Recycling

PART IX

RECOVERABLE RESOURCES OF METALLURGIC AND

ELECTROCHEMICAL INDUSTRY: TECHNOLOGICAL,

ENVIRONMENTAL AND ECONOMIC ASPECTS

THE SECOND INTERNATIONAL CONGRESS «NON-FERROUS METALS – 2010», SEPTEMBER 2–4, 2010, KRASNOYARSK, RUSSIA

The Second International Congress «Non-Ferrous Metals – 2010», September 2–4, Krasnoyarsk, Russia• Contents•

PART IX. RECOVERABLE RESOURCES OF METALLURGIC AND ELECTROCHEMICAL INDUSTRY: TECHNOLOGICAL, ENVIRONMENTAL AND ECONOMIC ASPECTS

The Approaches To The Increase of Iron-Containing Wastes Share . . . . . . . . . . . . . . . . . . . . . . . . . 480in The Raw Material BalanceN.I. Novikov

Comprehensive Anthropogenic Wastes Utilization of Southern Kuzbas Mining . . . . . . . . . . . . . . . 483and Smelting Complex, Problems and Perspectives. Exploring SituationF.I. Ivanov , E.V. Isakova , A.S. Golovko , V.A. Poluboyarov

Technological Studies of Slag Samples From Karsakpaysk Copper Smelter for Metal . . . . . . . . . . 488Re-Extraction Feasibility StudyS.G. Gritsay, G.I. Krivopustova, A.O. Teut, N.I. Utrobina

Sorption Recovery of PLatinum (II, IV) and Rhodium (III) From Chloride Solutions. . . . . . . . . . . 494of Spent CatalystsD.М. Кashirin, А.М. Мelnikov, О.N. Коnonova

Prospects of Use of The Nonferrous and Rare Metals Containing in Coals and Coal Ashes . . . . . . 503of Kuzbas for The Iron and Steel Industry of Siberian RegionV.A. Salikhov, E.S. Ljubushkina

Rhenium Extraction From Nickel-Based Complex Heat-Resistant Alloys . . . . . . . . . . . . . . . . . . . . 510A.G. Kasikov, A.M. Petrova, V.T. Kalinnikov

«EPOS-Process» – New Technology of Effective Ore-Smelting and Industrial . . . . . . . . . . . . . . . 515 Wastes Processing in Plasma Ore-Smelting Shaft FurnaceI.A. Bezrukov, S.N. Malyshev, O.B. Moiseyev, V.V. Pavlov, I.S. Parhomuk, А.P. Kuznetsov

Treatment of Radioactive Metallic Wastes by Melting . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 520I.E. Abroskin, U.N. Makaseev, A.C. Buynovsky, A.I. Abroskin, A.A. Chernoshchuk

Processing of Fluorine-Containing Waste Sand Middling Products of Aluminum Production . . ..525in Cement Industry B.P. Kulikov,V.D. Nikolaev,S.A.Ditrich, L.M. Larionov

Effective Strength Resource Saving Technology in Steel Industry Revisited . . . . . . . . . . . . . . . . . . 533N.I. Novikov

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The Second International Congress «Non-Ferrous Metals – 2010», Krasnoyarsk• Part IX • Recoverable Resources of Metallurgic and Electrochemical ...The Second International Congress «Non-Ferrous Metals – 2010», Krasnoyarsk• Part IX • Recoverable Resources of Metallurgic and Electrochemical ...

In ferrous metal industry annually millions of iron-containing wastes appear, such as oxide

scale, sullage, dust. About 45–50 % of iron, which wastes contain, go back to production mainly

through agglomerating and converter production, the rest 50–55 % are whether accumulated in

waste dumps and storage ponds, or lost outside the plants. Bringing back to the production

the iron, which wastes contain, is an important economic aim.

In metallurgical production iron-containing wastes formation take place at all stages of

production process, starting from preparing iron ore raw materials to production of finished

steel. The sources of iron-containing wastes formation and their volume are specified by the

example of convert production of Western Siberian Steelworks and are shown in the table 1.

Table 1

The sources and volume of iron-containing wastes at Western Siberian Steelworks

The sources of iron-containing wastes formation The volume of iron-containing

wastes, thousands

Converter production with productivity 7.5 million tones of steel,

Including

– oxide scale of continuous-casting machine

– gas-cutting machine tailing of continuous-casting machine

– gas-cleaning tailing

– gas-cleaning dust

86.7

20.4

22.5

12.5

31.25

Blast-furnace department, including

– tailing of aeration devices in the stockhouse

– gas-cleaning tailing

– flue dust

539.6

7.7

56.9

475.0

Agglomerative and preparation plant (minus sieve of sinter

less than 5 mm)

312.5

Oxide scale of mill products 203.75

Total 1142.55

The characteristic of iron-containing wastes on some parameters is shown in table 2.

Table 1

The characteristic of iron-containing wastes

Item of materials Iron content, % Fraction, mm Humidity, %

Oxide scale of continuous-casting machine 72.0 0.5–5.0 10–12

Tailing of gas-cutting machines 60.0 0.1–3.0 6–8

Continuous-casting machine

Flue dust 48.5 0.1–2.0 6–12

Tailing of converter department gas-cleaning 59.7 0.1–2.0 30–40

Tailing of blast-furnace department gas-cleaning 58.5 0.1–2.0 12–20

Oxide scale of rolling-mill department 69.0 0.5–5.0 16–12

Minus sieve of sinter 55.4 up to 5.0 up to 6.0

THE APPROACHES TO THE INCREASE OF IRON-CONTAINING WASTES SHARE IN THE RAW MATERIAL BALANCE

N.I. Novikov

Novokuznetsk branch of the institute state educational highest vocational institution,

Kemerovo State University, Novokuznetsk, Russia

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According to the experts’ estimation, a tone of iron, produced from industrial wastes is

5–7 times cheaper than a tone of iron, produced from original raw material (iron ore).

Iron ore mining costs rise every year because of geological conditions deteriorating, re-

moteness from steelworks and other factors.

In this respect, manufactures of iron have started to tackle more actively the problems of

using iron-containing wastes in production process. This works in the most active way at Mag-

nitigorsk steelworks (open joint-stock company MMK), open joint-stock company Ural Steel,

open joint-stock company Western Siberian steelworks (open joint-stock company ZSMK) and

a number of other enterprises. Iron-containing wastes share in the raw material balance is from

7 % to 13 % and it is not the maximum.

According to the experts’ estimation, the iron share in the raw material balance from

the volume of accumulated iron-containing wastes in enterprises’ dumps is quite possible to in-

crease 2 or 2.5 times for salvage of iron-containing wastes.

For efficient usage of these wastes in steel industry it is necessary to prepare them be-

forehand. The preparation includes tailing dehydration and homogenization. This requires the

building of a special department.

There are a lot of various ways of tailing utilization well-known worldwide. The easiest

and at the same time appropriate for steel industry option is the sequence of the following pro-

cedures: pulp thickening, drying on disk vacuum filters, then on drier drums to 6 % humidity.

Tailing drying is made through gas fuel burning, or with hot air from the hot-blast stove behind

the waste-heat boiler of the oxygen converter.

So, using the pre-prepared tailings in converter procedure is technologically possible and

economically reasonable, which is proved by local and foreign practice, including at ZSMK. At

the same time, conveying a large number of finely dispersed materials with the particles size

0.2–5.0 mm through the solids pipe will lead to significant increase in dust formation in the

convey areas and especially in the block of conveying the materials directly to the converter. On

this basis conveying iron-containing and carbon-bearing materials to the converter needs to be

made through the containers, similar to those used for loading of scrap metal to the converter.

The load is advised to be made in the department of iron-containing materials preparing.

Currently, the solution to the problem of more comprehensive utilization of iron-contain-

ing wastes is held by the lack of reliable and efficient technologies of recycling considerable

amounts of oxidized secondary raw materials which are poor-graded for agglomerating. Actu-

ally, only oxide scale is comprehensively utilized. For example, at open joint-stock company

MMK annually about 650,000 of oxide scale is formed and used in blast-furnace practice. On the

other hand, there is an option of more efficient using of oily oxide scale in converters, and it has

a number of technological, ecological and economical benefits.

Decrease in oxide scale, especially oily oxide scale (about 150,000 tones) from the agglomer-

ated charging material will contribute to saving fuel for the technological needs (60–70 kg/t of

agglomerate), increase in the equipment operating reliability and decrease in harmful wastes in

the atmosphere, as the oils in agglomeration process mainly sublime, but don’t burn. Then, during

agglogases cooling, oil vapor condensate on the equipment surface as solid particles, and partly go

to the atmosphere as oily fog and afterwards fall down and pollute the soil. Exclusion of oily oxide

scale through agglomeration from charging material will improve the operating of stoves, as the

permeability will increase and the moisture content of charging material will stabilize.

Oil content in oxide scale is not a big drawback when it is used with converter technol-

ogy, as oil is technological fuel and deoxidant and intensifies deoxidant processes in the ox-

ide scale. Certainly, one will have to deal with preparing oxide scale for melting in converters,

but the effect of bringing to production 150,000 tones of raw material annually, which is obvi-

ous in the iron ore raw material balance, will cover the costs. There are no difficulties with using

non-oily oxide scale and welding slag, which is proved by the practice of their efficient using in

2000–2005 at the Novokuznetsk steelworks. During the period at open joint stock company

ZSMK a version of converter process was applied with using a large amount of oxide scale (up to

15–20 % of hot metal weight), which is two or three times a s big as its regular share in converter

melting charging material.

Heat, which is necessary for the process for heating and deoxidation of oxide scale, is pro-

duced as a result of cast iron residual element oxidation and decrease in solid carbon-bearing

materials, mainly small-sized coke.

Technological process takes place in converter, equipped with blow-off device for both

joint and separated conveying of oxygen and azote. Cast iron is put into the converter, oxide

482


scale is loaded, also coal-bearing and slag-forming materials with simultaneous stirring the bath

through conveying blow-off mixture of oxygen and azote. The ratio of the blow-off mixture in-

gredients depends on the cast iron content and temperature, and also the oxide scale.

As a result of mixing and interaction of all the materials, loaded to the converter, iron oxide

deoxidation takes place and also transit of deoxidated iron into iron-carbon melt. The change

of conveying the blow-off mixture of oxygen and azote into the converter (deoxidating blow-

off) to oxygen blow-off for oxidating refining and heating of the metal, is made after recovery

of iron-containing, carbon-bearing and slag-forming materials and getting the optimal sharing

of recovered oxide scale iron between metal and slag. Achieving the optimal sharing allows the

process of metal oxidation refinement and heating to work smoothly and without wastes to the

given parameters on the temperature, constitution and liquid metal output.

Lately, at ZSMK the oxide scale share in the iron-containing raw material balance has de-

creased. In our point of view, this is due to several reasons. The first one is lack of complex

economic record of all the interconnected factors of oxide scale using in agglomerative, blast

furnace and converter industries. Some negative sides of oxide scale using in agglomerative and

blast furnace industries are mentioned above, but it is quite difficult to estimate them in terms

of economics.

The second one is that scrap metal replacement by oxide scale is connected to increase in

cast iron share in the metal charging material that leads to metal products costs rise. This should

be eliminated in the proper scrap metal pricing.

In Russia in the last few years, there has been a firm tendency of fall in scrap metal price

level compared to conversion pig iron. For example, in 1991 the cost of 1 tone of scrap metal in

proportion to 1 tone of conversion pig iron was 0.5, in 2000–0.23, in 2007–0.21. In the world

market scrap metal cost in proportion to conversion pig iron remains stable on the level of 0.8 to

the conversion pig iron cost. If we estimate the metal cost in the Russian market objectively, us-

ing oxide scale and other iron-containing wastes will become profitable for metallurgical plants,

as their retail price is much lower, then scrap metal price. This will allow to increase greatly their

share in the iron-containing raw material balance at a steelworks.

The manufactures are encouraged to increase the iron-containing wastes share in the raw

material balance at metallurgical plants by a significant decrease in scrap metal preparing and

recycling (by 15–25 %), which makes manufactures whether increase the cast iron expenditure

for steel smelting, or look for effective and appropriate ways of increasing the iron-containing

wastes share in the raw material balance.

To sum up, solutions to the problems with comprehensive utilization of steelworks wastes

and increasing the iron-containing wastes share in the raw material balance is not possible with-

out complex economical analysis and estimation of all energy, material and economic aspects

of various versions of their using.

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The south of Kuzbas is a top-ranked mining and smelting region of country and it has one

of the greatest differences between industrial economic potential, source of raw materials and

level of fundamental scientific researches of region issues. The priority research area does relate

to conservancy – a secondary raw materials of mining and smelting wastes utilization, i. e. a

scientific foundations of theirs recycling creation. This region is a coal mining leader in Russian

Federation (especially, its high-quality sorts) as well as metallurgy and heavy engineering industry

center, but the basic scientific researches in recycling carry out by certain small laboratories,

ingressed the chemistry of solids and mechanochemistry institute of SB RAS (in Novosibirsk)

and Kuzbas universities such as NBI KemSU, SibSU and KuzSTU.

The history of Kemerovo region’s industrial centers development related to tasks, that the

country had in 30–40 years of 20th century. At that time, the fastest achievement country’s and

region’s industry development on the strength of historical necessity (the industrial complex

creation, 2nd World War e. t.c) was requested, that in turn it imposed certain constraints on

conservancy and its ecological concepts adherence (on fact, the inobservance of theirs). The

big industrial enterprises distribution performed under the principles of profitability, such as

proximity the enterprises to resource origins and haulways. The infrastructure was created

according to the same principles. Violent maintenance of natural resources, first of all coal

and iron ore, led the region in short order to self-repair impossibility of environment and natural

resources. The number of wastes, accumulated on the Kuznetsk Basin’s territory, numbered a

hundreds of millions tonnes at the end of 20th century.

For example, during the 2001 year if was appeared 855812.319 thousands tonnes of non-

toxic industrial wastes, which included: dead-rocks – 836410.260 thousands tonnes, ash-and-

slad wastes – 2114.298 thousands tonnes, waste coal – 13620.780 thousands tonnes. Solid

domestic wastes are amount to 2 million tonnes. Waste piles, slurry ponds, tailing and rubbish

dumps occupy mare than 40 thousands hectares of land, excepted from usage [1]. Especial

unsightly situation has formed at the south of Kuzbas at region of towns Novokuznetsk and

Prokopyevsk, where the major industrial enterprises of coal mining, ferrous and nonferrous

metals and power industry are located. On the territory of these towns the solid wastes are

presented by large-tonnage wastes from metallurgical, heat-and-power engineering and coal-

processing industry alongside with colliery wastes that include all components that wasn’t ingress

of saleable coal, such as top covers, interburden layer, impregnations, coal slack and low-quality

carbons. The information about these wastes of Novokuznetsk is presented in table 1 [2].

Table 1

Large-tonnage wastes in Novokuznetsk at 1999

Item Amount,

tonnes

Reused,

tonnes

Located in storages,

tonnes

Slag 3029224 3310550 217096

Ash-and-slad wastes from heat

stations and boiler rooms

598000 33538 564462

Coal mining and coal-cleaning wastes 2148176 725491 1421685

Wet magnetic separation wastes 1089201 1089201

Solid domestic wastes 315000 315000

COMPREHENSIVE ANTHROPOGENIC WASTES UTILIZATION OF SOUTHERN KUZBAS MINING AND SMELTING COMPLEX,

PROBLEMS AND PERSPECTIVES. EXPLORING SITUATION

F.I. Ivanov 1, E.V. Isakova 1, A.S. Golovko 1, V.A. Poluboyarov 2

1 Novokuznetsk branch of the institute state educational highest vocational institution ,

Kemerovo State University, Novokuznetsk, Russia2 Chemistry of Solids and Mechanochemistry Institute of SB RAS, Novosibirsk, Russia

484


1. Slag

The main amount of slag is generated by metallurgical giants as public corporation «WSMW»

and «NKMW». Research group successfully examined a scientific bases development of slag’s

technological redistribution under the direction of doctor of technics, professor S.I. Pavlenko

at Siberian State Industrial University. Results, related to open-hearth slag processing practice

development, as dump likewise discharged wastes, were implemented on «NKMW» like

perspective alternative source of raw materials. It was started the special crushing and screening

plant to slag recycling. Non-magnetic fraction overworks into building materials; there are

also carry out the creating of cement-free bindings and concretes efforts [3–5]. Researches of

chemical and mineralogical composition that have executed under the direction of professor

S.I. Pavlenko are represented in table 2 and displaying the possibility of goal attainment.

Table 2

Results of open-hearth slag analysis (oxides, %)

Materials [SiO2] [Al2O3] [Fe2O3] [MnO] [CaO] [MgO] [SO3] [TiO2]

Discharged

open-hearth slag

15.16 10.65 17.82 5.08 27.2 14.99 0.67 1.06

Dump waste

open hearth slag

8.05 11.04 16.19 4.47 21.7 14.7 0.15 1.06

Losses by ignition for discharged slag amount 0.16 %, by [CaO] free 0.47 %, by [FeO] – 8.02 %.

Losses by ignition for dump waste slag amount 1.18 %, by [CaO] free 0.21 %, by [FeO] – 7.29 %.

2. Ash-and-slad wastes

At present ash-and-slad wastes, which arise on public corporations «Kuznetsk heat station»,

«KMW heat station» and «West-Siberian heat station», locate on their own sludge storages and

haven’t a wide application, however in research works by S.I. Arbuzov, V.A. Salihov an others [6–8]

economic-geological suitability of recycling theirs is validated. It based on content analysis a variety

of nonferrous and rare metals contained in carbons and ashes. Ash-and-slad wastes of heat stations

accumulate in amount of 2.6 million tonnes annually and produce the applied interest to metals

extraction. Table 3 summarizes the content of nonferrous and rare metals in carbons.

Table 3

The content of nonferrous and rare metals in Kuzbas coal

Metal Content in coal,

g/tonne

Concentration,

recommended to

evaluation, g/tonne

Maximum content

in ash, g/tonne

Conditions for

ores, %

Titanium 100–500 500 5600 10–15

Zirconium 100–500 500 3000 3

Copper to 15 100 3700 0.5

Plumbum to 25 50 4800 2

Zinc 10–300 100 16000 1

Barium 200 1000 5800 1

Vanadium to 50 100 5000 1

Tungsten to 3 100 1500 0.5–1

Beryllium to 1 100 430 0.3

Niobium 1–3 100 3000 0.1

Gallium 1–3 20 3000 0.04

Germanium to 1 10 2700 0.1

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3. Waste coal recycling

According State Standard 30772–2001 [9] waste coal is to be classified like anthropogenic

mineral raw materials and accumulation of theirs – like anthropogenic conglomerations of

carbonic row [10]. Usmanova T. V. and Rihvanov L. N. classify these conglomerations to the 3rd,

the 8th and the 10th types of anthropogenic coalfields [11].

Wastes of coal industry use to the best advantage on concentrating mill «Abashevskaya»,

where coal rocks of class 1–100 mm utilize partially in brick production, the rest of the amount

utilizes in road embankment, weirs filling and others objects. Idea, developed in [1] and called

«overadiabatic combustion of carbon-bearing raw materials and synthesis gas production» didn’t

meet with support in Kuzbas in spite of it obvious efficiency and ecological compatibility.

4. Solid domestic wastes

Collecting and storage problem of solid domestic wastes has been solved in Novokuznetsk.

It has been created highly mechanized high end factory that provides the accepting, preliminary

recycling and storage of solid domestic wastes.

5. Refractory wastes (that arise from blockwork repair of metallurgical furnaces and heat stations boilers) represent the particular interest for authors of this research paper

We propose to reuse all components of refractory wastes to remake refractory materials

applying nanostructured binding agent in compliance with basic idea of recycling.

The subject of inquiry is unshrinkable constructional refractory material and ceramics

from refractory wastes technology development for usage it in extreme conditions [12, 13]. This

technology uses attainments of mechanochemistry of solids in physical-chemical foundations

development to involve in refractory wastes processing the main metallurgical enterprises of

Kemerovo region.

Objectives: 1. Binding agent development on basis of mechanochemical technology,

including nanoparticles gaining, sized less than 100 nm (with specific surface more than 10 m 2/g).

2. Usage this binding agent for development the technology of gaining an unshrinkable material

patterns on basis of refractory wastes recycling and test operation in extreme conditions.

In the capacity of secondary raw materials there were used refractory wastes of «WSMW»

(Novokuznetsk): dinas, chamotte, electrocorundum (figure 1).

The trial designs prepare made on press «ДО-242» with pressure parameters fine-tuning in

different press molds of 380×125 mm and 150×120 mm size. It also made by vibrocompressive

method.

The checkout of trial designs ran at West-Siberian test center (accreditation certificate

№РОСС RU.0001.21 АЯ07) and at «WSMW» for full-scale testing in soaking pits during the one

year. The full-scale testing certificate is purely affirmative (table 4).

Table 4

Physical-mechanical results of fire brick, gained from refractory wastes by vibrocompressive method

Trial design Characteristics Unit of

measure

Checkout

method

Checkout

results

State Standard

requirements

Chamotte

brick

Refractoriness

Permanent linear shrinkage

oС

%

4069–69

5402–81

1540

–0.75

1580

0.5

Dinas brick Refractoriness


oС

%

4069–69

5402–81

1640

0.27

1580

0.5

Combined

brick

Refractoriness


oС

%

4069–69

5402–81

1550

–0.60

1580

0.5

Physical-mechanical checkout results of chamotte brick (figure 2), made from secondary

raw materials, conform to the State Standart specification 390–96 «Chamotte products».

Checkout results of dinas brick, made from the same (secondary raw) materials, practically

conform to the State Standart specification 4157–79 «Dinas products» by refractoriness, though

overtop it to a considerable extent by strength characteristics.

486


The first stage, related to refractory wastes recycling possibility research and industrial

refractories gaining using nanostructured binding agent, has finished successfully with support

of «Start» programme and «Small-scale enterprises in scientific and engineering business field

development promotion fund».

