0014 00 001

700

1

0.035

4.00e-0016.00e-0018.00e-0011.00e+0001.20e+0001.40e+0001.60e+0001.80e+0002.00e+0002.20e+0002.40e+0002 60e+000

600

2.60e+0002.80e+0003.00e+0003.20e+0003.40e+0003.60e+0003.80e+0004.00e+000

500

400


300

200

00


1

-200 -100 0 100 200 300 400 500 600 700 800 900 1000 1100

SECTOR 4 (SOUTH)PSEUDOSTATIC ANALYSIS


NOV. 2010 A13MEL



A


0.00e+0001.00e-0012.00e-001

60

1

0.0353.00e-0014.00e-0015.00e-0016.00e-0017.00e-0018.00e-0019.00e-0011.00e+0001.10e+0001.20e+0001.30e+0001 40 000

500

1.40e+0001.50e+0001.60e+0001.70e+0001.80e+0001.90e+0002.00e+000

400


300

00210

0


-100 -50 0 50 100 150 200 250 300 350 400 450 500 550 600 650 700 750 800 850 900

SECTOR 6 (WEST)PSEUDOSTATIC ANALYSIS


NOV. 2010 A14MEL



A

1


0.00e+0009.50e-0021.90e-0012 85 001

700

1 0.035

2.85e-0013.80e-0014.75e-0015.70e-0016.65e-0017.60e-0018.55e-0019.50e-0011.04e+0001.14e+0001.23e+0001 33 000

600

0 1.33e+0001.42e+0001.52e+0001.62e+0001.71e+0001.80e+0001.90e+000

500

400


300

200


-200 -100 0 100 200 300 400 500 600 700 800 900 1000

SECTOR 8 (NORTH)PSEUDOSTATIC ANALYSIS


NOV. 2010 A15MEL



A

KITSAULT MINE PROJECT ENVIRONMENTAL ASSESSMENT … · Figure 1: Site Location Map ... Preliminary Interramp Slope Design Curves: ... and current geologic solids provided by Avanti

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