Enrichment of tungsten ores. Maintaining the main method of enrichment of tungsten ores and the use of auxiliary dehydration processes in the technological scheme of the tungsten ore enrichment cycle
The main tungsten minerals are scheelite, hübnerite and wolframite. Depending on the type of minerals, ores can be divided into two types; scheelite and wolframite (huebnerite).Scheelite ores in Russia, as well as in some cases abroad, are enriched by flotation. In Russia, the process of flotation of scheelite ores in industrial scale carried out before the Second World War at the Tyrn-Auz factory. This plant processes very complex molybdenum-scheelite ores containing a number of calcium minerals (calcite, fluorite, apatite). Calcium minerals, like scheelite, float with oleic acid; the depression of calcite and fluorite is produced by stirring in a liquid glass solution without heating (long-term contact) or with heating, as at the Tyrn-Auz factory. Instead of oleic acid, fractions of tall oil are used, as well as acids from vegetable oils (reagents 708, 710, etc.) alone or in a mixture with oleic acid.
A typical flotation scheme for scheelite ore is shown in Fig. 38. Using this scheme, it is possible to remove calcite and fluorite and obtain tungsten trioxide-standard concentrates. However, apatite still remains in such quantity that the phosphorus content in the concentrate is higher than standard. Excess phosphorus is removed by dissolving apatite in weak hydrochloric acid. Acid consumption depends on the calcium carbonate content in the concentrate and is 0.5-5 g of acid per ton of WO3.
When leaching with acid, part of the scheelite, as well as powellite, is dissolved and then precipitated out of solution in the form of CaWO4 + CaMoO4 and other impurities. The resulting dirty sludge is then processed according to the I.N. method. Maslenitsky.
Due to the difficulty of obtaining quality tungsten concentrate, many factories abroad produce two products: a rich concentrate and a poor one for hydrometallurgical processing into calcium tungstate using the method developed in Mekhanobra I.N. Maslenitsky, - leaching with soda in an autoclave under pressure with transfer into solution in the form of CaWO4, followed by purification of the solution and precipitation of CaWO4. In some cases, with coarsely disseminated scheelite, finishing of flotation concentrates is carried out on tables.
From ores containing a significant amount of CaF2, extraction of scheelite by flotation has not been developed abroad. Such ores, for example in Sweden, are enriched on tables. Scheelite, entrained with fluorite in the flotation concentrate, is then separated from this concentrate on the table.
In Russian factories, scheelite ores are enriched by flotation, obtaining quality concentrates.
At the Tyrn-Auz plant, concentrates containing 6% WO3 are produced from ore containing 0.2% WO3 with a recovery of 82%. At the Chorukh-Dairon plant, with ore of the same VVO3 content, 72% WO3 is obtained in concentrates with an extraction of 78.4%; at the Koytash plant, with ore with 0.46% WO3 in concentrate, 72.6% WO3 is obtained with a WO3 recovery of 85.2%; at the Lyangarsky plant in ore 0.124%, in concentrates - 72% with extraction of 81.3% WO3. Additional recovery of poor products is possible by reducing losses in tailings. In all cases, if sulfides are present in the ore, they are separated before scheelite flotation.
The consumption of materials and energy is illustrated by the data below, kg/t:
Wolframite (Hübnerite) ores are enriched exclusively by gravity methods. Some ores with uneven and coarse-grained dissemination, such as Bukuki ore (Transbaikalia), can be pre-enriched in heavy suspensions, releasing about 60% waste rock with a particle size of 26+3 MM with a content of no more than 0.03% WO3.
However, with a relatively low productivity of factories (no more than 1000 tons/day), the first stage of enrichment is carried out in jigging machines, usually starting with a particle size of about 10 mm for coarsely disseminated ores. In new modern schemes, in addition to jiggers and tables, Humphrey screw separators are used, replacing part of the tables with them.