According to «Start» programme, to be entered into the 2nd stage of budgetary financing,

it’s necessary equivalent cofinancing of investor at the rate of 1.5 million rubles. Sponsors that

willing to affiliate with programme don’t satisfy the conditions of I. M. Bortnik fund.

This research has financed according to state contract № 02.513.11.3188 since April the

23 rd of 2007 in the network of «Researches and developments on priority ways of growth of

scientific-technical complex of Russia at 2007–2012 years» on topic called «Foundations of high-

effective methods of nanostructured nonshrinking corundum ceramics development and others

refractories that work in extreme exploitation conditions gaining based on binding material

from ultra- and nanodispersed powders gained by mechanochemical method».

Lot was won by programme of nanotechnology and it was one, scientists of Kuzbas

universities took the part in that.

The creation possibility of high-concentrated and, that is most importantly, cheap silica sols

demonstrated in patents: [1. «Method of alkali silicate gaining». Patent RF № 2187457 from 04 mar

2002. 2. «Method of nonfired building materials making». Patent RF № 2168481 from 31 jan 2001.

3. «Method of refracting masses gaining». Patent RF № 22143379 from 04 jun 2002. 4. «Method of

silicate bond gaining». Patent RF № 2144552 from 20 jan 2000 and others patents].

Authors of this research paper have explored the market of Kemerovo region by demand

for refractory materials. The basic providers of refractories to metallurgical enterprises of the

region are the close corporation «Magnesit» of Sverdlovsk region and some providers from

abroad, particularly from China.

Preliminary technical and economical assessments that have been traced with a glance of

nature-conservative and social significance of project are indicate of a considerable economical

advantages of refractory ceramic products gained from a secondary raw materials.

The basic consumers in Kemerovo region are: «WSMW», «NKMW», housing and communal

services of Kuzbas towns, Kuzbassenergo and others.

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487


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10. Коломенский Г. Ю., Техногенные месторождения угольного ряда /Г. Ю. Коло-

менский, Л. В. Гипич//Угольная база России. Т. 6. Основные закономерности углеобразо-

вания и размещения угленосности на территории России. – М.: Геоинформмарк, 2004. –

С. 519–540.

11. Усманова Т. В., Техногенные месторождения отходов горнорудных производств

Южной Сибири /Т. В. Усманова, Л. П. Рихванов//Горный журнал. Спец. выпуск. Цветные

металлы. – 2006. – № 4. – С. 29–31.

12. Огнеупорная безусадочная корундовая керамика на основе вяжущего из на-

нодисперсных порошков, полученных механохимическим способом /В. А. Полубояров,

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2007. – № 12. – С. 49–52.

13. Poluboyarov V. A., Korotaeva Z. A., Bebko A. N., Ivanov F. I. Influence of the

Nanostructure of Corundum Binder on the Strength of Nonshrinking Corundum Parts//Steel in

Translation. – 2009., Vol.39, № 2. – P.118–121.

488


Due to depletion of non-ferrous metal ore natural resources, enforcement of the require-

ments to storage conditions of wastes from metallurgical plants and rapid development of in-

dustrial technologies, industrial mineral deposits arouse interest. Economic motivations for in-

creased interest to industrial wastes are as follows:

– lower costs of mining, because it is not necessary to build mine;

– concentration of minerals is similar to concentration in natural deposits;

– decreasing of charges for environmental damages after retreatment of wastes due to

decreasing of their hazardous influence on the environment; after decreasing of con-

centration of components in the wastes down to safe, they can be used for production of

construction materials, and areas, used as wastes pits, may be restored.

The results of calculations of slag amount on the basis of geological-exploration studies of

waste pits from Karasakpayskiy copper plant are shown in table 1.

Table 1

Slag Amount and Concentration of Metals in Slag, Kept in Waste Pits of the Plant

Amount of waste slag

(thousand tons)

Metal concentration in slag

Аu, g/t

kg

Аg, g/t

t

Сu, %

thousand tons

Рb, %

thousand tons

Zn, %

thousand tons

2 885.7 0.22

634

10.25

29.6

0.91

26.3

0.29

8.3

0.36

10.5

It is known [1], that the way of cooling of melting products from non-ferrous metallurgy,

including copper matte and slag, has considerable influence on the level of the following metal

recovery by dressing methods. Thus, low decreasing of slag temperature facilitates separation

of metals, contained in slag (formation of secondary monominerals) and creates conditions for

growing of larger crystal grains. Magnetic and flotation separation of copper from the slag, which

has been cooled slowly, is more effective. And on the contrary, the process of quick cooling of

copper slag, which is similar to the examined one, is accompanied by formation of fine-grained,

cryptocrystalline structure, which is typical for hard melts. Separation of components in such a

material causes difficulties.

Laboratory technological researches

Material composition and physical characteristics of copper slag from waste pits of Karas-

akpaysiy copper plant have been studied for many years. Mineral composition has been studied

using ore microscope with high resolution. Several slag samples with different concentration of

copper (from 0.2 % up to 5.3 %) and other element have been received for examination, which

demonstrates that composition of the wastes differs greatly.

Sample materials consist mostly from vitreous slag particles, containing different inclu-

sions from metallurgical plant. Almost all vitreous slag contains particles of mono-magnetite

(of different size, saturation and structure), having size from 2–20 up to 150 μm and its ag-

gregates, including latticed-dendritic-skeletal aggregates of metacrystalline magnetite. 85 % of

TECHNOLOGICAL STUDIES OF SLAG SAMPLES FROM KARSAKPAYSK COPPER SMELTER FOR METAL

RE-EXTRACTION FEASIBILITY STUDY

S.G. Gritsay, G.I. Krivopustova, A.O. Teut, N.I. Utrobina

State Affiliate «The Eastern Mining and Metallurgical Research Institute for Non-ferrous Metals»

Republic State Affiliate «National Enterprise of Complex Processing for Mineral and Raw

Material of the Republic of Kazakhstan» (SA «VNIItsvetmet» RSA «NE CPMRM RK»),

Ust-Kamenogorsk, Republic of Kazakhstan

489


all the vitreous slag particles together with magnetite contain different products from metal-

lurgical plants mostly with copper-zinc composition (having low concentration of galena and

gold) in the form of pearlitic ball particles (having diameter from 2 up to 150 μm, sometimes

300–420 μm and bigger), presented as mixtures with different components, which are the fol-

lowing (in accordance with their prevalence rate): solid chalcocite-bornite solution, bornite,

chalcocite, sphalerite, chalcopyrite and sometimes eutectic mixture of galena and chalcocite

and also particles and discontinuous streaks of metallic copper and gold and silver (more rarely)

particles. In a circumferential direction from these pearlites and bigger inclusions, there may be

often seen thin discontinuous pectinal crusts, formed by thin metacrystalls of magnetite.

Sphalerite may be seen not only in inclusions of mixtures of different composition and with

solid chalcocite-bronite solution, but also in many particles of vitreous slag with magnetite in

the form of thin (1.5–6 μm in diameter) pearlitic inclusions of grey semi-transparent and hardly

determinable spalerite. Around these particles of sphalerite, there may be seen discontinuous

cover, represented by all copper-bearing components.

About 15 % from total amount of vitreous slag particles, containing copper-zinc compo-

nents (having size 10–200 μm), may be considered as free (from magnetite) slag particles. To-

gether with plated-latticed layers or sections consisting of pure slag, the main body of the mate-

rial consists from vitreous slag, impregnated by thin (1–6 μm) copper-zinc perlites and ultra-thin

(not more than 0.1–0.5 μm) and almost invisible through the microscope inclusions of all the

mentioned above copper-zinc components, which are absolutely inseparable from slag.

Large amount of magnetite, contained in major part of the material, may be explained

by very low effectiveness of magnetic separation during efforts of preliminary slag enriching.

There has been measured specific weight of slag, which is 2.8 g/cm 3 and bulk weight is

1.6–1.8 g/cm 3. High concentration of iron (from 9 % up to 36 %), which hardness is 6.0–6.5 ac-

cording to Mohs scale, and silicon (from 26 % up to 48 %) in copper slag is the main reason for

difficult crushing of the material.

Detailed chemical composition of one from the slag samples, analyzed in the laboratory, is

shown in nable 2.

Table 2

Analysis Results of Chemical Composition of Slag Laboratory Sample

Components Concentration, % Components Concentration, %

Zinc 4.71 Calcium oxide 4.33

Lead 0.82 Magnesium Oxide 1.14

Copper 1.25 Silicon dioxide 26.86

Gold, g/t 0.2 Barium oxide 1.67

Silver, g/t 9.2 Indium 0.0021

Iron 36.0 Cadmium 0.005

Pyrites sulfur 1.85 Bismuth >0.05

Aluminum trioxide 4.24 Antimony 0.003

Iron trioxide 46.19 Thallium 0.0013

The examination of ionic composition of liquid phase of slag slurry (table 3) shows that the

dissolution level of calcium and magnesium sulfates is very high.

Table 3

Ionic Composition of Liquid Phase of Slurry from Laboratory Sample of Copper Slag

Component рН Сusol. Zn Cа Fе Mg SO4 Рb

Original copper slag 8.25 0.05 0.11 58 0.057 9.4 62.9 <0.01

The results of analysis for states of the presented zinc, copper and lead compounds in slag

are shown in table 4. It has been specified that 60.87 % of copper in slag is represented by sulfate

forms, and 41.74 % from this amount is secondary copper minerals.

490


Table 4

Analysis Results of Balance Composition of Laboratory Slag Sample

Mineral

forms

Concen-

tration %

Distribu-

tion %

Mineral

forms

Concen-

tration %

Distribu-

tion %

Mineral

forms

Concen-

tration %

Distribu-

tion %

ZnCO3 0.11 2.34 Cufree 0.22 17.39 PbCO3 0.08 9.76

ZnSiO2 1.69 35.88 Cufixed 0.27 21.74 PbS 0.19 23.17

ZnS 2.91 61.78 Cusecon 0.52 41.74 Pbjaros. 0.55 67.07

CuS 0.24 19.13

Zntotal 4.71 100 Cutotal 1.25 100 Pbtotal 0.82 100

Table 5 shows distribution of elements according to the grain-size category of laboratory

slag sample. According to this table 91.6 % of copper concentrates in category of +0.074 mm,

however, taking into account characteristics and size of mineral copper-bearing formations for

flotation method of enriching there shall be used extra fine slag.

Table 5

Grain Composition and Distribution of Components According to Grain-size Category of the Analyzed Slag Sample

Grain-size category,

mm

Output,

%

Concentration, % Distribution, %

Cu SiO2 Fe Cu SiO2 Fe

–2 +0.63 43.93 1.18 46.84 9.37 41.51 44.68 43.22

–0.63+0.315 39.55 1.23 46.57 9.01 38.93 40.00 37.42

–0.315+0.074 7.33 1.40 46.18 10.36 8.22 7.35 7.98

–0.074+0.044 3.45 1.44 53.66 15.70 3.97 4.02 5.69

–0.044+0.020 2.15 1.66 50.98 13.42 2.86 2.38 3.03

–0.020 3.59 1.57 20.14 7.05 4.51 1.57 2.66

Total 100 1.25 46.05 9.52 100 100 100

Grinding coefficient of copper slag (in reference to copper-zinc ore) was 0.27.

During open laboratory experiment campaign, which included slag grinding, basic cop-

per floatation activities, control copper flotation, re-cleaning and re-grinding of tailings from

re-cleaning and concentrate from control flotation, there were selected the level of grinding,

reagents consumption and flotation time [2]. The experiments determined the following:

– grinding of slag shall be accomplished until 93 % of slag is in grain-size category of

–74 μm;

– рН of media shall be kept at the level of 10–11;

– total consumption of reagents was the following: butyl xanthogenate 500 g/t, OFMB

40 g/t;

– total flotation time is 13 min.;

– it is necessary to foresee regrinding of 98 % middlings up to grain-size category

–74 μm.

Table 6 shows the results of closed test, using selected parameters, according to the dia-

gram, shown in Figure 1.

Table 6

Technological Parameters of Laboratory Closed Test

Product Output,

%

Concentration, % Removal, %

Cu Fe Cu Fe

Copper concentrate 4.39 20.21 10.34 70.95 4.76

Tailings 95.61 0.38 9.48 29.05 95.24

Initial copper slag 100.00 1.25 9.52 100.00 100.00

491


Tailings

Copper slag with grain size of -2+0 mm

Grinding up to 93% – -0,074 mm category

Basic copper flotation t=5 min.

Copper recleaning t=2 min. Control flotation t=6 min.

Copper concentrate

t=70 min.

Kx-200 g/t

Kx – 100+100 g/t

CMAC – 300 g/t

OFMB – 20+10 g/t

Kx – 50+50 g/t

OFMB – 10 g/t

Fig. 1. Diagram of Closed Test for Slag Processing with Reagent Mode

Semi-industrial technology tests

For development of practical basis for industrial processing of copper slag, tests of labora-

tory-developed technology has been accomplished in the enriching plant. For semi-industrial

tests, there has been received more than 570 tons of slag from Karsakpayskiy copper plant, hav-

ing moisture content of 2.7 %.

The results of testing during receiving of raw material (table 7) shows considerable unho-

mogenuity of copper slag’s chemical composition, and high concentration of silicon dioxide in

wastes from copper smelting plants caused problem for its crushing and grinding.

During the tests, optimal variant of process flow-sheet for copper slag processing has been

determined (fig. 2), reagent mode of copper flotation has been specified and recycling water in-

fluence on technological parameters of the process has been studied. The results of tests are

shown in table 8.

Table 7

Chemical Composition of Industrial Slag Sample

Sample

No.

Concentration, %

Cu Fe Zn As Sb Pb Ca SiO2 S Au, g/t Ag, g/t

1 5.33 12.47 1.88 0.74 0.24 8.32 10.50 35.50 1.76 3.60 214.0

2 1.25 9.52 0.40 0.03 0.03 0.62 14.00 46.05 0.10 0.40 25.8

3 1.72 11.57 0.70 0.03 0.015 0.44 14.50 46.20 0.56 0.40 25.4

4 1.61 12.64 0.76 0.03 0.0027 0.84 14.00 46.53 0.34 0.30 32.9

5 0.74 11.12 0.23 0.03 0.0069 0.16 12.50 48.69 0.12 0.40 10.1

6 0.67 12.52 0.35 0.03 0.002 0.32 10.50 47.74 0.10 0.30 10.0

7 0.24 7.17 0.09 0.03 0.002 0.14 18.70 47.98 0.10 0.30 10.0

8 2.50 9.41 0.56 0.03 0.002 0.60 13.0 46.44 2.11 0.30 10.0

9 0.22 7.51 0.15 0.03 0.002 0.13 19.40 47.38 0.10 0.30 10.0

Portion

sample

1.52 10.37 0.54 0.10 0.03 1.18 14.22 46.01 0.01 0.59 33.5

Table 8

Process Balance of Semi-industrial Processing of Slag

Products Output,

%

Concentration, %, g/t Removal, %

Cu Zn Fe Au Ag Cu Zn Fe Au Ag

Initial slag 100.0 1.52 0.54 10.37 0.59 33.5 100 100 100 100 100

Copper

concentrate

4.94 20.03 2.09 12.05 3.70 353.0 65.1 19.1 5.7 31.0 52.0

Tailings 95.06 0.56 0.46 10.3 0.43 16.9 34.9 80.9 94.3 69.0 48.0

492


High concentration of iron in copper concentrate can be explained by its close associa-

tion with copper in secondary sulfide minerals, i. e. in bornite (Cu5FeS4), predominating in slag

composition.

Initial slag with grain size -16+0 mm

1 stage of grinding

Classification in classifying screen

Classification in hydrocyclone

2 stage of grinding

drain

drain

sands

sands

Basic copper flotation

Recleaning Control flotation

Classification in hydrocyclone

drainsands

Copper

concentrate

Tailings

BX – 200 g/t

OFMB – 30 g/t

BX – 100 g/t

OFMB – 10 g/t

regrinding

BX – 200 g/t

Fig. 2. Process Flow-sheet of Semi-industrial Tests

Process of copper slag enriching is accomplished using water closed circulation. Ionic com-

position of recycling water, which has been sampled during different stages of tests, is shown in

table 9.

Table 9

Ionic Composition of Recycling Water

Determined Ingredient Ionic Composition, mg/l, mg-eq/l

Beginning of tests Middle of tests End of tests

рН 7.8 7.95 7.45

Cu+2 0.7884 0.007 0.0065

Zn+2 0.7252 0.0277 0.1098

Pb+2 0.3716 0.013 0.198

Кх <0.5 30.5 <0.5

Total hardness 5.0 5.1 2.4

Mg+2 9.73 17.02 36.07

Ca+2 84.17 74.15 36.07

Oxidation 2.4 48 32

SO4–

433.31 690.5 456.77

Suspended elements 1700 1176 508

Dry residue 1088 1596 1182

Taking into account high hardness of slag due to its silicate basis, which considerably com-

plicated crushing and grinding, in laboratory tests for its desiliconization using clinkering by

ammonium fluoride have been accomplished [3]. It has been determined that slag clinkering by

this reagent during 2 hours under the temperature of 350 oC allows to reach the level of silicon

dioxide removal of more than 99 % [4], which is positive for mechanical characteristics of slag.

493


Summary

1. «Stale» slag of Karsakpayskiy copper plant has unhomogeneous composition and is diffi-

cult industrial raw material for production of commercial copper concentrate (20 % Cu, 2 % Zn,

12 % Fe, 3.7 g/t Au, 353.0 g/t Ag).

2. The suggested process flow-sheet of slag enriching showed that copper recovery into

concentrate under industrial conditions may be higher than 65 %, having commercial product

output of about 5 %. Herewith, copper and zinc concentration in flotation tailings is about 0.5 %

(of each one) and iron recovery into tailing is more than 94 %, zinc – 80 %.

3. During Process tests, possibility of application of complete water circulation without

negative influence on concentrate quality and on copper recovery level from slag into copper

concentrate has been determined.

4.The results of experiments in slag desiliconization during its clinkering with ammonium

fluoride show that it is possible to remove almost all silicon dioxide, which may considerably

facilitate its following enriching.

REFERENCES

1. Abdeyev M. A. Complex matte and converting. Alma-Ata: AGMNII AN KazSSR. 1962. –

p. 228

2. Gritsay S. G., Teut A. O. Development of process flow-sheet for copper recovery from

«stale» waste slag of Karsakpayskiy copper smelting plant by enriching methods//Equipment

and Technologies for enrichment of ore and non-ore materials. Materials from 6-th Internation-

al Scientific-Practical Conference. – Novosybirsk: Sybprint. 2008. p. 40–46

3. Diachenko A. N. Halogenammonia technology for processing of metallurgical slag//Non-

ferrous Metals. – 2005. – No. 5–6. – p. 71–74

4. Study of physical and chemical principles of industrial raw-materials desiliconization

during clinkering process with halides.//SAR Report (intermediate): 3.3.1–142-FI/VNIItsvet-

met; leader: Teut A. O., Ushakov N. N. – Ust-Kamenogorsk, 2009. – p. 32.

494


The present work is focused on sorption of platinum (II, IV) and rhodium (III) from chlo-

ride solutions on anion exchangers with different physical and chemical structure. The sorption

was carried out from solutions with HCl concentration 0.01–4.0 mol/L. Platinum and rhodium

concentrations in contacting solutions were 0.25–0.50 mmol/L. Sorption and kinetic properties

of the chosen anion exchangers were investigated, and the basic parameters (exchange capacity,

recovery degree, distribution coefficients, process rate, diffusion coefficients and half-exchange

times) were calculated. The highest selectivity to noble metals was demonstrated by complex-

ing anion exchanger Purolite S 985, strong base resin Purolite A 500 as well as weak base anion

exchanger AM-2B produced in Russia.

Introduction

The production of platinum group metals is steadily growing as their industrial application

becomes broader. However, since natural deposits of precious metals are being depleted, the tech-

nologies for noble metals recovery from different secondary raw materials (e. g. spent catalysts

from organic synthesis or motor-car catalysts, electronic scrap etc.) are becoming more impor-

tant [1, 2]. Low platinum and rhodium concentrations in these materials cause the application

of sorption methods for the recovery of noble metals. These methods are characterized by high

efficiency and ecological safety. The sorption recovery of platinum group metals is usually carried

out from solutions obtained by decomposition of noble metals-containing materials. The break-

down of these samples is carried out by different methods, such as dissolution in acids, chlorina-

tion, smelting and other methods [3–6]. As a result, these solutions contain complexes of plati-

num group metals with different stability and chemical inertness. Moreover, the platinum group

metals are affected by aquation and hydrolysis [2, 7–9]. The metal sorption from such solutions

can be complicated, and some valuable compounds can be lost. In view of that, the high selectiv-

ity of sorbents is very important. Usually, the nitrogen-containing anion exchangers with func-

tional groups of pyridine or amines are used for platinum group metals recovery from different

solutions [2, 10]. As a rule, these resins recover platinum and rhodium by means of ion exchange

between complex anions of Pt and Rh and mobile ions of the sorbent as well as by complexation

between nitrogen atoms of sorbents’ functional groups and atoms of noble metals [10].

To make the ion exchangers applicable in industrial scale, it is important to know not only

their sorption properties but kinetic characteristics as well, since the process rate is essential

for industrial applicability of developed method.

The present work is devoted to sorption recovery of platinum and rhodium at their simulta-

neous presence in chloride solutions on anion exchangers with different physical and chemical

structure.

Experimental

We have chosen for investigation some anion exchangers, which physical-chemical proper-

ties are summarized in table 1.

Before sorption all the anion exchangers were prepared according to the standard proce-

dures and then converted to chloride form.

