How is tungsten obtained? Development of technology for extracting tungsten from stale tailings of the Dzhidinsky MMC Technology of processing tungsten ores

The main tungsten minerals are scheelite, hübnerite and wolframite. Depending on the type of minerals, ores can be divided into two types; scheelite and wolframite (huebnerite).
Scheelite ores in Russia, as well as in some cases abroad, are enriched by flotation. In Russia, the process of flotation of scheelite ores on an industrial scale was carried out before the Second World War at the Tyrn-Auz factory. This plant processes very complex molybdenum-scheelite ores containing a number of calcium minerals (calcite, fluorite, apatite). Calcium minerals, like scheelite, float with oleic acid; the depression of calcite and fluorite is produced by stirring in a liquid glass solution without heating (long-term contact) or with heating, as at the Tyrn-Auz factory. Instead of oleic acid, fractions of tall oil are used, as well as acids from vegetable oils (reagents 708, 710, etc.) alone or in a mixture with oleic acid.

A typical flotation scheme for scheelite ore is shown in Fig. 38. Using this scheme, it is possible to remove calcite and fluorite and obtain tungsten trioxide-standard concentrates. However, apatite still remains in such quantity that the phosphorus content in the concentrate is higher than standard. Excess phosphorus is removed by dissolving apatite in weak hydrochloric acid. Acid consumption depends on the calcium carbonate content in the concentrate and is 0.5-5 g of acid per ton of WO3.
When leaching with acid, part of the scheelite, as well as powellite, is dissolved and then precipitated out of solution in the form of CaWO4 + CaMoO4 and other impurities. The resulting dirty sludge is then processed according to the I.N. method. Maslenitsky.
Due to the difficulty of obtaining quality tungsten concentrate, many factories abroad produce two products: a rich concentrate and a poor one for hydrometallurgical processing into calcium tungstate using the method developed in Mekhanobra I.N. Maslenitsky, - leaching with soda in an autoclave under pressure with transfer into solution in the form of CaWO4, followed by purification of the solution and precipitation of CaWO4. In some cases, with coarsely disseminated scheelite, finishing of flotation concentrates is carried out on tables.
From ores containing a significant amount of CaF2, extraction of scheelite by flotation has not been developed abroad. Such ores, for example in Sweden, are enriched on tables. Scheelite, entrained with fluorite in the flotation concentrate, is then separated from this concentrate on the table.
In Russian factories, scheelite ores are enriched by flotation, obtaining quality concentrates.
At the Tyrn-Auz plant, concentrates containing 6% WO3 are produced from ore containing 0.2% WO3 with a recovery of 82%. At the Chorukh-Dairon plant, with ore of the same VVO3 content, 72% WO3 is obtained in concentrates with an extraction of 78.4%; at the Koytash plant, with ore with 0.46% WO3 in concentrate, 72.6% WO3 is obtained with a WO3 recovery of 85.2%; at the Lyangarsky plant in ore 0.124%, in concentrates - 72% with extraction of 81.3% WO3. Additional recovery of poor products is possible by reducing losses in tailings. In all cases, if sulfides are present in the ore, they are separated before scheelite flotation.
The consumption of materials and energy is illustrated by the data below, kg/t:

Wolframite (Hübnerite) ores are enriched exclusively by gravity methods. Some ores with uneven and coarse-grained dissemination, such as Bukuki ore (Transbaikalia), can be pre-enriched in heavy suspensions, releasing about 60% waste rock with a particle size of 26+3 MM with a content of no more than 0.03% WO3.
However, with a relatively low productivity of factories (no more than 1000 tons/day), the first stage of enrichment is carried out in jigging machines, usually starting with a particle size of about 10 mm for coarsely disseminated ores. In new modern schemes, in addition to jiggers and tables, Humphrey screw separators are used, replacing part of the tables with them.
A progressive scheme for the enrichment of tungsten ores is shown in Fig. 39.
The finishing of tungsten concentrates depends on their composition.

Sulfides from concentrates thinner than 2 mm are separated by flotogravity: the concentrates, after mixing with acid and flotation reagents (xanthate, oils), are sent to a concentration table; The resulting CO2 concentrate is dried and subjected to magnetic separation. The coarse concentrate is pre-crushed. Sulfides are separated from fine concentrates from slurry tables by foam flotation.
If there are a lot of sulfides, it is advisable to separate them from the discharge of hydrocyclones (or classifier) ​​before enrichment on the tables. This will improve the conditions for the release of wolframite on tables and during concentrate finishing operations.
Typically, rough concentrates before finishing contain about 30% WO3 with recovery up to 85%. For illustration in table. 86 shows some data on factories.

During the gravitational enrichment of wolframite ores (Hübnerite, ferberite) from slurries thinner than 50 microns, the recovery is very low and the losses in the slurry part are significant (10-15% of the content in the ore).
From sludges, flotation with fatty acids at pH=10 can further extract WO3 into lean products containing 7-15% WO3. These products are suitable for hydrometallurgical processing.
Wolframite (Hübnerite) ores contain a certain amount of non-ferrous, rare and noble metals. Some of them pass during gravity enrichment into gravity concentrates and are transferred to finishing tailings. From sulfide finishing tailings, as well as from sludge, molybdenum, bismuth-lead, lead-copper-silver, zinc (they contain cadmium, indium) and pyrite concentrates can be isolated by selective flotation, and the tungsten product can also be isolated.

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The invention relates to a method complex processing tailings from the enrichment of tungsten-containing ores. The method includes their classification into small and large fractions, screw separation of the fine fraction to obtain a tungsten product and its cleaning. In this case, re-cleaning is carried out on a screw separator to obtain a rough tungsten concentrate, which is refined on concentration tables to obtain a gravity tungsten concentrate, which is subjected to flotation to obtain a high-grade conditioned tungsten concentrate and a sulfide-containing product. The tailings of the screw separator and the concentration table are combined and subjected to thickening. In this case, the waste obtained after thickening is fed to the classification of tailings for the enrichment of tungsten-containing ores, and the thickened product is subjected to enrichment in a screw separator to obtain secondary waste tailings and a tungsten product, which is sent for cleaning. The technical result is to increase the depth of processing of tungsten-containing ore tailings. 1 salary files, 1 table, 1 ill.

The invention relates to the beneficiation of minerals and can be used in the processing of tailings from the enrichment of tungsten-containing ores.

When processing tungsten-containing ores, as well as their tailings, gravitational, flotation, magnetic, as well as electrostatic, hydrometallurgical and other methods are used (see, for example, Bert P.O., with the participation of K. Mills. Gravity enrichment technology. Translated from English. - M.: Nedra, 1990). So, for preliminary concentration useful components(mineral raw materials) photometric and lumometric sorting are used (for example, the Mount Carbine and King Island processing plants), enrichment in difficult environments (for example, the Portuguese Panasquera factory and the English Hemerdan factory), jigging (especially poor raw materials), magnetic separation in a weak magnetic field (for example, for the separation of pyrite, pyrrhotite) or high-intensity magnetic separation (for the separation of wolframite and cassiterite).

For the processing of tungsten-containing sludge, the use of flotation is known, in particular wolframite in China and at the Canadian Mount Pleasade factory, and in some factories flotation has completely replaced gravity enrichment (for example, the Jokberg factories, Sweden and Mittersil, Austria).

It is also known to use screw separators and screw sluices for the enrichment of tungsten-containing ores, old dumps, stale tailings, and sludge.

For example, when processing old dumps of tungsten ore at the Cherdoyak factory (Kazakhstan), the initial dump material, after crushing and grinding to a size of 3 mm, was subjected to enrichment on jiggers, the under-sieve product of which was then cleaned on a concentration table. The technological scheme also included enrichment in screw separators, in which 75-77% of WO 3 was recovered with a yield of enrichment products of 25-30%. Screw separation made it possible to increase the extraction of WO 3 by 3-4% (see, for example, Anikin M.F., Ivanov V.D., Pevzner M.L. “Screw separators for ore dressing”, Moscow, Nedra publishing house ", 1970, 132 pp.).

