Selection, justification and calculation of tungsten-molybdenum ore processing technology. How tungsten is obtained Technology of processing tungsten ores

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. Weak use of resource-saving technologies contributes to big losses minerals in the extraction and enrichment of raw materials.

An analysis of the development of technology and technology for mineral processing over the past 10-15 years indicates significant achievements of domestic fundamental science in the field of knowledge of the basic phenomena and patterns in the separation of mineral complexes, which makes it possible to create highly efficient processes and technologies for primary processing 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 total mass raw materials supplied for enrichment.

The human impact on the natural environment in the process of economic activity is now becoming global in nature. In terms of the scale of extracted and transported rocks, transformation of 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, the accumulation of large amounts of waste on the Earth’s surface, in the atmosphere and hydrosphere, the gradual degradation of natural landscapes, a reduction in biodiversity, and a decrease 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 water, and 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 the wide range of minerals they contain (from main and associated components to the simplest building materials) make it difficult to calculate the total economic effect of their processing and determine an individual approach to the assessment of each tailings.

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 of the 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 the scientific direction of the Department of Mineral Processing and Environmental Engineering of the 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 paper topic No. 118 "Study of 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.

Installed quantitative characteristics stale tailings of ore processing of ores from the Dzhida VMC in terms of WO3 and sulfide sulfur content. 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 general content sulfur). 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, burial conditions, contamination with household waste, WO3 content in them and the degree of sulfur oxidation, industrial interest can only be top part emergency discharge tailings with resources of 1.0 million tons of sand and 1330 tons of WO3 with a WO3 content of 0.126%. 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 south-eastern 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. Consequently, the chemical composition of 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 colors (from dark green to light yellow and gray) and different material compositions (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 the selective effect of various types of 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 studying contrast), separation methods and characteristics are selected, their effectiveness is assessed, theoretical separation indices are determined, and a basic diagram of radiometric enrichment is developed, 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 application of multi-stage combined schemes for enriching the poor tungsten ores that ensure 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 has recently been paid to improving the technology for extracting tungsten minerals from sludge in two main areas: gravitational enrichment of sludge in modern multi-deck concentrators (similar to the enrichment of tin-containing sludge) with subsequent finishing of the concentrate by flotation and enrichment in 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.

The technological scheme of mineral processing, including two or more devices, embodies everything character traits 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%. results spectral analysis representative sample of the resulting concentrate, shown 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 to extract WO3 from HTO with a particle size of -3+0.5 mm is 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

Magnetic methods are widely used in the beneficiation of ferrous, non-ferrous and rare metal ores and in other areas of industry, including food. They are used for the enrichment of iron, manganese, copper-nickel tungsten ores, as well as for finishing concentrates of rare metal ores, regeneration of ferromagnetic weighting agents in installations for separation in heavy suspensions, for removing iron impurities from quartz sands, pyrite from coal, etc.

All minerals differ in specific magnetic susceptibility and for extraction weakly magnetic minerals fields with high magnetic characteristics are required in the working area of ​​the separator.

In ores of rare metals, in particular tungsten and niobium and tantalum, the main minerals in the form of wolframite and columbite-tantalite have magnetic properties and it is possible to use high-gradient magnetic separation with the extraction of ore minerals into the magnetic fraction.

Tests of tungsten and niobium-tantalum ore from the Spoikoininskoye and Orlovskoye deposits were carried out in the laboratory of magnetic enrichment methods at NPO ERGA. For dry magnetic separation, a roller separator SMVI manufactured by NPO ERGA was used.

The separation of tungsten and niobium-tantalum ore took place according to scheme No. 1. The results are presented in the table.

Based on the results of the work, the following conclusions can be drawn:

The content of useful components in the separation tailings is: WO3 according to the first separation scheme - 0.031±0.011%, according to the second - 0.048±0.013%; Ta 2 O 5 and Nb 2 O 5 -0.005±0.003%. This suggests that the induction in the working area of ​​the separator is sufficient to extract weakly magnetic minerals into the magnetic fraction and a magnetic separator of the SMVI type is suitable for obtaining waste tailings.

Tests of the magnetic separator SMVI were also carried out on baddeleyite ore in order to extract weakly magnetic iron minerals (hematite) into the tailings and purify the zirconium concentrate.

The result of separation was a decrease in the iron content in the non-magnetic product from 5.39% to 0.63% with a recovery of 93%. The zirconium content in the concentrate increased by 12%.

The separator operation diagram is shown in Fig. 1

The use of the SMVI magnetic separator has found wide application in the beneficiation of various ores. SMVI can serve both as the main enrichment equipment and as a finishing device for concentrates. This is confirmed by successful pilot tests of this equipment.

Main enrichment

For some beneficiation factories, in pre-beneficiation, first Xinhai will use moving screen jigger, and then enter into finishing operations.

Gravity enrichment

For wolframite gravity technology, Xinhai usually uses a gravity process that includes multi-stage jigging, multi-stage table and middling product regrinding. That is, after fine crushing, worthy ores, which, through the classification of a vibrating screen, carry out multi-stage jigging and produce coarse sand from jigging and gravity. Then the ballast products of the large class jigging will enter the mill for additional grinding. And the ballast products of the small class jigging will enter sorting through the classifications multi-stage table, then coarse sand is produced from gravity and from the table, then the tailings from the table will enter the tailings hopper, the middlings from the table are then returned to the regrinding cycle stage, and the gravity coarse sand from the jig and the table enters the finishing operation.