The initial platinum stock solution was prepared by dissolution of 1.0 g of metallic plati-

num in «aqua regia» (mass ratio of НСl and НNO3 was 3:1) with subsequent evaporation in

water bath in the presence of hydrochloric acid and distilled water, aiming to decompose the salt

of nitrosyl cations (NO)2 [PtCl6] and to withdraw the excess of nitric acid [7, 9]. The dry residue

SORPTION RECOVERY OF PLATINUM (II, IV) AND RHODIUM (III) FROM CHLORIDE SOLUTIONS

OF SPENT CATALYSTS

D.М. Кashirin, А.М. Мelnikov, О.N. Коnonova

Siberian Federal University, Krasnoyarsk, Russia

495


of H2 [PtCl6] was dissolved in hydrochloric acid. The content of platinum in this solution was

determined by gravimetric method with ammonium chloride [9].

The initial stock solution of rhodium was prepared by sintering of 0.5 g of metallic rhodium

with five-fold mass amount of BaO2 [9]. The content of rhodium in solution obtained was deter-

mined by gravimetric method with thiourea [9].

In this work we have investigated the freshly prepared chloride solutions simultaneously con-

taining platinum and rhodium. Pt and Rh concentrations in these solutions were 0.25–0.50 mmol/L,

and НСl concentration was varied from 0.01 to 4.0 mol/L.

The platinum and rhodium concentrations in solutions were determined by spectrophotomet-

rical method with SnCl2 [7, 9]. The concentrations and acidity of platinum and rhodium solutions

were taken with an intention to make the experiment closer to real industrial conditions [2, 3].

The sorption concentration of Pt (II, IV) and Rh (III) was studied as follows: sorbent mass –

0.1 g, volume of contacting solution – 10.0 mL, stirring at (20±1) oС. The equilibrium time was

about 24 h.

The efficiency of sorption recovery of platinum and rhodium on the chosen anion exchangers

was estimated by means of exchange capacity (EC, mmol/L), distribution coefficients (D, mL/g),

recovery degree (R, %) and separation factor (S). These values were calculated as follows:

( )0 pC C V

OE q− ⋅

= , (1)

where С0 and Ср are the initial and equilibrium molar concentrations of platinum or rhodium in

solution, respectively; V is the volume of contacting solution (mL) and q is the resin quantity (g).

p

OEDC

= ; (2)

( )

0

0

100%%

pC CR

C− ⋅

= ; (3)

Pt

Rh

DS

D= , (4)

where DPt and DRh are the distribution coefficients of platinum in the presence of rhodium and

of rhodium in the presence of platinum, respectively.

The kinetic behavior of platinum and rhodium sorption was investigated by «limited bath

method» with the corresponding criteria [11, 12]. The kinetic experiments were carried out under

the intensive stirring of solution (800 rev/min). The contact times of resins and solutions were

0.5; 1; 2; 3; 5; 15;30 and 45 min; 1; 3; 6 and 24 h. The radius of the resin grains was (0.5÷1.0)

mm. After a certain time period, the resins and solutions were quickly separated and the liquid

and solid phases were subjected to analysis. Then the concentrations of platinum and rhodium

were determined by spectrophotometrical method. Using the Boyd-Adams model, the interpreta-

tion of results on kinetics was carried out [11–13].

The exchange degree (F) was calculated from:

,tQFQ∞

= (5)

where Qt and Q∞ are the amounts (mmol) of platinum or rhodium to the time t (s) and to the

equilibrium time.

Table 1

Physical-chemical properties of ion exchangers investigated

Trade name

Resin type Functional groups Physical structure

EC to Cl –, mmol/g

Swelling grade, %

Purolite S 985

Complex-forming anion exchanger

Polyamine groups MP 10.0 42.0

Purolite A 500

Strong base anion exchanger

Quaternary ammonia base

MP 1.2 37.1

АМ-2B Weak base anion exchanger

Secondary and tertiary amines

MP 2.3 21.1

EC – exchange capacity; MP – macroporous

496


Then the kinetic curves were plotted on the coordinates F=ƒ(t) and the half-exchange

time of the kinetic process (t1/2) was determined at F=0.5.

After that, to confirm the applicability of the above model to the ion exchange kinetics in

the systems investigated, the dependences Bt=ƒ(t) were plotted. According to the Boyd’s model

[11, 12], the kinetic coefficient (B) was calculated from:

2 2(1,08) FB

t⋅= . (6)

If the process is controlled by gel diffusion, the function Bt=ƒ(t) should be linear.

The diffusion coefficients were calculated according to the equation:

2

21

24

sr

Dtπ

=⋅

, (7)

where is the diffusion coefficient in ion exchanger grain (cm 2/s); r is the radius of the resin

grain (cm); t1/2 is the half-exchange time (s) [11, 12].

All the results were subjected to statistical processing according to conventional proce-

dures. The average experimental error for three to four parallel runs was no higher than 6 %.

Results and discussion

The ionic states of platinum (II, IV) and rhodium (III) in chloride solutions are studied in

detail [2, 8, 9, 14, 15]. It is known that they depend on medium acidity and temperature. The

system Rh (III) – HCl – H2Ois characterized by various transformations of complexes as well as

by their aquation, hydrolysis, polymerization and isomerization.

The hexachloroplatinate (IV) complex PtCl6] 2– predominates in strong acidic media

(СHCl>3 mol/L).With the dilution of these solutions (increase in pH value), the complexes of

platinum (II) are formed, which co-exist in different ratios with chloride complexes of platinum

(IV). Moreover, with the decrease in solution’s acidity, the hydration and hydrolysis take place.

These effects lead to formation of various aquachloro – and aquahydroxo- complexes of plati-

num (II) and platinum (IV): [Pt (H2O)nCl4-n]n –2, [Pt (H2O)k (OH)mCl4m-k]k–2, [Pt (OH)nCl4-n] 2–,

[Pt (H2O)nCl6-n]n –2, [Pt (H2O)k (OH)mCl6m-k]k–2, [Pt (OH)mCl6-m] 2– (where n = 1, 2, k = 1, 2,

m changes from 1 to 6). Besides that, binuclear complexes [Pt2 (H2O)2(OH)7Cl] can be also

present in these systems [9, 15–17].

The hexachlororhodiate (III) complex [RhCl6]3 – prevails in solutions with hydrochlo-

ric acid concentration 6 mol/L or higher. With the decrease in acidity, the formation of

aquachlorocomplexes[RhCln (H2O)6-n] 3 –n (0<n<6) as well as of cis- and trans-isomerides

[RhCl6]3– and [Rh (H2O)Cl5] 2–is observed. The polymerization causes the formation of binuclear

complexes [Rh2Cl9] 3– [2, 17]. Therefore, the dominant forms of rhodium (III) in the investigated

systems are complexes[RhCl6]3– and [Rh (H2O)Cl5] 2– in 2М HCl solutions, whereas in solutions

with pH=3, the complexes[Rh (H2O)2Cl4] –and [Rh (H2O)3Cl3] 0 as well as cationic complexes[Rh

(H2O)4Cl2]+ and [Rh (H2O)5Cl] 2+ are also formed in these systems [2, 9, 16, 17].

Figure 1 contains the absorption spectrum of the initial platinum and rhodium solution at

their simultaneous presence in 0.01 M HCl. It

should be noted that absorption spectra of indi-

vidual solutions of noble metals, which we have

registered previously [18, 19], are in agree-

ment with the data [2, 9, 16, 17]. It can be seen

from Figure 1 that the absorption spectrum of

freshly prepared chloride solution of platinum

(II, IV) and rhodium (III) at their simultane-

ous presence reveals two absorption maxima

at 218 and 251 nm. This is in accordance with

works [2, 7]. The maximum at 218 nm indi-

cates a prevailing presence of complex [PtCl4] 2–

and cis- and trans-complexes of rhodium

(III) –[Rh (H2О)4Cl2]+ and [Rh (H2O)2Cl4] – in

solution of platinum (II). The second absorp-

tion maximum located at 251 nm corresponds

with hexachloro-complexes of platinum (II),

platinum (IV) and rhodium (III) [2].

Fig. 1. Absorption spectrum of chloride freshly

prepared solution of platinum (II, IV) and rho-

dium (III) at their simultaneous presence in

0.01 M HCl

497


We have studied in our previous works [18, 19] the sorption concentration of Pt and Rh

from individual chloride solutions (table 2). In contrast to platinum, rhodium is recovered to a

rather low extent by chosen anion exchangers. This is related to higher kinetic lability of plati-

num chloride complexes in comparison with rhodium complexes.

The sorption recovery of both metals at their simultaneous presence in dependence on solu-

tions acidity is presented in tables 3 and 4. It can be seen from these data that simultaneous recov-

ery of Pt and Rh on the same anion exchangers is higher as compared to their individual sorption.

It can be seen from the data that all the resins investigated recover more than 85 % of plati-

num from chloride solutions. However, rhodium is recovered significantly poorer that platinum,

with the exception of anion exchanger Purolite S 985. With the decrease in solution acidity, the

recovery degrees of Pt and Rh are growing. This phenomenon is caused by the structure of sor-

bents’ functional groups.

Table 2

Sorption concentration of Pt and Rh from individual chloride solutions (С0 (Pt)=С0 (Rh)=0.25 mmol/L; С (HCl) = 2.0 mol/L)

R, %

Purolite А 500 Purolite S 985 АМ-2B

Pt 85 95 94

Rh 20 82 45

Table 3

Sorption concentration of Pt in the presence of Rh from chloride solutions (С0 (Pt)=С0 (Rh)=0.25 mmol/L)

Trade Parameter С (HCl), mol/L

name 4.0 2.0 1.0 0.5 0.1 0.01

Purolite D, mL/g 632 835 844 1471 1735 3151

S 985 R, % 86.3 89.3 89.4 93.4 94.6 96.9

Purolite D, mL/g 583 802 900 1918 2093 5486

A 500 R, % 85.4 88.9 90.0 95.1 95.4 98.2

АМ- B D, mL/g 696 818 952 1342 2302 7071

R, % 87.4 89.1 90.5 93.2 95.8 98.6

Table 4

Sorption concentration of Rh in the presence of Pt from chloride solutions (С0 (Pt)=С0 (Rh)=0.25 mmol/L)

Trade Parameter С (HCl), mol/L

name 4.0 2.0 1.0 0.5 0.1 0.01

Purolite D, mL/g 1143 1554 1584 1594 1635 3024

S 985 R, % 91.9 93.9 94.0 94.1 94.2 96.8

Purolite D, mL/g 119 120 125 169 181 223

A 500 R, % 54.4 54.5 55.5 62.8 64.4 69.1

АМ-2B D, mL/g 99 146 226 235 303 355

R, % 50.7 59.4 69.3 70.2 75.2 78.0

It is known [10], that the ability of complexing ion exchangers (AM-2B and Purolite S 985)

to formation of resin complexes can be estimated by degree of protonation of their functional

groups. With the increase in pH of contacting solution, the deprotonation of nitrogen atoms of

aminogroups is growing, i. e. their ability to complexation increases. Therefore, in strong acidic

media (CHCl = 4 mol/L) the functional groups are completely protonated, and the weak base an-

ion exchangers recover chloride complexes of platinum (II, IV) and rhodium (III) only through

the ion exchange mechanism. With the decrease in solution acidity, when deprotonation of

functional groups takes place, the additional complexation occurs beside the anion exchange

between platinum or rhodium and nitrogen atoms of functional groups. The strong base anion

498


exchangers with quaternary ammonia base groups, unlike the weak base resins, do not possess

the complexation ability. The recovery of platinum group metals on these sorbents practically

does not depend on pH of contacting solution. In strong acidic media, the competing effect

between complex anions of platinum or rhodium and chloride ions can take place during their

sorption on strong base resins. Accordingly, the lesser the concentration of chloride ions is in

solution, the weaker is their competing effect. This causes the higher recovery degree of Pt and

Rh on anion exchanger Purolite A 500.

Moreover, the lower recovery of rhodium (III) in comparison with platinum (II, IV) can be

explained by greater kinetic inertness of its chloride complexes, especially in weak acidic media,

owing to presence of neutral and charged aquacomplexes[Rh (H2O)3Cl3] 0, [Rh (H2O)2Cl4]–,

[Rh (H2O)Cl5] 2 – [2].

To determine the mechanism of sorption concentration of Pt and Rh on chosen anion ex-

changers, we have registered the Raman spectra of these sorbents in initial chloride form and

after saturation by solutions of noble metals. Figure 2 contains the Raman spectra for anion

exchanger Purolite S 985.

The comparison of resin spectra in initial form and after saturation with platinum in 0.01 M

HCl (fig. 2 a) shows that two bands appear at 311 and 342 сm–1. They correspond to vibrations

of N→Ptt bond, which is characteristic for complexation in the resin phase [20]. Moreover,

these bands characterize vibrations of anions [PtCl4] 2 – and [PtCl6] 2 – in the anion exchanger

phase, sorbed during the ion exchange [10, 20]. Comparing the Raman spectra of platinum in

1.0 M HCl and in 0.01 M HCl between themselves, the notable smoothing of band intensity is

observed at 311 and 342 сm-1 in case of strong acidic solution. These changes can be attributed

to protonation of nitrogen atoms of functional groups in anion exchanger Purolite S 985.

Fig. 2. Raman spectra of anion exchanger Purolite S 985 saturated with platinum (a)

and rhodium (b)

The Raman spectrum of this sorbent saturated with rhodium solution contains two evident

bands at 343 and 315 сm–1, which correspond to vibrations of N→Rh bond in the resin [10,

20]. In addition, the band at 343 сm–1 characterizes vibrations of complex anion[RhCl6]3–sorbed

through ion exchange mechanism (fig. 2 b). Therefore, the presented data show the mixed sorp-

tion mechanism of platinum (II, IV) and rhodium (III) chloride complexes on anion exchanger

Purolite S 985. The sorption proceeds according to ion exchange (eq. 9) and to complexation

mechanism (eq. 10). The strong base anion exchanger Purolite A 500 recovers Pt and Rh in con-

formity with anion exchange (eq. 9). The weak base anion exchanger AM-2B recovers rhodium

through complexation (eq. 10), whereas platinum is sorbed by mixed mechanism (eq. 9 and 10).

, (9)

, (10)

where Ме=Pt (Rh), L=Cl –, n=4, 6 (for Pt) and 6 (for Rh), m=2 (for Pt) и 3 (for Rh).

The isotherms of platinum (II, IV) and rhodium (III) sorption were plotted for the anion

exchangers investigated. These curves for the resin Purolite S 985 are presented in Figure 3. It is

known [10, 11, 13] that their shape is an evidence of sorption selectivity.

499


It can be seen from figure 3 that isotherms are convex curves, indicating the selectivity of

resins during ion exchange. Such isotherms are classified to Langmuir isotherms described as

follows:

,1

EC ECK Ceq⋅

K Ceq⋅= +∞ + (11)

where EC and EC∞ are the equilibrium exchange capacity and the maximal exchange capacity

of the resin to Pt or Rh, respectively, mmol/g; Сeq is the equilibrium Pt or Rh concentration,

mmol/L; К is the apparent constant of ion exchange equilibrium, L/mmol.

Fig. 3. Sorption isotherms of Pt in the presence of Rh (a) and Rh in the presence of Pt (b) on

anion exchanger Purolite S 985 from chloride solutions (C (Pt)=C (Rh)=0.5 mmol/L)

By transforming the equation (11) to the linear form:

1 1 1 1

EC EC EC K Ceq= + ⋅

⋅∞ ∞, (12)

we calculated ion exchange equilibrium constants as well as values of maximal equilibrium ex-

change capacity and determination coefficients (R 2), which are summarized in tables 4 and 5.

Table 4

Linear correlation of obtained sorption isotherms of platinum in the presence of rhodium from chloride solutions and parameters of Langmuir equation

Trade name C (HCl)=2.0 mol/L C (HCl)=0.01 mol/L

ОЕ∞ К R 2 ОЕ∞ К R 2

Purolite А-500 0.33 2.05 0.962 0.41 6.25 0.991

Purolite S-985 0.45 1.71 0.992 0.36 6.33 0.983

АМ-2B 0.67 1.36 0.970 0.38 8.12 0.993

Table 5

Linear correlation of obtained sorption isotherms of rhodium in the presenceof platinum from chloride solutions and parameters of Langmuir equation

Trade name C (HCl)=2.0 mol/L C (HCl)=0.01 mol/L

ОЕ∞ К R 2 ОЕ∞ К R 2

Purolite А-500 0.09 1.24 0.945 0.11 1.43 0.947

Purolite S-985 0.49 2.33 0.997 0.32 5.35 0.999

АМ-2B 0.17 1.12 0.964 0.16 1.40 0.986

It can be seen from tables 4 and 5 that the apparent constants of ion exchange equilibrium

on anion exchangers investigated correlate with the selectivity of sorbents.

Table 6 contains the separation factors of platinum and rhodium on the resins investigated.

These values are more than 1, i. e. the separation of noble metals can be carried out during their

recovery from freshly prepared chloride solutions.

500


Table 6

Separation factors of platinum and rhodium

С (HCl) Purolite А-500 Purolite S-985 АМ-2B

S 2.0 5.51 1.20 4.17

0.01 6.46 1.90 5.25

Therefore, all the anion exchangers investigated reveal a high affinity to platinum (II, IV)

and rhodium (III) at their simultaneous presence, especially complexing resin Purolite S 985,

which possesses the best sorption properties.

Further we have studied kinetics of Pt and Rh sorption at their simultaneous presence from

chloride solutions. The dependences of concentration process rate on time are represented in

figure 4 for anion exchanger Purolite S 985.

Fig. 4. Kinetic curves of concentration process rate for Pt in the presence of Rh (a) and Rh

in the presence of Pt (b) on anion exchanger Purolite S 985 in dependence

on HCl concentration (С (Pt)=С (Rh)=0.25 mmol/L)

It can be seen from the figure 4 that the resin Purolite S 985 possesses good kinetic proper-

ties, since the process rate is high. Over a period of 20 min, the resins investigated are saturated

with platinum to 50–81 % and with rhodium – to 36–80 % of their total exchange capacity. The

rate of sorption concentration is higher in weak acidic solutions. This fact is in accordance with

the supposed process mechanism. It should be noted that the sorption rate of platinum and

rhodium from strong acidic solutions is also high. Further we have determined the kinetics type

during sorption of platinum and rhodium using the Boyd – Adams model with the correspond-

ing criteria [13]. Figure 5 contains the dependences Bt=f(t), which are consistent with the

above-mentioned model for gel kinetics of ion exchange. It can be seen that these dependences

are linear for all the resins investigated. It means that the whole sorption process is controlled

by interdiffusion of the ions exchanged in a resin grain.

Fig. 5. Kinetic dependences of Bt function on time t for anion exchanger Purolite S 985 accord-

ing to the Boyd – Adams model for gel kinetics and effect of HCl concentration: (a) sorption of

Pt in the presence of Rh; (b) sorption of Rh in the presence of Pt

(С (Pt)=С (Rh)=0.25 mmol/L)

501


Table 6

Kinetic parameters of sorption concentration of Pt (II, IV) and Rh (III) at their simultaneous presence from chloride solutions

Trade name C (HCl) t1/2, s D⋅10 8, сm 2/s υ⋅10 5, mmol/g⋅sPurolite А 500 Pt 2.0 427 3.71 1.60

0.1 356 4.45 2.57

0.01 306 5.18 2.80

Rh 2.0 1363 1.16 0.79

0.1 905 1.75 0.90

0.01 658 2.41 1.23

Purolite S 985 Pt 2.0 308 5.15 3.45

0.1 206 7.69 4.08

0.01 174 9.11 6.41

Rh 2.0 363 4.37 2.61

0.1 258 6.14 3.56

0.01 182 8.71 3.76

АМ-2B Pt 2.0 531 7.64 1.84

0.1 338 12.00 2.04

0.01 307 13.21 2.98

Rh 2.0 1501 2.70 0.83

0.1 517 7.85 1.62

0.01 488 8.31 2.02

It can be seen from table 6 that average diffusion coefficients are on the level of 10–8 cm 2/s

and the sorption rate is on the level of 10–5 mmol/g⋅s. With the increase in acid concentration in

contacting solution, the half-exchange time is shorter, whereas the values of diffusion coeffi-

cients become greater for all the sorbents investigated. In case of rhodium recovery, the diffu-

sion coefficients values are bigger and the half-exchange times are shorter than for platinum

sorption. Therefore, the kinetic parameters comply with our perception of the resins selectivity.

Based on the results obtained, the investigated anion exchangers can be recommended for

recovery of chloride complexes of platinum (II, IV) and rhodium (III).

REFERENCES

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ovskiy Obrazovetelny Zhurnal. 1999. V.11. P. 45–49.

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Moscow: Editorial URSS, 2003. – 592 p.

3. Spektor, O. V.; Ryumin, A. I.; Pochekutova, M. G. Methods for recovery of platinum group

metals from spent catalysts.//Tsvetnye Metally. 1998. V.7. P. 31–39.

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and Rh (III) using anion-exchange solid-phase extraction cartridges (SPE)//Solvent Extraction

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Nauka, 1991. – 248 p.

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and Platinum. Oxford: Pergamon Press, 1967. – pp. 155–180.

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imiya, 1980. – 356 p.

502


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platinum group metals. Minsk: Izdatelstvo Universitetskoe, 1990. – 241 p.

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503


Development of scientific and technical progress, introduction of scientific achieve-

ments in various areas of the industry (aviation, space etc.) have led to increase in the require-

ment for many nonferrous and rare metals, but they are mostly claimed in an iron and steel in-

dustry where it is made about 70 various metals, and also in a machine-building complex. Within

last decade world extraction has increased on: nioby – 35 %, tantalum – 15 %, titan – 32 %, cop-

per and antimony – 30 %, molybdenum – 24 %, tin and bauxites – on 10 % [1, 2, 3].

Placing of manufactures on extraction of metals and the extraction is defined by a number

of factors: geological (metalloheric territory specialisation), economic (demand for mineral raw

materials, conditions of working out of deposits), historical (formation of an infrastructure of

mining and mountain-metallurgical areas), political (basically, an export-import policy) and

social (preservation and creation of workplaces by means of government support, i. e. special

state social programs).