A progressive scheme for the enrichment of tungsten ores is shown in Fig. 39.
The finishing of tungsten concentrates depends on their composition.
Sulfides from concentrates thinner than 2 mm are separated by flotogravity: the concentrates, after mixing with acid and flotation reagents (xanthate, oils), are sent to a concentration table; The resulting CO2 concentrate is dried and subjected to magnetic separation. The coarse concentrate is pre-crushed. Sulfides are separated from fine concentrates from slurry tables by foam flotation.
If there are a lot of sulfides, it is advisable to separate them from the discharge of hydrocyclones (or classifier) before enrichment on the tables. This will improve the conditions for the release of wolframite on tables and during concentrate finishing operations.
Typically, rough concentrates before finishing contain about 30% WO3 with recovery up to 85%. For illustration in table. 86 shows some data on factories.
During the gravitational enrichment of wolframite ores (Hübnerite, ferberite) from slurries thinner than 50 microns, the recovery is very low and the losses in the slurry part are significant (10-15% of the content in the ore).
From sludges, flotation with fatty acids at pH=10 can further extract WO3 into lean products containing 7-15% WO3. These products are suitable for hydrometallurgical processing.
Wolframite (Hübnerite) ores contain a certain amount of non-ferrous, rare and noble metals. Some of them pass during gravity enrichment into gravity concentrates and are transferred to finishing tailings. From sulfide finishing tailings, as well as from sludge, molybdenum, bismuth-lead, lead-copper-silver, zinc (they contain cadmium, indium) and pyrite concentrates can be isolated by selective flotation, and the tungsten product can also be isolated.
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Cassiterite SnO 2– the main industrial mineral of tin, which is present in tin-bearing placers and bedrock ores. The tin content in it is 78.8%. Cassiterite has a density of 6900...7100 kg/t and a hardness of 6...7. The main impurities in cassiterite are iron, tantalum, niobium, as well as titanium, manganese, pigs, silicon, tungsten, etc. physicochemical characteristics cassiterite, for example, magnetic susceptibility, and its flotation activity.
Stannin Cu 2 S FeS SnS 4- tin sulfide mineral, although it is the most common mineral after cassiterite, has no industrial significance, firstly, because it has a low tin content (27...29.5%), and secondly, the presence of copper and iron sulfides in it complicates the metallurgical processing of concentrates and, thirdly, the proximity of the flotation properties of the bed to sulfides makes separation during flotation difficult. The composition of tin concentrates obtained at concentration plants is different. From rich tin placers, gravity concentrates containing about 60% tin are isolated, and slurry concentrates obtained by both gravity and flotation methods can contain from 15 to 5% tin.
Tin deposits are divided into placer and bedrock deposits. Alluvial Tin deposits are the main source of world tin production. Placers contain about 75% of the world's tin reserves. Indigenous Tin deposits have a complex material composition, depending on which they are divided into quartz-cassiterite, sulfide-quartz-cassiterite and sulfide-cassiterite.
Quartz-cassiterite ores are usually complex tin-tungsten ores. Cassiterite in these ores is represented by large-, medium- and finely disseminated crystals in quartz (from 0.1 to 1 mm m more). In addition to quartz and cassiterite, these ores typically contain feldspar, tourmaline, micas, wolframite or scheelite, and sulfides. Sulfide-cassiterite ores are dominated by sulfides - pyrite, pyrrhotite, arsenopyrite, galena, sphalerite and stanine. Also contains iron minerals, chlorite and tourmaline.
Tin placers and ores are enriched mainly by gravity methods using jigging machines, concentration tables, screw separators and sluices. Placers are usually much easier to enrich by gravity methods than ores from primary deposits, because they do not require expensive crushing and grinding processes. Finishing of rough gravity concentrates is carried out using magnetic, electrical and other methods.
