Maintenance of the main method of enrichment of tungsten ores and the use of auxiliary dehydration processes in the technological scheme approx. Industrial production of tungsten The main ores of tungsten and their enrichment
Introduction
1 . 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. The need to use
1.3. Technogenic mineral formation of the Dzhida VMK
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 of the Dzhida VMC
2.1. Geological sampling and evaluation of tungsten distribution
2.2. The 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 initial size
2.3.3. Gravity analysis
2.3.4. Magnetic analysis
3. Development of a technological scheme
3.1. Technological testing of different gravity devices during the enrichment of stale tailings of various sizes
3.2. Optimization of the GR processing scheme
3.3. Semi-industrial testing of the developed technological scheme for the enrichment of general relativity and industrial plant
Introduction to work
Mineral enrichment sciences are primarily aimed at developing the theoretical foundations of mineral separation processes and creating enrichment apparatuses, at revealing the relationship between the distribution patterns 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 reduction in resource consumption in recent years, the depletion of mineral resources is one of the most important problems in Russia. Weak 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 of domestic fundamental science in the field of understanding 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 consequently, to 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 consumption, 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 needs of the industry for a particular metal, which led to the irrational use of natural mineral resources and an increase in the cost of 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, starting from 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 extraction of non-ferrous and ferrous metal ores, mining and chemical raw materials, deterioration in the quality of processed ores and, as a result, the involvement in processing of refractory 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 refractory 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 of economic activity is now becoming global. In terms of the scale of extracted and transported rocks, the transformation of the relief, the impact on the redistribution and dynamics of surface and groundwater, the activation of geochemical transport, etc. this activity is comparable to geological processes.
The unprecedented scale of recoverable mineral resources leads to their rapid depletion, the accumulation of a large amount of waste on the Earth's surface, in the atmosphere and hydrosphere, the gradual degradation of natural landscapes, the reduction of biodiversity, the decrease in the natural potential of territories and their life-supporting functions.
Waste storage facilities for ore processing are objects of increased environmental hazard due to their negative impact on the air basin, underground and surface waters, and soil cover over vast areas. Along with this, tailings are poorly explored man-made deposits, the use of which will provide additional
sources of ore and mineral raw materials with a significant reduction in the scale of disturbance of the geological environment in the region.
The 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 the 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 assessing each tailing.
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 in the processing of refractory ores and man-made deposits, the environmentally aggravated situation in the mining regions and the state of technology, technology and organization of the primary processing of mineral raw materials.
The issues of using wastes from the enrichment of polymetallic, gold-bearing and rare metals have both economic and environmental aspects.
V.A. Chanturia, V.Z. Kozin, V.M. Avdokhin, SB. Leonov, L.A. Barsky, A.A. Abramov, V.I. Karmazin, S.I. Mitrofanov and others.
An important part of the overall strategy of the mining industry, incl. tungsten, is the growth in the use of ore processing waste as additional sources of ore and mineral raw materials, with a significant reduction in the extent 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 is a detailed mineralogical and technological study of each specific,
individual technogenic deposit, the results of which will allow developing 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 Engineering Ecology 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 its integrated use, taking into account environmental problems in complex industrial systems ” and the film theme No. 118 “Research on the washability of stale tailings of the Dzhida VMK”.
Objective- scientifically substantiate, develop and test
rational technological methods of enrichment of stale
The following tasks were solved in the work:
Estimate the distribution of tungsten over the entire space of the main
technogenic formation of the Dzhida VMK;
to study the material composition of the stale tailings of the Dzhizhinsky VMK;
to investigate the contrast of stale tailings in the original size according to the content of W and S (II);
to investigate the gravitational washability of the stale tailings of the Dzhida VMK 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 from the OTO of the Dzhida VMK;
to conduct semi-industrial tests of the developed scheme for extracting W from stale tailings of the FESCO;
To develop a scheme of a chain of apparatus for the industrial processing of stale tailings of the Dzhida VMK.
To perform the research, a representative technological sample of stale tailings of the Dzhida VMK was used.
When solving the formulated problems, the following research methods: spectral, optical, chemical, mineralogical, phase, gravitational and magnetic methods for analyzing the material composition and technological properties of the initial mineral raw materials and enrichment products.
The following are defended main scientific provisions:
The patterns of distribution of the initial technogenic mineral raw materials and tungsten by size classes are established. The necessity of primary (preliminary) classification by size 3 mm is proved.
Quantitative characteristics of stale tailings of ore-dressing of ores of the Dzhida VMK have been established in terms of the content of WO3 and sulfide sulfur. It is proved that the original mineral raw materials belong to the category of non-contrast ores. A significant and reliable correlation between the contents of WO3 and S (II) was revealed.
Quantitative patterns of gravitational enrichment of stale tailings of the Dzhida VMK have been established. It has been proven that for the source material of any size, an effective method for extracting W is gravity enrichment. Predictive technological indicators of gravitational enrichment of initial mineral raw materials are determined in different size.
Quantitative regularities in the distribution of stale tailings of the Dzhida VMK ore concentration by fractions of different specific magnetic susceptibility have been established. The successive use of magnetic and centrifugal separation has been proven to improve the quality of crude W-containing products. Technological modes of magnetic separation have been optimized.
The material composition of mineral raw materials
When examining a secondary tailing dump (emergency dump tailing dump (HAS)) 35 furrow samples were taken from the pits and strippings along the slopes of the dumps; the total length of the furrows is 46 m. The pits and strippings are located in 6 exploration lines, spaced 40-100 m apart from each other; the distance between the pits (cleanings) in the exploration lines is from 30-40 to 100-150 m. All lithological varieties of sands have been tested. The samples were analyzed for the content of W03 and S (II) . In this area, 13 samples were taken from pits 1.0 m deep. 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 the content of WO3 and sulfur are given in Table. 2.1. Table 2.1 - The content of WO3 and sulfide sulfur in private samples of XAS It can be seen that the content of WO3 varies between 0.05-0.09%, with the exception of sample M-16, taken 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 tailing dump, 11 samples of various lithological differences were taken. The content of WO3 and S (II), depending on the origin of the sands, varies in a wide range: from 0.09 to 0.29% and from 0.78 to 5.8%, respectively. Elevated WO3 contents are characteristic of medium-coarse-grained sand varieties. The content of S (VI) is 80 - 82% of the total content of S, but in some samples, mainly with low contents of tungsten trioxide and total sulfur, it decreases to 30%.
The reserves of the deposit can be estimated as resources of category Pj (see Table 2.2). In the upper part of the length of the pit, they vary in a wide range: from 0.7 to 9.0 m, so the average content of controlled components is calculated taking into account the parameters of the pits. In our opinion, based on the above characteristics, taking into account the composition of stale tailings, their safety, conditions of occurrence, contamination with household waste, the content of WO3 in them and the degree of sulfur oxidation, only the upper part of the emergency tailing dump with resources of 1.0 million tons of sands and 1330 tons of WO3 with a WO3 content of 0.126%. Their location in close proximity to the projected processing plant (250-300 m) favors their transportation. The lower part of the emergency tailing dump is to be disposed of as part of the environmental rehabilitation program for the city of Zakamensk.