Disadvantages technological scheme processing of old dumps is a high load at the head of the process for the jigging operation, insufficiently high WO 3 extraction and a significant yield of enrichment products.

There is a known method for the associated production of tungsten concentrate by processing molybdenite flotation tailings (Climax Molybdenum factory, Canada). Tailings containing tungsten are separated using screw separation into tungsten waste sludge (light fraction), primary wolframite - cassiterite concentrate. The latter is subjected to hydrocyclonation and the sludge discharge is sent to the waste tailings, and the sand fraction is sent to the flotation separation of pyrite concentrate containing 50% S (sulfides) and discharged into the waste tailings. The chamber product of sulfide flotation is purified using screw separation and/or cones to obtain waste pyrite-containing tailings and wolframite-cassiterite concentrate, which is processed on concentration tables. In this case, wolframite-cassiterite concentrate and waste tailings are obtained. After dehydration, the crude concentrate is cleaned sequentially by purifying it from iron using magnetic separation, removing monazite from it by flotation (phosphate flotation) and then dewatering, drying, classifying and separating using stage magnetic separation into a concentrate containing 65% WO 3 after stage I and 68% WO 3 after stage II. A non-magnetic product is also obtained - tin (cassiterite) concentrate containing ~35% tin.

This processing method has disadvantages - complexity and multi-stage nature, as well as high energy intensity.

There is a known method for additional extraction of tungsten from gravity enrichment tailings (Boulder plant, USA). Gravity enrichment tailings are further crushed and deslimed in a classifier, the sands of which are separated using hydraulic classifiers. The resulting classes are enriched separately on concentration tables. Coarse tailings are returned to the grinding cycle, and fine tailings are thickened and re-enriched on slurry tables to produce a finished concentrate, middlings sent to regrinding, and tailings sent to flotation. The main flotation concentrate is subjected to one cleaning. The original ore contains 0.3-0.5% WO 3; Tungsten recovery reaches 97%, with about 70% of tungsten recovered by flotation. However, the tungsten content in the flotation concentrate is low (about 10% WO 3) (see, Polkin S.I., Adamov E.V. Enrichment of non-ferrous metal ores. Textbook for universities. M., Nedra, 1983, 213 pp.)

The disadvantages of the technological scheme for processing tailings from gravity enrichment are the high load at the head of the process on the enrichment operation on concentration tables, multi-operation, and low quality of the resulting concentrate.

There is a known method for processing scheelite-containing tailings in order to remove hazardous materials from them and process non-hazardous and ore minerals using an improved separation process (KR 20030089109, CHAE et al., 11/21/2003). The method includes the stages of homogenizing mixing of scheelite-containing tailings, introduction of the pulp into the reactor, “filtration” of the pulp using a screen to remove various foreign materials, subsequent separation of the pulp by screw separation, thickening and dehydration of non-metallic minerals to produce a cake, drying the cake in a rotary dryer, crushing the dry cake using a hammer crusher operating in a closed cycle with a screen, separation of crushed minerals using a “micron” separator into fractions of fine and coarse grains (granules), as well as magnetic separation of the coarse-grained fraction to obtain magnetic minerals and a non-magnetic fraction containing scheelite. The disadvantage of this method is the multi-operation nature and the use of energy-intensive drying of the wet cake.

There is a known method for additional extraction of tungsten from the waste tailings of the concentrating plant of the Ingichki mine (see A.B. Ezhkov, Kh.T. Sharipov, K.L. Belkov “Involvement in the processing of stale tungsten-containing tailings of the Ingichki mine.” Abstracts of reports of the III Congress of concentrators of the CIS countries, vol.1, MISiS, M., 2001). The method includes preparing the pulp and desliming it in a hydrocyclone (removal class - 0.05 mm), subsequent separation of the deslimed pulp on a cone separator, two-stage re-cleaning of the cone separator concentrate on concentration tables to obtain a concentrate containing 20.6% WO 3 with an average recovery 29.06%. The disadvantages of this method are the low quality of the resulting concentrate and insufficiently high WO 3 extraction.

The results of research on the gravitational enrichment of tailings from the Ingichkinsky enrichment plant are described (see S.V. Rudnev, V.A. Potapov, N.V. Salikhova, A.A. Kanzel “Research on the selection of the optimal technological scheme for the gravitational enrichment of man-made formations at the Ingichkinsky enrichment plant "//Mining Bulletin of Uzbekistan, 2008, No. 3).

The closest to a patented technical solution is a method for extracting tungsten from stale tailings of the enrichment of tungsten-containing ores (Artemova O.S. Development of a technology for extracting tungsten from stale tailings of the Dzhida VMC. Abstract of the thesis. Candidate of Technical Sciences, Irkutsk State Technical University, Irkutsk, 2004 - prototype).

The technology for extracting tungsten from stale tailings using this method includes the operations of obtaining rough tungsten-containing concentrate and middling product, gold-bearing product and secondary waste tailings using gravitational methods of wet enrichment - screw and centrifugal separation - and subsequent finishing of the resulting rough concentrate and middling product using gravitational (centrifugal) enrichment and magnetic separation to obtain a conditioned tungsten concentrate containing 62.7% WO 3 with a recovery of 49.9% WO 3 .

According to this method, stale tailings are subjected to primary classification with the release of 44.5% of the mass. into secondary tailings in the form of a +3 mm fraction. The tailings fraction with a particle size of -3 mm is divided into classes -0.5 and +0.5 mm, and from the latter, coarse concentrate and tailings are obtained using screw separation. The -0.5 mm fraction is divided into classes -0.1 and +0.1 mm. From the +0.1 mm class, a coarse concentrate is separated using centrifugal separation, which, like the coarse concentrate of screw separation, is subjected to centrifugal separation to obtain rough tungsten concentrate and a gold-containing product. The tailings of screw and centrifugal separation are further crushed to -0.1 mm in a closed cycle with classification and then divided into classes -0.1+0.02 and -0.02 mm. The -0.02 mm grade is removed from the process as secondary tailings. Class -0.1+0.02 mm is enriched by centrifugal separation to produce secondary tailings and tungsten middlings, sent for finishing by magnetic separation along with the centrifugal separation concentrate, ground to a size of -0.1 mm. In this case, tungsten concentrate (magnetic fraction) and middling product (non-magnetic fraction) are obtained. The latter is subjected to magnetic separation II with the release of a non-magnetic fraction into secondary tailings and tungsten concentrate (magnetic fraction), which is enriched sequentially by centrifugal, magnetic and again centrifugal separation to obtain standard tungsten concentrate containing 62.7% WO 3 with a yield of 0.14 % and recovery 49.9%. In this case, the tailings of centrifugal separations and the non-magnetic fraction are sent to secondary tailings, the total yield of which at the stage of finishing the rough tungsten concentrate is 3.28% with a content of 2.1% WO 3.

The disadvantages of this method are the multi-operational nature of the technological process, which includes 6 classification operations, 2 additional grinding operations, as well as 5 centrifugal and 3 magnetic separation operations using relatively expensive equipment. At the same time, finishing the crude tungsten concentrate to the required standard is associated with the production of secondary waste tailings with a relatively high tungsten content (2.1% WO 3).

The objective of the present invention is to improve the method of processing enrichment tailings, including stale tailings from the enrichment of tungsten-containing ores, to obtain high-grade tungsten concentrate and an associated sulfide-containing product while reducing the tungsten content in the secondary waste tailings.

The patented method for complex processing of tailings from the enrichment of tungsten-containing ores includes classification of tailings into small and large fractions, screw separation of the fine fraction to obtain a tungsten product, re-cleaning of the tungsten product, and finishing to obtain high-grade tungsten concentrate, sulfide-containing product and secondary waste tailings.