Prominent aspect for development of a mineral-raw-material base (MRB) of metal miner-

als are factors of placing of metallurgical manufacture: raw (location of manufactures near to

sources of mineral raw materials), fuel and energy (affinity of manufacture to cheap sources of

the electric power) etc.

Now the great value gets the consumer factor (capacity of commodity markets), and also the

transport factor. The important tendency of placement of the metallurgical enterprises is gravita-

tion to the centers of consumption of the metals, having the corresponding infrastructure, the

prepared qualified manpower etc. Besides, the establishment of close connections of the metallur-

gical enterprises with their clients, coordination in manufacture and sale planning is observed.

For the nonferrous metallurgy enterprises it is now noticed worldwide strengthening ten-

dency of the power orientation in the enterprises placing that can be explained by the develop-

ment of metallurgy of light nonferrous metals and rare metals. Besides, it is observed a shift of

metallurgical manufacture into developing countries which have more mineral raw materials.

As a whole, as a result of the theoretical analysis of MRB rare and nonferrous metals, it is

possible to draw following conclusions:

– Metal minerals, including nonferrous and rare metals, are claimed in many industries of

the world;

– Consumption of nonferrous and rare metals grows all over the world, thus in developing

countries the bottom stages of a production cycle prevail, and in economically developed coun-

tries – high stages (including manufacture of rare metals);

– Economically developed countries pursue a policy of preservation of own stocks of min-

eral raw materials, increase import; (for example, Belgium produces metals, without having its

own MRB).

– The considerable volume in manufacture of nonferrous and rare metals occupies sec-

ondary raw materials, thus secondary raw materials and metal semiproducts export not only

developing countries, but also the economically developed ones (export of semiproducts of alu-

minium and titan to the developed countries is connected with high power consumption of their

manufacture);

– Formal security by mineral raw materials of the majority of nonferrous and rare metals is

high (tens and hundreds years), and real security, taking into account growth of extraction and

consumption of metals and deterioration of mountain-geological and economic-geographical

service conditions of ore deposits, is much less.

All factors set forth above and tendencies of development of mountain-metallurgical man-

ufacture are of great influence on development of MRB valuable rare and nonferrous metals

both in the world and in the Russian Federation.

PROSPECTS OF USE OF THE NONFERROUS AND RARE METALS CONTAINING IN COALS AND COAL ASHES OF KUZBAS

FOR THE IRON AND STEEL INDUSTRY OF SIBERIAN REGION

V.A. Salikhov, E.S. LjubushkinaNovokuznetsk branch of the institute state educational highest vocational institution


504


Analysis MRB of nonferrous and rare metals in the Russian Federation shows that the Rus-

sian Federation, despite the crisis phenomena of a transition period, on the former takes leading

places on stocks, extraction and consumption of the majority of nonferrous metals. At the same

time it is necessary to notice that almost all deposits of nonferrous metals of Russia differ from

geografo-economic conditions of working out that reduces profitability of their development.

As consequence, the share of active stocks decreases, because of the high prices for the electric

power and transport transportations, more than half of reconnoitered deposits of nonferrous

and rare metals are unprofitable. Thus in connection with gradual lifting of the domestic in-

dustry by 2020 in the country growth of consumption of nonferrous metals (approximately in

1.5–2 times) [1] is predicted.

Analysis MRB of rare metals in the Russian Federation shows that consumption of rare

metals all over the world actively grows, and it will be obligatory to grow in Russia; Russian MRB

possesses large stocks of almost all rare metals, but in a qualitative sense requires improvement;

developments MRB of strategic rare metals demand interests of national safety. To the list of

strategic kinds of the mineral raw materials, confirmed by the Order of the Government of the

Russian Federation № 50 from 1/16/1996, from among rare metals are mentioned Li, Be, Nb,

Ta, TRY, Zr, Ge, Re, Sc [3].

Thus, carried out analysis MRB of nonferrous and rare metals in the world and in the Rus-

sian Federation confirms an urgency of use of the valuable nonferrous and rare metals contain-

ing in a waste of mineral raw materials. In the Kemerovo region considerable volumes of a waste

of the mineral production which great part is made by ashes-slag waste of coals are saved up.

They correspond to definition of technogenic deposits, as large-tonnage congestions of a waste

of extraction and processing of mineral raw materials which can be used with economic effect.

Thus it is possible to form such deposits by kinds of passing useful components taking into ac-

count their structure and concentration.

The complex geologo-economic estimation of deposits based on the account of the basic

and passing useful components, essentially raises economic potential of the reconnoitered stocks,

allows to conduct their rational working out, and also promotes introduction in operation of re-

munerative deposits. Now technologies of extraction from mineral raw materials and a waste

of many valuable metals, the last laboratory and semiindustrial tests are developed. A number

of metals is taken in small amounts in the industrial way (Ge, V, Ti, Zr). For example, the bra-

zilite for zirconium production is taken in small volumes from Hibin apatite deposits nepheline, grothite for getting titanic products. As a whole, complex working out is spent insufficiently, and

an accumulated waste of extraction and processing of mineral raw materials, i. e. technogenic de-

posits are also used in small volumes. It can be explained by insufficient financing of research and

development, the high cost price of experimental technologies, ecological danger of developed

methods on extraction of metals. At the same time, decrease in profitable stocks of ore miner-

als assumes necessity of extraction of metals from complex ores, and also their extraction from

technogenic deposits. Efficiency of working out of technogenic deposits, according to the author,

is possible to estimate by means of the differential rent I (in comparison with ore deposits) and II

(in comparison of methods of extraction of metals from a waste of mineral raw materials), and

also by means of the dynamic rent (the additional income during the late periods of time).

Thus the main methodological principles at an economic estimation of technogenic depos-

its are principles of systematic and integrated approach. The principle of systematic allows to

consider not only geological, but also technological, and also social and economic and ecologi-

cal aspects of realisation of projects on extraction of metals, for example from a waste of coal

production. The integrated approach principle allows to carry out researches on the basis of

mountain-geological and social and economic methods. The great value has a dynamic prin-

ciple, and also a principle of priority development. The dynamic principle provides regular trac-

ing of a situation, especially external social and economic and ecological factors, for acceptance

of adequate decisions. The principle of priority development allows to predict situations in the

markets of mineral raw materials for 10–20 years.

One of perspective directions of providing of an iron and steel industry with valuable and

extremely scarce metals is use of coals and a coal waste which are estimated as a potential raw-

material base of an iron and steel industry. For example, ashes-slag waste of the USA could

provide not less than half of annual requirement of the country in such elements, as As, Be, Co,

Ga, Ge, Hf, Nb, Se, Sr, Те, Tl, Y, and also in rare-earth elements. From this waste even at modern

technologies such metals, as Al, Cd, Ga, Ge, Fe, Mo, Ti, V and Zn can be taken out. Separate coals

with raised (more than 5 gr/ton) the content of germanium are considered as ore. [4].

505


The Kuznetsk basin is one of the largest coal fields of the world with the reconnoitered

stocks of coals about 60 billion ton, including 30 billion ton coked coals and about 20 billion ton

coals of especially valuable marks; annual extraction makes more than 100 million ton. The es-

timation of the maintenance of valuable nonferrous and rare metals in coal deposits on geologo-

economic region of Kuzbas, made on the basis of the received by the author data, and also on the

basis of the theoretical analysis of works of other researchers, has shown the following:

1) On deposits of Kuzbass there are some anomalies of several valuable rare and nonfer-

rous metals (such as: Ti, Zr, V, W, Y, Co, Ga, Ge, Nb, Be) which are more often connected with

tectonic breaks of layers, especially if layers are located near intrusive files.

2) Basically, the raised concentration of these metals are observed in southern and southeast

geologo-economic region (Bajdaevsky, Bachatsky, Mrassky, Tom’-Usinsk, Uskatsky, Osinovsky).

3) Practical interest, from the point of view of concentration of the metals, recommended

to an estimation, can represent the titan and zirconium, and also a number of nonferrous and

rare metals (Sr, Zn, Pb, Cu, V). Under a condition of complex extraction some interest also rep-

resent Ga, Ge, Be, Nb, W.

4) Actually nonferrous metals form, as a rule, insignificant anomalies in zones of tectonic bro-

ken layers, which more often connected with located hypsometric lower ore bodies of polymetallic

deposits. The raised concentration on layers are formed only by aluminium (to 10 % and more).

5) Average maintenance in coals of rare and rare-earth metals are close to clarke mainte-

nances or below it; practical interest (at selective working off and complex extraction) is repre-

sented by the anomalies of metals dated for zones of tectonic breaks.

6) a number of nonferrous metals (copper, zinc, lead), forming on separate coal deposits

on layers of anomaly of hydrothermal genesis can be taken together with rare metals, and alu-

minium (in the presence of requirements and technologies) – separately.

7) Special practical interest is represented by ashes-slag waste of processing of coals in which

the maintenance of valuable nonferrous and rare metals can increase to 10 times and more.

8) In Kuznetsk coals steady concentration of precious metals (gold, platinum and silver)

which would represent practical interest are not revealed.

9) The maintenance of radioactive elements (uranium and thorium) in Kuznetsk coals is

much lower, than in other coal basins, but it is necessary to reveal their abnormal concentration

representing ecological danger.

10) The maintenance of some toxic metals (mercury, chrome, arsenic, etc.) in Kuznetsk

coals does not exceed 2–3 g/t, as a rule, and more often it is much lower; their extraction can be

made together with valuable nonferrous and rare metals, regarding ecological reasons.

Thus, the greatest interest for practical use represent titan and zirconium, and taking into

account complex extraction – zinc, barium and vanadium, occasionally – manganese and yt-

trium, and in perspective – strontium, niobium, gallium and germanium. Toxic and potentially

toxic elements (arsenic, beryllium, manganese, nickel, lead, chrome, etc.) are revealed, but

their maintenance is much less than maximum concentration limit, therefore their passing ex-

traction is expedient.

Practical interest to complex use of mineral raw materials is supported recently with per-

fection and creation of new technological schemes of enrichment and extraction of metals. For

example, there is variety of the technological methods allowing to utilize effectively some kinds

of ashes-slag waste for the purpose of extraction of some useful components from them. Thus,

in the middle of the XX-th century they extracted aluminium and iron. The aluminium mainte-

nance in coals exceeds 10 % that makes possible its industrial extraction. The magnetic fraction

contains from 30 to 60 % of iron, and also in small amounts – cobalt, nickel, titan and other valu-

able nonferrous and rare metals [4].

In the Kemerovo region the large metallurgical enterprises of black and nonferrous metal-

lurgy (ZSMK, NKMK, Novokuznetsk aluminium and «Kuznetsk ferroalloys» factories) are lo-

cated. Considering poor quality and great labour expenses while enriching local aluminium ores

(bauxites and nepheline syenite), aluminium can be considered perspective additional mineral

raw material. Low security of Siberian region with profitable iron-ore makes extraction of iron

and other ferrous metals from ashes-slag waste deposits perspective. [5].

Directly from coals in the course of coking or from harms of ablation at power-generating

coal burning they receive germanium. The basic source of its obtaining in the USSR was coal.

Germanium is made basically on chemical-recovery factories from Donetsk coals or from ger-

manium-rich power-generating coal. The technology of complex extraction germanium, gallium

and other rare elements from cindery ablations is well fulfilled. Extraction of scandium from

506


coals, very expensive metal with small volumes of extraction is perspective both in Russia and in

the world, however the process of preliminary enrichment of ashes is extremely complicated:

primary opening and metal transfer in a solution with an exit 60–80 % is possible only in an

autoclave that makes the process considerably more expensive.

One of effective ways of decision of this problem is sorption leaching of ashes-slag waste.

Similarly, i. e. with the help of leaching (for example, thermochlorination), it is possible to take

gold, lithium, vanadium, tungsten, yttrium, rare-earth and other elements. So, from ashes-slag

waste of power-generating brown coals it is taken to 40–67 % of the titan, 45–77 % of a beryl-

lium, 70–87 % of copper, 50–81 % of manganese, 74–84 % of arsenic, 48–60 % of vanadium and

62–83 % of gallium [4].

Predesigns show that on one pilot production on extraction of metals from ashes by a

method of thermochlorination it is possible to process 2 thousand ton ashes within a year and

to receive nearby 10 ton of titan, 10 ton zirconium, 1 ton vanadium and 100 kg of gallium. Thus

the net profit will make about 30 million rbl. Shop on processing of ashes-slag waste can be a

part of a mine, and the project can work on the basis of cooperation of tehnologo-economic rela-

tions between the coal-mining, processing, fuel and energy and metallurgical enterprises, i. e.

to include consumers of coal and suppliers of ashes [5].

Calculation of stability of the project shows, that even capital and operational expenses in-

crease to 100 % (≈ 10–15 % a year) and profit reduces also to 2 times, the project pays off within

5 years.

In 2006 the rise in prices for nonferrous and rare metals has considerably outstripped in-

crease in cost of capital and operational expenses (table 1).

Table 1

Change of the prices for metals (2000–2006)

Mineral raw materials and repartition

products

Unit Price, US dollar

2000 2006

Tungsten kg 10 45

Vanadium kg 10 40

Aluminium metal kg 1.5 2.5–3

Strontium metal kg 60 120

Copper refined kg 2–2.5 7–11

Nickel kg 5–8 30–35

Tin kg 5.5–7 10–15

Lead kg 0.6–0.9 1.5–2

Zinc kg 1–1.3 3.5–4

Molybdenum kg 5–10 80

titan spongy kg 10 20

Zirconium spongy kg 25 30

Nioby kg 60–75 230–240

Gallium kg 380–400 1200

Germanium kg 825–1300 2500

The prices (in 2006 in comparison with 2000) on aluminium, lead, tin, cobalt have in-

creased 2 times, on zinc, nickel, gallium in 3 times, on vanadium and tungsten 4 times, and on

copper 5 times. By December, 2008 there was a sharp reduction of prices on nonferrous metals

(1.5–2 times), the prices for the majority of rare metals till August, 2009 remained steady and

had tendencies to grow, but then the prices have decreased (> 50 %). By the end of 2009 the

price for rare metals have returned on level of August, 2009; the price on nonferrous metals in

2009, basically, had tendencies to decrease, the insignificant rise in prices is noted by the end of

the year. This data confirms unstable character of the market of valuable nonferrous and rare

metals. At the same time dynamics of the prices is in the limits providing stability of the project

of valuable metals extraction.

507


Operation of technogenic deposits allows to save considerable financial means (hundred mil-

lion rbl. and more) on carrying out geological surveyance and building of the mining enterprises

(indirect effect), and also improves an ecological situation. Ashes deposit areas of fuel-power sta-

tions in Kuzbas occupies several thousand hectare, and their volume is about 40 million ton. Tak-

ing into account high cost of farmland of 1 hectare (> 1 million rbl.) and expenses for storage

of sailings ≈ 20 rbl./t a year, ecological effect is considerable and reaches the sum of hundreds

million rbl.

Thus, realisation of the project of valuable metals extraction from ashes-slag coal waste,

leads to considerable total economic effect for the region.

Besides, low expenses of processing ash disposal area to get concentrate (according to the

available data about 1000 rbl./t), preferential taxes on subsoil use, on profit, VAT decrease, the

high prices for metals and value factor of metals extraction from a concentrate (≈ 1), allow to

lower size of the minimum maintenance, i. e. take f larger spectrum of valuable nonferrous and

rare metals.

In Kuzbas annually collects to 15 million ton of ashes-slag waste (including metallurgical

and other enterprises one), where out of 2.6 million ton of an annual exit of ashes and slag of

fuel-power stations, 2.4 million ton by the way of hydroremoval goes to sailings in the form of

ashes-slag mixes, (now it is accumulated ≈ 40 million ton of them and it is possible to use not

less than 20 million ton in concrete and solutions.). With the account of low maintenances of the

majority of nonferrous and rare metals in coals (basically these are epigenetic and, less often,

syngenetic anomalies), extraction of valuable nonferrous and rare metals from ashes-slag waste

of fuel-power stations represents practical interest, where their stocks can make thousands ton,

ten thousands ton and more (table 2). Valuable metals can be taken also from ashes of ablation,

where their maintenance 2–3 times more than in ash disposal area.

Table 2

Maintenances of some nonferrous and rare metals in Kuzbass coals

Metal Maintenance in

coals, g/t

Concentration

recommended to an

estimation, g/t

Maximum

maintenance in

ashes, g/t

Standards for

ores, %

Titan 100–500 500 5600 10–15

Zirconium 100–500 500 3000 3

Copper up to 15 100 3700 0.5

Lead up to 25 50 4800 2

Zinc 10–300 100 16000 1

Barium 200 1000 5800 1

Vanadium up to 50 100 5000 1

Tungsten up to 3 100 1500 0.5–1

Strontium 100–500 1000 2300 5

Nioby 1–3 100 3000 0.1

Gallium 1–3 20 3000 0.04

Germanium up to 1 10 2700 0.1

Considering high demand of the industry of the Russian Federation for titan and Zirco-

nium (accordingly: 600–700 and 100 thousand ton a year) and low level of extraction (5 and

3 thousand ton a year), carried out on Lovozersky and Kovdorsky ore-dressing and processing

enterprise (Murmansk area) and, hence, a high share of import – it is necessary to consider ex-

pediency of passing extraction of these scarce metals on the basis of available skilled technolo-

gies of their extraction from ash-slag waste (for example, leaching, thermochlorination, etc.),

which are less power-hungry in comparison with traditional ones (processing titan – and zirco-

nium rich silicates). Besides, considering requirement of nuclear technics, electronics in Sr, Y,

Ga and Ge the careful estimation of these and other rare elements in ashes dump of fuel-power

stations, with prospect to use these metals in the future, considering technological possibilities

of their extraction. Also complex extraction together with rare metals and polymetals, such as

Cu, Pb and Zn. is regarded perspective.

508


The further development of the industry of Kuzbas, especially metallurgy and machine-

building complex, the increase of local MRB demand some black, nonferrous and rare metals.

Competitiveness of the mountain-metallurgical companies is defined by having their own MRB,

production assortment, consolidation of separate manufactures. Technogenic deposits can be

potential MRB of many valuable nonferrous and rare metals.

In the Kemerovo region branches of manufacture of a mountain-metallurgical complex of

the industry are traditionally developed. The metallurgical enterprises – the Western-Siberian

and Novokuznetsk metallurgical industrial complexes (WSMK and NKMK), Novokuznetsk alu-

minium factory, Open Society «Kuznetsk ferroalloys» bring the considerable contribution to na-

tional economy and areas. Black and nonferrous metallurgy provide about 40 % of industrial

output of Kuzbas; 20 % of tax revenues in the regional budget and more than 40 % of currency re-

ceipts. In nonferrous metallurgy Novokuznetsk aluminium factory, the fifth enterprise by size in

the country, provides more than 90 % of all production in area. The coal-mining companies, the

mining enterprises for extraction and processing of ores black, nonferrous and precious metals

work actively. Today the Kemerovo region provides more than 50 % of a national coal mining

(including 79 % – coked marks); more than 50 % of export of coals (basically, coked marks).

At the same time it is necessary to note instability of development of Kuzbas economy. The

basic sources of the Kemerovo region budget financing – the tax to incomes of physical persons

(approximately 25 %), payments for using natural resources (more than 10 %), the profit tax

(more than 10 %) and the tax to property (about 10 %). To lower the budget deficiency it is nec-

essary to increase incomes by one and a half time.

Now manufacture growth occurs, basically, in so-called corporate sectors of economy

which include the large mountain-metallurgical companies (Evrazholding, SibCEK, «Russian

aluminium and others). At the same time slump in production in mechanical engineering and in

other not corporate sectors of economy (light, food-processing industry, agriculture etc.) is

observed. In area economy the share of small enterprises (especially high technological ones)

is insignificant, volume of investments into this sector of economy is insignificant too. The Ke-

merovo region lags behind the next Novosibirsk and Tomsk areas in sphere of innovative activ-

ity of economy. Expenses for technological innovations in Kuzbas three – ten times less than in

other industrial region analogues.

One of variants of development of the high technological small enterprises is creation of

compact manufactures on processing of technogenic waste and extraction from them valuable

metals. These manufactures can be created as a part of the mountain companies on the basis of

cooperation of tehnologo-economic relations between the power and coal-mining enterprises.

Thus, small enterprises will get financial support for innovative transformations, i. e. for intro-

duction of methods of extraction of valuable metals from technogenic deposits on the basis of

scientific-technical progress. The manufacture diversification will provide additional profit to

the coal-mining enterprises, i. e. will raise their financial stability in the conditions of stable

work in difficult mountain-geological conditions.

Out of ashes dump of fuel-power station accumulated within a year at the power enterpris-

es of Kuzbas area it is possible to take not less than 100 ton Ti, Zr, Sr,> 10 ton V and Ga,> 1 ton

Nb, Ge and of some other rare metals. Considering that fact that for extraction of valuable met-

als can be used not less than 20 million ton accumulated in the Kemerovo region of ashes dump

of fuel-power station, it is possible to assume that volumes of extracted metals from them will be

10 times more. Comparing predicted volumes of extraction of valuable metals from ashes-slag

waste of the coals accumulated on Kuzbas area territories, with requirements of the industry

of the Russian Federation (table 2), it is possible to assume that these metals can be claimed,

first of all, in the industry of the Kemerovo region and Siberian region. For example, according

to the experts, predicted stocks of rare metals in complex apatite-magnetitovyh and rare metal ores and their waste in Murmansk area (a leading potential source of strategic raw materials

of Russia) make billions ton (ores Ti), millions ton (ores Zr), considerable stocks of Ta, Nb and

other rare metals. Their annual extraction can make from tens (Ta, Nb) and thousands ton (Li)

to tens (Zr) and hundreds thousand ton (Ti). Thus, look-ahead indicators of stocks and extrac-

tion of rare metals here are several times above, than in ashes dump of fuel-power station of the

Kemerovo region.