5 samples were taken on the deposit area. The interval between sampling points is 1000-1250 m. Samples were taken for the entire thickness of the layer, analyzed for the content of WO3, Ptot and S (II) (see Table 2.3). Table 2.3 - The content of WO3 and sulfur in individual ATO samples From the results of the analyzes it can be seen that the content of WO3 is low, varies from 0.04 to 0.10%. The average content of S (II) is 0.12% and is of no practical interest. The work carried out does not allow us to consider the secondary alluvial tailing dump as a potential industrial facility. However, as a source of environmental pollution, these formations are subject to disposal. The main tailing dump (MTF) has been explored along parallel exploration lines oriented along the azimuth of 120 and located 160 - 180 m apart. 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, the bulldozer passed trenches to a depth of 3-5 m, from which pits were driven 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 driven through 20 - 50 m in the central part of the deposit and after 100 m - on the southeastern 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 NTO along the distribution border is 1015 thousand m2 (101.5 ha); along the long axis (along the valley of the river Barun-Naryn) it is extended for 1580 m, in the transverse direction (near the dam) its width is 1050 m. Consequently, one pit illuminates an area of 12850 m, which is equivalent to an average network of 130x100 m. all workings); the area of the exploration network averaged 90x100 m2. On the extreme southeastern flank, at the site of a former settling pond in the area of development of fine-grained sediments - silts, 12 pits (15% of the total) 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 uneven-grained 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 passed (10% of the number of workings in the technogenic deposit) and the average area of the exploration network was 145x100 m. of the tested section at the man-caused deposit is 4.3 m, including on uneven-grained sands -5.2 m, silty sands -2.1 m, silts -1.3 m. - 1115 m near the upper part of the dam, up to 1146 - 148 m in the central part and up to 1130-1135 m on the southeastern flank. In total, 60 - 65% of the capacity of the technogenic deposit has been tested. Trenches, pits, clearings and digs are documented in M 1:50 -1:100 and tested with a furrow with a section of 0.1x0.05 m2 (1999) and 0.05x0.05 m2 (2000). The length of furrow samples was 1 m, weight 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 detailing areas and individual sections outside the network, it was 459 m. The mass of the samples taken was 5 tons.
The samples together with the passport (breed characteristic, sample number, production and performer) were packed in polyethylene and then cloth 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 the methods of NS AM. The correctness of the analyzes was 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 individual technological samples taken at the OTO are given in Appendix 1. The main (OTO) and two side tailings (KhAT and ATO) of the Dzhida VMK were statistically compared in terms of WO3 content using Student's t-test (see Appendix 2) . With a confidence level of 95%, the following was established: - no significant statistical difference in WO3 content between private samples of side tailings; - average results of OTO sampling in terms of WO3 content in 1999 and 2000. belong to the same general population. Consequently, the chemical composition of the main tailing dump changes insignificantly over time under the influence of external influences. All stocks of GRT can be processed using a single technology.; - the average results of testing the main and secondary tailings in terms of WO3 content significantly differ from each other. Therefore, the development of a local enrichment technology is required to involve minerals from side tailings.
Technological properties of mineral raw materials
According to the granular composition, the sediments are divided into three types of sediments: inequigranular sands; silty sands (silty); silts. There are gradual transitions between these types of precipitation. More distinct boundaries are observed in the thickness of the section. They are caused by the alternation of sediments of different size composition, different colors (from dark green to light yellow and gray) and different material composition (quartz-feldspar non-metallic part and sulfide with magnetite, hematite, hydroxides of iron and manganese). The entire sequence is layered - from finely to coarsely layered; the latter is more characteristic of coarse-grained deposits or interlayers of essentially sulfide mineralization. Fine-grained (silty, silty fractions, or layers composed of dark-colored - amphibole, hematite, goethite) usually form thin (the first cm - mm) layers. The occurrence of the entire sequence of sediments is subhorizontal with a predominant dip of 1-5 in the northern points. Inequigranular sands are located in the northwestern and central parts of the OTO, which is due to their sedimentation near the source of discharge - the pulp conduit. The width of the strip of uneven-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 grain composition is variable - from fine-grained to coarse-grained varieties up to gravelstone lenses with a thickness of 5-20 cm and a length of up to 10-15 m. Silty (silty) sands stand out in the form of a layer 7-10 m thick (horizontal thickness, outcrop 110-120 m ). They lie under uneven-grained sands. In the section, they are a layered stratum of gray, greenish-gray color with alternating fine-grained sands with interlayers 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 compose the southeastern part of the OTO and are represented by finer particles of enrichment wastes of dark gray, dark green, bluish-green color with interlayers of grayish-yellow sands. The main feature of their structure is a more uniform, more massive texture with less pronounced and less clearly expressed 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 classes are given in Table. 2.8. According to the granulometric analysis, the bulk of the OTO sample material (about 58%) has a particle size of -1 + 0.25 mm, 17% each fall into large (-3 + 1 mm) and small (-0.25 + 0.1) mm classes. The proportion of material with a particle size of less than 0.1 mm is about 8%, of which half (4.13%) falls on the sludge class -0.044 + 0 mm. Tungsten is characterized by a slight fluctuation in the 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 the size class -0 .1+0.044 mm. In the slime class -0.044+0 mm, the tungsten content is reduced to 0.19%. Huebnerite accumulation occurs only in small-sized material, that is, in the -0.1 + 0.044 mm class. 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 mineral raw materials OTO by size classes and histograms of the absolute and relative distribution of W by size classes of mineral raw materials OTO are shown in Fig. 2.2. and 2.3. In table. 2.9 shows data on impregnation of hubnerite and scheelite in mineral raw materials OTO of initial size and crushed to - 0.5 mm.
In the class -5 + 3 mm of the original mineral raw material, there are no grains of pobnerite and scheelite, as well as intergrowths. In the -3+1 mm class, the content of free grains of scheelite and hübnerite is quite high (37.2% and 36.1%, respectively). In the -1 + 0.5 mm class, both mineral forms of tungsten are present in almost equal amounts, 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 boundary and the content of free grains of scheelite and hübnerite and their intergrowths is practically the same in it. Based on the data in Table. 2.9, it can be concluded that it is necessary to classify the deslimed mineral raw materials OTO according to the size of 0.1 mm and separate enrichment of the resulting classes. From a large class, it is necessary to separate free grains into a concentrate, and tailings containing intergrowths must be subjected to regrinding. Crushed and de-sludged tailings should be combined with de-sludged grade -0.1+0.044 of the original mineral raw materials and sent to gravity operation II in order to extract fine grains of scheelite and pobnerite into middlings.