Enrichment on sluices is used when the cassiterite grain size is more than 0.2 mm, because smaller grains are poorly captured on the sluices and their extraction does not exceed 50...60%. More efficient devices are jigging machines, which are installed for primary enrichment and allow the extraction of up to 90% of cassiterite. Finishing of coarse concentrates is carried out on concentration tables (Fig. 217).
Fig. 217. Scheme of enrichment of tin placers
Primary enrichment of placers is also carried out on dredges, including sea dredges, where drum screens with holes of 6...25 mm in size are installed to wash sand, depending on the distribution of cassiterite according to the classes of sand size and washability. Jigging machines are used to enrich the under-screen product of screens. various designs usually with an artificial bed. Gateways are also installed. Primary concentrates are subjected to cleaning operations on jigging machines. Finishing is usually carried out at onshore finishing installations. The recovery of cassiterite from placers is usually 90...95%.
Enrichment of indigenous tin ores, characterized by the complexity of the material composition and uneven dissemination of cassiterite, is carried out according to more complex multi-stage schemes using not only gravitational methods, but also flotogravity, flotation, and magnetic separation.
When preparing tin ores for beneficiation, it is necessary to take into account the ability of cassiterite to sludge due to its size. More than 70% of tin losses during enrichment are due to sludged cassiterite, which is carried away with the drains of gravity devices. Therefore, the grinding of tin ores is carried out in rod mills, which operate in a closed cycle with screens. At some factories, enrichment in heavy suspensions is used at the head of the process, which makes it possible to separate up to 30...35% of the host rock minerals into the tailings, reduce grinding costs and increase tin extraction.
To isolate coarse-grained cossiterite at the head of the process, jigging is used with a feed size ranging from 2...3 to 15...20 mm. Sometimes, instead of jigging machines, when the material size is minus 3+ 0.1 mm, screw separators are installed, and when enriching material with a size of 2...0.1 mm, concentration tables are used.
For ores with uneven dissemination of cassiterite, multi-stage schemes are used with sequential grinding of not only tailings, but also poor concentrates and middlings. In tin ore, which is enriched according to the scheme presented in Fig. 218, cassiterite has a particle size of 0.01 to 3 mm.
Rice. 218. Scheme of gravity enrichment of primary tin ores
The ore also contains iron oxides, sulfides (arsenopyrite, chalcopyrite, pyrite, stanine, galena), and wolframite. The nonmetallic part is represented by quartz, tourmaline, chlorite, sericite and fluorite.
The first stage of enrichment is carried out in jigging machines at an ore size of 90% minus 10 mm with the release of coarse tin concentrate. Then, after additional grinding of the tailings of the first stage of enrichment and hydraulic classification according to equal incidence, enrichment is carried out on concentration tables. The tin concentrate obtained according to this scheme contains 19...20% tin with an extraction of 70...85% and is sent for finishing.
During finishing, sulfide minerals and host rock minerals are removed from coarse tin concentrates, which makes it possible to increase the tin content to standard levels.
Coarsely disseminated sulfide minerals with a particle size of 2...4 mm are removed by flotogravity on concentration tables, before which the concentrates are treated with sulfuric acid (1.2...1.5 kg/t), xanthate (0.5 kg/t) and kerosene (1...2 kg/t). T).
Cassiterite is extracted from gravity enrichment sludge by flotation using selective collecting reagents and depressants. For ores of complex mineral composition containing significant amounts of tourmaline and iron hydroxides, the use of fatty acid collectors makes it possible to obtain poor tin concentrates containing no more than 2...3% tin. Therefore, when flotating cassiterite, selective collectors such as Asparal-F or aerosol -22 (succinamates), phosphonic acids and the IM-50 reagent (alkylhydroxamic acids and their salts) are used. Liquid glass and oxalic acid are used to depress minerals in host rocks.