Cleaning

In the wolframite finishing operation, a combined flotation and gravity enrichment technology or a combined flotation technology - gravity and magnetic enrichment is usually used. At the same time, returns the accompanying item.

The finishing operation usually uses a combined method of flotation and enrichment table and washing of sulfur pyrites through flotation. At the same time, we can enter into the flotation separation of sulfur pyrites. After this, wolframite concentrates are produced, if wolframite concentrates contain scheelite and cassiterite, then wolframite concentrates, scheelite concentrates and cassiterite concentrates are produced through a combined flotation and gravity enrichment technology or a combined gravitational and magnetic flotation technology enrichment.

Fine sludge treatment

The processing method for fine sludge in Xinhai is usually as follows: firstly, desulfurization is carried out, then, according to the properties of the fine sludge and material, gravity, flotation, magnetic and electrical enrichment technology is used, or a combined beneficiation technology of several technologies is used to return tungsten ore, and at the same time time will carry out the utilization of associated ore minerals.

Practical examples

The Xinhai wolframite object was taken as an example; the size distribution of the ore of this mine was inhomogeneous, and the ore was very heavily sludged. The initial technological scheme used by the beneficiation plant, which includes pre-concentration crushing, gravity and refining, due to a number of technological defects, resulted in huge losses of small-grade tungsten ores, high beneficiation costs, such as the poor state of comprehensive beneficiation indicators. in order to improve the sorting status of wolframite, this obage fabric authorized Xinhai to carry out technical reconstruction tasks. After careful research on the properties of ore and beneficiation technology of this factory, Xinhai optimized the technology for beneficiation of wolframite of this factory and added fine sludge processing technology. and ultimately obtain ideal enrichment rates. The enrichment indicator of the factory before and after the transformation is as follows:

After the transformation, the extraction of tungsten ore increased significantly. And mitigated the effects of fine sludge on the wolframite sorting process, achieved good recovery rate, effectively improved the economic efficiency of the factory.

Tungsten minerals, ores and concentrates

Tungsten is a rare element, its average content in earth's crust 10-4% (by mass). About 15 tungsten minerals are known, however practical significance only minerals of the wolframite and scheelite group have.

Wolframite (Fe, Mn)WO4 is an isomorphic mixture (solid solution) of iron and manganese tungstates. If the mineral contains more than 80% iron tungstate, the mineral is called ferberite; if manganese tungstate predominates (more than 80%), it is called hübnerite. Mixtures lying in composition between these limits are called wolframites. Minerals of the wolframite group are colored black or brown and have a high density (7D-7.9 g/cm3) and a hardness of 5-5.5 on the mineralogical scale. The mineral contains 76.3-76.8% W03. Wolframite is weakly magnetic.

Scheelite CaWOA is calcium tungstate. The color of the mineral is white, gray, yellow, brown. Density 5.9-6.1 g/cm3, hardness on the mineralogical scale 4.5-5. Scheelite often contains an isomorphic admixture of powellite - CaMoO4. When irradiated with ultraviolet rays, scheelite fluoresces with blue light. When the molybdenum content is more than 1%, the fluorescence becomes yellow. Scheelite is non-magnetic.

Tungsten ores are usually low in tungsten. The minimum W03 content in ores at which their exploitation is profitable is currently 0.14-0.15% for large deposits and 0.4-0.5% for small deposits.

Along with tungsten minerals, molybdenite, cassiterite, pyrite, arsenopyrite, chalcopyrite, tantalite or columbite, etc. are found in ores.

Based on the mineralogical composition, two types of deposits are distinguished - wolframite and scheelite, and based on the shape of ore formations - vein and contact types.

In vein deposits, tungsten minerals mostly occur in quartz veins small thickness (0.3-1 m). The contact type of deposits is associated with contact zones of granite rocks and limestones. They are characterized by deposits of sheelite-bearing skarn (skarns are silicified limestones). Skarn-type ores include the largest Tyrn-Auz deposit in the USSR in the North Caucasus. When vein deposits are weathered, wolframite and scheelite accumulate, forming placers. In the latter, wolframite is often combined with cassiterite.

Tungsten ores are enriched, producing standard concentrates containing 55-65% W03. A high degree of enrichment of wolframite ores is achieved using various methods: gravity, flotation, magnetic and electrostatic separation.

When enriching scheelite ores, gravity-flotation or pure flotation schemes are used.

The extraction of tungsten into standard concentrates during the enrichment of tungsten ores ranges from 65-70% to 85-90%.

When enriching ores of complex composition or difficult to enrich, it is sometimes economically advantageous to remove middling products containing 10-20% W03 from the enrichment cycle for chemical (hydrometallurgical) processing, which results in the production of “artificial scheelite” or technical tungsten trioxide. Such combined schemes ensure high extraction of tungsten from ores.