The method differs in that the resulting tungsten product is subjected to re-cleaning on a screw separator to obtain rough concentrate and tailings, the rough concentrate is subjected to finishing on concentration tables to obtain gravitational tungsten concentrate and tailings. The tailings of the concentration table and the cleaning screw separator are combined and subjected to thickening, then the thickening discharge is fed to the classification stage at the head of the technological scheme, and the thickened product is subjected to enrichment on a screw separator to obtain secondary tailings and a tungsten product, which is sent for cleaning. The gravity tungsten concentrate is subjected to flotation to obtain a high-grade tungsten concentrate (62% WO 3) and a sulfide-containing product, which is processed by known methods.

The method can be characterized by the fact that the tailings are classified into fractions, mainly with a size of +8 mm and -8 mm.

The technical result of the patented method is to increase the depth of processing while reducing the number of technological operations and the load on them due to the separation at the head of the process of the bulk of the initial tailings (more than 90%) into secondary waste tailings, using energy-saving screw separation technology that is simpler in design and operation. This allows you to dramatically reduce the load on subsequent enrichment operations, as well as capital expenditures and operating costs, which ensures optimization of the enrichment process.

The effectiveness of the patented method is shown using the example of complex processing of tailings from the Ingichkinsky enrichment plant (see drawing).

Processing begins with the classification of tailings into small and large fractions with the separation of secondary waste tailings in the form of a large fraction. The fine fraction of the tailings is subjected to screw separation with the separation of the bulk of the original tailings (more than 90%) at the head of the technological process into secondary dump tailings. This allows for a correspondingly dramatic reduction in downstream workload, capital costs and operating costs.

The resulting tungsten product is subjected to re-cleaning using a screw separator to obtain rough concentrate and tailings. The rough concentrate is subjected to finishing on concentration tables to obtain gravity tungsten concentrate and tailings.

The tailings of the concentration table and the cleaning screw separator are combined and subjected to thickening, for example, in a thickener, mechanical classifier, hydrocyclone and other devices. The condensation discharge is fed to the classification stage at the head of the technological scheme, and the condensed product is subjected to enrichment in a screw separator to obtain secondary tailings and a tungsten product, which is sent for cleaning.

The gravity tungsten concentrate is brought by flotation to a high-grade tungsten concentrate (62% WO 3) to obtain a sulfide-containing product.

Thus, high-grade (62% WO 3 ) conditioned tungsten concentrate is isolated from tungsten-containing tailings upon achieving a relatively high WO 3 extraction of ~49% and a relatively low tungsten content (0.04% WO 3 ) in the secondary waste tailings.

The resulting sulfide-containing product is processed in a known way, for example, is used to produce sulfuric acid and sulfur, and is also used as a corrective additive in the production of cements.

High-grade tungsten concentrate is a highly liquid commercial product.

As follows from the results of implementing the patented method using the example of stale tailings from the enrichment of tungsten-containing ores from the Ingichkinsky concentrating plant, its effectiveness is shown in comparison with the prototype method (see table). Provides additional production of sulfide-containing product, reducing the volume of fresh water consumed by creating a water cycle. It creates the possibility of processing significantly poorer tailings (0.09% WO 3), a significant reduction in the tungsten content in secondary waste tailings (up to 0.04% WO 3). In addition, the number of technological operations has been reduced and the load on most of them has been reduced due to the separation at the head of the technological process of the bulk of the initial tailings (more than 90%) into secondary waste tailings, using a simpler and less energy-intensive screw separation technology, which reduces capital costs for the purchase of equipment and operating costs.

1. A method for complex processing of tailings from the enrichment of tungsten-containing ores, including their classification into small and large fractions, screw separation of the fine fraction to produce a tungsten product, its re-cleaning and finishing to produce high-grade tungsten concentrate, sulfide-containing product and secondary waste tailings, characterized in that the resulting after screw separation, the tungsten product is subjected to re-cleaning on a screw separator to obtain rough tungsten concentrate, the resulting rough tungsten concentrate is subjected to finishing on concentration tables to obtain gravity tungsten concentrate, which is subjected to flotation to obtain high-grade conditioned tungsten concentrate and sulfide-containing product, tailings of the screw separator and concentration table are combined and subjected to thickening, the resulting waste after thickening is fed to the classification of tailings for the enrichment of tungsten-containing ores, and the thickened product is subjected to enrichment in a screw separator to obtain secondary waste tailings and a tungsten product, which is sent for cleaning.

Introduction

1 . The importance of technogenic mineral raw materials

1.1. Mineral resources of the ore industry in the Russian Federation and the tungsten sub-industry

1.2. Technogenic mineral formations. Classification. Need for use

1.3. Technogenic mineral formation of the Dzhida VMC

1.4. Goals and objectives of the study. Research methods. Provisions for defense

2. Study of the material composition and technological properties of stale tailings from the Dzhidinsky MMC

2.1. Geological testing and evaluation of tungsten distribution

2.2. Material composition of mineral raw materials

2.3. Technological properties of mineral raw materials

2.3.1. Grading

2.3.2. Study of the possibility of radiometric separation of mineral raw materials in the original size

2.3.3. Gravity analysis

2.3.4. Magnetic analysis

3. Development of a technological scheme

3.1. Technological testing of various gravity devices for the enrichment of stale tailings of various sizes

3.2. Optimization of the general waste processing scheme

3.3. Pilot testing of the developed technological scheme for the enrichment of general waste and an industrial plant

Introduction to the work

The sciences of mineral processing are primarily aimed at developing theoretical foundations mineral separation processes and the creation of enrichment apparatuses, to reveal the relationship between the patterns of distribution of components and separation conditions in enrichment products in order to increase the selectivity and speed of separation, its efficiency and economy, and environmental safety.

Despite significant mineral reserves and a decline in last years resource consumption, depletion of mineral resources is one of the most important problems in Russia. Poor use of resource-saving technologies contributes to large losses of minerals during the extraction and enrichment of raw materials.

An analysis of the development of equipment and technology for mineral processing over the past 10-15 years indicates significant achievements in the domestic fundamental science in the field of knowledge of the main phenomena and patterns in the separation of mineral complexes, which makes it possible to create highly efficient processes and technologies for the primary processing of ores of complex material composition and, as a result, provide the metallurgical industry with the necessary range and quality of concentrates. At the same time, in our country, in comparison with developed foreign countries, there is still a significant lag in the development of the machine-building base for the production of main and auxiliary enrichment equipment, in its quality, metal intensity, energy intensity and wear resistance.

In addition, due to the departmental affiliation of mining and processing enterprises, complex raw materials were processed only taking into account the necessary industry needs for a specific metal, which led to the irrational use of natural mineral resources and increased costs for waste storage. Currently accumulated

more than 12 billion tons of waste, the content of valuable components in which in some cases exceeds their content in natural deposits.

In addition to the above negative trends, since the 90s, the environmental situation at mining and processing enterprises has sharply worsened (in a number of regions, threatening the existence of not only biota, but also humans), there has been a progressive decline in the production of non-ferrous and ferrous metal ores, mining and chemical raw materials, deterioration in the quality of processed ores and, as a consequence, the involvement in the processing of difficult-to-process ores of complex material composition, characterized by a low content of valuable components, fine dissemination and similar technological properties of minerals. Thus, over the past 20 years, the content of non-ferrous metals in ores has decreased by 1.3-1.5 times, iron by 1.25 times, gold by 1.2 times, the share of difficult ores and coal has increased from 15% to 40% of the total mass of raw materials supplied for enrichment.

Human impact on the natural environment in the process economic activity is now becoming global. In terms of the scale of extracted and transported rocks, transformation of the relief, impact on the redistribution and dynamics of surface and groundwater, activation of geochemical transfer, etc. this activity is comparable to geological processes.

The unprecedented scale of extracted mineral resources leads to their rapid depletion and accumulation on the Earth's surface, in the atmosphere and hydrosphere large number waste, gradual degradation natural landscapes, reduction in biodiversity, reduction in the natural potential of territories and their life-supporting functions.