The metals received from ashes-slag waste of coals of Kuzbas, can be claimed in electro-

technical mechanical engineering of the Kemerovo region, and also on machine-building, fer-

roalloy and other enterprises of Sibirian Federal Units (Novosibirsk and Irkutsk areas, Krasno-

yarsk region). A number of valuable, scarce metals (such as Ti, Zr, Ga, Ge) can be claimed in

509


other Russian markets and, probably, on foreign markets. Profit from metal realisation it is pos-

sible to define, taking into account conditions of sales «spot» and «future», under the formula:

P = (А1 × Pr 1 + А2 ×Pr 2 + Аn× Pr n) – E, (1)

Where:

P – profit on realisation of metals, rbl.;

А1 – Аn – annual manufacture of grades of metals, ton or kg;

Pr 1 – Prn – the high-quality prices per unit of output, rbl.;

E – expenses for manufacture and realisation of metals, rbl.;

Volumes of annual manufacture of metals are defined on the basis of marketing researches

of internal and foreign metals markets.

Thus, it is possible to consider the ashes-slag waste of the coals which have been accumu-

lated in Kuzbas, as potentially perspective mineralno-raw-material base for an iron and steel in-

dustry. It is especially important at increase of demand for metals on a foreign market, and in

the near future –on internal one.

REFERENCES

1. Kozlovskij E. A., Malyutin JU. S.mineralno in economy of Russia//Markshejderija and

minerals deposit development. 2002. – № 2. – With. 8–28, 2002. – № 3. – p. 6–18.

2. Novikov A. A., Blagutin J. L., Pinchuk A. V.problem of strengthening and expansions

of a mineralno-raw-material base of nonferrous metallurgy in Russia//Mountain magazine. –

2003. – № 10. – p. 58–62.

3. The Mineralno-raw-material base of rare metals in Russia: condition and ways of

development/M. F. Komin, T. J. Usova, T. I. Zueva, D. S. Kljucharev, etc.//Investigation and pro-

tection of subsoil. – 2004. – № 11. – p. 32–37.

4. Rare elements in coals of Kuznetsk basin/Arbuzov S. I., V. V. A. A., Rihvanov L. P. – Ke-

merovo, 1999. – 248 p.

5. Salikhov V. A. Scientific bases and perfection of a geologo-economic estimation of pass-

ing useful components of coal deposits (on an example of Kuzbas)/V. A. Salikhov; SibSIU. –

2 edit.,. – Kemerovo: Kuzbasvuzizdat, 2008. – 249 p.

510


Rhenium recycling has recently become a burning issue. This is due to many factors, the prin-

cipal one being runaway prices resulting from disbalance of supply-demand at the rhenium mar-

ket. Peaking at 2008 (USD 10560\kg) [1], the Re prices remain fairly stable (6600 $/кг) [1],

notwithstanding the ongoing economic crisis.

In Russia, there is a special interest in rhenium owing to its loss as the result of the USSR

disintegration. Nowadays, both the mineral sources and production capacities are in Kazakh-

stan, Uzbekistan and Armenia, and the Russian machine building totally depends on imported

rhenium, needing it for the strategic aircraft- and spacecraft industries. Lacking commercial-

ly developed raw material resources, we are turning to Re-enriched (up to 7–9 %) [2] wastes

of complex heat-resistant nickel alloys (HRNA). According to VIAM [3], the HRNA annually

dumped in Russia amount to 25–35 t. Even with an averaged rhenium content of 2.5 %, this may

yield about 700 kg rhenium. Moreover, there are other valuable and expensive, rare and non-

ferrous metals (W, Mo, Nb, Ta, Ni, Co, Cr, etc) contained in multicomponent alloys, that may

also be regenerated to achieve a really in-depth processing of the starting materials.

As of today, we can enumerate several technologies of rhenium recycling from alloys [4].

However, according to A. Lipmann, an authority in the «minor-metal» issues [5], the industrial

application of the methods for HNRA processing is limited due to their high cost (1 kg of sec-

ondary rhenium costs ~3000 USD\kg) and low effectiveness (the yield is ~70 %). Hence is the

necessity to develop effectual technologies of rhenium recovery from HRNA.

This work deals with rhenium recycling from the following products:

– From grinding wastes of HRNA parts of the ZhS-32 type (fig. 1), representing a fine-

grained powder (<0.16 mm, >95 mass/ %);

– From HRNA lump debris (up to 5 kg) of the ZhS-32 alloy (fig. 2).

The grinding of HRNA lump debris

As multicomponent products, HRNA are separated using pyrometallurgical methods [4].

The problem in this case consists in grinding the high-strength lumps (for instance, blades and

other elements of turbine engines weighing several kg). Proceeding from own experience of hard

nickel alloys processing [4], we proposed high-temperature alloying of the lumps with granulat-

ed aluminium. Melting at 1500–1700 oС in an induction furnace for 0.5 h yielded a homogeneous

alloy of a 4-kg charge with a mass ratio of m (alloy) : m (Al)=5:1. The nickel base of HRNA was

converted to Ni3Al aluminide characterized by brittleness at low-temperatures, which permitted

to grind the alloy after cooling in standard equipment (jaw crusher, disc attritor). Both the grain-

size composition of the resulting alloy and the chemical composition of starting and dispersed

rhenium-containing alloy wastes are presented, respectively, in tables 1 and 2.

RHENIUM EXTRACTION FROM NICKEL-BASED COMPLEX HEAT-RESISTANT ALLOYS

A.G. Kasikov, A.M. Petrova, V.T. Kalinnikov

I.V. Tananaev Institute of chemistry and technology of rare elements

and mineral raw materials of Kola science centre of RAS, Apatity, Russia

Fig. 1. Rhenium-containing grinding

wastes of HRNA

Fig. 2. HRNA lump debris (ZhS-32 alloy)

511


Leaching of HRNA wastes

Analysis of technological solutions proposed for the processing of rhenium-containing al-

loy wastes (for details, see [4]) has revealed that irrespective of the decomposition method,

rhenium can be extracted by two ways:

– by directly passing rhenium from wastes to solution (using, for instance, oxidizing distillation

of Re2O7 and trapping the sublimate to solution; by oxidizing-thermal decomposition followed by

leaching of the cake with water; by electrochemical oxidizing or acid leaching in oxidizing media);

– leaching of the alloy base metals and concentrating rhenium in the residue, where-

from it is extracted (using electrochemical and acid methods to dissolve the base)

We have tested acid leaching of rhenium-containing alloy wastes both in oxidizing environ-

ment and in the absence of an oxidizer.

The leaching of fine-grained products in laboratory conditions was performed in 2-litre, me-

chanically agitated constant-temperature open glass reactors on the basis of IKA Werke elements.

Table 1

Grain size of the ground product obtained after melting of lump wastes with aluminium

Fraction, mcm 250–300 200–250 160–200 125–160 71–125 40–71 -40

Content, mass % 13.06 9.54 10.45 13.80 15.00 16.25 21.90

Table 2

Starting product composition

Product Content of elements, mass %

Ni Co Cr Mo W Re Al Ta Nb

ZhS-32 alloy 61.0 9.0 4.9 1.0 8.5 2.6 5.9 2.1 1.6

Grinding wastes 61.0 6.8 3.4 0.9 4.3 2.3 18.2 1.6 1.1

Ground alloy with Al 48.0 6.3 3.1 0.9 6.5 1.5 32.0 1.1 0.7

The leaching of rhenium-containing wastes in oxidizing conditions was performed by 1.5–

6 mole/l H2SO4 solutions at S:L=1:10–20 and a temperature of 70–85 oС. Since earlier it has been

established that adding a peroxide oxidizer at the initial stage is undesirable due to a highly in-

tensive interaction between the active fine-grained materials and H2SO4 with H2 evolution and

self-heating of the reaction mixture to 70–90 oС, the first 2–3 hours the process was carried out

without an oxidizer. After the reaction of the metal components with H2SO4 was largely over, the

reaction mixture was gradually supplied with an oxidizer (for 1.5–2 h). At different times, the oxi-

dizers were H2O2, Na2S2O8, (NH4)2S2O8, K2S2O8 solutions. Satisfactory results were obtained only

with the first two reagents. In the latter two cases, the components extraction to solution was in-

sufficient either due to the formation of nickel-ammonia alums or low solubility of potassium

perrhenate. The concentrated oxidizing solution was fed to the reactor either in discrete portions

every 10 minutes, proceeding from Voxid.:Vleach sol.=1:5, or continuously, drop-wise, at a rate main-

taining the assigned redox potential value (RDP). The continuous feeding was performed using a

Masterflex C/L microdosing peristaltic pump. The RDP was controlled by measuring it in the reac-

tion mixture relatively the chlorine-silver-saturated electrode at a Tsch-300 digital voltammeter.

The acids tested for the leaching of the alloy base were hydrochloric and sulphuric ones.

The process occurred similarly to that described above, until the reaction between the metallic

components and mineral acid (for 2–3 h) was over. In view of the vigorous H2 evolution in the

case of fine-grained alloy and HCl interaction, the material was fed in batches.

For non-ferrous metals, the resulting solutions were analyzed by the AAS method at a Shimad-

zu ICPE-9000 atom-emission spectrometer. Rhenium and refractory metals were analyzed by the

AAS IPC method at a Plasma-400 atomic-emission spectrometer with induction-bound plasma.

It was established that in the case of an oxidizer-free acid leaching, the nickel base dissolves,

without much rhenium passing to the solution (table 3). The solid phase diminishes by 82 mass %,

which makes it possible to concentrate rhenium in undissolved residue, where its content may in-

crease to 9.3 mass %. It should be noted that hydrochloric acid is preferable for base leaching,

because there arise no problems with subsequent separation of non-ferrous metals in chloride me-

dia. From a solid concentrate containing Re, Nb, Ta, and W, rhenium can be extracted in the form

of Re2O7 by either hydrometallurgical methods or by high-temperature distillation.

512


Table 3

The level of some elements recovery to solution in an oxidizer-free leachedground alloy with Al. S:L=1:15, Т=80 oС, τ =3 h

СAcid, mole/l Acid Element recovery, %

Ni Mo Re

1.5 H2SO4 89.8 – 1.5

2.5 H2SO4 94.2 11.5 2.0

3.0 H2SO4 95.4 13.0 2.1

4.0 H2SO4 92.5 20.0 5.6

5.0 H2SO4 96.2 30.1 8.0

3.0 HCl 55.2 20.2 0.3

6.0 HCl 83.2 – 0.8

An oxidizer, added to acidic leaching, noticeably promotes the rhenium recovery. Since

Re is better isolated in sulphuric acid, the leaching of alloy wastes was performed by H2SO4.

table 4 presents the results of rhenium, and some other metals, extraction from grinding wastes

using different oxidizers in concentrated solutions. Evidently, the most convenient of them is

hydrogen peroxide, effectively converting rhenium to solution without introducing impurities.

Concentrated H2SO4(10N) solutions are necessary to dissolve the non-ferrous metal base to ob-

tain leaching solutions with high enough acid contents, from which rhenium is extracted by the

method described in [6].

Тable 4

The level of element recovery to solution during the oxidizing sulphuric-acid leaching of grinding wastes of the ZhS-32 type

Oxidizer,

Vоxid: Vsol=1:5

Element recovery, %

Ni Co Cr Re Mo

HNO3 99.9 99.9 98.3 98.6 77.2

H2O2 99.8 99.8 98.2 99.8 80.0

Na2S2O8 98.8 – – 70.0 66.9

Special experiments on optimizing of the oxidizer consumption have revealed that a practi-

cally total rhenium extraction can be achieved by adding hydrogen peroxide in amounts neces-

sary to maintain ROP at 0.55–0.75 V for 2–3 hours, which noticeably reduced the H2O2 con-

sumption compared with similar experiments.

Rhenium solvent extraction from waste leaching solutions

For selective solvent extraction of rhenium from HRNA waste leaching we tested, on both

laboratory and larger, scales the extraction with secondary octyl alcohol [6]. This extractant is

selective vis- -vis rhenium in a sulphuric acid medium in the presence of molybdenum and other

non-ferrous metals [7]. Having a high capacity in terms of rhenium (~100 g/l), it can isolate

this element from fairly concentrated solutions.

In laboratory experiments, extraction from leaching solutions was carried out in glass sepa-

rating funnels with a volume of 0.25–0.5 l at a room temperature (20±2 oС) and under mechani-

cal agitation. The extractant was domestic, «pure» brand 2-octanol, available and inexpensive.

To prevent additional extraction of the acid from leaching solution, the extractant was prelimi-

narily saturated with H2SO4. The time of contact between the phases was 5 min, the O:W ra-

tio varied between 1–5:1, depending on operation.

According to laboratory data, extracting from a solution after polishing wastes leaching con-

taining, g/l: Ni – 36.0; Со – 5.7; Сr – 3.6; Мо – 0.7; W – 0.01; Re – 2.4 and 4.2 mole/l H2SO4 in

one stage at О:W=1:1 yielded 97.9 % rhenium at a 22.8 % co-extraction of Мо. After washing

with water at О:W=2:1, the bulk of Mo, not more than 20 % Re and practically the entire of

co-extracted H2SO4 passed to the scrubbing water. After re-extracting with a 3 mole/l NH4OH

solution at О:W=1:1, we obtained a solution containing 1.9 g/l Re and 0.04 g/l Мо, which

means a 79.0 % level of rhenium extraction.

513


Similar results were obtained in the course of an experiment with leaching solutions of

ground alloy with added Al. The leaching solution had a concentration of, g/l: Ni – 21.8; Co –

3.0; Al – 15; Cr – 1.4; Mo – 0.36; W – 0.25; Re – 0.75 and 3.7 mole/l H2SO4. In one stage at

О:W=1:1, the extraction to the organic phase was 97.3 % Re and only 19.4 % Mo, with 61.9 %

of the latter and the bulk of H2SO4 recovered in a one-step scrubbing at О: W =5:1, at an insig-

nificant removing of rhenium. So, re-extracting with 3 mole/l NH4OH resulted in a purified rhe-

nium-containing solution with rhenium and molybdenum concentrations of 0.5 g/l and 0.03 g/l,

respectively. The Re extraction to ammonia re-extract was ~80 %.

Apparently rhenium losses with scrubbing water can be eliminated by returning it to the

process, for instance, to the extraction stage or to the leaching solutions preparation.

The extraction process was organized in a laboratory cascade of mixer-settler extractors pro-

duced at ICTREMRM RAS in counter-flow regime, with returning of scrubbing water to extrac-

tion. This allowed to improve rhenium extraction from Re-containing alloy leaching solution.

As a starting solution in 2-octanol-based, large-scale laboratory experiments, we uses the av-

eraged solution from grinding waste leaching, which contained, g/l: Ni – 30.2; Co – 4.2; Cr – 3.1;

Mo – 0.6; W – 0.01; Re – 2.0 and 4.5 mole/l H2SO4. Extraction was carried out at 3 cascade steps at

a О:W=1:2, whereupon the extract was transferred to scrubbing with an weak acidified aqueous

solution (one cascade step) at О:W=5:1. Re-extraction occurred at a 3 mole/l NH4OH at О:W=3:1

(2 cascade steps). The amount of solution tested was 3.2 L. The obtained ammonia re-extract con-

tained 12 g/l rhenium, from which distilling yielded ammonium perrhenate purified from main im-

purities. The results of spectral analysis of the salt produced are demonstrated in table 5.

Тable 5

Spectral analysis data for ammonium perrenate obtained by solvent extraction from HRNA grinding wastes

Impurity Ni Co Al Cr Fe Mo W Тa Nb

Content, w/o 0.001 <0.003 ≥0.01 <0.003 ≥0.01 <0.003 <0.01 <0.03 <0.01

It can be stated that rhenium can be selectively isolated from a solution of complex salt

composition by using liquid extraction yielding purified ammonium perrhenate.

Based on the new findings and experience gained from liquid extraction of molybdenum

and H2SO4 from complex-salt solutions [8–10], we have proposed a basic flowsheet for in-depth

processing of rhenium-containing alloy wastes (fig. 4), whereby pure rhenium and molybde-

num salts, rare and non-ferrous metal concentrates, saving a part of H2SO4, returned to the stage

of waste leaching.

Fig. 4. Basic flowsheet of in-depth processing of rhenium-containing wastes

514


The work has been performed with the support of the RAS Presidium «Support for Innovations and Developments» programme.

REFERENCES

1. Metal Price History Charts. Rhenium/Режим доступа: [http://www.catalysts.basf.com

20.05.2010]

2. Kablov Е. N., Petrushin N. V., Bronfin M. B., Alekseev A. A. Peculiarities of monocrystal-

line heat-resistant nickel rhenium-doped alloys.//Metally. – 2006. – № 5. P. – 47–57.

3. Paretsky V. M., Besser A. D., Gedgagov E. I. Ways of increasing rhenium extraction from

ore and technogenic materials/Tsvetnye Metally. – 2008. – № 10. – P. 17–21.

4. Kasikov A. G., Petrova A. M. Rhenium recycling from heat-resistant and special alloy

wastes//Tekhnologiya Metallov. – 2010. – № 2. – P.2–12.

5. Lipmann A. Rhenium 2009 and beyond. 2009. Feb./A. Lipmann, M. Husakiewicz.

Режим доступа: [http://www.lipmann.co.uk 20.05.2010]

6. RF Patent 2330900, MPK S22V 61/00, S22V 3/26. A method for rhenium (VII) ex-

traction from an acid solution/A. G. Kasikov, A. M. Petrova (RF).– № 2006142845/02; Appl.

04.12.06; Publ. 10.08.08. Bull. № 22.

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8, RF patent 2159293 RF, MPK7 S22V 3/20, S01V 17/90. A method for the processing of

solutions containing sulphuric acid and non-ferrous metals/G. P. Miroevsky, K. A. Demidov et

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9. A. G. Kasikov, A. M. Petrova. Sulphuric and hydrochloric acid extraction by high-molecu-

lar aliphatic spirits of different structures//ZhPrKh – 2008. – V. 81. № 12. – P. 1966–1970.

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als in refractory metal technologies. – M.: «Nauka», 1988. – 240 p.

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On operating plasma ore-smelting shaft furnace advantages of use of technology «EPOS-

process» are proved. Achievement of high economy at processing of some ores and industrial

wastes as alternative ore-smelting furnace which work under the traditional scheme is proved.

In recent years we have developed and technically tested on several new production units [1–7] technology «EPOS-process» – deoxidizing metals from ores and industrial wastes using a new generation of electric ore-smelting shaft furnace, heaters with plasma-burner of special construction.

Based on proposals of a number of large Russian enterprises for treatment of ores and industrial

wastes, for 2009–2010, were carried out numerical and experimental work on developing the

technology of extracting metals from ore deposits the CHEK-SU, industrial waste of a number

of mines and enterprises of Kuzbass, the Urals, as well as ore deposits in Georgia and Ukraine.

Completed work has shown that «EPOS-process» perfectly suitable for processing a wide range of

ore minerals, industrial wastes, metallurgical and extractive industries enterprises. Depending on

the composition of the original product, techno – commercial characteristics «EPOS-process» to

reach tens of percent – up to 2.5 or more times, and the organization of plant costs half the price.

A general view of the electric RSHPP-1,5 I1 (power 1,5 MW capacity 1,0 ton silicomanganese or 4.5 tons of melt per hour. Furnace built by us in the of Novokuznetsk city. The first ore was melted in April 2009). Picture of the arc flame (heating area) and plasmatron are shown in Figure 1–3. Through a series of research works in recent years, the design of plasma ore-smelting shaft

furnace has undergone significant improvements, while remaining nevertheless a fundamentally

new and the base for development,

Results of scientific, experimental and technical works was presented at scientific

conferences [2, 4–6]. These technologies were discussed with leading experts Russia (including –

with the Director General of the Ural Institute of Metals, tsp.-corr. Academy of Sciences, prof.

Smirnov, LA in 2010, heads of leading schools in the area of manganese ferroalloys). Estimates

of many reputable independent experts can confidently assert that such a construction scheme

of the furnace and smelting technology silicomanganese applied for the first time, has clear

advantages and has no analogues.

Fig. 1. General view of plasma shaft furnace

for technology «EPOS-process»

«EPOS-PROCESS» – NEW TECHNOLOGY OF EFFECTIVE ORE-SMELTING

AND INDUSTRIAL WASTES PROCESSING IN PLASMA ORE-SMELTING SHAFT FURNACE

I.A. Bezrukov, S.N. Malyshev, O.B. Moiseyev, V.V. Pavlov, I.S. Parhomuk, А.P. Kuznetsov

«EPOS», NSTU, Novosibirsk, Russia

516


Fig. 2. The plasmatron and form a working arc torch provided

by the system of the plasma arc

A comparative study and calculation options furnace of different furnace construction, we

are concluded that there exist substantial preference of shaft type furnace, as prototypes for

future high-power electrical thermal units for the deoxidizing processes taking place to replace

the existing generation of furnaces, and developed a special «perpetual» plasmatron for realize

the process. Many experts still not appreciated the advantages of the «EPOS-process». Have in

mind the pre-existing scheme construction of plasmatron and furnaces, their low resources,

their shortcomings (including low efficiency), limiting their scope – they insist on unpromising

plasma-furnaces, on the development of domestic ferroalloy industry in the traditional, outmoded

ways. This in the future, predestination its backwardness and noncompetitive. Therefore, we

once again stop at the description of the technical solutions.

For the first time realized the scheme of the process and furnace design of the plasma

torch, operating under a layer of charge and in touch with her. Arc burns from the surface of the

coaxial electrode charge, without the hearth electrode. For the first time realized the scheme

of controlled recirculation of hot dusty raw gas entering the plasma torch. Graphite Electrode

Capacity-part without stopping the process. The plasma torch has an unlimited resource. The

geometry of the plasma torch is controlled during the melting process. Chemical and thermal

energy reductant fully used in the furnace in the smelting process. Emissions of gas and dust in

the gas cleaning system are small (diagram in fig. 4).