The state standard (GOST 213-73) provides for the content of W03 in tungsten concentrates of the 1st grade not lower than 65%, of the 2nd grade - not lower than 60%. The content of impurities P, S, As, Sn, Cu, Pb, Sb, Bi is limited in them, ranging from hundredths of a percent to 1.0%, depending on the grade and purpose of the concentrate.

The explored reserves of tungsten as of 1981 are estimated at 2903 thousand tons, of which 1360 thousand tons are in China. The USSR, Canada, Australia, the USA, South and North Korea, Bolivia, Brazil, and Portugal have significant reserves. Production of tungsten concentrates in capitalist and developing countries in the period 1971 - 1985. fluctuated between 20 - 25 thousand tons (in terms of metal content).

Methods for processing tungsten concentrates

The main product of direct processing of tungsten concentrates (in addition to ferrotungsten smelted for the needs of ferrous metallurgy) is tungsten trioxide. It serves as the starting material for tungsten and tungsten carbide - the main component of hard alloys.

Production schemes for processing tungsten concentrates are divided into two groups depending on the adopted decomposition method:

Tungsten concentrates are sintered with soda or treated with aqueous soda solutions in autoclaves. Tungsten concentrates are sometimes decomposed with aqueous solutions of sodium hydroxide.

Concentrates are decomposed with acids.

In cases where alkaline reagents are used for decomposition, solutions of sodium tungstate are obtained, from which, after purification from impurities, the final products are produced - ammonium paratungstate (PVA) or tungstic acid. 24

When the concentrate is decomposed with acids, a precipitate of technical tungstic acid is obtained, which is purified from impurities in subsequent operations.

Decomposition of tungsten concentrates. alkaline reagents Sintering with Na2C03

Sintering of wolframite with Na2C03. The interaction of wolframite with soda in the presence of oxygen actively occurs at 800-900 C and is described by the following reactions: 2FeW04 + 2Na2C03 + l/202 = 2Na2W04 + Fe203 + 2C02; (l) 3MnW04 + 3Na2C03 + l/202 = 3Na2W04 + Mn304 + 3C02. (2)

These reactions occur with a large decrease in the Gibbs energy and are practically irreversible. With the ratio in wolframite FeO:MnO = i:i AG°1001C = -260 kJ/mol. With an excess of Na2C03 in the charge of 10-15% above the stoichiometric amount, complete decomposition of the concentrate is achieved. To accelerate the oxidation of iron and manganese, 1-4% nitrate is sometimes added to the mixture.

Sintering of wolframite with Na2C03 at domestic enterprises is carried out in tubular rotary kilns lined with fireclay bricks. In order to avoid melting of the charge and the formation of accretions (accumulations) in zones of the furnace with a lower temperature, tailings from the leaching of cakes (containing iron and manganese oxides) are added to the charge, reducing the W03 content in it to 20-22%.

A furnace with a length of 20 m and an outer diameter of 2.2 m at a rotation speed of 0.4 rpm and an inclination angle of 3 has a charge capacity of 25 tons/day.

The components of the charge (crushed concentrate, Na2C03, saltpeter) are fed from bins into a screw mixer using automatic scales. The charge enters the furnace hopper, from which it is fed into the furnace. Upon exiting the furnace, cake pieces pass through crushing rolls and a wet grinding mill, from which the pulp is directed to a higher laminator (Fig. 1).

Sintering of scheelite with Na2C03. At temperatures of 800-900 C, the interaction of scheelite with Na2C03 can proceed through two reactions:

CaW04 + Na2CQ3 Na2W04 + CaC03; (1.3)

CaW04 + Na2C03 *=*■ Na2W04 + CaO + C02. (1.4)

Both reactions proceed with a relatively small change in the Gibbs energy.

Reaction (1.4) occurs to a noticeable extent above 850 C, when decomposition of CaCO3 is observed. The presence of calcium oxide in the cake leads, when the cake is leached with water, to the formation of slightly soluble calcium tungstate, which reduces the extraction of tungsten into the solution:

Na2W04 + Ca(OH)2 = CaW04 + 2NaOH. (1.5)

With a large excess of Na2C03 in the charge, this reaction is significantly suppressed by the interaction of Na2C04 with Ca(OH)2 with the formation of CaCO3.

To reduce the consumption of Na2C03 and prevent the formation of free calcium oxide, quartz sand is added to the charge to bind calcium oxide into poorly soluble silicates:

2CaW04 + 2Na2C03 + Si02 = 2Na2W04 + Ca2Si04 + 2C02;(l.6) AG°100IC = -106.5 kJ.

Still, in this case, to ensure a high degree of tungsten extraction into the solution, it is necessary to introduce a significant excess of Na2C03 into the charge (50-100% of the stoichiometric amount).

Sintering of the scheelite concentrate charge with Na2C03 and quartz sand is carried out in drum furnaces, as described above for wolframite at 850-900 °C. To prevent melting, leaching dumps (containing mainly calcium silicate) are added to the charge to reduce the W03 content to 20-22%.

Leaching of soda speco. When cakes are leached with water, sodium tungstate and soluble impurity salts (Na2Si03, Na2HP04, Na2HAs04, Na2Mo04, Na2S04), as well as excess Na2C03, pass into the solution. Leaching is carried out at 80-90 °C in steel reactors with mechanical stirring, operating in hierarchical conditions.