2.3.2 Study of the possibility of radiometric separation of mineral raw materials in the initial size Radiometric separation is a process of large-sized 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. More than twenty methods of radiometric enrichment are known; the most promising of them are X-ray radiometric, X-ray luminescent, radio resonance, photometric, autoradiometric and neutron absorption. With the help of radiometric methods, the following technological problems are solved: preliminary enrichment with the removal of waste rock from the ore; selection of technological varieties, varieties with subsequent enrichment according to separate schemes; isolation of products suitable for chemical and metallurgical processing. The assessment of radiometric washability includes two stages: the study of the properties of ores and the experimental determination of the technological parameters of enrichment. At the first stage, the following main properties are studied: the content of valuable and harmful components, particle size distribution, single- and multi-component contrast of the ore. At this stage, the fundamental possibility of using radiometric enrichment is established, the limiting separation indicators are determined (at the contrast study stage), separation methods and signs are selected, their effectiveness is evaluated, theoretical separation indicators are determined, and a schematic diagram of radiometric enrichment is developed, taking into account the specifics of the subsequent processing technology. At the second stage, the modes and practical results of separation are determined, enlarged laboratory tests of the radiometric enrichment scheme are carried out, a rational version of the scheme is selected 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 case, the mass, size and number of technological samples are set depending on the properties of the ore, the structural features of the deposit and the 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 the method of radiometric enrichment 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 GR processing scheme
In connection with the involvement 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 in San Francisco was devoted to the problems of improving the technology of enrichment of low-grade ores. An analysis of the technological schemes of operating enterprises showed that various methods of preliminary concentration have become widespread in ore preparation: photometric sorting, preliminary jigging, enrichment in heavy media, wet and dry magnetic separation. In particular, photometric sorting is effectively used at one of the largest suppliers of tungsten products - at Mount Corbine in Australia, which processes ores with a tungsten content of 0.09% at large Chinese factories - Taishan and Xihuashan.
For preliminary concentration of ore components in heavy media, highly efficient Dinavirpul devices from Sala (Sweden) are used. According to this technology, the material is classified and the +0.5 mm class is enriched in a heavy medium, represented by a mixture of ferrosilicon. Some factories use dry and wet magnetic separation as pre-concentration. So, at the Emerson plant in the USA, wet magnetic separation is used to separate pyrrhotite and magnetite contained in the ore, and at the Uyudag plant in Turkey, grade - 10 mm is subjected to dry grinding and magnetic separation in separators with low magnetic intensity to separate magnetite, and then enriched in separators with high tension in order to separate the garnet. Further enrichment includes bench concentration, flotation gravity and scheelite flotation. An example of the use of multi-stage combined schemes for the enrichment of poor tungsten ores, which ensure the production of high-quality concentrates, are the technological schemes used at factories in the PRC. So, at the Taishan plant with a capacity of 3000 tons / day for ore, wolframite-scheelite material with a tungsten content of 0.25% is processed. The original ore is subjected to manual and photometric sorting with the removal of 55% of waste rock to the dump. Further enrichment is carried out on jigging machines and concentration tables. The obtained rough gravity concentrates are adjusted by the methods of flotation gravity and flotation. The factories of Xihuashan, which processes ores with a wolframite to scheelite ratio of 10:1, use a similar gravity cycle. The draft gravity concentrate is fed to flotation gravity and flotation, due to which sulfides are removed. Next, wet magnetic separation of the chamber product is carried out in order to isolate wolframite and rare earth minerals. The magnetic fraction is sent to electrostatic separation and then wolframite flotation. The non-magnetic fraction enters the flotation of sulphides, and the flotation tails are subjected to magnetic separation to obtain scheelite and cassiterite-wolframite concentrates. The total content of WO3 is 65% with an extraction of 85%.
There is an increase in the use of the flotation process in combination with the chemical refinement of the resulting poor concentrates. In Canada, at the Mount Pleasant plant for the enrichment of complex tungsten-molybdenum ores, a flotation technology has been adopted, including flotation of sulfides, molybdenite and wolframite. In the main sulfide flotation, copper, molybdenum, lead, and zinc are recovered. The concentrate is cleaned, finely ground, subjected to steaming and conditioning with sodium sulfide. Molybdenum concentrate is cleaned and subjected to acid leaching. Sulfide flotation tailings are treated with sodium fluorosilicone 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 plant (Canada), the scheelite flotation process is complicated by the presence of talc in the ore, therefore, a primary talc flotation cycle is introduced, then copper minerals and pyrrhotite are flotation. The flotation tailings are subjected to gravity enrichment to obtain two tungsten concentrates. Gravity tailings are sent to the scheelite flotation cycle, and the resulting flotation concentrate is treated with hydrochloric acid. At the Ikssheberg plant (Sweden), the replacement of the gravity-flotation scheme with a purely flotation one made it possible to obtain a scheelite concentrate with a content of 68-70% WO3 with a recovery of 90% (according to the gravity-flotation scheme, the recovery was 50%) . Recently, much attention has been paid to improving the technology of extracting tungsten minerals from sludge in two main areas: gravitational sludge enrichment in modern multi-deck concentrators (similar to tin-containing sludge enrichment) with subsequent refinement of the concentrate by flotation and enrichment in wet magnetic separators with a high magnetic field strength (for wolframite slimes).
An example of the use of combined technology are factories in China. The technology includes slime thickening to 25-30% solids, sulphide flotation, tailings enrichment in centrifugal separators. The crude concentrate obtained (WO3 content 24.3% with a recovery of 55.8%) is fed to wolframite flotation using organophosphorus acid as a collector. The flotation concentrate containing 45% WO3 is subjected to wet magnetic separation to obtain wolframite and tin concentrates. According to this technology, a wolframite concentrate with a content of 61.3% WO3 is obtained from sludge with a content of 0.3-0.4% WO3 with a recovery of 61.6%. 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. So, at the factory Kuda (Japan), when enriching complex ores, 6 marketable products are obtained. In order to determine the possibility of additional extraction of useful components from stale tailings in the mid-90s. in TsNIGRI, a technological sample with a tungsten trioxide content of 0.1% was studied. 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. The conducted studies have shown that for the successful extraction of tungsten trioxide, significant costs will be required for regrinding tailings, and at this stage, their involvement in processing is not promising.
The technological scheme of mineral processing, which includes two or more devices, embodies all the characteristic features of a complex object, and the optimization of the technological scheme can, apparently, be the main task of system analysis. In solving this problem, almost all the previously considered modeling and optimization methods can be used. However, the structure of concentrator circuits is so complex that additional optimization techniques need to be considered. Indeed, for a circuit consisting of at least 10-12 devices, it is difficult to implement a conventional factorial experiment or to carry out multiple nonlinear statistical processing. Currently, several ways to optimize circuits are outlined, an evolutionary way of summarizing the accumulated experience and taking a step in the successful direction of changing the circuit.