The ability to create conditions for the proper of deoxidation processes in the solid phase is

an advantage by shaft-type furnace. You may receive additional savings of energy through heat

transfer from flue gases to downloadable raw, due to saving of material by reducing the amount

of dust, by making full use of the chemical energy of gases through the mine to work properly

with feedstock.

Experimentally confirmed that plasma properly structured and controlled will be operate

exactly in the selected area of the furnace, will increase the percentage of minerals extracted

from ore to 90–95 % of the original, which makes plasma mine rehabilitation process, with its

correct understanding and managing one of the most promising in the field of ore processing

and disposal of industrial wastes.

Fig. 3. Photo appearance RSHPP-1, 5I1

for the technology of «EPOS-process»

Fig. 4. General view and arrangement of

equipment in the shop shaft furnaces

517


Fig. 5. Symbolic circuit «Automated Control System» management screen

and remote control RSHPP-1,5

Briefly recall the features of «EPOS-process»:

– Essence of technology EPOS-process:

The plasma torches are used, allowing to work under a layer of the loaded material (batch),

with changeable geometry of a plasma torch. Plasmatrons work on the hot crude gases submitted

at once from the furnace without temperature restrictions.

– The internal geometry (form) of the furnace is specially designed under EPOS-process

technology.

– As the basic deoxidizers the controllable and operated environment is applied: hydrogen

and carbon monoxide.

– EPOS-process takes place in the absence of an additional superfluous oxidizer, on an exit

from the furnace – gas СО2 and steams Н2О.

– The high furnace shaft is used. Drying, preliminary heating and deoxidizing of solid phase

(without fusion) processes are occur in the shaft. Conditions for correct courses of deoxidation

processes in a solid phase are created.

– It is applied gas cycling: gas is got from a shaft, from furnace roof and it goes back to the

furnace through plasmatron and through other pipelines. It provides full use of deoxidation

abilities and reserved thermal energy gases from furnace atmosphere.

– «The self-sufficient» briquette is used, that is briquette contains an ore material, a

carbonaceous deoxidizer and other necessary blending ratio, sufficient for a complete deoxidized

of components of ore.

As we pointed out earlier, and this is confirmed by long unsuccessful attempts by our

competitors to repeat our results in the 2009–10 year, the technology is not have insignificant

details and dimensions, in particular, the problem of plasma furnaces may be in the wrong

scheme, inadequate resources, technical complexity and relatively low efficiency of steel-torches

previous constructions. Traditionally used plasmtrons, used in the usual way, give the opposite

effect and discredit the technology of plasma ore-smelting shaft furnace.

This issue, we paid special attention, and today there are simple and reliable plasmatrons,

including – Coaxial, with controlled shape of the plasma torch with graphite electrodes, working

with efficiencies over 97 %. Plasmatron has no limited on resource work. Plasmatron does not

contaminate the liquid melt copper and other materials that are not contained in the furnace feed.

Plasmatron allow the furnace to work relentlessly throughout the campaign, to routine maintenance

and repairs to the furnace [7–12]. To improve efficiency we have organized a special way the

geometry of the arc flame zone (heating area) and descent of the furnace feed to protect noncooled

lining was carried out by it. It is believed that the use of plasma torches – furnace roof, furnace shaft,

plasmatrons to be water-cooled, water-cooled as are the basic elements of powerful arc furnace. This

misconception comes from the misunderstanding of physical and chemical processes in the shaft

and in the working chamber technology deoxidizing and melting going in plasma shaft furnace.

Water-cooled in all parts of the furnace, without regeneration, making the technology a competitive

disadvantage not only with the shaft-scheme, but even with a normal arc furnace.

Quality of the work of the unit also provides the correct briquette, as we wrote in detail

previously [2, 3, 6]. The use of «EPOS-process» reduces the dozens of times carryover material

from the furnace, reduces Dust and Gas Cleaning system requirements, the annual dust emissions

may be about 9–10 tonnes for the program output to 45000 tons silicomanganese per year

(at the 5 furnaces specified capacity). The correct mode allows reduced power consumption by

518


the furnace more than half, reducing the specific power consumption and overall energy costs-up

to 2–2.5 times. All these solutions have been fully incorporated in the draft electric RSHPP-1,5I1.

A major new ambitious project ferroalloy production from ore is the project of processing

ores CHECK-SU: Kuzbass, Krasnoyarsk, Usinskoe field. Reserves ores in Usinskoe (more than

98.5 million tons, total Russian reserves – about 148.2 million tons) and its complexity (the

presence of more than 92 million tons of carbonate and 5.7 million tons – of oxidized ores),

with collective processing of concentrate per year 727.14 thousand tons, with the amount of

manganese 205.41 thousand tons (mass fraction of manganese in the collective concentrate

28.25 %) makes it relevant to new and modern approaches. If use of traditional open arc

furnaces in this project will require about 360 MVA installed capacity furnaces. Thus, the

question of cost optimization comes to the fore.

A special feature of manganese Usinskoye raw materials is the high content of phosphorus,

constituting 0.23 per cent in the oxidized and 0.15 per cent in the carbonate ore. Today there is

a project ferroalloy production in the traditional manner, which provides the product with

Table 1

Comparative evaluation of traditional technology ore-smelting furnace and technologies «EPOS – process» for ore processing «CHEK-SU» ferromanganese

and silicomanganese production

Parameter CHEK-SU «EPOS-process» Effect Effect, times

1 Number of furnaces 4×27.6 МВА 7

4×63 МВА

2 Total power, MW 199 112.7 -86.3 1.77

3 Installed capacity MVA 362.4 140.875 221.525 2.57

4 Electricity consumption, MWh 1623987 919322.6 -704664

Energy consumption, MWh

5 ton alloy 7.04 4.14 -2.9 1.7

6 per ton of manganese 9.57 5.15 -4.42 1.86

7 per ton of silicon 63.54 25.02 -38.52 2.54

Consumption charge, tons

8 Concentrates 727 727

9 Coke 135.32 81.86 -53.46 1.65

10 Quartzite 100.91 0 -100.91

11 Scale 3 9.5 6.5 0.32

12 Electrode mass 14.26

13 Furnace gas (dry), million Nm3 304.45 205.62 -98.83 1.48

14 Slag, thousand tons 301.84 159.28 -142.56 1.9

Output thousand tons

15 Silicomanganese 144.26 142.36

Mn 102.89 110.68 7.79

P 0.48 0.064 -0.416

Р, % 0.33 0.045 7.33

C, % 1.62 1.7

16 Ferromanganese/silicomanganese 86.35 79.63

Mn 67.35 67.74 0.39

P 0.36 0.026 -0.334

Р, % 0.42 0.032 13.13

C, % 6.47 1.68

17 Associated alloy thousand tons 23.82 61.65

Mn 5.13 12.4

Р 0.26 1.06

Ratio. retrieval, max., %

18 Mn 85.09 92.87 1.09

19 Si 28.31 81.28 2.87

Р 18.09 91.23

519


phosphorus 0.33 and 0.42 %. According to the experience of ferro-alloys, based on the mature

market needs, we believe that the product with phosphorus 0.33 and 0.42 % will be in demand.

The project may be already be very costly only on energy (not to mention the other features) and

may be unprofitable in the future.

We have done preliminary development of a modified process «EPOS-process», under the

above mentioned materials, which provides, in contrast to the framework of the project phosphorus

concentration in the ferroalloy 0.33 and 0.42 %, phosphorus content 0.06–0.02 % or less.

Estimates show that the process of technology «EPOS-process» is realized at nearly three

times less than the installed capacity of equipment, and wasting power – almost twice less than

the traditional scheme of processing (table 1). This could fundamentally change the technology

the better.

The experimental data are processed briquettes for the project «CHEK-SU showed positive results

of the implementation process« EPOS-process »in the project, compared to conventional ovens.

By authors also performed calculations of the technology «EPOS-process» for the processing

of alloys without manganese – ferrosilicon, ferrochromium, ferrovanadium, ferrotitanium and

other ferroalloys. Each of these processes have significant individual characteristics that must

be taken into account in the concept of plasma shaft furnace. Common to all processes is the

possibility of significant savings in energy costs in the production of ferro alloys, increase in

utilization deoxidizer up to two times, a drastic reduction of the material of the project,

infrastructure costs, as well as multiple decrease dust and gas emissions.

Currently implemented practical work on developing the technology for industrial

ferrosilicon, preparation works on ferrotitanium.

Scheme and design of shaft plasma furnace to implement «EPOS-process» in 2009. received

a positive conclusion of examination of industrial safety.

REFERENCES

1. Безруков И. А., Помещиков А. Г. Новые разработки НПП «ЭПОС». «Электрометал-

лургия», 2008. № 7. С. 46.

2. Помещиков А. Г., Павлов В. В., Моисеев О. Б., Малышев С. Н., Безруков И. А. Получе-

ние железа и ферросплавов способом водородно-углеродного восстановления в шахтной

плазменной печи. Труды конференции с международным участием «Перспективы развития

технологий переработки вторичных ресурсов в Кузбассе. Экологические, экономические

и социальные аспекты». Кемеровский государственный университет. 6–9 октября 2009 г.

3. В. В. Павлов, А. Г. Помещиков, И. А. Безруков, С. Н. Малышев. Плазменная шахтная

руднотермическая печь нового поколения. М. – Электрометаллургия//2010, № 1. – С. 13–17.

4. Безруков И. А., Малышев С. Н., Кузнецов А. П., Пархомук И. С. Исследование харак-

теристик электродугового плазмотрона коаксиального типа. НАУКА. ТЕХНОЛОГИИ. ИН-

НОВАЦИИ//Материалы всероссийской научной конференции молодых ученых в 7-и ча-

стях. Новосибирск: Изд-во НГТУ, 2008. Часть 3–212 с.

5. И. А. Безруков, С. Н. Малышев, А. П. Кузнецов, И. С. Пархомук. Экспериментальные

исследования характеристик электродугового плазмотрона коаксиального типа. Труды

международной научно-технической конференции «Проблемы электротехники, электроэ-

нергетики и электротехнологии», Тольятти, 12–15 мая 2009 г. – Тольятти: Изд-во ТГУ, 2009.

6. И. А. Безруков, С. Н. Малышев, А. П. Кузнецов, И. С. Пархомук, М. Н. Соколовский,

Е. П. Демиденко. Плазменные шахтные печи для получения стали и ферросплавов из руды. Тру-

ды международной научно-технической конференции «Проблемы электротехники, электроэ-

нергетики и электротехнологии», Тольятти, 12–15 мая 2009 г. – Тольятти: Изд-во ТГУ, 2009.

7. Безруков И. А., Алиферов А. И. Многодуговая плавильная электропечь. М. – Элек-

трометаллургия//2004, № 4. – С. 10–14.

8. Устройство для ведения плавки с жидким стартом. Патент РК № 13361.

от 05.11.2001.Безруков И. А. и др.

9. Устройство для ведения плавки с жидким стартом Патент РК № 13844. от 05.11.2001.

Безруков И. А. и др.

10. Способ ведения плавки с жидким стартом и устройство для его осуществления

Патент РК № 13726. о т 05.11.2001. Безруков И. А. и др.

11. Способ ведения плавки с жидким стартом и устройство для его осуществления

Патент РК № 14141. от 01.10.2001. Кислов А. П. Безруков И. А., Алиферов А. И.

12. И. А. Безруков. Способ ведение плавки и устройство для его осуществления. Па-

тент № 2361375. от 26.11.2007 г.

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An attempt to associate experimental experience of treatment radioactive metallic wastes by melting, which was accumulated on two largest factories, was made in this article. Availably machinery of substance responsible for radioactive pollution transition to slag was analyzed.

For more than seventy years of nuclear industry history, collected the fair quantity of ra-

dioactive wastes (RAW), and the problem of their utilization became very important. The com-

plication of solving this problem is that the radioactive wastes are different in composition and

properties. Perhaps, the biggest problem of RAW treatment is extreme complexity, and because

of this the high prices of offering technological decision. That’s why the land disposal is one of

the most cheap problem decisions.

Radioactive metallic wastes (RAMW) are one of the radioactive wastes. Like the typical

RAW the metallic ones are generated on all stages of nuclear-fuel cycle. They are worked out

components of equipment, underground leaching pipes, constructive components of fuel as-

sembly. May be, total quantity of such wastes type in Russia is more than hundred thousand

tones. [3–5] This problem is still not solved abroad, in particular on Kazaatomprom manufac-

tures. Such wastes types are basically washed off up to indexes which are more than allowed

radiation level, and this metal is not allowed even for limited reuse. Such RAMW are stored at

closed industrial platforms. Processes of the open or deep waste burial though are relatively

cheap methods do not lead to the problem solution, considering overall dimensions of RAW, and

as consequence to the necessity of using considerable areas for burial grounds.

The cases of RAMW plundering are known, when the polluted scrap metal simply was

taken to points of reception of metal [1, 2]. Solving this problem at the expense of protection

of ranges of a burial place is expensive because of their considerable extent [5]. There are also

other methods of processing RAMW.

On the majority of the enterprises of Rosatom, the surfaces wash method is used. The pro-

cess essence is that the polluted detail is processed by various active agents, such as acids, alka-

lis, surfactant [19]. Additionally activation of a surface at the expense of mechanical influence

by abrasive adaptation is used. The other variant is washing by use pulsating apparatus where

process interaction of the polluted surface with a solution is intensified for the account pulsating

contact of a washed surface and a washing solution [8].

The other method of improving the washing process is overlapping of ultrasonic fluctua-

tions.

Ativation of a surface at the expense of electrochemical influence is also interesting. [6, 7].

For this moment considerable experience in carrying out of this process is stored at the

Rosatom enterprises, however there is a number of unsolved problems:

– The process is very complex in the presence of RAMW having big overall dimensions,

or various forms of scrap metal.

– The process is unproductive, and shows low enough factors of clearing.

– In case of achievement of low indicators on the activity, allowing to use the received

metal in a national economy without restrictions [21, 22], additional remelt stage is

necessary.

– Significant amounts of a liquid radioactive waste (LRW) are formed.

The offered way of processing RAMW by melting, not only free from above described meth-

od’s disadvantages, but also allows to increase considerably technical and economic indicators

of process of the recycling by reusing over 90 % of cleared scrap metal.

TREATMENT OF RADIOACTIVE METALLIC WASTES BY MELTING

I.E. Abroskin1, U.N. Makaseev2, A.C. Buynovsky2, A.I. Abroskin2, A.A. Chernoshchuk2

1 SibUniversal Ltd., Novosibirsk, Russia2 Seversk Technical Institute of National Research Federal Nuclear University, Russia

521


The description of technological process

The main objective of offered clearing was removal of the substances responsible for radio-

active pollution, from surface and volume of metal components.

The various forms of metal breakage are fragmenting preliminary, and then loaded into

the induction furnace, where they are remelting.

The induction furnace is an electrothermal installation for fusion of materials with use

of induction heating. In the industry are applied basically induction crucible furnaces and in-

duction channel furnaces.

The crucible induction furnace (CIF) consists of inductor, which is the solenoid executed

from a copper water cooled tube, and crucible which produced from ceramic materials, depend-

ing on properties of melt, or from graphite and steel in special cases.

Lacks of such furnaces are:

– Rather low temperature of the slags, being at a mirror of melt for the purpose of its tech-

nological processing, slag in CIF is warmed up from metal, therefore its temperature is

always low;

– Rather low fettling firmness at high melt temperature.

– Heat change presence (sharp fluctuations of furring temperature at full metal plum)

[12].

Processed metal was formed during works on reception of ceramic uranium fuel (UO2 tab-

lets), and as cycle of manufacturing of fuel assembly. The share of uranium activity is about

60 %, thus it is main radionuclide responsible for radioactive pollution, the others of 40 % are

brought by products of radioactive decay of uranium. Activity of a breakage is: on α activity to

100 particles/sm 2, on β activity to 20000 particles/sm 2.

Uranium is presumably presented on a surface in a kind of oxide substance. From the

diagramme of uranium condition, one can see that uranium does not form intermetallic sub-

stance in so small concentration that’s why all present metallic uranium in the course of heating

will be easily oxidized by free oxygen to octaoxide threeuranium U3O8. Uranium on a surface

can be also present in the form of sulphatic, nitrate, and other substance if the processed mate-

rial was formed as a result of work of sorption or extraction processes or during reception of

uranium oxide through sedimentation of ammonium deuranium, so-called the SOAD process.

However all substances are oxidized and decayed during heating, giving oxides [13, 14].

The essence of technological process consists in redistribution of the substances respon-

sible for radioactive pollution, and their transition from volume to a surface of the melt metal.

Thus, radioactive components emerge in the top zone.

The given process can be caused by two reasons:

As it was specified earlier, uranium on a surface is in a kind of oxide substance. The minimum

oxide density is characteristic for β-UO3 and is 7.15 g/sm3, the given phase is steady to tempera-

ture 600 oC, further passes into form U3O8. Density of other U-O systems are presented in table 1.

Table 1

Density of U-O systems

Systems Density г/см 3

UO 13.63

UO2 10.96

U O9 11.16

U3O9 8.39

αUO3 8.34

Density of uranium particles is more than that of stainless steel, consequently, the par-

ticles should go down in the fused metal however the carried out researches have shown that

substance responsible for radioactive pollution, move upwards in volume of the fused metal.

The reason of such behaviour can be the fact that proceeding from the diagramme of conditions

and experimental data, all uranium oxides at temperature above 1200 oC are passed into accord-

ing to the reaction

UxOy → XUO2 + (y – 2⋅x)/2O2 (1)

522


Free oxygen promotes carrying the uranium deoxide particles from the melt volume onto

the surface.

The other reason is that the above described effect of displacement, can be the mechanical

mixing. As it is known, in the course of CIF work, metal mixes up intensively that can lead to

carrying out of particles onto the melt surface.

In both cases, keeping of uranium deoxide particles on a melt surface, presumably, occurs

thanks to forces of a superficial tension. However in case of keeping the fused bath for a long

time, gradual subsidence of uranium deoxide particles is possible.

Partial interaction of uranium substances with free O2, C and N2, being in melt, can be the

other reason of carrying particles in the top area. Such process is characteristic at metallother-

mic uranium restoration. As a result of interaction substances of variable structure – uranium

oxicarbonitrides are formed. The given substances are characterized by low density, so they are

capable to emerge in metal volume. [10, 13].

Such behavior of uranium substances in melt has allowed to organized the process of re-

melting in the induction furnace with prompting «pseudo slag» [20]. However deviations of the

received ingots from demanded norms have repeatedly been found out, during the treatment.

Partial subsidence of uranium oxide particles and their fastening in melt vollume could be a

reason of such deviation.

For an exception of such deviation of the processed metal from demanded norms, the deci-

sion to introduce a complex of the substances promoting fastening of materials, responsible for

radioactive pollution, on the melt surface was accepted.

The process with gumboils application cared out in the arc steel-smelting furnace (ASF),

because of the complexity of the process organization with slag application in ASF.

ASF is the electric furnace in which the thermal effect of an electric arch for fusion of metals

and other materials is used. Steel fusion is conducted in the working space limited by the dome-

shaped arch from above, spherical pallet from below and walls from the sides. Pod and wall fire-

resistant masonry concluded into a metal casing from outside. Through symmetrically located

three apertures in the arch current-carrying electrodes are introduced into working space.

The scrap metal was loaded above with the charging tub help. The one third part of scrap

metal party was loaded manually, to save the fettling. Release of ready steel and slag was carried

out through steel-tapping hole and trench by an inclination of working space.

Slag serves like an environment where metal containing inclusions removed leave as a

result of chemical reaction or dissolution and per se it is a system where necessary regulation of

the impurity maintenance is carried out. Also slag carries out a number of auxiliary functions,

for example, protects the fused metal from direct atmospheric oxidation, prevents formation of

a pipe defect and internal shrinkage. All it means that, varying slag structure, it is possible to in-

fluence on the ingot chemical compound and structure [12].

Among all variants of the slag systems, the mix of calcium fluoride (CaF2) and calcium oxide

(CaO) in the 3:2 ratio has appeared to be the most suitable. The entered gumboil weight is about

5 % from the weight of loaded metal. It is necessary to notice that the gumboil was introduced dur-

ing the process into already fused metal. After slag building up and 30–60 minutes equalizing, slag

was downloaded, carrying away with itself the substances responsible for radioactive pollution.

Now the basic question there is a mechanism of retention of radioactive substances in slag.

On one of the theories, slag mechanically keeps uranium oxides particles or oxicarboni-

trides and promotes better separation of the substances responsible for radioactive pollution.

On another, the uranium is kept by slag because of the intermediate substances formation. How-

ever it is an absolutely authentic fact that the radioactive substances pass into a slag structure

and are well separate from metal.

After slag separation the fused metal was poured into forms and cooled. The weight of the

metal downloaded with slag, makes about 8–10 % of weights.

Micro photos of the slag surface, made on an electronic microscope, are presented in fig-

ures 1 and 2.

Such process realisation at Ulbinsky metallurgical combine has allowed to treat about

800 tons of the polluted metal scrap. The important thing is that the all received ingots had fol-

lowing indicators: on volume: alpha activity less than 1 particle/sm2, beta activity is less than

20 particles/sm2. Volume activity did not exceed 300 Bk/kg. Appearance of ready ingots is pre-

sented in figure 3.

523


After the slag research, following indicators of activity have been received: surface activity

from1500 to 2000 particles/sm 2⋅min, volume one from 35 to 60 kBk/kg.

There were no deviations from indicators, during the installation operating time. All re-

ceived ingots corresponded to demanded norms so, the received metal is suitable for applica-

tion in a national economy without restrictions [21, 22].

REFERENCES

1. http://forum-msk.org/material/lenty/332827.html.

2. http://www.school12-murmansk51.narod.ru/radia.htm.

3. http://vybory.org/articles/1234.html.