Concentrates with soda:

Elevator feeding concentrate to the mill; 2 - ball mill operating in a closed cycle with an air separator; 3 - auger; 4 - air separator; 5 - bag filter; 6 - automatic weighing dispensers; 7 - transport screw; 8 - screw mixer; 9 - charge hopper; 10 - feeder;

Drum oven; 12 - roll crusher; 13 - rod mill - lixiviant; 14 - reactor with stirrer

Wild mode, or drum rotating leaches of continuous operation. The latter are filled with crushing rods to crush pieces of cake.

The recovery of tungsten from the sinter into the solution is 98-99%. Strong solutions contain 150-200 g/l W03.

Autoclave is the only way to decompose tungsten concentrates

The autoclave-soda method was proposed and developed in the USSR1 in relation to the processing of scheelite concentrates and industrial products. Currently, the method is used at a number of domestic factories and in foreign countries.

The decomposition of scheelite with Na2C03 solutions is based on the exchange reaction

CaW04CrB)+Na2C03(pacTB)^Na2W04(pacTB)+CaC03(TB). (1.7)

At 200-225 °C and a corresponding excess of Na2C03, depending on the composition of the concentrate, decomposition proceeds with sufficient speed and completeness. The concentration equilibrium constants of reaction (1.7) are small, increase with temperature and depend on the soda equivalent (i.e., the number of moles of Na2C03 per 1 mole of CaW04).

With a soda equivalent of 1 and 2 at 225 C, the equilibrium constant (Kc = C / C cq) is 1.56 and

0.99 respectively. It follows from this that at 225 C the minimum required soda equivalent is 2 (i.e., the excess of Na2C03 is 100%). The real excess of Na2C03 is higher, since as equilibrium is approached the rate of the process slows down. For scheelite concentrates containing 45-55% W03 at 225 C, a soda equivalent of 2.6-3 is required. For industrial products containing 15-20% W03, 4-4.5 moles of Na2C03 per 1 mole of CaW04 are required.

The CaCO3 films formed on scheelite particles are porous and up to a thickness of 0.1-0.13 mm, their influence on the rate of decomposition of scheelite by Na2C03 solutions was not detected. With intense stirring, the rate of the process is determined by the rate of the chemical stage, which is confirmed by the high value of the apparent activation energy E = 75+84 kJ/mol. However, if the mixing speed is insufficient (which

Occurs in horizontal rotating autoclaves), an intermediate regime is realized: the rate of the process is determined by both the rate of supply of the reagent to the surface and the rate of chemical interaction.

0.2 0.3 0, it 0.5 0.5 0.7 0.8 Ш gШШУШгС031

As can be seen from Fig. 2, the specific reaction rate decreases approximately inversely with the increase in the ratio of molar concentrations of Na2W04:Na2C03 in solution. This

Cassock. Fig. 2. Dependence of the specific rate of decomposition of scheelite by soda solution in autoclave j on the molar ratio of Na2W04/Na2C03 concentrations in the solution at

Determines the need for a significant excess of Na2C03 against the minimum required, determined by the value of the equilibrium constant. To reduce the consumption of Na2C03, two-stage countercurrent leaching is carried out. In this case, the tailings after the first leaching, which contain little tungsten (15-20% of the original), are treated with a fresh solution containing a large excess of Na2C03. The resulting solution, which is recycled, enters the first stage of leaching.

Decomposition with Na2C03 solutions in autoclaves is also used for wolframite concentrates, but the reaction in this case is more complicated, as it is accompanied by hydrolytic decomposition of iron carbonate (manganese carbonate is only partially hydrolyzed). The decomposition of wolframite at 200-225 °C can be represented by the following reactions:

MnW04(TB)+Na2C03(paCT)^MiiC03(TB)+Na2W04(paCTB); (1.8)

FeW04(TB)+NaC03(pacT)*=iFeC03(TB)+Na2W04(paCTB); (1.9)

FeC03 + HjO^FeO + H2C03; (1.10)

Na2C03 + H2C03 = 2NaHC03. (l.ll)

The resulting iron oxide FeO at 200-225 °C undergoes a transformation according to the reaction:

3FeO + H20 = Fe304 + H2.

The formation of sodium bicarbonate leads to a decrease in the concentration of Na2C03 in the solution and requires a large excess of the reagent.

To achieve satisfactory decomposition rates of wolframite concentrates, it is necessary to finely grind them and increase the consumption of Na2C03 to 3.5-4.5 g-eq, depending on the composition of the concentrate. High-manganese wolframites are more difficult to decompose.

Adding NaOH or CaO to the autoclave pulp (which leads to causticization of Na2C03) improves the degree of decomposition.

The rate of decomposition of wolframite can be increased by introducing oxygen (air) into the autoclave pulp, which oxidizes Fe (II) and Mil (II), which leads to the destruction of the crystal lattice of the mineral on the reacting surface.