Ore processing waste storage facilities are objects of increased environmental hazard due to their negative impact on the air basin, ground and surface waters, soil cover over vast areas. Along with this, tailings dumps are little-studied technogenic deposits, the use of which will make it possible to obtain additional

sources of ore and mineral raw materials with a significant reduction in the scale of disturbance of the geological environment in the region.

Production of products from technogenic deposits, as a rule, is several times cheaper than from raw materials specially mined for this purpose, and is characterized by a quick return on investment. However, the complex chemical, mineralogical and granulometric composition of tailings, as well as a wide range of minerals contained in them (from main and associated components to the simplest building materials) make it difficult to calculate the total economic effect from their processing and determine individual approach to the assessment of each tailings dump.

Consequently, at the moment a number of insoluble contradictions have emerged between the change in the nature of the mineral resource base, i.e. the need to involve difficult-to-process ores and technogenic deposits in the processing, the environmentally aggravated situation in mining regions and the state of technology, technology and organization of primary processing of mineral raw materials.

The issues of using waste from the enrichment of polymetallic, gold-containing and rare metals have both economic and environmental aspects.

In achieving the current level of development of the theory and practice of processing tailings from the enrichment of non-ferrous, rare and precious metal ores huge contribution contributed by V.A. Chanturia, V.Z. Kozin, V.M. Avdokhin, SB. Leonov, L.A. Barsky, A.A. Abramov, V.I. Karmazin, SI. Mitrofanov and others.

Important integral part general strategy of the ore industry, incl. tungsten, is the increased use of ore processing waste as additional sources of ore and mineral raw materials, with a significant reduction in the scale of disturbance of the geological environment in the region and the negative impact on all components environment.

In the field of using ore processing waste, the most important thing is a detailed mineralogical and technological study of each specific

an individual technogenic deposit, the results of which will allow the development of an effective and environmentally friendly technology for the industrial development of an additional source of ore and mineral raw materials.

The problems considered in the dissertation work were solved in accordance with scientific direction Department of Mineral Processing and Environmental Engineering of Irkutsk State technical university on the topic “Fundamental and technological research in the field of processing of mineral and technogenic raw materials for the purpose of their integrated use, taking into account environmental problems in complex industrial systems” and x/d topic No. 118 “Research on the beneficiation of stale tailings of the Dzhida VMC.”

Goal of the work- scientifically substantiate, develop and test
rational technological methods for enriching stale

The following tasks were solved in the work:

Evaluate the distribution of tungsten throughout the entire space of the main
technogenic education of the Dzhida VMC;

study the material composition of the stale tailings of the Dzhizhinsky VMC;

study the contrast of stale tailings in the original size in terms of the content of W and S (II);

to study the gravitational enrichment of stale tailings of the Dzhida VMC in various sizes;

determine the feasibility of using magnetic enrichment to improve the quality of crude tungsten-containing concentrates;

to optimize the technological scheme for the enrichment of technogenic raw materials of the general waste treatment plant of the Dzhida VMC;

conduct pilot tests of the developed scheme for extracting W from the stale tailings of DVMC;

To develop a circuit diagram of devices for the industrial processing of stale tailings from the Dzhida VMC.

To carry out the research, a representative technological sample of stale tailings from the Dzhida VMC was used.

When solving the formulated problems, the following were used research methods: spectral, optical, chemical, mineralogical, phase, gravitational and magnetic methods for analyzing the material composition and technological properties of initial mineral raw materials and enrichment products.

The following are submitted for defense: basic scientific principles:

The patterns of distribution of initial technogenic mineral raw materials and tungsten by size classes have been established. The need for primary (preliminary) classification by size of 3 mm has been proven.

The quantitative characteristics of the stale ore dressing tailings of the Dzhidinsky VMC in terms of WO3 and sulfide sulfur content have been established. It has been proven that the initial mineral raw materials belong to the category of non-contrasting ores. A reliable and reliable correlation between the contents of WO3 and S (II) was revealed.

Quantitative patterns of gravitational enrichment of stale tailings from the Dzhida VMC have been established. It has been proven that for source material of any size, an effective method for extracting W is gravitational enrichment. Forecast technological indicators of gravitational enrichment of initial mineral raw materials have been determined V of various sizes.

Quantitative patterns of distribution of stale ore dressing tailings of the Dzhida VMC into fractions of different specific magnetic susceptibility have been established. The effectiveness of the sequential use of magnetic and centrifugal separation has been proven to improve the quality of rough W-containing products. The technological modes of magnetic separation have been optimized.

Material composition of mineral raw materials

When examining a secondary tailings dump (emergency discharge tailings dump (EDT)), 35 furrow samples were taken from pits and clearings along the slopes of the dumps; the total length of the furrows is 46 m. ​​The pits and clearings are located in 6 exploration lines, spaced 40-100 m from each other; the distance between pits (clearings) in exploration lines is from 30-40 to 100-150 m. All lithological varieties of sands were tested. Samples were analyzed for W03 and S(II) content. In this area, 13 samples were taken from pits with a depth of 1.0 m. The distance between the lines is about 200 m, between the workings - from 40 to 100 m (depending on the distribution of the same type of lithological layer). The results of sample analyzes for WO3 and sulfur content are given in table. 2.1. Table 2.1 - Content of WO3 and sulfide sulfur in private samples of CAS It can be seen that the content of WO3 ranges from 0.05-0.09%, with the exception of sample M-16, selected from medium-grained gray sands. In the same sample, high concentrations of S (II) were found - 4.23% and 3.67%. For individual samples (M-8, M-18), a high content of S sulfate was noted (20-30% of the total sulfur content). In the upper part of the emergency discharge tailings dump, 11 samples of various lithological varieties were taken. The content of WO3 and S (II), depending on the origin of the sands, varies over a wide range: from 0.09 to 0.29% and from 0.78 to 5.8%, respectively. Elevated WO3 contents are typical for medium-to-coarse-grained sand varieties. The S(VI) content is 80 - 82% of the total S content, but in individual samples, predominantly with low contents of tungsten trioxide and total sulfur, it decreases to 30%.

The deposit's reserves can be assessed as Pj category resources (see Table 2.2). Along the upper part, the length of the pit varies in a wide range: from 0.7 to 9.0 m, therefore the average content of controlled components is calculated taking into account the parameters of the pits. In our opinion, based on the given characteristics, taking into account the composition of stale tailings, their preservation, conditions of occurrence, contamination household waste, their WO3 content and the degree of sulfur oxidation, only the upper part of the emergency discharge tailings with resources of 1.0 million tons of sand and 1330 tons of WO3 with a WO3 content of 0.126% may be of industrial interest. Their location in close proximity to the designed enrichment plant (250-300 m) is favorable for their transportation. The lower part of the emergency discharge tailings dump is subject to disposal as part of the environmental rehabilitation program for the city of Zakamensk.

5 samples were taken from the deposit area. The interval between sampling points is 1000-1250 m. Samples were taken over the entire thickness of the layer and analyzed for the content of WO3, Btot and S (II) (see Table 2.3). Table 2.3 - Content of WO3 and sulfur in private ATO samples From the analysis results it is clear that the content of WO3 is low, varying from 0.04 to 0.10%. The average S(II) content is 0.12% and is of no practical interest. The work carried out does not allow us to consider the by-product alluvial tailings dump as a potential industrial facility. However, as a source of environmental pollution, these formations must be disposed of. The main tailings dump (MTD) was explored along parallel exploration lines oriented at azimuth 120 and located 160 - 180 m from each other. The exploration lines are oriented across the strike of the dam and the slurry pipeline, through which the ore tailings were discharged, deposited subparallel to the dam crest. Thus, the exploration lines were also oriented across the bedding of technogenic deposits. Along the exploration lines, a bulldozer drove trenches to a depth of 3-5 m, from which pits were drilled to a depth of 1 to 4 m. The depth of the trenches and pits was limited by the stability of the walls of the workings. The pits in the trenches were made through 20 - 50 m in the central part of the deposit and through 100 m - on the south-eastern flank, on the area of ​​​​the former settling pond (now dried up), from which water was supplied to the processing plants during the operation of the plant.