Before cassiterite flotation, material with a particle size of minus 10...15 microns is removed from the sludge, then sulfide flotation is carried out, from the tails of which at pH 5 with the supply of oxalic acid, liquid glass and the Asparal-F reagent (140...150 g/t), supplied to cassiterite floats as a collector (Fig. 219). The resulting flotation concentrate contains up to 12% tin with extraction from the operation up to 70...75% tin.
Sometimes Bartles-Moseley orbital locks and Bartles-Crosbelt concentrators are used to extract cassiterite from slurries. The rough concentrates obtained on these devices, containing 1...2.5% tin, are sent for finishing to slurry concentration tables to obtain commercial slurry tin concentrates.
Tungsten in ores is represented by a wider range of minerals having industrial value than tin. Of the 22 tungsten minerals currently known, four are the main ones: wolframite (Fe,Mn)WO 4(density 6700...7500 kg/m 3), hübnerite MnWO 4(density 7100 kg/m 3), ferberite FeWO 4(density 7500 kg/m 3) and sheelite CAWO 4(density 5800...6200 kg/m3). In addition to these minerals, molybdoscheelite, which is scheelite and an isomorphic admixture of molybdenum (6...16%), is of practical importance. Wolframite, Huebnerite and Ferberite are weakly magnetic minerals, they contain magnesium, calcium, tantalum and niobium as impurities. Wolframite is often found in ores together with cassiterite, molybdenite and sulfide minerals.
Industrial types of tungsten-containing ores include vein quartz-wolframite and quartz-cassiterite-wolframite, stockwork, skarn and placer. In the deposits vein type contains wolframite, hübnerite and scheelite, as well as molybdenum minerals, pyrite, chalcopyrite, tin, arsenic, bismuth and gold minerals. IN stockwork In deposits, the tungsten content is 5...10 times lower than in vein deposits, but they have large reserves. IN skarn The ores, along with tungsten, represented mainly by scheelite, contain molybdenum and tin. Alluvial tungsten deposits have small reserves, but play a significant role in tungsten mining. The industrial content of tungsten trioxide in placers (0.03...0.1%) is significantly lower than in bedrock ores, but their development is much simpler and more economically profitable. These placers, along with wolframite and scheelite, also contain cassiterite.
The quality of tungsten concentrates depends on the material composition of the ore being processed and the requirements that are placed on them when used in various industries. So, to produce ferrotungsten, the concentrate must contain at least 63% WO 3, wolframite-huebnerite concentrate for the production of hard alloys must contain at least 60% WO 3. Scheelite concentrates typically contain 55% WO 3. The main harmful impurities in tungsten concentrates are silica, phosphorus, sulfur, arsenic, tin, copper, lead, antimony and bismuth.
Tungsten placers and ores are enriched, like tin, in two stages - primary gravitational enrichment and finishing of rough concentrates using various methods. With a low content of tungsten trioxide in the ore (0.1...0.8%) and high requirements for the quality of concentrates, the total degree of enrichment ranges from 300 to 600. This degree of enrichment can only be achieved by combining various methods, from gravity to flotation.
In addition, wolframite placers and bedrock ores usually contain other heavy minerals (cassiterite, tantalite-columbite, magnetite, sulfides), therefore, during primary gravity enrichment, a collective concentrate containing from 5 to 20% WO 3 is released. When finishing these collective concentrates, conditioned monomineral concentrates are obtained, for which flotogravity and sulfide flotation, magnetic separation of magnetite and wolframite are used. It is also possible to use electrical separation, enrichment on concentration tables, and even flotation of minerals from displacement rocks.
The high density of tungsten minerals makes it possible to effectively use gravitational enrichment methods for their extraction: in heavy suspensions, on jigging machines, concentration tables, screw and jet separators. During enrichment and especially during finishing of collective gravity concentrates, magnetic separation is widely used. Wolframite has magnetic properties and therefore separates in a strong magnetic field, for example, from non-magnetic cassiterite.