Semi-industrial testing of the developed technological scheme for the enrichment of general relativity and 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 shown in fig. 3.4 and 3.5 and in table. 3.6. It can be seen that the yield of the conditioned concentrate is 0.14%, the content is 62.7% with the extraction of WO3 49.875%. The results of spectral analysis of a representative sample of the obtained concentrate, are given in table. 3.7, confirm that the W-concentrate of the III magnetic separation is conditioned and corresponds to the grade KVG (T) of GOST 213-73 "Technical requirements (composition,%) for tungsten concentrates obtained from tungsten-containing ores". Therefore, the developed technological scheme for the extraction of W from the stale tailings of the Dzhida VMK ore beneficiation can be recommended for industrial use and the stale tailings are transferred into additional industrial mineral raw materials of the Dzhida VMK.
For the industrial processing of stale tailings according to the developed technology at Q = 400 t/h, a list of equipment has been developed, which is given in class -0.1 mm must be carried out on a KNELSON centrifugal separator with periodic discharge of the concentrate. Thus, it has been established that the most effective way to extract WO3 from RTO 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 crushed to -0.1 mm tailings of primary enrichment - centrifugal separation. The essential features of the technology for processing stale tailings of the Dzhida VMK are as follows: 1. A narrow classification of the feed sent for primary enrichment and refinement is necessary; 2. An individual approach is required when choosing the method of primary enrichment of classes of various sizes; 3. Obtaining tailings is possible with the primary enrichment of the finest feed (-0.1 + 0.02 mm); 4. Use of hydrocyclone operations to combine dehydration and sizing 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 conditioned concentrate that meets the requirements of GOST 213-73.
Kiselev, Mikhail Yurievich
Tungsten minerals, ores and concentrates
Tungsten is a rare element, its average content in the earth's crust is Yu-4% (by mass). About 15 minerals of tungsten are known, however, only minerals of the wolframite group and scheelite are of practical importance.
Wolframite (Fe, Mn)WO4 is an isomorphic mixture (solid solution) of iron and manganese tungstates. If there is more than 80% iron tungstate in the mineral, the mineral is called ferberite, in the case of the predominance of manganese tungstate (more than 80%) - 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 according to the mineralogical scale 4.5-5. Scheelite often contains an isomorphic admixture of powellite, CaMo04. When irradiated with ultraviolet rays, scheelite fluoresces blue - blue light. At a molybdenum content of more than 1%, fluorescence becomes yellow. Scheelite is non-magnetic.
Tungsten ores are usually poor in tungsten. The minimum content of W03 in ores, at which their exploitation is profitable, is currently 0.14-0.15% for large and 0.4-0.5% for small deposits.
Together with tungsten minerals, molybdenite, cassiterite, pyrite, arsenopyrite, chalcopyrite, tantalite or columbite, etc. are found in ores.
According to the mineralogical composition, two types of deposits are distinguished - wolframite and scheelite, and according to the shape of ore formations - vein and contact types.
In vein deposits, tungsten minerals mostly occur in quartz veins of small thickness (0.3-1 m). The contact type of deposits is associated with zones of contact between granite rocks and limestones. They are characterized by deposits of scheelite-bearing skarn (skarns are silicified limestones). The skarn-type ores include the Tyrny-Auzskoye deposit, the largest in the USSR, in the North Caucasus. During the weathering of vein deposits, wolframite and scheelite accumulate, forming placers. In the latter, wolframite is often combined with cassiterite.
Tungsten ores are enriched to obtain 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 purely flotation schemes are used.
The extraction of tungsten into conditioned 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 middlings with a content of 10–20% W03 from the enrichment cycle for chemical (hydrometallurgical) processing, as a result of which "artificial scheelite" or technical tungsten trioxide is obtained. Such combined schemes provide a 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 less than 65%, the 2nd grade - not less than 60%. They limit the content of impurities P, S, As, Sn, Cu, Pb, Sb, Bi in the range 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 the PRC. 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 within 20 - 25 thousand tons (in terms of metal content).
Methods for processing tungsten concentrates
The main product of the 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 constituent of hard alloys.
Production schemes for the processing of tungsten concentrates are divided into two groups depending on the accepted method of decomposition:
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 by acids.
In cases where alkaline reagents are used for decomposition, solutions of sodium tungstate are obtained, from which, after purification from impurities, end products are produced - ammonium paratungstate (PVA) or tungstic acid. 24
When the concentrate is decomposed by acids, precipitation of technical tungstic acid is obtained, which is purified from impurities in subsequent operations.
Decomposition of tungsten concentrates. alkaline reagents Sintering with Na2C03
Sintering wolframite with Na2C03. The interaction of wolframite with soda in the presence of oxygen proceeds actively 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 proceed with a large loss of 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% in excess of the stoichiometric amount, complete decomposition of the concentrate is achieved. To accelerate the oxidation of iron and manganese, sometimes 1-4% nitrate is added to the charge.
Sintering wolframite with Na2C03 at domestic enterprises is carried out in tubular rotary kilns lined with fireclay bricks. In order to avoid the melting of the charge and the formation of deposits (growths) in the 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 content of W03 in it to 20-22%.
The furnace, 20 m long and with an outer diameter of 2.2 m, at a rotation speed of 0.4 rpm and an inclination of 3, has a capacity of 25 t/day in terms of charge.
The components of the charge (crushed concentrate, Na2C03, saltpeter) are fed from the hoppers to the screw mixer using automatic scales. The mixture enters the furnace hopper, from which it is fed into the furnace. After exiting the kiln, the sinter pieces pass through the crushing rolls and the wet grinding mill, from which the pulp is sent to the upper polisher (Fig. 1).
Scheelite sintering with Na2C03. At temperatures of 800-900 C, the interaction of scheelite with Na2C03 can proceed according to two reactions:
CaW04 + Na2CQ3 Na2W04 + CaCO3; (1.3)
CaW04 + Na2C03 *=*■ Na2W04 + CaO + C02. (1.4)
Both reactions proceed with a relatively small change in the Gibbs energy.
Reaction (1.4) proceeds to an appreciable extent above 850 C, when decomposition of CaCO3 is observed. The presence of calcium oxide in the sinter leads, when the sinter is leached with water, to the formation of poorly soluble calcium tungstate, which reduces the extraction of tungsten into solution:
Na2W04 + Ca(OH)2 = CaW04 + 2NaOH. (1.5)
With a large excess of Na2CO3 in the charge, this reaction is largely suppressed by the interaction of Na2CO4 with Ca(OH)2 to form CaCO3.