4. http://eco.rian.ru/business/20100203/207554308.html.

5. http://www.nuclearpolicy.ru/publications/rezonans/rez7.shtml.

6. Бойко В. И., Колпаков Г. Н., Колпакова Н. А., Комаров Е. А., Кузов В.А,. Хвостов В. И.

Способ электролитической дезакцивации металлических отходов Патент на изобре-

тение № 2328050. Зарегистр. 2 июня 2008 г № заявки 2006100787). Опубл. 27.06.2008.

Бюл.№ 18.

7. Способ электролитической дезактивации сталей в растворах хлоридов щелочных

металлов. Патент Японии № 57–76500 МПК G 21 F 9/28. Заявл. 10.10.80 № 151411.

Fig. 1. Micro photo of slag surface, lower part

Fig. 3. Appearance of the received ingots

Fig. 2. Micro photo of slag surface, head part

524


8. Способ удаления радиоактивных отложений. Патент Великобритании

GB № 1142776.

9. Стерлин Я. М. Металлургия урана. М.: Атомиздат, 1962.

10. Металлургия ядерного горючего. В. С. Емельянов, А. И. Евстюхин. Учебник для

вузов. Изд. 2. М., Атомиздат, 1968, 484 стр.

11. Электрошлаковый переплав. Латаш Ю. В., Медовар Б. И. Изд-во «Металлургия»,

1970 с. 240.

12. Электрошлаковый переплав. Докуорт У. Э., Хойл Дж. Пер. с англ., М., «Металлур-

гия», 1973, 192 с.

13. Галкин Н. П. и др. Технология урана. М. Атомиздат 1964 г. 398 с.

14. Б. В. Громов. Введение в химическую технологию урана, «Атомиздат», М. 1978.

15. Металлургия цветных меллов. Уткин Н. И. Учебник для техникумов. М., «Метал-

лургия», 1985, 440 с.

16. Нержавеющая сталь. Бородулин Г. М., Мошкевич Е. И. М., «Металлургия», 1973,

320 с.

17. Фиргер И. В. Термическая обработка сплавов: Справочник.- Л.: Машиностроение.

Ленинградское отд-ние, 1982.- 304., ил. (серия справочников для рабочих).

18. Извлечение металлов и неорганических соединений из отходов. Справ.: изд. Сит-

тиг М.,/Пер. с англ. Под ред. Эмануэля Н. М. М., «Металлургия», 1985, 408 с.

19. Скачек М. А. Обращение с отработавшим ядерным топливом и радиоактивными

отходами АЭС: учебное пособие для вузов – М.: Издательский дом МЭИ, 2007–448 с.: ил.

20. Способ переработки металлических отходов содержащих радионуклиды. Патент

РФ №: 2268515

21. СП 2.6.1.799–99 – Основные санитарные правила обеспечения радиационной

безопасности.

22. НРБ – 99. Нормы радиационной безопасости.

525


Production of aluminum by electrolysis of cryolite-alumina melts involves a complex of in-

terrelated, series-parallel processes, each of them is specified by a certain technological level of

development. Permanent progressive advance of improving methods and means of process con-

duct notwithstanding, numerous engineering problems have yet to be solved. These problems are

due to lack of theoretically and economically substantiated engineering solutions and a powerful

leverage to improve production efficiency.

Evolution of aluminum industry at the present stage involves development and implementa-

tion of resource-saving and environmentally friendly technologies to process secondary resources

and technogenic wastes. Engineering solutions implemented to reduce material and labor ex-

penditures in operating production, unmarketable wastes and middling product efficiently in-

volved into processing shall increase competitiveness, economic attractiveness and environmen-

tal safety of aluminum business.

Formation and accumulation of fine fluorocarbon-containing wastes is a serious problem

for aluminum smelters. Annually 4 Siberian smelters (IrkAZ, BrAZ, KrAZ, NkAZ) store in set-

tling ponds more than 70 thousand tons of wastes, mirabilite formation not included. The set-

tling ponds of the smelters are on the verge of being filled. Remaining period of their service

life is at OJSC «RUSAL-Bratsk» 6–8 years, at OJSC «IrkAZ-SUAL» 1–2 years. To build new settling

ponds is conjectural for two reasons: first, it is the problems with land allocation because there

are no free areas in the immediate vicinity of the smelters; second, it is considerable financial

expenditures amounting to hundreds of million rubles.

These reasons make the issues of managing aluminum production wastes processing top

priority.

Changes in electrolytic processes of aluminum production made most developments relat-

ed to recycling of valuable components (Na, Al, F) to electrolysis process lose their applicability.

Bath acidification in aluminum cells changed the balance of structural consumption of fluorine

and sodium compounds. Apparent now is the disproportion between increased yield of alkaline

regeneration cryolite (due to increasing HF concentration in electrolysis gases from acid bath)

and its limited use in electrolysis of aluminum.

Today economically viable and environmentally sound is to process in large scale fluoro-

carbon-containing wastes by outside consumers. This can be use of wastes in ferrous metallurgy,

to make building materials, in cement industry or to process into alumina as part of sintering ag-

gregate. The part of wastes and middling products to be processed to be recycled into aluminum

process is small.

This conclusion is based on two circumstances:

а) Mammoth stock of accumulated wastes, and their annual increase by tens of thousands

tons make their processing into aluminum fluoride or acid fluorides, i. e. into products in de-

mand by aluminum smelters technically difficult and economically inadvisable. Technical dif-

ficulties are related to the necessity to process wastes by multiconversion processes (employing

hydrofluoric or sulfuric acids).

б) Utilization of wastes in related sectors can make the balance between the generation

stock and waste processing negative and radically improve the environmental situation. Waste

utilization scale in ferrous metallurgy, cement industry, in alumina production can amount to

tens of thousands annually because annual production of cast iron, steel, cement, alumina is

measured in millions of tons. Expenses to prepare wastes for related sectors are minimal and all

too often are limited just to drying.

Limited Liability Company Trading House «Baikalsky Aluminii» initiated a project to pro-

cess fine fluorocarbon wastes of aluminum production in cement industry. The project is based

PROCESSING OF FLUORINE-CONTAINING WASTES AND MIDDLING PRODUCTS OF ALUMINUM PRODUCTION

IN CEMENT INDUSTRY

B.P. Kulikov 1, V.D. Nikolaev 1, S.A. Ditrich 2, L.M. Larionov 1

1 Limited Liability Company Trading House «Baikalsky Aluminii», Irkutsk, Russia2 OJSC «RUSAL Bratsk», Bratsk, Russia

526


on a process employing fluorine-containing wastes of aluminum production as mineralizer for

production of Portland cement clinker [1, 2].

The essence of the process is to replace conventionally used calcium-fluoride-based mineral-

izers for wastes and middling products of aluminum production. This solves two problems at the

same time: considerable volume of fluorocarbon-containing wastes is utilized, and cement indus-

try receives adequate replacement of natural mineralizer based on fluorite ore or concentrate. In

addition the aluminum-smelter-wastes-based mineralizer can be efficiently used in cement plants,

which earlier did not use mineralizers. In this case its use will reduce the primary cost of products

due to reduced fuel consumption and/or increased output of clinker sintering furnaces.

Currently most cement plants in Russia do not use mineralizers. The reasons is remote-

ness of calcium fluoride sources from potential consumers, because major deposits of fluorite

ore are in Siberian and Far Easter regions and in Mongolia. In the European part fluorite is not

produced in industrial scale, because of inconsiderable reserve and low content of base material

CaF2 in the ore. High cost of fluorite raw plus cost of transportation to the nearest manufactur-

ers in Transbaikal krai to the central and western part of Russia make use of calcium fluoride in

European cement plant unprofitable.

Owing to low cost of electrode scrap and reduced transport expenses technogenic alumi-

num-smelter-wastes-based mineralizers minimize expenses of cement producers. The trans-

portation costs are low because the electrode scrap is supplied to Siberian cement plants from

Bratsk Aluminum Smelter, for the European cement plants from Volgograd Aluminum Smelter.

Certification of fluorocarbon-containing wastes

Certification of fluorocarbon-containing wastes of aluminum production transferred

them into secondary raw category – «Electrode scrap for cement industry». Specifications

ТУ 1789–001–53364274–2009 were developed for the electrode scrap and included into the

State Register of Standards: registration № 003667, code ОКП 191483. Protocol of Sanitar-

yand Healthcare Inspection of Rospotrebnadzor was issued for Specifications ТУ 1789–001–

53364274–2009. Rospotrebnadzor also issued a positive conclusion of toxicological examina-

tion of electrode scrap.

Electrode scrap for cement industry shall meet norms and regulations presented in table 1.

Table 1

Requirement specifications for electrode scrap

Index Norm

Weight fraction of fluorine, %, not less than 15.0

Weight fraction of carbon, %, not more than 50.0

Weight ratio of sodium to fluorine, %, not more than 0.8

Weight fraction of absorbed moisture, %, not more than 20.0

Subject to these requirement specifications are not only wastes, but also some fluorine-

containing middling products of aluminum production, viz.: carbon dust, crushed bath mate-

rial, recovered cryolite, etc.

Mineralizing properties of electrode scrap

Studies were carried out by thermogravimetry, high-temperature radiography, X-ray phase

and chemical analyses. Thermogravimetric analysis was carried out with STA 449 Jupiter syn-

chronous thermal analysis instrument. Samples were heated to the temperature from 25 oC to

1480 oC at the rate of 8 degrees С/min with deficient oxidizer (5 % oxygen) to model the at-

mosphere in commercial clinker sintering furnace. Sample 1 is the mixture of raw slurry with

water at the ratio of 1:0.2; sample 2 is the mixture of raw slurry with fluorocarbon-containing

wastes of aluminum production (electrode scrap) and water at the ratio of 1:0.0054:0.4. With

this addition of mineralizer fluorine content in the raw slurry (sample 2) was 0.1 % weight. The

weight of first sample was 118.4 mg, weight of the second sample was 170.2 mg. Studies were

performed in alumina crucibles. Qualitative and quantitative composition of gas products of

thermolysis during the experiments were monitored with quadrupole mass-spectrometer Aelos

with electron impact energy 70 eV.

527


Initial slurry components used as to produce cement clinker were raw materials of OJSC «Vol-

skcement» with the following composition: limestone – 79.5 %, clay – 15.7 %, butts – 3.5 %, high-

aluminate clay – 1.3 %. Chemical composition and module characteristics of raw material are

given in table 2. Phase composition of the raw slurry is represented by calcium carbonate CaCO3,

quartz SiO2, hematite Fe2O3, heulandite Na2Ca(Al2Si6)16⋅5H2O and mica KMg3 (AlSi3O10)⋅(OH)2.

Table 2

Characteristics of raw slurry of OJSC«Volskcement»

SiO2 Al2O3 Fe2O3 CaO MgO SO3 Na2O K2O Cl –

TiO2 P2O5 Mn2O3 Кн n p W

13.77 3.23 3.32 42.84 1.15 0.51 0.29 0.60 0.009 0.15 0.21 0.020 0.94 2.10 0.97 42.4

Notes: 1) Sc – saturation coefficient is an indicator specifying incomplete saturation of silica

with calcium oxide in clinker formation process:

CaO – (1.65Al2O3 + 0.35Fe2O3)

Sc = ———————————————; 2.8SiO2

; 2) n is the silica modulus: % SiO2/( % Al2O3 + % Fe2O3);

3) p is the alumina (aluminate) module: % Al2O3/ % Fe2O3;

4) W is the water content in the slurry, % weight.

Thermogams of sintering of individual raw slurry and slurry with electrode scrap are giv-

en in fig. 1, 1 а and 2, 2 а.

To study effect of sodium fluoroaluminate (cryolite and chiolite) on sintering of raw slurry

additional studies were carried out. With this aim the raw slurry was added in 10-fold quantity of

fluorine-containing mineralizer, which maintained fluorine content in the slurry ~1.0 %. Mineral-

izer dose was increased because when fluorine content in the samples is about ~0.1 % low inten-

sity of analytical lines denies identification of fluorine-containing compounds. The slurry with in-

Fig. 1. Thermogams of sintering of individual

raw slurry

Fig. 1а

Fig. 2. Thermogams of sintering of raw slurry

with electrode scrap

Fig. 2а

CaCO3 decarbonization with carbon dioxide liberation (mass number 44)

528


creased dose of mineralizer was sintered at 1000 °C, 1100 °C, after that the cooled sinters were

analyzed with automatic X-ray diffractometer D 8 ADVANCE with copper anode (fig. 3, 4).

Thermal radiography was carried out with Shimatzu XRD 700, Bragg-Brentano focusing with

monochromator with diffracted beam in Cu-emission with sample heating rate 10 degrees С/min in

air within range 500÷1200oC. The composition of examined samples was analogous to samples 1,

2 of thermogravimetric analysis (without addition of water). The X-ray spectrum was recorded

upon achievement of preset temperature with subsequent heating of the same sample to higher

temperature under the following conditions: voltage 40 kV, current 50.0 mA, scan space 10–60 de-

grees, scanning pitch 0.1 degrees, scanning speed 1.5 degrees/min. Diffraction patterns were re-

corded at constant temperature. Diffraction patterns were recorded at temperatures 25, 500, 600,

700, 800, 900, 1000, 1100, and 1200 °C. Variation of intensity of most characteristic analytical

lines of components of raw slurry and sintering products is given in fig. 5, 6 and in table 3.

Table 3

Temperature dependence of analytical lines’ intensity, pulse/s

t,oC

Individual raw slurry

СаСО3 СаО α-SiO2 Fe2O3 C3S β-С2S С3А С4АF

3.03 2.405 3.34 2.53 1.764 2.40 2.70 2.64

25 14616 1525 292

500 11260 642 228

600 9704 670 232

700 6262 1462 636 202 338

800 5668 582 174 106 104 406

900 4660 466 138 160 170 136

1000 3540 230 124 198 194 270 186

1100 3146 144 88 186 208 242 266

1200 2316 92 98 236 218 410 292

Raw material slurry with added electrode scrap (0.1 % по F)

25 14834 906 268

500 12254 866 242

600 9944 738 244

700 9316 776 246 396 190

800 5158 618 214 426 254

900 4790 434 180 188 188 266 224

1000 4046 204 136 242 222 292 274

1100 3060 94 114 256 240 332 294

1200 2096 – 94 290 278 464 336

Fig. 3. X-ray radiograph of products of raw

slurry sintered with electrode scrap,

1 % F, t=1000 oC

Fig. 4. X-ray radiograph of products of raw

slurry sintered with electrode scrap,

1 % F, t=1100 oC

529


Discussion of experimental results

ThermogravimetryIn sintering thermogram of individual raw slurry (fig. 1) endothermic effect in range

35÷165 oC, accompanied by loss of sample weight by 20.8 %, is due to water evaporation.

In range 600÷900 oC endothermic effect with sample weight loss by 27.0 % (32.0 mg) is due to

decarbonization CaCO3 with liberation of CO2, which is proved by the signal from the mass-

spectrometer (mass number 44) (fig. 1 а).

In range 900÷1450 oC one endothermic effect observed at 1310 oC is caused by melting of

the reaction mixture. Slight endothermic deviation is related to overlapping endothermic and

exothermic effects, the latter dominate and are caused by synthesis of base clinker minerals.

Sintering of raw slurry with fluorine-containing mineralizer in range 260÷890 oC is

attended by loss of sample weight by 20.5 % (34.9 mg) and several thermal effects (fig. 2).

The first thermal effect in temperature range 600 oC is exothermic and is caused by carbon burn-

ing from the fluorocarbon-containing mineralizer. The following endothermic effect is related

to decomposition of calcium carbonate. The endoeffect of thermal dissociation of calcium car-

bonate is partly overlapped by exoeffects caused by burning of carbon and resins, and formation

of tricalcium aluminate С3А and tetracalcium alumoferrite С4АF. Both effects are accompanied

by liberation of СО2 into the gas phase which is proved by the signal of the mass-spectrometer

(mass number 44) (fig. 2 а). It should be noted that oxidation of carbon which is part of the

mineralizer releases 0.95 mg of СО2, while the total amount of released СО2 is 32.85 mg.

In range 890÷1450 oC the 2.1 % (3.6 mg) loss of weight by the sample is accompanied by

several thermal effects (fig. 2). Basic endothermic processes in this temperature range are con-

nected with formation of the liquid phase, polymorphic transformations of silicon dioxide and belit

С2S. Exothermic effects are caused by synthesis of base clinker compounds: C4AF, С3А, C2S, C3S.

Comparison of thermograms of raw slurry sintering with fluorine-containing mineralizer

(fig. 2) and without it (fig. 1) shows that addition of fluorine in the amount of 0.1 % weight inten-

sifies thermal dissociation of calcium carbonate. Decarbonization of CaCO3 in the presence of min-

eralizer starts at 500÷550 oC, while without it dissociation starts at 680 °C. Intensity of СО2 lib-

eration with mineralizer is maximum at 720 oC, without mineralizer – at 850 oC. СО2 liberation

from the slurry with mineralizer is stepwise, with different intensity at different temperatures.

So, addition of electrode scrap to the raw slurry shifts the beginning of CaCO3 decarbon-

ization two the lower temperature range with positive effect on the subsequent synthesis of

clinker compounds.

High-temperature radiogaphyFrom figures 5 and 6 and table 3, presenting temperature dependencies of intensity of

analytical lines of initial components and sintering products of individual raw slurry and slurry

with addition of electrode scrap, it is apparent that samples have no visible changes up to 600 oC,

the composition of samples is identical to the initial one. Decreasing intensity of basic line of

CaCO3 is indicative of the start of limestone decomposition process.

Fig. 5. Temperature dependence of intensity of

analytical lines of initial components of indi-

vidual raw slurry and slurry with addition of

electrode scrap (0.1 % in terms of fluorine)

Fig. 6. Temperature dependence of intensity

of analytical lines of sintering products of indi-

vidual raw slurry and slurry with addition of

electrode scrap (0.1 % in terms of fluorine)

J pulse/с

t,oC

0

200

400

600

800

1000

1200

1400

1600

25 500 600 700 800 900 1000 1100 1200

CaCO3 with scrap CaO with scrap

α-SiO2 with scrap Fe2O3 with scrap

CaCO3 raw CaO raw

α-SiO2 raw Fe2O3 raw

050

100150200250300350400450500

600 700 800 900 1000 1100 1200C3A with scrap C4AF with scrap

C3S with scrap β-C2S with scrap

C3A raw C4AF raw

C3S raw β-C2S raw

J pulse/s

t,oC

530


At 700 oC it is obvious that concentrations of elementary oxides of silicon, calcium, iron tend

to decrease, while the tricalcium aluminate С3 А and alumoferrite С4АF tend to increase. To eval-

uate content of calcium carbonate and calcium, silicon and iron content we used the strongest

lines free from overlapping, therefore dependence of their intensities on temperature is obvious.

At 700 oC both samples start solid-phase reactions between calcium oxide and dehydration prod-

ucts of clay minerals to produce calcium silicates which is proved by emergence on thermograms

of lines 2.89, 2.78 and 2.75 . The content of clinker phases is not sufficient to accurately identify

them, as the said lines belong both to alit and belit, while the weaker lines specific for these com-

pounds did not emerge yet. The thermogram of individual slurry exhibits clear lines of calcium

oxide 2.403 and 1.699 . In the sample with addition of electrode scrap no calcium oxide lines

are observed, which is probably related to interaction of СаО with aluminum oxide.

Thermograms of both samples produced at temperature 800 oC have no СаСО3 lines, and

the analytical line of free calcium oxide reaches its maximum. The sample of individual slurry at

this temperature exhibits clear lines of alit С3S, tricalcium aluminate C3 A and belit β-С2 S. Be-

sides the said phases the sample with electrode scrap addition has ferrite phase C4 AF.

Further increase of temperature reduces intensity of analytical lines of initial components

of raw slurry СaCO3, Al2O3, SiO2, Fe2O3 (fig. 5) and increases the quantity of clinker com-

pounds in sintering products (fig. 6). Intensity of lines C3А, С3S, β-С2S, С4АF in slurry sintered

with с electrode scrap is 13–25 % higher than in sintering of individual raw slurry, which proves

the intensifying properties of the mineralizer.

So, experiments proved that the solid-phase sintering processes in the samples under study

start as early as 700–800 oC, and interaction of clay materials with calcium oxide with electrode

scrap addition is higher than in mineralizer-free slurry.

Interaction of electrode scrap with raw slurryBehavior of basic compounds of fluorine-containing mineralizer during sintering of clinker

aggregate produced by thermogravimetric and X-ray phase analysis is presented by the follow-

ing basic equations (table4).

Table 4

Interaction of electrode scrap components with raw slurry components

Interaction Temperature range, oC Reaction

3Na5Al3F14 = 5Na3AlF6 + 4AlF3 725 1

Na3AlF6 = 3NaF + AlF3 850–950 2

C + O2 = CO2 + Q 450–950 3

2AlF3 + 3CaO = 3CaF2 + Al2O3 750–1000 4

2NaF + К2O + 2CaF2 = 2КCaF3 + Na2O 1000–1200 5

11CaO+7Al2O3+CaF2 = CaO ⋅ 11Al2O3 ⋅ 7CaF2 1000–1150 6

AlO3 + 3CaO = 3CaO ⋅ Al2O3 1000–1250 7

2CaF2 + SiO2 = 2CaO + SiF4 1200–1300 8

At 725 oC chiolite incongruently melts to produce cryolite and a melt containing cryolite

and aluminum fluoride [3]. Cryolite is thermally more stable as compared to chiolite, its melting

temperature is 1008 °C [4]. However, in range 850–950 oC cryolite thermally dissociates to form

sodium fluoride and aluminum fluoride (reaction 2). Dissociation of cryolite is intensified by in-

teraction of highly active calcium oxide formed by decomposition of limestone and aluminum

fluoride in reaction 4. At 1000 oC all aluminum fluoride transforms into calcium fluoride and

aluminum oxide.