Secondary steam

Cassock. 3. Autoclave installation with a horizontally rotating autoclave: 1 - autoclave; 2 - loading pipe for pulp (steam is also introduced through it); 3 - pulp pump; 4 - pressure gauge; 5 - reactor-pulp heater; 6 - self-evaporator; 7 - droplet separator; 8 - pulp input into the self-evaporator; 9 - bumper made of armored steel; 10 - pipe for pulp removal; 11 - pulp collection

Leaching is carried out in steel horizontal rotating autoclaves heated with live steam (Fig. 3) and continuous vertical autoclaves with pulp mixing using bubbling steam. Approximate process mode: temperature 225 pressure in the autoclave ~2.5 MPa, T:L ratio = 1:(3.5*4), duration at each stage 2-4 hours.

Figure 4 shows a diagram of a battery of autoclaves. The initial autoclave pulp, heated by steam to 80-100 °C, is pumped into autoclaves, in which it is heated to

Secondary steam

Rve. 4. Scheme of a continuous autoclave installation: 1 - reactor for heating the initial pulp; 2 - piston pump; 3 - autoclave; 4 - throttle; 5 - self-evaporator; 6 - pulp collector

200-225 °C with live steam. During continuous operation, the pressure in the autoclave is maintained by releasing the pulp through a choke (a calibrated carbide washer). The pulp enters a self-evaporator - a vessel under a pressure of 0.15-0.2 MPa, where the pulp is rapidly cooled due to intense evaporation. The advantages of autoclave-soda decomposition of scheelite concentrates before sintering are the elimination of the furnace process and a slightly lower content of impurities in tungsten solutions (especially phosphorus and arsenic).

The disadvantages of this method include the high consumption of Na2C03. A high concentration of excess Na2C03 (80-120 g/l) entails an increased consumption of acids to neutralize solutions and, accordingly, high costs for the disposal of waste solutions.

Decomposition of tungstate conce n irate solutions and sodium hydroxide

Sodium hydroxide solutions decompose wolframite according to the exchange reaction:

Me WC>4 + 2Na0Hi=tNa2W04 + Me(0 H)2, (1.13)

Where Me is iron, manganese.

The value of the concentration constant of this reaction Kc = 2 at temperatures of 90, 120 and 150 °C is 0.68, respectively; 2.23 and 2.27.

Complete decomposition (98-99%) is achieved by treating the finely ground concentrate with a 25-40% sodium hydroxide solution at 110-120 °C. The required excess of alkali is 50% or higher. The decomposition is carried out in sealed steel reactors equipped with stirrers. Passing air into the solution accelerates the process due to the oxidation of iron (II) hydroxide Fe(OH)2 into hydrated iron (III) oxide Fe2O3-NH20 and manganese (II) hydroxide Mn(OH)2 into hydrated manganese oxide (IV) Mn02-lH20 .

The use of decomposition with alkali solutions is advisable only for high-grade wolframite concentrates (65-70% W02) with a small content of silica and silicates. When processing low-grade concentrates, highly contaminated solutions and difficult-to-filter sediments are obtained.

Processing of sodium tungstate solutions

Solutions of sodium tungstate containing 80-150 g/l W03, in order to obtain tungsten trioxide of the required purity, have so far been predominantly processed according to the traditional scheme, which includes: purification from compounds of impurity elements (Si, P, As, F, Mo); deposition

Calcium tungsten (artificial scheelite) followed by its decomposition with acids and the production of technical tungstic acid; dissolving tungstic acid in ammonia water, followed by evaporation of the solution and crystallization of ammonium paratungstate (PVA); calcination of PVA to obtain pure tungsten trioxide.

The main disadvantage of the scheme is that it is multi-stage, most operations are carried out in a periodic manner, and the duration of a number of stages. An extraction and ion exchange technology for converting Na2W04 solutions into (NH4)2W04 solutions has been developed and is already used at some enterprises. Below we briefly review the main stages of the traditional scheme and new extraction and ion exchange technology options.

Cleaning from impurities

Silicon removal. When the Si02 content in solutions exceeds 0.1% of the W03 content, preliminary removal of silicon is necessary. Purification is based on the hydrolytic decomposition of Na2Si03 by boiling a solution neutralized to pH = 8*9 with the release of silicic acid.

The solutions are neutralized with hydrochloric acid, which is added in a thin stream with stirring (to avoid local peroxidation) to the heated solution of sodium tungstate.

Purification from phosphorus and arsenic. To remove phosphate and arsenate ions, the method of precipitation of ammonium-magnesium salts Mg(NH4)P04 6H20 and Mg(NH4)AsC)4 6H20 is used. The solubility of these salts in water at 20 C is 0.058 and 0.038%, respectively. In the presence of excess Mg2+ and NH4 ions, solubility is lower.

Precipitation of phosphorus and arsenic impurities is carried out in the cold:

Na2HP04 + MgCl2 + NH4OH = Mg(NH4)P04 + 2NaCl +

Na2HAsQ4 + MgCl2 + NH4OH = Mg(NH4)AsQ4 + 2NaCl +

After standing for a long time (48 hours), crystalline precipitates of ammonium-magnesium salts fall out of the solution.

Purification from fluoride ions. With a high fluorite content in the initial concentrate, the content of fluoride ions reaches 5 g/l. Solutions are purified from fluoride ions by precipitation with magnesium fluoride from a neutralized solution to which MgCl2 is added. Fluorine removal can be combined with hydrolytic separation of silicic acid.