To reduce the consumption of Na2C03 and prevent the formation of free calcium oxide, quartz sand is added to the mixture to bind calcium oxide into insoluble silicates:
2CaW04 + 2Na2C03 + Si02 = 2Na2W04 + Ca2Si04 + 2C02;(l.6) AG°100IC = -106.5 kJ.
Nevertheless, in this case, too, to ensure a high degree of tungsten extraction into the solution, a significant excess of Na2CO3 (50–100% of the stoichiometric amount) must be introduced into the charge.
The 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 at the rate of reducing the content of W03 to 20-22%.
Leaching of soda specks. When cakes are leached with water, sodium tungstate and soluble salts of impurities (Na2Si03, Na2HP04, Na2HAs04, Na2Mo04, Na2S04), as well as an excess of Na2C03, pass into the solution. Leaching is carried out at 80-90 ° C in steel reactors with mechanical agitation, operating in hierio-
Concentrates with soda:
Elevator feeding the 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 weight dispensers; 7 - conveying auger; 8 - screw mixer; 9 - charge hopper; 10 - feeder;
Drum oven; 12 - roll crusher; 13 - rod mill-leacher; 14 - reactor with stirrer
Wild mode, or continuous drum rotary lixiviators. The latter are filled with crushing rods for crushing pieces of cake.
The extraction of tungsten from the sinter into the solution is 98-99%. Strong solutions contain 150-200 g/l W03.
Autoclave o-c One method of decomposition of tungsten concentrates
The autoclave-soda method was proposed and developed in the USSR1 in relation to the processing of scheelite concentrates and middlings. Currently, the method is used in 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 the 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 actual excess of Na2C03 is higher, since the rate of the process slows down as equilibrium is approached. For scheelite concentrates with a content of 45-55% W03 at 225 C, a soda equivalent of 2.6-3 is required. For middlings containing 15-20% W03, 4-4.5 moles of Na2C03 per 1 mole of CaW04 are required.
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 scheelite decomposition by Na2CO3 solutions was not found. With intensive 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, in case of insufficient stirring speed (which
Occurs in horizontal rotating autoclaves), an intermediate regime is realized: the rate of the process is determined both by 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
As can be seen from Fig. 2, the specific reaction rate decreases approximately in inverse proportion to the increase in the ratio of molar concentrations of Na2W04:Na2C03 in solution. This is
Ryas. Fig. 2. Dependence of the specific rate of scheelite decomposition by a soda solution in an autoclave j on the molar ratio of Na2W04/Na2C03 concentrations in the solution at
Causes 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, a two-stage countercurrent leaching is carried out. In this case, the tailings after the first leaching, in which there is little tungsten (15-20% of the original), are treated with a fresh solution containing a large excess of Na2C03. The resulting solution, which is circulating, enters the first stage of leaching.
Decomposition with Na2C03 solutions in autoclaves is also used for wolframite concentrates, however, the reaction in this case is more complicated, since 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 + H2CO3; (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 Na2CO3 in the solution and requires a large excess of the reagent.
To achieve satisfactory decomposition of wolframite concentrates, it is necessary to grind them finely 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.
The addition of NaOH or CaO to the autoclaved slurry (which leads to causticization of Na2C03) improves the degree of decomposition.
The decomposition rate 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
Ryas. 3. Autoclave unit with a horizontally rotating autoclave: 1 - autoclave; 2 - loading pipe for the pulp (steam is introduced through it); 3 - pulp pump; 4 - pressure gauge; 5 - pulp reactor-heater; 6 - self-evaporator; 7 - drop separator; 8 - pulp input into the self-evaporator; 9 - chipper made of armored steel; 10 - pipe for pulp removal; 11 - pulp collector
Leaching is carried out in steel horizontal rotating autoclaves heated with live steam (Fig. 3) and vertical continuous autoclaves with stirring of the pulp with bubbling steam. Approximate process mode: temperature 225 pressure in the autoclave ~ 2.5 MPa, ratio T: W = 1: (3.5 * 4), duration at each stage 2-4 hours.
Figure 4 shows a diagram of an autoclave battery. The initial autoclave pulp, heated by steam to 80-100 °C, is pumped into autoclaves, where it is heated to
secondary steam
Ditch. Fig. 4. Scheme of a continuous autoclave plant: 1 - reactor for heating the initial pulp; 2 - piston pump; 3 - autoclave; 4 - throttle; 5 - self-evaporator; 6 - pulp collector
200-225 °C live steam. In continuous operation, the pressure in the autoclave is maintained by discharging the slurry through a throttle (calibrated carbide washer). The pulp enters the self-evaporator - a vessel under pressure of 0.15-0.2 MPa, where the pulp is rapidly cooled due to intensive evaporation. The advantages of autoclave-soda decomposition of scheelite concentrates before sintering are the exclusion of the furnace process and a somewhat lower content of impurities in tungsten solutions (especially phosphorus and arsenic).
The disadvantages of the method include a large consumption of Na2C03. A high concentration of excess Na2C03 (80-120 g/l) entails an increased consumption of acids for the neutralization of solutions and, accordingly, high costs for the disposal of waste solutions.
Decomposition of tungstate conc.
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 equal to 0.68, respectively; 2.23 and 2.27.
Complete decomposition (98-99%) is achieved by treating the finely divided concentrate with 25-40% sodium hydroxide solution at 110-120°C. The required excess of alkali is 50% or more. The decomposition is carried out in steel sealed reactors equipped with stirrers. The passage of air into the solution accelerates the process due to the oxidation of iron (II) hydroxide Fe (OH) 2 into hydrated iron (III) oxide Fe203-«H20 and manganese (II) hydroxide Mn (OH) 2 into hydrated manganese (IV) oxide Mn02-lH20 .
The use of decomposition with alkali solutions is advisable only for high-grade wolframite concentrates (65-70% W02) with a small amount of silica and silicate impurities. When processing low-grade concentrates, highly contaminated solutions and hard-to-filter precipitates 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 mainly processed according to the traditional scheme, which includes: purification from compounds of impurity elements (Si, P, As, F, Mo); precipitation
Calcium tungsten mag (artificial scheelite) with its subsequent decomposition with acids and obtaining technical tungstic acid; dissolution of 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 drawback of the scheme is its multi-stage nature, carrying out most of the operations in a periodic mode, and the duration of a number of redistributions. An extraction and ion-exchange technology for converting Na2W04 solutions into (NH4)2W04 solutions has been developed and is already being used at some enterprises. The main redistributions of the traditional scheme and new extraction and ion-exchange variants of the technology are briefly considered below.
Purification of impurities
Silicon cleaning. When the content of Si02 in solutions exceeds 0.1% of the content of W03, preliminary purification from 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, added in a thin stream with stirring (to avoid local peroxidation) to a heated solution of sodium tungstate.
Purification of 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 an excess of Mg2+ and NH4 ions, the solubility is lower.
The 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 a long standing (48 hours), crystalline precipitates of ammonium-magnesium salts precipitate from the solution.