The area of ​​the OTO along the distribution boundary is 1015 thousand m (101.5 hectares); along the long axis (along the valley of the Barun-Naryn river) it extends for 1580 m, in the transverse direction (near the dam) its width is 1050 m. In this area, 78 pits were made from pre-created trenches in five main exploration lines. Consequently, one pit illuminates an area of ​​12,850 m, which is equivalent to an average network of 130x100 m. In the central part of the field, represented by sands of different grains, in the area where slurry lines are located on an area of ​​530 thousand m (52% of the TMO area), 58 pits and one well (75% all workings); The exploration network area averaged 90x100 m2. On the extreme southeastern flank, on the site of a former settling pond in the area of ​​development of fine-grained sediments - silts, 12 pits (15% of the total number) were drilled, characterizing an area of ​​​​about 370 thousand m (37% of the total area of ​​the technogenic deposit); the average network area here was 310x100 m2. In the area of ​​transition from heterogeneous sands to silts, composed of silty sands, on an area of ​​about 115 thousand m (11% of the area of ​​the technogenic deposit), 8 pits were drilled (10% of the number of workings in the technogenic deposit) and the average area of ​​the exploration network was 145x100 m. Average length the sampled section at the technogenic deposit is 4.3 m, including for different-grained sands - 5.2 m, silty sands - 2.1 m, silts - 1.3 m. Absolute marks The modern topography of the surface of the technogenic deposit in the tested sections varies from 1110-1115 m near the upper part of the dam, to 1146-148 m in the central part and 1130-1135 m on the southeastern flank. In total, 60 - 65% of the capacity of the technogenic deposit has been tested. Trenches, pits, strippings and burials were documented in M ​​1:50 -1:100 and tested with a furrow with a cross section of 0.1x0.05 m2 (1999) and 0.05x0.05 m2 (2000). The length of the furrow samples was 1 m, the weight was 10 - 12 kg in 1999. and 4 - 6 kg in 2000. The total length of the tested intervals in the exploration lines was 338 m, in general, taking into account the areas of detailing and individual sections outside the network - 459 m. The weight of the samples taken was 5 tons.

The samples, together with a passport (characteristics of the rock, sample number, production and performer) were packaged in plastic and then fabric bags and sent to the RAC of the Republic of Buryatia, where they were weighed, dried, analyzed for the content of W03, and S (II) according to NS AM methods. The accuracy of the analyzes is confirmed by the comparability of the results of ordinary, group (RAC analyses) and technological (TsNIGRI and VIMS analyses) samples. The results of the analysis of private technological samples taken at the OTO are given in Appendix 1. The main (OTO) and two secondary tailings dumps (KhAT and ATO) of the Dzhida VMC were statistically compared in terms of WO3 content using the Student's t test (see Appendix 2). With a confidence probability of 95% it was established: - no significant statistical difference in WO3 content between private samples of side tailings; - average results of OTO testing for WO3 content in 1999 and 2000. belong to the same general population. Hence, chemical composition the main tailings pond changes insignificantly over time under the influence external influences. All general waste reserves can be processed using a single technology.; - average sampling results of the main and side tailings dumps in terms of WO3 content differ significantly from each other. Consequently, to involve mineral raw materials from side tailings, the development of local enrichment technology is required.

Technological properties of mineral raw materials

Based on their granular composition, sediments are divided into three types of sediments: heterogeneous sands; silty sands (silty); silts There are gradual transitions between these types of sediments. Clearer boundaries are observed in the thickness of the section. They are caused by the alternation of sediments of different grain compositions, different color(from dark green to light yellow and gray) and different material composition (quartz-feldspathic nonmetallic part and sulfide with magnetite, hematite, hydroxides of iron and manganese). The entire thickness is layered - from fine to coarsely layered; the latter is more typical for coarse-grained varieties of sediments or layers of significant sulfide mineralization. Fine-grained (silty, silt fractions, or layers composed of dark-colored materials - amphibole, hematite, goethite) usually form thin (a few cm - mm) layers. The occurrence of the entire thickness of sediments is subhorizontal with a predominant fall of 1-5 in the northern directions. Sands of different grains are located in the northwestern and central parts of the OTO, which is due to their sedimentation near the source of discharge - the pulp pipeline. The width of the strip of different-grained sands is 400-500 m; along the strike they occupy the entire width of the valley - 900-1000 m. The color of the sands is gray-yellow, yellow-green. The granular composition is variable - from fine-grained to coarse-grained varieties up to lenses of gravelstones 5-20 cm thick and up to 10-15 m long. Silty (silty) sands stand out in the form of a layer 7-10 m thick (horizontal thickness, outcrop 110-120 m ). They lie under heterogeneous sands. In cross-section they represent a layered formation of gray, greenish-gray color with alternation of fine-grained sands with layers of silt. The volume of silts in the section of silty sands increases in the southeast direction, where silts make up the main part of the section.

Silts make up the southeastern part of the OTO and are represented by finer particles of enrichment waste of dark gray, dark green, bluish-green color with layers of grayish-yellow sand. The main feature of their structure is a more uniform, more massive texture with less frequent and less clearly defined layering. The silts are underlain by silty sands and lie on the base of the bed - alluvial-deluvial deposits. The granulometric characteristics of OTO mineral raw materials with the distribution of gold, tungsten, lead, zinc, copper, fluorite (calcium and fluorine) by size class are given in Table. 2.8. According to granulometric analysis, the bulk of the OTO sample material (about 58%) has a particle size of -1 + 0.25 mm, 17% each is coarse (-3 + 1 mm) and small (-0.25 + 0.1) mm classes. The share of material with a particle size of less than 0.1 mm is about 8%, of which half (4.13%) is of the slurry class - 0.044 + 0 mm. Tungsten is characterized by a slight fluctuation in content in size classes from -3 +1 mm to -0.25+0.1 mm (0.04-0.05%) and a sharp increase (up to 0.38%) in size class -0 .1+0.044 mm. In the slurry class -0.044+0 mm, the tungsten content is reduced to 0.19%. The accumulation of hübnerite occurs only in small-sized material, that is, in the class -0.1 + 0.044 mm. Thus, 25.28% of tungsten is concentrated in the -0.1+0.044 mm class with an output of this class of about 4% and 37.58% in the -0.1+0 mm class with an output of this class of 8.37%. Differential and integral histograms of the distribution of particles of GTO mineral raw materials by size class and histograms of the absolute and relative distribution of W by size class of GTO mineral raw materials are presented in Fig. 2.2. and 2.3. In table Table 2.9 shows data on the dissemination of hübnerite and scheelite in the OTO mineral raw material of the original size and crushed to - 0.5 mm.

In the -5+3 mm class of initial mineral raw materials there are no pobnerite and scheelite grains, as well as intergrowths. In the -3+1 mm class, the content of free scheelite and hübnerite grains is quite large (37.2% and 36.1%, respectively). In the -1+0.5 mm class, both mineral forms of tungsten are present in almost equal quantities, both in the form of free grains and in the form of intergrowths. In thin classes -0.5+0.25, -0.25+0.125, -0.125+0.063, -0.063+0 mm, the content of free grains of scheelite and hübnerite is significantly higher than the content of intergrowths (the content of intergrowths varies from 11.9 to 3. 0%) The size class -1+0.5 mm is limiting and in it the content of free grains of scheelite and hübnerite and their intergrowths is almost the same. Based on the data in table. 2.9, we can conclude that it is necessary to classify delimed mineral raw materials OTO according to a particle size of 0.1 mm and separate enrichment of the resulting classes. From the large class, it is necessary to separate the free grains into a concentrate, and the tailings containing splices must be subjected to further grinding. The crushed and deslimed tailings should be combined with the deslimed class -0.1+0.044 of the initial mineral raw materials and sent to gravity operation II in order to extract fine grains of scheelite and pobnerite into the middling product.