The original tungsten ore, like tin ore, is crushed to a size of minus 12+ 6 mm and enriched by jigging, where coarse wolframite and part of the tailings with a waste content of tungsten trioxide are isolated. After jigging, the ore is crushed into rod mills, in which it is crushed to a particle size of minus 2+ 0.5 mm. To avoid excessive sludge formation, grinding is carried out in two stages. After grinding, the ore is subjected to hydraulic classification with the separation of sludge and enrichment of sand fractions on concentration tables. The industrial products and tailings obtained on the tables are further crushed and sent to the concentration tables. The tailings are also successively further crushed and enriched on concentration tables. Enrichment practice shows that the extraction of wolframite, hübnerite and ferberite by gravitational methods reaches 85%, while scheelite, inclined to sludge, is extracted by gravitational methods only by 55...70%.
When enriching finely disseminated wolframite ores containing only 0.05...0.1% tungsten trioxide, flotation is used.
Flotation is especially widely used to extract scheelite from skarn ores, which contain calcite, dolomite, fluorite and barite, floated by the same collectors as scheelite.
Collectors during flotation of scheelite ores are fatty acid oleic type, which is used at a temperature of at least 18...20°C in the form of an emulsion prepared in soft water. Often, before entering the process, oleic acid is saponified in a hot solution of soda ash at a ratio of 1:2. Instead of oleic acid, tall oil, naphthenic acids, etc. are also used.
It is very difficult to separate scheelite from alkaline earth metal minerals containing calcium, barium and iron oxides by flotation. Scheelite, fluorite, apatite and calcite contain calcium cations in the crystal lattice, which provide chemical sorption of the fatty acid collector. Therefore, selective flotation of these minerals from scheelite is possible within narrow pH limits using depressants such as liquid glass, sodium fluorosilicone, soda, sulfuric and hydrofluoric acid.
The depressive effect of liquid glass during flotation of calcium-containing minerals with oleic acid is the desorption of calcium soaps formed on the surface of the minerals. In this case, the floatability of scheelite does not change, but the floatability of other calcium-containing minerals sharply deteriorates. Increasing the temperature to 80...85°C reduces the contact time of the pulp with the liquid glass solution from 16 hours to 30...60 minutes. Liquid glass consumption is about 0.7 kg/t. The process of selective scheelite flotation, shown in Fig. 220, using a steaming process with liquid glass, is called the Petrov method.
Rice. 220. Scheme of flotation of scheelite from tungsten-molybdenum ores using
finishing according to Petrov's method
The concentrate of the main scheelite flotation, which is carried out at a temperature of 20°C in the presence of oleic acid, contains 4...6% tungsten trioxide and 38...45% calcium oxide in the form of calcite, fluorite and apatite. Before steaming, the concentrate is thickened to 50...60% solid. Steaming is carried out sequentially in two vats in a 3% solution of liquid glass at a temperature of 80...85°C for 30...60 minutes. After steaming, cleaning operations are carried out at a temperature of 20...25°C. The resulting scheelite concentrate can contain up to 63...66% tungsten trioxide with its recovery being 82...83%.
There are several ways to obtain it; the first stage is ore enrichment, separating valuable components from the main mass - waste rock. Enrichment methods are common for heavy ores and metals: grinding and flotation with subsequent operations - magnetic separation (for tungsten ores) and oxidative roasting.The resulting concentrate is most often sintered with an excess of soda to convert tungsten into a soluble compound - sodium wolframite. Another method of obtaining this substance is leaching; tungsten is extracted soda solution under pressure and at elevated temperature (the process takes place in an autoclave), followed by neutralization and precipitation in the form of artificial scheelite, i.e. calcium tungstate. The desire to obtain tungstate is explained by the fact that it is relatively simple to produce, in just two stages:
CaWO4 → H2WO4 or (NH4)2WO4 → WO3,
tungsten oxide, purified from most of the impurities, can be isolated.