Purification from fluoride ions. With a high content of fluorite in the original 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. Purification of fluorine can be combined with hydrolytic isolation of silicic acid.
Molybdenum cleaning. Solutions of sodium tungstate" must be purified from 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, in the processing of scheelite-powellite Tyrny-Auzsky concentrates), the isolation of molybdenum is of particular importance, since it is aimed at obtaining a molybdenum chemical concentrate.
A common method is to precipitate the sparingly soluble molybdenum trisulfide MoS3 from a solution.
It is known that when sodium sulfide is added to solutions of tungstate or sodium 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 much larger than for Na2W04(^^0 » Kzr). Therefore, if an amount of Na2S is added to the solution, sufficient only for interaction with Na2Mo04 (with a slight excess), then molybdenum sulfosalt is predominantly formed. With the 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 co-precipitates. By dissolving the sulfide precipitate in a soda solution and re-precipitating molybdenum trisulfide, a molybdenum concentrate is obtained with a W03 content of not more than 2% with a loss of tungsten 0.3-0.5% of the initial amount.
After partial oxidative roasting of the precipitate of molybdenum trisulfide (at 450-500 ° C), a molybdenum chemical concentrate is obtained with a content of 50-52% molybdenum.
The disadvantage of the method of precipitation of molybdenum in the composition of trisulfide is the release of hydrogen sulfide according to reaction (1.17), which requires expenses for the neutralization of gases (they use the absorption of H2S in a scrubber irrigated with a sodium hydroxide solution). The selection of molybdenum trisulfide is carried out from a solution heated to 75-80 C. The operation is carried out in sealed steel reactors, gummed or coated with acid-resistant enamel. The trisulfide precipitates are separated from the solution by filtration on a filter press.
Obtaining tungstic acid from solutions of sodium tungstate
Tungstic acid can be directly isolated from a solution of sodium tungstate with hydrochloric or nitric acid. However, this method is rarely used due to the difficulty of washing precipitates from sodium ions, the content of which in tungsten trioxide is limited.
For the most part, calcium tungstate is initially precipitated from the solution, which is then decomposed with acids. Calcium tungstate is precipitated by adding a CaCl2 solution heated to 80-90 C to a sodium tungstate solution with 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 liquor, which ensures their low content in tungstic acid. 99-99.5% W precipitates from the solution, mother solutions contain 0.05-0.07 g/l W03. The CaW04 precipitate washed with water in the form of a paste or pulp enters for decomposition with hydrochloric acid when heated to 90 °:
CaW04 + 2HC1 = H2W04i + CaCl2. (1.18)
During decomposition, a high final acidity of the pulp is maintained (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). Precipitates of tungstic acid require thorough washing from impurities (especially from calcium salts
and sodium). In recent years, continuous washing of tungstic acid in pulsating columns has been mastered, which greatly simplified the operation.
At one of the enterprises in the USSR, when processing sodium tungstate solutions, instead of hydrochloric acid, nitric acid is used to neutralize the solutions and decompose CaW04 precipitates, and the precipitation of the latter is carried out by introducing Ca(N03)2 into the solutions. In this case, the nitric acid mother liquors are disposed of, obtaining nitrate salts used as fertilizer.
Purification of technical tungstic acid and obtaining 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, the production of tungsten requires trioxide of a higher purity with a total impurity content of no more than 0.05%.
The ammonia method for purifying tungstic acid is generally accepted. It is easily soluble 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, alkali metal salts.
From the ammonia solution, as a result of evaporation and subsequent cooling, a crystalline precipitate of PVA is isolated:
Evaporation
12(NH4)2W04 * (NH4)10H2W12O42 4Н20 + 14NH3 +
In industrial practice, the composition of PVA is often written in the oxide form: 5(NH4)20-12W03-5H20, which does not reflect its chemical nature as an isopoly acid salt.
Evaporation is carried out in batch or continuous devices made of stainless steel. Usually 75-80% of tungsten is isolated into crystals. Deeper crystallization is undesirable in order to avoid contamination of the crystals with impurities. It is significant that most of the molybdenum impurity (70-80%) remains in the mother liquor. 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 with cold water and dried.
Tungsten trioxide is obtained by thermal decomposition of tungstic acid or PVA:
H2W04 \u003d "W03 + H20;
(NH4) 10H2W12O42 4H20 = 12W03 + 10NH3 + 10H20. (1.20)
Calcination is carried out in rotary electric furnaces with a pipe made of heat-resistant steel 20X23H18. The calcination mode depends on the purpose of tungsten trioxide, the required size of its particles. So, to obtain tungsten wire grade VA (see below), PVA is calcined at 500-550 ° C, wire grades VCh and VT (tungsten without additives) - at 800-850 ° C.
Tungstic acid is calcined at 750-850 °C. Tungsten trioxide derived from PVA has larger particles than trioxide derived from tungstic acid. In tungsten trioxide, intended for the production of tungsten, the content of W03 must be at least 99.95% for the production of hard alloys - at least 99.9%.
Extraction and ion-exchange methods for processing solutions of sodium tungstate
The processing of sodium tungstate solutions is greatly simplified when tungsten is extracted from solutions by extraction with an organic extractant, followed by re-extraction from the organic phase with an ammonia solution with separation of PVA from an ammonia solution.
Since in a wide range of pH=7.5+2.0 tungsten is found in solutions in the form of polymeric anions, anion-exchange extractants are used for extraction: salts of amines or quaternary ammonium bases. In particular, the sulfate salt of trioctylamine (i?3NH)HS04 (where R is С8Н17) is used in industrial practice. The highest rates of tungsten extraction are observed at pH=2*4.
Extraction is described by the equation:
4 (i? 3NH) HS04 (opr) + H2 \ U120 * "(aq) + 2H + (aq) ї \u003d ї
Ї \u003d ї (D3GSh) 4H4 \ U12O40 (org) + 4H80; (aq.). (l.2l)
The amine is dissolved in kerosene, to which a technical mixture of polyhydric alcohols (C7 - C9) is added to prevent the precipitation of a solid phase (due to the low solubility of amine salts in kerosene). The approximate composition of the organic phase: amines 10%, alcohols 15%, kerosene - the rest.
Solutions purified from mrlibden, 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. The extraction is carried out in mixer-settler type apparatuses or in pulsating columns with packing.
The advantages of extraction processing of sodium tungstate solutions are obvious: the number of operations of the technological scheme is reduced, it is possible to carry out a continuous process for obtaining ammonium tungstate solutions from sodium tungstate solutions, and production areas are reduced.
Wastewater from the extraction process may contain an admixture of 80-100 mg/l of amines, as well as impurities of higher alcohols and kerosene. To remove these environmentally harmful impurities, froth flotation and adsorption on activated carbon are used.
Extraction technology is used at foreign enterprises and is also implemented at domestic plants.