Sodium fluoride formed by thermal dissociation of cryolite starts interacting with К2O and

CaF2 at temperatures above 1000 oC (reaction 5). Prior to this interaction sodium fluoride acts as

mineralizer more efficiently, at that, than CaF2 [5]. In temperature range 1000–1200 oC sodium flu-

oride is replaced by potassium fluoride which interacts with CaF2 to form КСaF3 (reaction 5, fig. 3,

4). In addition, complex compound CaO⋅11Al2O3⋅7CaF2 forms in the same temperature range.

Highly active aluminum oxide forming from aluminum fluoride is bound into tricalcium

aluminate С3А (reaction 7). Carbon which is part of aluminum production wastes burns in tem-

perature range 450–950 oC to form carbon dioxide (reaction 3).

During sintering of raw slurry a part of calcium fluoride interacts with silicon oxide to form

highly active CaO and gaseous silicon tetrafluoride (reaction 8).

531


Industrial testsTests conducted formed theoretical basis to industrially approve the developed technol-

ogy to use aluminum-production-wastes-based fluorine-containing mineralizers to produce

Portland cement clinker. Electrode scrap was tested in 6 plants: Angarsky, Timlyuysky, Altaisky,

Notroitsky, Volsky, Iskitimsky cement plants. Initially the technology was industrially elaborat-

ed in OJSC«Angarskcement».

To produce cement OJSC«Angarskcement» uses unconventional raw materials – crystal-

line limestone and fuel ash, contrary to conventional – limestone and clay. This has negative im-

pact on clinker formation processes and the quality of clinker produced. Production of Port-

land cement clinker from unconventional raw material necessitates use of mineralizer, for this

OJSC«Angarskcement» used natural calcium fluoride. Pilot tests to replace calcium fluoride by

electrode scrap helped elaborate basic process variables of raw slurry sintering to produce clin-

ker with high physical-mechanical characteristics.

In other cement plants electrode scrap was tested to replace expensive calcium fluoride

and to intensify clinker sintering in the plants which did not use mineralizers. In the latter case

the electrode scrap acted as an additional component of raw slurry.

After successful tests of electrode scrap in OJSC«Iskitimsky cement», the management of

the cement plant made a decision to expand the test scale; for this purpose Bratsk Aluminum

Smelter supplied 1000 tons of mineralizer.

After industrial tests Volsky and Novotroitsky cement plants also made decisions to enlarge

the test scale.

Industrial tests found that:

1. Optimum addition of electrode scrap into raw slurry is 0.10–0.25 % in terms of fluorine.

2. Electrode scrap improves spreadability of raw slurry by 5–15 %, depending on its rheo-

logical properties. This makes possible to reduce by several percent moisture content in the raw

slurry and reduce fuel consumption to sinter the clinker.

3. Carbon which is part of the electrode scrap acts as a burning addition intensifying pro-

cesses in the medium-temperature zone of the sintering furnace.

4. Addition of electrode scrap to raw slurry with its invariable moisture content reduces specif-

ic fuel consumption by 4–6 kg/1 t of clinker and increases output of sintering furnaces by 2–4 %.

5. Clinker produced with addition of electrode scrap exhibits food structure of basic clinker

minerals and grindability (fig. 7).

6. In some instances addition of elec-

trode scrap was found to strengthen protec-

tive coating on the refractory lining of the

tubular rotary kiln.

7. Electrode scrap does not affect building-

engineering properties and physical-mechani-

cal characteristics of the cement produced.

8. Among negative aspects of electrode

scrap application is addition of sodium to raw

slurry which is objectionable. Addition of elec-

trode scrap to the raw slurry in the amount of

0.10÷0.15 % weight in terms of fluorine in-

creases sodium content 0.06÷0.10 % weight

which is admissible for most cement plants.

Mastering industrial production of Portland cement clinker with electrode scrapPilot tests of new mineralizer demonstrated that application of electrode scrap does not

change cement production process in OJSC «Angarskcement» and does not reduce qualitative

characteristics of the clinker produced. This makes possible to recommend this addition to be

used as a mineralizer to sinter Portland cement clinker.

In 2005 OJSC «Angarskcement» continued testing electrode scrap. In one year 7.7 thousand

tons of slurry were processed to produce 457.7 thousand tons of cement clinker. The quality of

clinker was satisfactory: its activity varied from 45 MPa to 48 MPa which is consistent to the activ-

ity of clinker with addition of electrode scrap. Annual economic benefit was 2.31 million rubles.

In 2007 slurry was supplied to OJSC «Angarskcement» by Irkutsky Aluminum Smelter. In

the course of one year more than 14 thousand tons of electrode scrap were supplied to be pro-

cessed in 2007–2008.

Fig. 7. Portland cement clinker

microstructure

532


In 2009 a work package was realized to enlarge the scale of electrode scrap application

from Bratsk Aluminum Smelter to OJSC «Angarskcement»:

– «Extraction of Electrode scrap from reclaimed Settling Pond № 2 at OJSC «Bratsk-RUSAL»

project has been developed;

– long-term (10 years) contract to extract electrode scrap in the amount up to 400 thou-

sand tons from OJSC «Bratsk-RUSAL» and supply to OJSC «Angarskcement», to clean

settling pond № 2 of OJSC «Bratsk-RUSAL» to restore the newly produced aluminum

production wastes;

– amount of electrode scrap delivery to OJSC «Angarskcement» in 2009 was 11400 t.

In 2009 conditioned electrode scrap was first delivered to Altaisky cement plant. After suc-

cessful pilot tests in 2010 the plant completely turned to electrode scrap instead of earlier use

fluorite ore.

REFERENCES

1. Patent 2383506 РФ, С 04 В 7/42. Method of Protland Cement Production./Kulikov B. P.,

Nikolaev M. D., Kuznetsov A. A., Pigarev M. N. Priority of 30.09.2008. Publ. 10.03.2010. Bul. № 7.

2. Kulikov B. P., Barinov V. V., Nikolaev M. D. et al. Development and Implementation of

Clinker Sintering Technology Employing Technogenic Mineralizer on the Basis Of Secondary

Fluorine-Containing Resource of Aluminum Production.//Abstract of presentation at PetroCem

conference 2010. S. Petersburg. С. 80–81.

3. Toritsky I. A., Zheleznov V. A. Metallurgy of Aluminum. Reference Guide. Chap-

ter XIV./М.: Metallurgiya. 1977. 392 p.

4. Non-ferrous metallurgist’s reference book. Production of Aluminum./М.: Metallurgiya,

1971. 560 p.

5. Volkonsky B. V., Konovalov N. F., Makashev S.D Mineralizers in Cement Industry./М.,

Stroiizdat, 1964. 200 p.

533


Material and energy consumption in metal products manufacture within Russian

companies is 20–25 % and 80–90 % respectively higher in comparison with developed western

countries [1]. The most important tendency for reduction of energy and material consumption is

resource saving technology.

Under present-day economic conditions due to commercializing it is necessary to find

effective source of materials, process fuel and all kinds of energy, widely introduce modern

resource saving technologies. In ferrous metal industry it is especially important for steelmaking

process stage, where the best converter process technology alternative with wide processing

mixture ratio of metal stock and process fuel is to be tried out.

At Zapsibmetkombinat (JSC «ZSMK») the number of resource saving technologies were

developed, passed industrial tests and were accepted into production in partnership with the

author.

For example, at JSC «ZSMK» for the first time in native ferrous metal industry there

was developed and passed industrial test steelmaking practice by 100 % smelting burdening

using metal scrap in a converter with a capacity of 130 t (KKZ-1). The main purpose of this

technology introduction is material resources economizing and foremost economizing of high-

priced steel-making iron. In 2006 at JSC «ZSMK» conditions the cost of 1 t of metal scrap was

a bit more then 60 % from the cost of 1 t of liquid steel-making iron. With such price ratio for

the basic kinds of materials for smelting burdening there is no need to give any other reasons.

Although it is necessary to define such aspect as process fuel expenses. Surely process fuel

expenses for 1 t of melted steel by 100 % smelting burdening using metal scrap are higher then

by traditional burdening – hot metal and metal scrap. In order to minimize the expenses for

melted steel production using this processing technology to get the necessary temperature in a

bath furnace where melting takes place not only normal oxygen lancing through the form is used

but metal scrap preheating is used, coal (steaming coal, is chipper) is added, about 40–60 kg

per tone of melting charge and small coke after the coking, with worsen furnace burden grain

fineness for blast furnace smelting and has a negative impact on blast-furnace run (reduces blast

furnace technical-and-economic indexes by iron smelting).

According to this steelmaking technology to provide melting charge melt in a converter and

normal slag adjustment metal scrap is filled into the converter in portions in definite periods of

time – it provides the desired temperature.

Based on the calculation process fuel unit costs for 1 t of melted steel by converter burdening

with metal scrap in comparison with traditional converter burdening smelting using hot metal

and metal scrap is 62–68 % higher.

However production calculations for 1 t of melted steel by 100 % metal scrap charging

and traditional converter burdening show that production cost for 1 t of steel by 100 % usage

of metal scrap is 11–13 % lower. Based on this fact one comes to a conclusion that the 100 %

usage of metal scrap in converter melting charge is a working resource saving technology in

steel industry. We consider that to become widely used in economics the viewed technology

needs to be worked out in more details as far as the technological aspects is concerned and is

to get an integrated economic appraisal for all energy-material-economic aspects in different

alternatives.

Process fuel costs reduce for melted steel production is up-to-date not only by 100 %

burdening using metal scrap. Heat-transfer efficiency in converter smelting is defined mostly

by the way of heat-transfer entering into the converter, the entering time depending on the

smelting period and quality of materials used as a heat-transfer.

EFFECTIVE STRENGTH RESOURCE SAVINGTECHNOLOGY IN STEEL INDUSTRY REVISITED

N.I. Novikov

Novokuznetsk branch of the institute state educational highest vocational institution


534


At JSC «ZSMK» the technology of heat-transfer entering through a blowout pipe and

through a conveying pipe is worked out. In the first case heat-transfers were loaded along the

bath lancing. In the second case – before the beginning of bath lancing (metal scrap preheating)

and using combined process [2].

Under present-day conditions the best practice is metal scrap preheating in converter. As

heat-transfers within this technology own-produced coke and different coal ranks from Kuzbass

coal companies are used. Yearly time history for coke and coal consumption rate is shown on

figure 1.

19,2 18,917,5 17 17,5

19,621,4

0

5

10

15

20

25

2001 2002 2003 2004 2005 2006 2007

Years Years

8,26

9,74

6,267,3

7,5

6,75

8

0

2

4

6

8

10

12

2001 2002 2003 2004 2005 2006 2007

Co

al

con

sum

pti

on

ra

te, k

g/

t

Co

ke c

on

sum

pti

on

ra

te, k

g/

t

а) b)

Fig. 1. Changes for coke and coal consumption rate, years 2001–2007:

a) in an oxygen-converter plant (OCP) № 1; b) in OCP № 2

The measures pointed above led to reduction of high-priced cast iron consumption rate and

to achievement of the necessary sensible heat in converter bath (according to the steelmaking

practice). Changes in cast iron consumption rate per tone of melted steel is given on drawing 2.

853,2

810,1

801,8

810,4 808,5

806,6

807800,2

770

780

790

800

810

820

830

840

850

860

1999 2000 2001 2002 2003 2004 2005 2006

834,4

800

807,8

805,4

829,5

815,4

811,6808

780

790

800

810

820

830

840

1999 2000 2001 2002 2003 2004 2005 2006

Ca

st i

ron

co

nsu

mp

tio

n r

ate

, k

g/

t

Ca

st i

ron

co

nsu

mp

tio

n r

ate

, k

g/

t

Years Yearsа) b)

Fig. 2. Changes in cast iron consumption rate, years 1999–2006:

a) in OCP № 1; b) in OCP № 2

Operation by reduced cast iron flow using heat-transfers did not lead not only to oxidation

enrichment of metal and dross more then it is regulated in a process message (drawing 3) but to

reducing of good steel make.

Fig. 3. Changes in FeO containing in dross, years 1999–2006:

a)in OCP № 1; b) in OCP № 2

18,919,1

19,3

18,117,9

17,717,4

17,2

16

16,5

17

17,5

18

18,5

19

19,5

1999 2000 2001 2002 2003 2004 2005 2006

19,3

20,7

21,4

20,4

21,5

20,9

19,8

19,4

18

18,5

19

19,5

20

20,5

21

21,5

22

1999 2000 2001 2002 2003 2004 2005 2006

Fe

O c

on

tain

ing

in

dro

ss, %

Fe

O c

on

tain

ing

in

dro

ss, %

Years Yearsа) b)

535


The other perspective and significant resource saving technology for company economics is

the steel continuous casting. Although this technology is not new in the world iron-and-

steel industry, but nevertheless it is relevant for Russian companies and especially for West-

Siberian region. For example on one of the leading companies in iron-and-steel industry in

Russia – JSC «West Siberian Metallurgic Plant» (JSC «ZSMK») which is within «Evrazgruop»

Holding and has a 45 year history the continuous casting is used only for about 5–6 years.

Nowadays at the plant using a continuous-casting machine (CCM) about 32–38 % of melted

steel is ladled out (CCM production capacity lets to ladle out about 50–55 %) and it is the fact

to think about. The company conversion to 100 % continuous casting using the CCM is now

stopped and this problem solving within the plant is postponed for some indefinite time (long

term prospects). Meanwhile both world practice and technical-economic performance on JSC

«ZSMK» in the last 5 years show the undeniable advantage of using the CCM for steel ladle out.

The information given in table 1 proves this fact.

Table 1

Production cost per 1 t for steel products of different types, made of rolled billet and continuous cast steel billet from dead-melted steel (DMS) and low alloy steel (LAS), RUB

List of items produced Rolled billet Continuous cast steel billet Difference

Rod DMS 7459.02 7142.73 316.26

Rod LAS 8090.93 7635.61 455.32

Angle bar DMS 7417.17 7100.88 316.29

Angle bar LAS 8001.25 7587.94 413.31

Channel bar DMS 7435.33 7062.11 373.22

Channel bar LAS 8021.10 7547.60 474.61

Beam DMS 7507.36 7203.32 304.04

Beam LAS 8092.06 7692.40 399.66

Average production cost, RUB. 7753.03 7371.56 381.47

Additionally after putting the CCM into operation the plant got the opportunity to widen

greatly the output products grade composition and therefore to move from production of common

steel types (unkilled steel, semikilled steel) to production of higher quality steel types (dead-

melted steel and low alloy steel). Steel products made of high quality steel are of high demand

on market and the economic effect for the company is higher. Changes in grade composition of

steel produced at ZSMK in 2005–2009 in % are given in table 2.

Table 2

Changes in grade composition of melted steel produced on OAO « ZSMK» in 2005–2009, %

Steel type 2005 2006 2007 2008 2009

Unkilled 16.5 9.9 8.3 5.8 2.0

Semikilled 76.9 53.2 49.0 41.0 15.2

Dead-melted 1.1 30.5 34.8 40.6 64.5

Low alloy and alloy 5.5 6.4 7.9 12.6 18.3

Strong rising of high quality steel types production volume in the last years was caused by

slab caster setting in operation, which in short time has been brought to the designed capacity

(2.5 MIO t slab continuous casting per year).

Based on successful operation of billet and slab CCM one can consider the complete change-

over of the OCP-1 for continuous casting in the future without building the third CCM, but it will take

to modernization of the working CCMs, which is much (40–45 %) cheaper then to build the third

CCM. According to plant specialists’ estimates increase in active CCMs productivity for about 25–

30 % can be provided by using practically tested innovations and by increase of process organization

level. So, in a billet CCM in 2008 running downtime due to the lack of melted steel was 25 days or

8.5 % from time rating. Significant reserve for productivity increase is in casting sequence increasing,

which nowadays on ZSMK is only 8–9 melting operations and on foreign companies (Japan, France,

India, Korea) it is more then 100 melting operations per production run.

536


Continuous casting application in steel industry leads to expenses economy due to

technological conversion elimination. By an ordinary ladle out (ingots technology) the

technological conversion has the following structure – converter – ladle to mold casting –

stripper – soaking pit station – blooming-billet mill and the structure by continuous casting is

the following: converter – ladle furnace – CCM – rolling mill. Expenses (labor cost, process

fuel, electrical energy, production labor hours, upkeep expenses etc.) per each technological

conversion grow, accordingly the more technological conversions we have the more the

production cost of metal products is. In addition to this, continuous casting application provides

metal yield increase due to discards minimization and to process fuel cost saving, which is

spent for ingot heating before rolling; one gets labor hour savings that gives the opportunity

to raise production; it provides quality improvement for metal products and finally it leads

to diminution in working hours and to labor conditions improvement (labor becomes more

automated). To make a conclusion we can say that the change from ordinary steel ladle out to

continuous casting provides substantial improvements both for economic and ecological indexes

not only in steel melting production but also for the steel company in general that leads to its

cost effectiveness and turn out products competitiveness increase.

One of the effective resource saving technologies in steelmaking is the application of used

tyres as a process fuel within an oxygen-converter process.

As a fuel, rubber component has evident advantages in comparison with the best coal types;

low ash content – 2–3 %, almost zero humidity and extremely high combustion value – 33.5–

37.7 МJ/kg. Converter configuration and the present converter smelting technology let to load the

whole ties into the converter providing their quick and complete burning in bulk oxygen atmosphere

(99.5 %) by high temperature (1300–1500 oC). Availability of strong gas-cleaning units with water

recycling closed cycle almost entirely removes any environmental pollution probability.

The developed technology supposes loading of 6–10 tyres for smelting or 3.3–5.5 kg/t

steel in 160 t converter (OCP-1) together with metal scrap and burning them additionally to coal

rank TOM which has the following characteristics: grain-size – 13–15 mm; humidity – not more

then 8 %; ash content – not more then 13 %; volatile – not more then 17 %; combustion value –

25.1–28.1 MJ/kg.

Technical analyses results of tyres after metal cord removal are given in table 3. The lowest

combustion value calculated to working mass is 33.5–35.6 MJ/kg.

Tyres organic part on an average consists of working/organic mass elements %: carbon –

85.5/88.0; hydrogen – 8.00/8.20; oxygen – 2.30/2.40; nitrogen – 0.40/0.40.

Table 3

Technical analyses results for tyre casting, %

Tyres type Humidity,

%

Ash content,

%

Volatile

content, %

Containing, %

sulphur unvarying carbon

Medium-sized tires

260×508 0.46 3.15 71.2 1.32 27.76

280×508 0.55 2.63 70.6 1.40 28.47

260×508 Р 0.61 3.30 70.0 1.42 28.83

320×508 0.53 2.38 69.9 1.63 29.23

Small-sized tires

840×15 0.40 2.10 71.0 1.70 28.28

205/70×14 1.00 3.33 70.5 1.62 28.23

Limit value 0.40–1.00 2.10–3.30 69.9–71.2 1.32–1.70 27.76–29.23

During industry research there were estimated technological indexes for converter smelting

and dust-gas burst into the atmosphere, including dust level, sulfur oxide and carbon content.

The obtained results showed that by tyres application there is no negative influence on smelting

characteristics production date and on quality of the steel made (table 4). Steel sulphur content

was not stated.

The used tyres application admitted to increase heat arrival at a converter at the expense of

fuel burning heat increase and improvement of heat transmission conditions owing to increase

of flame brightness under soot burning that provided fuel consumption decrease in the period

of iron bar preheating.

537


It is established that tyres input into a converter has not influenced dust content in released into

the atmosphere gases. High temperature and oxygen excess in a converter bath promoted full

tyres burning and their thermal decomposition products that excluded atmosphere pollution

with incomplete fuel burning products (soot, carbon oxide, hydrocarbons). SО2 concentration in

gases released into the atmosphere was not higher than in a common coal using technology.

Thus tyres are valuable heat energy source for high temperature metallurgical processes.

The worked out technology of steel smelting in oxygen converters with used tires application is

low cost-based, ecologically permissible and is the recycling mass method that helps to solve

wastes utilization problems in developed metallurgical industrial regions.

The recourse-saving steel industry technologies considered in this article provide an

enterprise with essential improvements of technical economic rates that increase its financial

stability and competetiveness both in foreign and home market.

REFERENCES

1. Nikolaev A. L. and others. Iron-containing waste materials – raw materials for steel in-

dustry/ A. L. Nikolaev, A. A. Nikolaev, V. N. Jurchenko, N. I. Novikov, A. G. Zaraeva//Regional

conference transactions, Novokuznetsk, 9–11 october 2003/Under the editorship of F. I. Ivanov

and V. K. Butorin: NFI KemGU. – Novokuznetsk, 2003. P. 12–14 Works

2. JSC «West Siberian Metallurgical Plant». – Access mode: www.zsmk.ru. – 27.01.2010.

Table 4

Averaged technical economic work rates of 160 t converters with used tires application

Rates Technology

Worked-out common

Number of meltings 100 100

Charges, kg/t of fluid steel

Cast iron 773.3 775.1

Iron bar 313.6 312.0

Limestone 53.7 54.0

Used tires 2.8 –Coal 14.2 18.5

Job duration, minutes:

preheating 5.19 6.0

blowing 21.5 21.6

Oxygen total makeup, m 3/melting 9703 9695

Cast iron composition, %:

Si 0.35 0.38

Мn 0.43 0.42

S 0.016 0.016

Р 0.17 0.16

Cast iron temperature, oС 1326 1330

Metal composition at the first turndown, %:

С 0.13 0.10

Мn 0.16 0.15

S 0.021 0.020

Р 0.015 0.016

Metal temperature at the first turndown, oС 1620 1610

Slag composition at the first turndown, %:

FеОgeneral 22.8 24.1

MgO 3.5 3.5

Slag basic capacity 3.7 3.6

Melted steel yield, % 92.0 92.0

Recycling

Documents

raw material

karsakpaysk

hrna lump

secondary

kemerovo state

conversion

oily oxide

southern kuzbas