Molybdenum removal. Solutions of sodium tungstate must be cleaned of molybdenum if its content exceeds 0.1% of the W03 content (i.e. 0.1-0.2 t/l). At a molybdenum concentration of 5-10 g/l ( for example, when processing scheelite-powellite Tyrny-Auz concentrates), the release of molybdenum becomes special meaning, since it aims to obtain molybdenum chemical concentrate.

A common method is the precipitation of poorly soluble molybdenum trisulfide MoS3 from solution.

It is known that when sodium sulphide is added to solutions of sodium tungstate or molybdate, sulfosalts Na23S4 or oxosulfosalts Na23Sx04_x (where E is Mo or W) are formed:

Na2304 + 4NaHS = Na23S4 + 4NaOH. (1.16)

The equilibrium constant of this reaction for Na2Mo04 is significantly greater than for Na2W04(^^0 » Kzg). Therefore, if an amount of Na2S is added to the solution only sufficient to react with Na2Mo04 (with a slight excess), then molybdenum sulfosalt is predominantly formed. Upon subsequent acidification of the solution to pH = 2.5 * 3.0, the sulfosalt is destroyed with the release of molybdenum trisulfide:

Na2MoS4 + 2HC1 = MoS3 j + 2NaCl + H2S. (1.17)

Oxosulfosalts decompose with the release of oxosulfides (for example, MoSjO, etc.). Together with molybdenum trisulfide, a certain amount of tungsten trisulfide is coprecipitated. By dissolving the sulfide precipitate in a soda solution and re-precipitating molybdenum trisulfide, a molybdenum concentrate is obtained with a W03 content of no more than 2% with a loss of tungsten of 0.3-0.5% of the original amount.

After partial oxidative roasting of the precipitate - molybdenum trisulfide (at 450-500 °C) a molybdenum chemical concentrate containing 50-52% molybdenum is obtained.

The disadvantage of the method of deposition of molybdenum in the composition of trisulfide is the release of hydrogen sulfide by reaction (1.17), which requires costs for gas neutralization (they use the absorption of H2S in a scrubber irrigated with a solution of sodium hydroxide). Isolation of molybdenum trisulfide is carried out from a solution heated to 75-80 C. The operation is carried out in sealed steel reactors, rubberized or coated with acid-resistant enamel. Trisulfide precipitates are separated from the solution by filtration on a filter press.

Preparation of tungstic acid from solutions of sodium tungstate

Tungstic acid can be directly isolated from a solution of sodium tungstate with hydrochloric or nitric acids. However, this method is rarely used due to the difficulties of washing sediments from sodium ions, the content of which in tungsten trioxide is limited.

For the most part, calcium tungstate is initially precipitated from solution, which is then decomposed by acids. Calcium tungstate is precipitated by adding a CaC12 solution to a sodium tungstate solution heated to 80-90 C at a residual alkalinity of the solution of 0.3-0.7%. In this case, a white, finely crystalline, easily settled precipitate falls out; sodium ions remain in the mother solution, which ensures their low content in tungstic acid. 99-99.5% W is precipitated from the solution; mother liquors contain 0.05-0.07 g/l W03. The CaW04 precipitate washed with water in the form of a paste or pulp is sent for decomposition with hydrochloric acid when heated to 90°:

CaW04 + 2HC1 = H2W04i + CaCl2. (1.18)

During decomposition, the final acidity of the pulp is maintained high (90-100 g/l HCI), which ensures the separation of tungstic acid from impurities of phosphorus, arsenic and partly molybdenum compounds (molybdic acid dissolves in hydrochloric acid). Tungstic acid deposits require careful washing to remove impurities (especially calcium salts).

and sodium). In recent years, continuous washing of tungstic acid in pulsation columns has been developed, which has significantly simplified the operation.

At one of the enterprises in the USSR, when processing solutions of sodium tungstate, instead of hydrochloric acid, nitric acid is used to neutralize solutions and decompose CaW04 precipitates, and the precipitation of the latter is carried out by introducing Ca(N03)2 into solutions. In this case, nitrate mother liquors are utilized to obtain nitrate salts used as fertilizer.

Purification of technical tungstic acid and production of W03

Technical tungstic acid obtained by the method described above contains 0.2-0.3% impurities. As a result of acid calcination at 500-600 C, tungsten trioxide is obtained, suitable for the production of hard alloys based on tungsten carbide. However, for the production of tungsten, trioxide of higher purity is required with a total impurity content of no more than 0.05%.

The ammonia method for purifying tungstic acid is generally accepted. It easily dissolves in ammonia water, while most of the impurities remain in the sediment: silica, iron and manganese hydroxides and calcium (in the form of CaW04). However, ammonia solutions may contain an admixture of molybdenum and alkali metal salts.

A crystalline precipitate of PVA is isolated from the ammonia solution as a result of evaporation and subsequent cooling:

Evaporation

12(NH4)2W04 * (NH4)10H2W12O42 4H20 + 14NH3 +

In industrial practice, the composition of PVA is often written in oxide form: 5(NH4)20-12W03-5H20, which does not reflect its chemical nature as an isopolyacid salt.