The use of ion-exchange resins is a direction of the scheme for processing sodium tungstate solutions that competes with extraction. For this purpose, low-basic anion exchangers containing amine groups (often 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 the total 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 by the equations, respectively:
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 enterprises of the USSR. 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 resin bed in a continuous mode. A complicating circumstance is the partial isolation of PVA crystals at the stage of elution, which requires their separation from the resin particles. As a result of elution, solutions containing 150–170 g/l W03 are obtained, which are fed to the evaporation and crystallization of PVA.
The disadvantage of ion-exchange technology compared to extraction is the unfavorable kinetics (contact time 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 with acids
In industrial practice, mainly in the processing of 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 much higher than the stoichiometrically required one (250–300%) due to the inhibition of the process by tungstic acid films 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 of various deposits differ in reactivity).
A single treatment does not always lead to a 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 with a content of 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. So, at 20 C in 270 g/l HC1, the solubilities of H2Mo04 and H2WO4 are 182 and 0.03 g/l, respectively. Despite this, complete separation of molybdenum is not achieved. Precipitates of tungstic acid contain 0.2-0.3% molybdenum, which cannot be extracted by re-treatment with hydrochloric acid.
The acid method differs from the alkaline methods of scheelite decomposition by a smaller number of operations of the technological scheme. However, when processing concentrates with a relatively low content of W03 (50-55%) with a significant content of impurities, in order to obtain conditioned ammonium paratungstate, two or three ammonia purifications of tungstic acid have to be carried out, which is uneconomical. Therefore, decomposition with hydrochloric acid is mostly used in the processing of rich and pure scheelite concentrates.
The disadvantages of the method of decomposition 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 the light of the tasks of creating waste-free technologies, the nitric acid method of decomposition of scheelite concentrates is of interest. In this case, the mother solutions are easy to dispose of, obtaining nitrate salts.
Tungsten ores in our country were processed at large GOKs (Orlovsky, Lermontovsky, Tyrnauzsky, Primorsky, Dzhidinsky VMK) according to the now classic technological schemes with multi-stage grinding and enrichment of material divided into narrow size classes, as a rule, in two cycles: primary gravitational enrichment and fine-tuning of rough concentrates by various methods. This is due to the low content of tungsten in the processed ores (0.1-0.8% WO3) and high quality requirements for concentrates. Primary enrichment for coarsely disseminated ores (minus 12+6 mm) was carried out by jigging, and for medium-, fine- and finely disseminated ores (minus 2+0.04 mm) screw apparatuses of various modifications and sizes were used.
In 2001, the Dzhida tungsten-molybdenum plant (Buryatia, Zakamensk) ceased its activity, having accumulated after it the Barun-Naryn technogenic tungsten deposit, multimillion in terms of sand volume. Since 2011, Zakamensk CJSC has been processing this deposit 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), middlings regrinding and flotation of the bulk gravity concentrate to obtain a KVGF grade concentrate. During operation, a number of factors were noted in the operation of Knelson concentrators that negatively affect 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 capacities and the increased cost of electricity, this factor is of particular importance;
Low degree of extraction of tungsten minerals into gravity concentrate (about 60% of the operation);
The complexity of this equipment in operation: with fluctuations in the material composition of the enriched raw materials, centrifugal concentrators require intervention in the process and operational settings (changes in the pressure of the fluidizing water, the speed of rotation of the enrichment bowl), which leads to fluctuations in the quality characteristics of the obtained gravity concentrates;
Significant remoteness of the manufacturer and, as a result, a long waiting time for spare parts.
In search of an alternative method of gravitational concentration, Spirit conducted laboratory tests of the technology screw separation using industrial screw separators SVM-750 and SVSH-750 manufactured by LLC PK Spirit. Enrichment took place in two operations: main and control with the receipt of three enrichment products - concentrate, middlings and tailings. All enrichment products obtained as a result of the experiment were analyzed in the laboratory of ZAO Zakamensk. The best results are presented in table. one.
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 step was to conduct semi-industrial tests on the existing enrichment scheme. A pilot semi-industrial plant was assembled 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 operating enrichment plant, and sampling was carried out directly from the enrichment process without stopping the operation of the equipment. Indicators of semi-industrial tests are presented in table. 2.
Table 2. Results of comparative semi-industrial tests of screw apparatuses and centrifugal concentratorsknelson
|
Indicators |
Source food |
Concentrate |
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Recovery, % |
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The results show that the enrichment of sands is more efficient on screw apparatus than on centrifugal concentrators. This translates into a lower concentrate yield (16.87% versus 32.26%) with an increase in recovery (83.13% versus 67.74%) into tungsten mineral concentrate. This results in a higher quality WO3 concentrate (0.9% versus 0.42%),
Vladivostok
annotation
In this paper, technologies for the enrichment of scheelite and wolframite are considered.
The technology of enrichment of tungsten ores includes: preliminary concentration, enrichment of crushed products of preliminary concentration to obtain collective (rough) concentrates and their refinement.
Keywords
Scheelite ore, wolframite ore, heavy medium separation, jigging, gravity method, electromagnetic separation, flotation.
1. Introduction 4
2. Preconcentration 5
3. Technology of beneficiation of wolframite ores 6
4. Technology of enrichment of Scheelite ores 9
5. Conclusion 12
References 13
Introduction
Tungsten is a silver-white metal with high hardness and a boiling point of about 5500°C.
The Russian Federation has large explored reserves. Its tungsten ore potential is estimated at 2.6 million tons of tungsten trioxide, in which the proven reserves are 1.7 million tons, or 35% of those in the world.
Fields under development in Primorsky Krai: Vostok-2, OJSC Primorsky GOK (1.503%); Lermontovskoye, AOOT Lermontovskaya GRK (2.462%).
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).
When processing tungsten-containing ores, gravity, flotation, magnetic, as well as electrostatic, hydrometallurgical and other methods are used.
preliminary concentration.
The cheapest and at the same time highly productive methods of preconcentration are gravitational ones, such as heavy media separation and jigging.
Heavy media separation makes it possible to stabilize the quality of the food entering the main processing cycles, to separate not only the waste product, but also to separate the ore into rich coarsely disseminated and poor finely disseminated ore, often requiring fundamentally different processing schemes, since they differ markedly in material composition. The process is characterized by the highest density separation accuracy compared to other gravity methods, which makes it possible to obtain a high recovery of a valuable component with a minimum concentrate yield. When enriching ore in heavy suspensions, a difference in the densities of the separated pieces of 0.1 g/m3 is sufficient. This method can be successfully applied to coarsely disseminated wolframite and scheelite-quartz ores. The results of studies on the enrichment of tungsten ores from the Pun-les-Vignes (France) and Borralha (Portugal) deposits under industrial conditions showed that the results obtained using enrichment in heavy suspensions are much better than when enriched only on jigging machines - into a heavy fraction recovery was more than 93% of the ore.