2.3.2 Study of the possibility of radiometric separation of mineral raw materials in the original size Radiometric separation is the process of large-piece separation of ores according to the content of valuable components, based on selective influence various types radiation on the properties of minerals and chemical elements. Over twenty methods of radiometric enrichment are known; the most promising of them are X-ray radiometric, X-ray luminescence, radio resonance, photometric, autoradiometric and neutron absorption. Using radiometric methods, the following technological problems are solved: preliminary enrichment with the removal of waste rock from ore; selection of technological varieties, varieties with subsequent enrichment according to separate schemes; selection of products suitable for chemical and metallurgical processing. Assessment of radiometric enrichment includes two stages: studying the properties of ores and experimental determination technological indicators enrichment. At the first stage, the following basic properties are studied: the content of valuable and harmful components, particle size distribution, single- and multi-component contrast of ore. At this stage, the fundamental possibility of using radiometric enrichment is established, the maximum separation indices are determined (at the stage of contrast study), separation methods and characteristics are selected, their effectiveness is assessed, theoretical separation indices are determined, and schematic diagram radiometric enrichment taking into account the features of subsequent processing technology. At the second stage, the modes and practical results separation, conduct large-scale laboratory tests of the radiometric enrichment scheme, select a rational version of the scheme based on a technical and economic comparison of the combined technology (with radiometric separation at the beginning of the process) with the basic (traditional) technology.

In each specific case, the mass, size and number of technological samples are determined depending on the properties of the ore, the structural features of the deposit and methods of its exploration. The content of valuable components and the uniformity of their distribution in the ore mass are the determining factors in the use of radiometric enrichment. The choice of radiometric enrichment method is influenced by the presence of impurity elements isomorphically associated with useful minerals and in some cases playing the role of indicators, as well as the content of harmful impurities, which can also be used for these purposes.

Optimization of the general waste processing scheme

In connection with the involvement in industrial exploitation of low-grade ores with a tungsten content of 0.3-0.4%, in recent years multi-stage combined enrichment schemes based on a combination of gravity, flotation, magnetic and electrical separation, chemical finishing of low-grade flotation concentrates, etc. have become widespread. . A special International Congress in 1982 from San Francisco. An analysis of the technological schemes of existing enterprises showed that during ore preparation, various methods of preliminary concentration have become widespread: photometric sorting, preliminary jigging, enrichment in heavy environments, wet and dry magnetic separation. In particular, photometric sorting is effectively used at one of the largest suppliers of tungsten products - at the Mount Corbijn plant in Australia, which processes ores with a tungsten content of 0.09% at large factories in China - Taishan and Xihuashan.

For the preliminary concentration of ore components in heavy media, highly efficient Dinavirpul devices from Sala (Sweden) are used. Using this technology, the material is classified and the +0.5 mm class is enriched in a heavy environment represented by a ferrosilicon mixture. Some factories use dry and wet magnetic separation as pre-concentration. Thus, at the Emerson plant in the USA, wet magnetic separation is used to separate the pyrrhotite and magnetite contained in the ore, and at the Uyudag plant in Turkey, class - 10 mm is subjected to dry grinding and magnetic separation in separators with low magnetic intensity to isolate magnetite, and then enriched in high tension separators to separate the garnet. Further enrichment includes table concentration, flotogravity and scheelite flotation. An example of the use of multi-stage combined schemes for the enrichment of low-grade tungsten ores, ensuring the production of high-quality concentrates, are the technological schemes used in Chinese factories. Thus, at the Taishan factory with a capacity of 3000 tons/day of ore, wolframite-scheelite material with a tungsten content of 0.25% is processed. The original ore is subjected to manual and photometric sorting with 55% of waste rock removed to the dump. Further enrichment is carried out on jigging machines and concentration tables. The resulting rough gravity concentrates are finished using flotogravity and flotation methods. Xihuashan, which processes ore with a 10:1 ratio of wolframite to scheelite, uses a similar gravity cycle. The crude gravity concentrate is sent to flotogravity and flotation, through which sulfides are removed. Next, wet magnetic separation of the chamber product is carried out to isolate wolframite and rare earth minerals. The magnetic fraction is sent to electrostatic separation and then flotation of wolframite. The non-magnetic fraction is fed to sulfide flotation, and the flotation tailings are subjected to magnetic separation to produce scheelite and cassiterite-wolframite concentrates. The total WO3 content is 65% with a recovery of 85%.

There has been an increase in the use of the flotation process in combination with chemical finishing of the resulting poor concentrates. In Canada, at the Mount Pleasant plant, flotation technology has been adopted for the beneficiation of complex tungsten-molybdenum ores, including the flotation of sulfides, molybdenite and wolframite. In the main sulfide flotation, copper, molybdenum, lead, and zinc are recovered. The concentrate is cleaned, further crushed, steamed and conditioned with sodium sulfide. The molybdenum concentrate is purified and subjected to acid leaching. Sulfide flotation tailings are treated with sodium fluoride to depress gangue minerals and wolframite is floated with organophosphorus acid, followed by leaching of the resulting wolframite concentrate with sulfuric acid. At the Kantung factory (Canada), the scheelite flotation process is complicated by the presence of talc in the ore, so a primary talc flotation cycle was introduced, then copper minerals and pyrrhotite are floated. The flotation tailings are subjected to gravity enrichment to produce two tungsten concentrates. Gravity tailings are sent to the scheelite flotation cycle, and the resulting flotation concentrate is processed hydrochloric acid. At the Ixsjöberg factory (Sweden), replacing the gravity-flotation scheme with a purely flotation scheme made it possible to obtain scheelite concentrate containing 68-70% WO3 with a recovery of 90% (according to the gravity-flotation scheme, the recovery was 50%). Much attention in Lately is focused on improving extraction technology tungsten minerals from sludge in two main directions: gravitational enrichment of sludge on modern multi-deck concentrators (similar to the enrichment of tin-containing sludge) with subsequent finishing of the concentrate by flotation and enrichment on wet magnetic separators with high tension magnetic field(for wolframite sludge).

An example of the use of combined technology is factories in China. The technology includes sludge thickening to 25-30% solids, sulfide flotation, tailings enrichment in centrifugal separators. The resulting rough concentrate (WO3 content 24.3% with recovery 55.8%) is sent to wolframite flotation using organophosphorus acid as a collector. Flotation concentrate containing 45% WO3 is subjected to wet magnetic separation to obtain wolframite and tin concentrates. Using this technology, wolframite concentrate containing 61.3% WO3 with a recovery of 61.6% is obtained from sludge containing 0.3-0.4% WO3. Thus, technological schemes for the enrichment of tungsten ores are aimed at increasing the complexity of the use of raw materials and separating all associated valuable components into independent types of products. Thus, at the Kuda factory (Japan), when enriching complex ores, 6 commercial products are obtained. In order to determine the possibility of additional extraction of useful components from stale enrichment tailings in the mid-90s. TsNIGRI studied a technological sample containing 0.1% tungsten trioxide. It has been established that the main valuable component in the tailings is tungsten. The content of non-ferrous metals is quite low: copper 0.01-0.03; lead - 0.09-0.2; zinc -0.06-0.15%, gold and silver were not found in the sample. Studies have shown that successful extraction of tungsten trioxide will require significant costs for regrinding tailings and at this stage involving them in processing is not promising.

A technological scheme for the enrichment of minerals, including two or more devices, embodies all the characteristic features of a complex object, and optimization of the technological scheme can apparently constitute the main task of system analysis. Almost all previously discussed modeling and optimization methods can be used to solve this problem. However, the structure of concentrator plant circuits is so complex that it is necessary to consider additional methods optimization. Indeed, for a circuit consisting of at least 10-12 devices, it is difficult to implement a conventional factorial experiment or carry out multiple nonlinear statistical processing. Currently, several ways to optimize circuits are being outlined - an evolutionary path to generalize the accumulated experience and take a step in the successful direction of changing the circuit.