Evaporation is carried out in periodic or continuous devices made of stainless steel. Typically, 75-80% tungsten is separated into crystals. It is undesirable to carry out deeper crystallization to avoid contamination of the crystals with impurities. It is significant that most of the molybdenum impurity (70-80%) remains in the mother solution. From the mother liquor, enriched with impurities, tungsten is precipitated in the form of CaW04 or H2W04, which is returned to the appropriate stages of the production scheme.

PVA crystals are squeezed out on a filter, then in a centrifuge, washed cold water and dry.

Tungsten trioxide is obtained by thermal decomposition of tungstic acid or PVA:

H2W04 = "W03 + H20;

(NH4)10H2W12O42 4H20 = 12W03 + 10NH3 + 10H20. (1.20)

Calcination is carried out in rotating electric furnaces with a pipe made of heat-resistant steel 20Х23Н18. The calcination mode depends on the purpose of tungsten trioxide and the required size of its particles. Thus, to obtain VA tungsten wire (see below), PVA is calcined at 500-550 °C, HF and VT grade wires (tungsten without additives) - at 800-850 °C.

Tungstic acid is calcined at 750-850 °C. Tungsten trioxide made from PVA has larger particles than trioxide made from tungstic acid. In tungsten trioxide intended for the production of tungsten, the W03 content must be at least 99.95%; for the production of hard alloys - at least 99.9%.

Extraction and ion exchange methods for processing sodium tungstate solutions

The processing of sodium tungstate solutions is significantly simplified by extracting tungsten from solutions by extraction with an organic extractant, followed by re-extraction from the organic phase with an ammonia solution with the separation of PVA from the ammonia solution.

Since tungsten is found in solutions in the form of polymer anions in a wide range of pH = 7.5 + 2.0, anion-exchange extractants are used for extraction: salts of amines or quaternary ammonium bases. In particular, in industrial practice, trioctylamine sulfate salt (i?3NH)HS04 (where R is C8H17) is used. The highest rates of tungsten extraction are observed at pH=2*4.

Extraction is described by the equation:

4(i?3NH)HS04(opr) + Н2\У120*"(aq) + 2Н+(aq)ї=ї

Ї=ї(Д3ГШ)4Н4\У12О40(org) + 4Н80;(aq). (l.2l)

The amine is dissolved in kerosene, to which a technical mixture of polyhydric alcohols (C7 - C9) is added to prevent the release of the solid phase (due to the low solubility of amine salts in kerosene). Approximate composition of the organic phase: amines 10%, alcohols 15%, kerosene - the rest.

Solutions purified from m-libdenum, as well as impurities of phosphorus, arsenic, silicon and fluorine are sent for extraction.

Tungsten is re-extracted from the organic phase with ammonia water (3-4% NH3), obtaining solutions of ammonium tungstate, from which PVA is isolated by evaporation and crystallization. Extraction is carried out in mixer-settler type devices or in pulsation columns with packing.

The advantages of extraction processing of sodium tungstate solutions are obvious: the number of operations in the technological scheme is reduced, the possibility of carrying out a continuous process for obtaining ammonium tungstate solutions from sodium tungstate solutions is created, and production space is reduced.

Extraction wastewater may contain an admixture of 80-100 mg/l of amines, as well as admixtures of higher alcohols and kerosene. To remove these environmentally harmful impurities, foam flotation and adsorption on activated carbon are used.

Extraction technology is used at foreign enterprises and is also implemented at domestic factories.

The use of ion exchange resins is a competing direction with extraction in the scheme for processing sodium tungstate solutions. For this purpose, low-basic anion exchangers containing amine groups (usually tertiary amines) or amphoteric resins (ampholytes) containing carboxyl and amine groups are used. At pH = 2.5 + 3.5, tungsten polyanions are sorbed on resins, and for some resins full capacity is 1700-1900 mg W03 per 1 g of resin. In the case of resin in the 8C>5~ form, sorption and elution are described respectively by the equations:

2tf2S04 + H4W12044; 5^«4H4W12O40 + 2SOf; (1.22)

I?4H4WI2O40 + 24NH4OH = 12(NH4)2W04 + 4DON + 12H20. (l.23)

The ion exchange method was developed and applied at one of the USSR enterprises. The required contact time of the resin with the solution is 8-12 hours. The process is carried out in a cascade of ion exchange columns with a suspended layer of resin in continuous mode. A difficult circumstance is the partial separation of PVA crystals at the elution stage, which requires their separation from the resin particles. As a result of elution, solutions containing 150-170 g/l W03 are obtained, which are sent to the evaporation and crystallization of PVA.

The disadvantage of ion exchange technology compared to extraction is unfavorable kinetics (contact duration 8-12 hours versus 5-10 minutes for extraction). At the same time, the advantages of ion exchangers include the absence of waste solutions containing organic impurities, as well as the fire safety and non-toxicity of resins.

Decomposition of scheelite concentrates by acids

In industrial practice, mainly when processing high-grade scheelite concentrates (70-75% W03), direct decomposition of scheelite with hydrochloric acid is used.

Decomposition reaction:

CaW04 + 2HC1 = W03H20 + CoCl2 (1.24)

Almost irreversible. However, the acid consumption is significantly higher than the stoichiometrically required one (250-300%) due to the inhibition of the process by films of tungstic acid on scheelite particles.