Jigging in comparison with heavy-medium enrichment, it requires less capital expenditures, allows enriching the material in a wide range of density and fineness. Large-sized jigging is widely used in the enrichment of large- and medium-disseminated ores that do not require fine grinding. The use of jigging is preferable when enriching carbonate and silicate ores of skarn, vein deposits, while the value of the contrast ratio of ores in terms of gravitational composition should exceed one.
Technology of beneficiation of wolframite ores
The high specific gravity of tungsten minerals and the coarse-grained structure of wolframite ores make it possible to widely use gravity processes in their enrichment. To obtain high technological indicators, it is necessary to combine apparatuses with different separating characteristics in the gravitational scheme, in which each previous operation in relation to the next one is, as it were, preparatory, improving the enrichment of the material. A schematic diagram of the enrichment of wolframite ores is shown in fig. one.
Jigging is used starting from the size at which tailings can be identified. This operation is also used for separating coarsely disseminated tungsten concentrates with subsequent regrinding and enrichment of jigging tailings. The basis for choosing the scheme of jigging and the size of the enriched material are the data obtained by separating the density of the material with a size of 25 mm. If the ores are finely disseminated and preliminary studies show that large-sized enrichment and jigging are unacceptable for them, then the ore is enriched in suspension-carrying flows of small thickness, which include enrichment on screw separators, jet chutes, cone separators, locks, concentration tables. With staged grinding and staged enrichment of ore, the extraction of wolframite into rough concentrates is more complete. Rough wolframite gravity concentrates are brought to standard according to developed schemes using wet and dry enrichment methods.
Rich wolframite concentrates are enriched by electromagnetic separation, while the electromagnetic fraction can be contaminated with iron zinc blende, bismuth minerals and partially arsenic (arsenopyrite, scorodite). To remove them, magnetizing roasting is used, which increases the magnetic susceptibility of iron sulfides, and at the same time, sulfur and arsenic, which are harmful to tungsten concentrates, are removed in the form of gaseous oxides. Wolframite (hubnerite) is additionally extracted from sludge by flotation using fatty acid collectors and the addition of neutral oils. Rough gravitational concentrates are relatively easy to bring to standard using electrical methods of enrichment. Flotation and flotation gravity are carried out with the supply of xanthate and blowing agent in a slightly alkaline or slightly acidic medium. If the concentrates are contaminated with quartz and light minerals, then after flotation they are subjected to recleaning on concentration tables.
Similar information.
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, and also 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 Tyrny-Auz factory. This factory processes very complex molybdenum-scheelite ores containing a number of calcium minerals (calcite, fluorite, apatite). Calcium minerals, like scheelite, are floated with oleic acid, the depression of calcite and fluorite is produced by mixing in a liquid glass solution without heating (long contact) or with heating, as at the Tyrny-Auz factory. Instead of oleic acid, tall oil fractions are used, as well as acids from vegetable oils (reagents 708, 710, etc.) alone or in a mixture with oleic acid.
A typical scheme of scheelite ore flotation is given in fig. 38. According to this scheme, it is possible to remove calcite and fluorite and obtain concentrates that are conditioned in terms of tungsten trioxide. Ho apatite still remains in such quantity that the phosphorus content in the concentrate is above the standards. Excess phosphorus is removed by dissolving apatite in weak hydrochloric acid. The consumption of acid depends on the content of calcium carbonate in the concentrate and is 0.5-5 g of acid per ton of WO3.
In acid leaching, part of the scheelite, as well as powellite, is dissolved and then precipitated from solution in the form of CaWO4 + CaMoO4 and other impurities. The resulting dirty sediment is then processed according to the method of I.N. Maslenitsky.
Due to the difficulty of obtaining a conditioned tungsten concentrate, many factories abroad produce two products: a rich concentrate and a poor one for hydrometallurgical processing into calcium tungstate according to the method developed in Mekhanobre I.N. Maslenitsky, - leaching with soda in an autoclave under pressure with transfer to a 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, the extraction of scheelite abroad by flotation has not been mastered. Such ores, for example in Sweden, are enriched on tables. Scheelite entrained with fluorite in the flotation concentrate is then recovered from this concentrate on a table.
At factories in Russia, scheelite ores are enriched by flotation, obtaining conditioned concentrates.
At the Tyrny-Auz plant, ore with a content of 0.2% WO3 is used to produce concentrates with a content of 6о% WO3 with an extraction of 82%. At the Chorukh-Dairon plant, with the same ore in terms of VVO3 content, 72% WO3 is obtained in concentrates with an extraction of 78.4%; at the Koitash plant, with ore with 0.46% WO3 in concentrate, 72.6% WO3 is obtained with a WO3 recovery of 85.2%; at the Lyangar plant in ore 0.124%, in concentrates - 72% with an extraction of 81.3% WO3. Additional separation of poor products is possible by reducing losses in the tailings. In all cases, if sulfides are present in the ore, they are isolated 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 the Bukuki ore (Transbaikalia), can be pre-enriched in heavy suspensions, separating about 60% of waste rock at a fineness of -26 + 3 MM with a content of no more than 0.03% WO3.
However, with a relatively low productivity of factories (not more than 1000 tons / day), the first stage of enrichment is carried out in jigging machines, usually starting from a particle size of about 10 mm with coarsely disseminated ores. In new modern schemes, in addition to jigging machines and tables, Humphrey screw separators are used, replacing some of the tables with them.
The progressive scheme of enrichment of tungsten ores is given in fig. 39.
Finishing of tungsten concentrates depends on their composition.
Sulfides from concentrates thinner than 2 mm are isolated by flotation gravity: concentrates after mixing with acid and flotation reagents (xanthate, oils) are sent to a concentration table; the resulting CO table concentrate is dried and subjected to magnetic separation. The coarse-grained concentrate is pre-crushed. Sulfides from fine concentrates from slurry tables are isolated by froth flotation.
If there are a lot of sulfides, it is advisable to separate them from the hydrocyclone drain (or classifier) before enrichment on the tables. This will improve the conditions for separating wolframite on the tables and during concentrate finishing operations.
Typically, coarse concentrates prior to finishing contain about 30% WO3 with recovery up to 85%. For illustration in table. 86 shows some data on factories.
During gravitational enrichment of wolframite ores (hubnerite, ferberite) from slimes thinner than 50 microns, the extraction is very low and losses in the slime part are significant (10-15% of the content in the ore).
From sludges by flotation with fatty acids at pH=10, additional WO3 can be recovered 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 precious metals. Some of them pass during gravitational enrichment into gravitational concentrates and are transferred to finishing tailings. Molybdenum, bismuth-lead, lead-copper-silver, zinc (they contain cadmium, indium) and pyrite concentrates can be isolated by selective flotation from sulfide tailings, as well as from sludge, and the tungsten product can also be additionally isolated.
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