Pilot testing of the developed technological scheme for the enrichment of general waste and an industrial plant

The tests were carried out in October-November 2003. During the tests, 15 tons of initial mineral raw materials were processed in 24 hours. The results of testing the developed technological scheme are presented in Fig. 3.4 and 3.5 and in table. 3.6. It can be seen that the yield of the standard concentrate is 0.14%, the content is 62.7% with a WO3 recovery of 49.875%. The results of spectral analysis of a representative sample of the obtained concentrate are given in table. 3.7, confirm that W-concentrate III of magnetic separation is standard and complies with the KVG (T) grade of GOST 213-73 “Technical requirements (composition,%) for tungsten concentrates obtained from tungsten-containing ores.” Consequently, the developed technological scheme for the extraction of W from the stale tailings of the ore processing of the Dzhidinsky VMC can be recommended for industrial use and the stale tailings are converted into additional industrial mineral raw materials of the Dzhidinsky VMC.

For the industrial processing of stale tailings using the developed technology at Q = 400 t/h, a list of equipment has been developed, given in To carry out an enrichment operation with a particle size of +0.1 mm, it is recommended to install a KNELSON centrifugal separator with continuous unloading of the concentrate, while for centrifugal enrichment class -0.1 mm must be carried out on a KNELSON centrifugal separator with periodic unloading of the concentrate. Thus, it has been established that the most effective way extraction of WO3 from general waste with a particle size of -3+0.5 mm is carried out by screw separation; from size classes -0.5+0.1 and -0.1+0 mm and primary enrichment tailings crushed to -0.1 mm - centrifugal separation. The essential features of the technology for processing stale tailings from the Dzhida VMC are as follows: 1. A narrow classification of the feed directed to primary enrichment and finishing is necessary; 2. An individual approach is required when choosing a method for primary enrichment of classes of different sizes; 3. Obtaining waste tailings is possible with the primary enrichment of the finest feed (-0.1+0.02mm); 4. Use of hydrocycloning operations to combine dewatering and size separation operations. The drain contains particles with a particle size of -0.02 mm; 5. Compact arrangement of equipment. 6. Profitability of the technological scheme (APPENDIX 4), the final product is a standard concentrate that meets the requirements of GOST 213-73.

Kiselev, Mikhail Yurievich

Tungsten is the most refractory metal, melting point 3380°C. And this determines its scope. It is also impossible to build electronics without tungsten; even the filament in a light bulb is tungsten.

And, naturally, the properties of the metal also determine the difficulties in obtaining it...

First, you need to find ore. These are just two minerals - scheelite (calcium tungstate CaWO 4) and wolframite (iron and manganese tungstate - FeWO 4 or MnWO 4). The latter has been known since the 16th century under the name "wolf's foam" - "Spuma lupi" in Latin, or "Wolf Rahm" in German. This mineral accompanies tin ores and interferes with the smelting of tin, turning it into slag. Therefore, it is possible to find it already in antiquity. Rich tungsten ores usually contain 0.2 - 2% tungsten. Tungsten was actually discovered in 1781.

However, this is the easiest thing to find in tungsten mining.
Next, the ore needs to be enriched. There are a bunch of methods and they are all quite complex. First of all, of course. Then - magnetic separation (if we have wolframite with iron tungstate). Next is gravitational separation, because the metal is very heavy and the ore can be washed, much like when mining gold. Nowadays they still use electrostatic separation, but it is unlikely that the method will be useful to the endangered person.

So, we have separated the ore from the gangue. If we have scheelite (CaWO 4), then we can skip the next step, but if we have wolframite, then we need to turn it into scheelite. To do this, tungsten is extracted with a soda solution under pressure and at elevated temperatures (the process takes place in an autoclave), followed by neutralization and precipitation in the form of artificial scheelite, i.e. calcium tungstate.
It is also possible to sinter wolframite with an excess of soda, then we obtain tungstate not of calcium, but of sodium, which for our purposes is not so significant (4FeWO 4 + 4Na 2 CO 3 + O 2 = 4Na 2 WO 4 + 2Fe 2 O 3 + 4CO 2).

The next two stages are leaching with water CaWO 4 -> H 2 WO 4 and decomposition with hot acid.
You can take different acids - hydrochloric (Na 2 WO 4 + 2HCl = H 2 WO 4 + 2NaCl) or nitric.
As a result, tungsten acid is isolated. The latter is calcined or dissolved in an aqueous solution of NH 3, from which paratungstate is crystallized by evaporation.
As a result, it is possible to obtain the main raw material for the production of tungsten - WO 3 trioxide with good purity.

Of course, there is also a method for producing WO 3 using chlorides, when tungsten concentrate is treated with chlorine at elevated temperatures, but this method will not be simple for the outsider.

Tungsten oxides can be used in metallurgy as an alloying additive.

So, we have tungsten trioxide and there is only one step left - reduction to metal.
There are two methods here - reduction with hydrogen and reduction with carbon. In the second case, the coal and the impurities it always contains react with tungsten, forming carbides and other compounds. Therefore, tungsten comes out “dirty”, brittle, and for electronics it is pure that is very desirable, because having only 0.1% iron, tungsten becomes brittle and it is impossible to draw the thinnest wire for incandescent filaments from it.
The coal process also has one more drawback - high temperature: 1300 – 1400°C.

However, production with hydrogen reduction is also not a gift.
The reduction process takes place in special tube furnaces, heated in such a way that as it moves through the tube, the “boat” of WO3 passes through several temperature zones. A stream of dry hydrogen comes towards it. Recovery occurs in both “cold” (450...600°C) and “hot” (750...1100°C) zones; in “cold” ones – to the lower oxide WO 2, then – to the elemental metal. Depending on the temperature and duration of the reaction in the “hot” zone, the purity and grain size of the powdered tungsten released on the walls of the “boat” change.

So, we have obtained pure tungsten metal in the form of a tiny powder.
But this is not yet an ingot of metal from which something can be made. The metal is produced by powder metallurgy. That is, it is first pressed, sintered in a hydrogen atmosphere at a temperature of 1200-1300 °C, then passed through it electricity. The metal is heated to 3000 °C, and sintering occurs into a monolithic material.

However, we rather need not ingots or even rods, but thin tungsten wire.
As you yourself understand, here again everything is not so simple.
Wire drawing is carried out at a temperature of 1000°C at the beginning of the process and 400-600°C at the end. In this case, not only the wire, but also the die is heated. Heating is carried out by a gas burner flame or an electric heater.
In this case, after drawing, the tungsten wire is coated with graphite lubricant. The surface of the wire must be cleaned. Cleaning is carried out using annealing, chemical or electrolytic etching, and electrolytic polishing.

As you can see, the task of producing a simple tungsten filament is not as simple as it seems. And only the basic methods are described here; there are probably a lot of pitfalls there.
And, of course, even now tungsten is not a cheap metal. Now one kilogram of tungsten costs more than $50, the same molybdenum is almost two times cheaper.

Actually, there are several uses for tungsten.
Of course, the main ones are radio and electrical engineering, where tungsten wire goes.

The next one is the production of alloy steels, which are distinguished by their particular hardness, elasticity and strength. Added together with chromium to iron, it produces so-called high-speed steels, which retain their hardness and sharpness even when heated. They are used to make cutters, drills, milling cutters, as well as other cutting and drilling tools (in general, drilling tools contain a lot of tungsten).
Tungsten-rhenium alloys are interesting - they are used to make high-temperature thermocouples that operate at temperatures above 2000°C, although only in an inert environment.

Well, and one more thing interesting application- These are tungsten welding electrodes for electric welding. Such electrodes are non-consumable and it is necessary to supply additional metal wire to the welding site to provide a weld pool. Tungsten electrodes are used in argon arc welding - for welding non-ferrous metals such as molybdenum, titanium, nickel, as well as high-alloy steels.