The decomposition is carried out in sealed reactors with stirrers, lined with acid-resistant enamel and heated through a steam jacket. The process is carried out at 100-110 C. The duration of decomposition varies from 4-6 to 12 hours, which depends on the degree of grinding, as well as the origin of the concentrate (scheelites from different deposits differ in reactivity).

A single treatment does not always lead to complete opening. In this case, after dissolving tungstic acid in ammonia water, the residue is re-treated with hydrochloric acid.

During the decomposition of scheelite-powellite concentrates containing 4-5% molybdenum, most of the molybdenum passes into the hydrochloric acid solution, which is explained by the high solubility of molybdic acid in hydrochloric acid. Thus, at 20 C in 270 g/l HC1, the solubilities of H2Mo04 and H2W04 are 182 and 0.03 g/l, respectively. Despite this, complete separation of molybdenum is not achieved. Tungstic acid precipitates contain 0.2-0.3% molybdenum, which cannot be extracted by repeated treatment with hydrochloric acid.

The acid method differs from the alkaline methods of scheelite decomposition in a smaller number of operations in the technological scheme. However, when processing concentrates with a relatively low content of W03 (50-55%) with a significant content of impurities, to obtain standard paravol-ammonium framate, it is necessary to carry out two or three ammonia purifications of tungstic acid, which is uneconomical. Therefore, decomposition with hydrochloric acid is mostly used in the processing of rich and pure scheelite concentrates.

The disadvantages of the decomposition method with hydrochloric acid are the high consumption of acid, the large volume of waste solutions of calcium chloride and the complexity of their disposal.

In light of the challenges of creating waste-free technologies, the nitrate method of decomposition of scheelite concentrates is of interest. In this case, mother solutions can be easily disposed of to obtain nitrate salts.

Tungsten ores in our country were processed at large mining and processing plants (Orlovsky, Lermontovsky, Tyrnauzsky, Primorsky, Dzhidinsky VMK) according to classic technological schemes with multi-stage grinding and enrichment of the material, divided into narrow size classes, usually in two cycles: primary gravity enrichment and finishing of rough concentrates using various methods. This is explained by the low tungsten content in the processed ores (0.1-0.8% WO3) and high requirements for the quality of concentrates. Primary enrichment for coarsely disseminated ores (minus 12+6 mm) was carried out by jigging, and for medium-, finely and finely disseminated ores (minus 2+0.04 mm) screw devices of various modifications and sizes were used.

In 2001, the Dzhidinsky tungsten-molybdenum plant (Buryatia, Zakamensk) ceased its activities, having accumulated a multimillion-dollar volume of sand in the Barun-Narynskoye technogenic tungsten deposit. Since 2011, this deposit has been processed by ZAO Zakamensk at a modular processing plant.

The technological scheme was based on enrichment in two stages on Knelson centrifugal concentrators (CVD-42 for the main operation and CVD-20 for cleaning), additional grinding of middlings and flotation of the collective gravity concentrate to produce KVGF grade concentrate. During operation, a number of factors were noted in the operation of Knelson concentrators that negatively affected the economic performance of sand processing, namely:

High operating costs, incl. energy costs and the cost of spare parts, which, given the remoteness of production from generating facilities and the increased cost of electricity, this factor becomes especially important;

Low degree of extraction of tungsten minerals into gravity concentrate (about 60% from the operation);

The complexity of this equipment in operation: when the material composition of the enriched raw material fluctuates, centrifugal concentrators require intervention in the process and prompt adjustment (changes in the pressure of the burning water, the rotation speed of the enrichment bowl), which leads to fluctuations in the quality characteristics of the resulting gravity concentrates;

Considerable distance from the manufacturer and, as a result, long waiting times for spare parts.

Looking for alternative method gravitational concentration, the Spirit company conducted laboratory tests of the technology screw separation using industrial screw separators SVM-750 and SVSh-750 produced by PC Spirit LLC. Enrichment took place in two operations: main and control, producing three enrichment products - concentrate, middlings and tailings. All enrichment products obtained as a result of the experiment were analyzed in the laboratory of JSC Zakamensk. The best results are presented in table. 1.

Table 1. Results of screw separation in laboratory conditions

The data obtained showed the possibility of using screw separators instead of Knelson concentrators in the primary enrichment operation.

The next stage was to conduct pilot tests on the existing enrichment circuit. An experimental semi-industrial installation was installed with screw devices SVSh-2-750, which were installed in parallel with Knelson CVD-42 concentrators. Enrichment was carried out in one operation, the resulting products were sent further according to the scheme of the existing enrichment plant, and sampling was carried out directly from the enrichment process without stopping the operation of the equipment. The indicators of pilot tests are presented in table. 2.

Table 2. Results of comparative pilot tests of screw devices and centrifugal concentratorsKnelson

The results show that sand enrichment occurs more efficiently using screw devices than centrifugal concentrators. This translates into lower concentrate yield (16.87% vs. 32.26%) with increased recovery (83.13% vs. 67.74%) in the tungsten mineral concentrate. This results in a higher quality WO3 concentrate (0.9% versus 0.42%),