As you can see, tungsten production is not for ancient times.
And why is tungsten there?
Tungsten can only be obtained with the construction of electrical engineering - with the help of electrical engineering and for electrical engineering.
No electricity means no tungsten, but you don’t need it either.

Tungsten minerals and ores

Of the tungsten minerals, the minerals of the wolframite and scheelite group are of practical importance.

Wolframite (xFeWO4 yMnWO4) is an isomorphic mixture of iron and manganese tungstates. If a mineral contains more than 80% iron, the mineral is called ferberite. If the mineral contains more than 80% manganese, then the mineral is called hubernite.

Scheelite CaWO4 is almost pure calcium tungstate.

Tungsten ores contain small amounts of tungsten. The minimum WO3 content at which their processing is advisable. is 0.14-0.15% for large deposits and 0.4-0.5% for small deposits. In ores, tungsten is accompanied by tin in the form of cassiterite, as well as the minerals molybdenum, bismuth, arsenic and copper. The main gangue rock is silica.

Tungsten ores undergo beneficiation. Wolframite ores are enriched using the gravity method, and scheelite ores are enriched by flotation.

Tungsten ore enrichment schemes are varied and complex. They combine gravitational enrichment with magnetic separation, flotation gravity and flotation. By combining various enrichment methods, concentrates containing up to 55-72% WO3 are obtained from ores. The extraction of tungsten from ore into concentrate is 82-90%.

The composition of tungsten concentrates varies within the following limits,%: WO3-40-72; MnO-0.008-18; SiO2-5-10; Mo-0.008-0.25; S-0.5-4; Sn-0.03-1.5; As-0.01-0.05; P-0.01-0.11; Cu-0.1-0.22.

Technological schemes for processing tungsten concentrates are divided into two groups: alkaline and acidic.

Methods for processing tungsten concentrates

Regardless of the method of processing wolframite and scheelite concentrates, the first stage of their processing is opening, which is the transformation of tungsten minerals into easily soluble chemical compounds.

Wolframite concentrates are opened by sintering or fusion with soda at a temperature of 800-900°C, which is based on chemical reactions:

4FeWO4 + 4Na2CO3 + O2 = 4Na2WO4 + 2Fe2O3 +4CO2 (1)

6MnWO4 + 6Na2CO3 + O2 = 6Na2WO4 + 2Mn3O4 +6CO2 (2)

When sintering scheelite concentrates at a temperature of 800-900°C, the following reactions occur:

CaWO4 + Na2CO3 = Na2WO4+ CaCO3 (3)

CaWO4 + Na2CO3 = Na2WO4+ CaO + CO2 (4)

In order to reduce soda consumption and prevent the formation of free calcium oxide, silica is added to the charge to bind calcium oxide into a sparingly soluble silicate:

2CaWO4 + 2Na2CO3 + SiO2 = 2Na2WO4+ Ca2SiO4 + CO2 (5)

Sintering of scheelite concentrate with soda and silica is carried out in drum furnaces at a temperature of 850-900°C.

The resulting cake (alloy) is leached with water. During leaching, sodium tungstate Na2WO4 and soluble impurities (Na2SiO3, Na2HPO4, Na2AsO4, Na2MoO4, Na2SO4) and excess soda pass into the solution. Leaching is carried out at a temperature of 80-90°C in steel reactors with mechanical stirring, operating in batch mode, or in continuous drum rotary kilns. The recovery of tungsten into the solution is 98-99%. The solution after leaching contains 150-200 g/l WO3. The solution is filtered, and after separating the solid residue, it is sent for purification from silicon, arsenic, phosphorus and molybdenum.

Purification from silicon is based on the hydrolytic decomposition of Na2SiO3 by boiling a solution neutralized at pH = 8-9. Neutralization of excess soda in the solution is carried out with hydrochloric acid. As a result of hydrolysis, slightly soluble silicic acid is formed:

Na2SiO3 + 2H2O = 2NaOH + H2SiO3 (6)

To remove phosphorus and arsenic, the method of precipitation of phosphate and arsenate ions in the form of poorly soluble ammonium-magnesium salts is used:

Na2HPO4 + MgCl2+ NH4OH = Mg(NH4)PO4 + 2NaCl + H2O (7)

Na2HAsO4 + MgCl2+ NH4OH = Mg(NH4)AsO4 + 2NaCl + H2O (8)

Purification from molybdenum is based on the decomposition of molybdenum sulfosalt, which is formed when sodium sulfide is added to a solution of sodium tungstate:

Na2MoO4 + 4NaHS = Na2MoS4 + 4NaOH (9)

Upon subsequent acidification of the solution to pH = 2.5-3.0, the sulfosalt is destroyed with the release of slightly soluble molybdenum trisulfide:

Na2MoS4 + 2HCl = MoS3 + 2NaCl + H2S (10)

Calcium tungstate is first precipitated from a purified solution of sodium tungstate using CaCl2:

Na2WO4 + CaCl2 = CaWO4 + 2NaCl. (eleven)

The reaction is carried out in a boiling solution containing 0.3-0.5% alkali

while stirring with a mechanical stirrer. The washed sediment of calcium tungstate in the form of a pulp or paste is subjected to decomposition with hydrochloric acid:

CaWO4 + 2HCl = H2WO4 + CaCl2 (12)

During decomposition, the high acidity of the pulp is maintained at about 90-120 g/l HCl, which ensures the separation of impurities of phosphorus, arsenic and partly molybdenum, which are soluble in hydrochloric acid, from the tungstic acid sediment.

Tungstic acid can also be obtained from a purified solution of sodium tungstate by direct precipitation with hydrochloric acid. When the solution is acidified with hydrochloric acid, H2WO4 precipitates as a result of hydrolysis of sodium tungstate:

Na2WO4 + 2H2O = 2NaOH + H2WO4 (11)

The alkali formed as a result of the hydrolysis reaction reacts with hydrochloric acid:

2NaOH + 2HCl = 2NaCl + 2H2O (12)

The addition of reactions (8.11) and (8.12) gives the total reaction of precipitation of tungstic acid with hydrochloric acid:

Na2WO4 + 2HCl = 2NaCl + H2WO4 (13)

However, in this case, great difficulties arise in washing the sediment from sodium ions. Therefore, at present, the latter method of tungstic acid deposition is used very rarely.

The technical tungstic acid obtained by precipitation contains impurities and therefore needs to be purified.

The most widely used method is the ammonia method for purifying technical tungsten acid. It is based on the fact that tungstic acid is highly soluble in ammonia solutions, while a significant part of the impurities it contains are insoluble in ammonia solutions:

H2WO4 + 2NH4OH = (NH4)2WO4 + 2H2O (14)

Ammonia solutions of tungstic acid may contain impurities of molybdenum and alkali metal salts.

Deeper cleaning is achieved by isolating large crystals of ammonium paratungstate from the ammonia solution, which are obtained by evaporating the solution:

12(NH4)2WO4 = (NH4)10W12O41 5H2O + 14NH3 + 2H2O (15)

tungsten acid anhydride precipitation

Deeper crystallization is impractical to avoid contamination of the crystals with impurities. From the mother liquor, enriched with impurities, tungsten is precipitated in the form of CaWO4 or H2WO4 and returned to the previous stages.

Paratungstate crystals are squeezed out on filters, then in a centrifuge, washed with cold water and dried.

Tungsten oxide WO3 is obtained by calcining tungstic acid or paratungstate in a rotating tubular furnace with a stainless steel pipe and heated by electricity at a temperature of 500-850oC:

H2WO4 = WO3 + H2O (16)

(NH4)10W12O41 5H2O = 12WO3 + 10NH3 +10H2O (17)

In tungsten trioxide intended for the production of tungsten, the WO3 content must be no lower than 99.95%, and for the production of hard alloys - no lower than 99.9%