Let's look at another way to obtain tungsten oxide - through chlorides. Tungsten concentrate at elevated temperatures they are treated with chlorine gas. The resulting tungsten chlorides are quite easily separated from the chlorides of other metals by sublimation, using the temperature difference at which these substances transform into a vapor state. The resulting tungsten chlorides can be converted into oxide, or they can be processed directly into elemental metal.
Converting oxides or chlorides into metal is the next stage in tungsten production. The best reducing agent for tungsten oxide is hydrogen. Reduction with hydrogen produces the purest tungsten metal. The reduction process takes place in tube furnaces, heated in such a way that as it moves through the tube, the WO3 “boat” passes through several temperature zones. A stream of dry hydrogen comes towards it. Recovery occurs in both “cold” (450...600°C) and “hot” (750...1100°C) zones; in “cold” ones - to the lower oxide WO2, then to the elemental metal. Depending on the temperature and duration of the reaction in the “hot” zone, the purity and grain size of the powdered tungsten released on the walls of the “boat” change.
Reduction can occur not only under the influence of hydrogen. In practice, coal is often used. The use of a solid reducing agent somewhat simplifies production, but in this case a higher temperature is required - up to 1300...1400°C. In addition, coal and the impurities it always contains react with tungsten to form carbides and other compounds. This leads to metal contamination. Meanwhile, electrical engineering needs very pure tungsten. Just 0.1% iron makes tungsten brittle and unsuitable for making the finest wire.
The production of tungsten from chlorides is based on the process of pyrolysis. Tungsten forms several compounds with chlorine. With the help of excess chlorine, all of them can be converted into a higher chloride - WCl6, which decomposes into tungsten and chlorine at 1600°C. In the presence of hydrogen this process it's already underway at 1000°C.
This is how metal tungsten is obtained, but not compact, but in the form of a powder, which is then pressed in a stream of hydrogen at high temperature. At the first stage of pressing (when heated to 1100...1300°C), a porous, brittle ingot is formed. Pressing continues at an even higher temperature, almost reaching the melting point of tungsten at the end. Under these conditions, the metal gradually becomes solid, acquires a fibrous structure, and with it ductility and malleability. Further...
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Navoi Mining and Metallurgical Plant
Navoi State Mining Institute
"Chemical and Metallurgical" Faculty"
Department of Metallurgy
Explanatory note
for final qualifying work
on the topic of: “Selection, justification and calculation of tungsten-molybdenum ore processing technology”
Graduate: K. Sayfiddinov
Navoi-2014
- Introduction
- 1. General information on methods of beneficiation of tungsten ores
- 2. Enrichment of molybdenum-tungsten ores
- 2. Technological section
- 2.1 Calculation of crushing scheme with equipment selection
- 2.2 Calculation of the grinding scheme
- 2.3 Selection and calculation of semi-autogenous grinding mills
- List of used literature
Introduction
Minerals are the basis National economy, and there is not a single industry where minerals or their processed products are not used.
Significant mineral reserves in many deposits of Uzbekistan make it possible to build large, highly mechanized mining, processing and metallurgical enterprises that extract and process many hundreds of millions of tons of minerals with high technical and economic indicators.
The mining industry deals with solid minerals from which, with modern technology, it is advisable to extract metals or other minerals. The main conditions for the development of mineral deposits are increasing their extraction from the subsoil and complex use. This is due to:
- significant material and labor costs during exploration and industrial development of new deposits;
- the growing need of various sectors of the national economy for almost all mineral components that make up the ore;
- the need to create waste-free technology and thereby preventing contamination environment production waste.
For these reasons, the possibility of industrial use of a deposit is determined not only by the value and content of the mineral, its reserves, geographical location, conditions of production and transportation, other economic and political factors, but also the presence effective technology processing of mined ores.
1. General information about methods of beneficiation of tungsten ores
Tungsten ores are enriched, as a rule, in two stages - primary gravity enrichment and finishing of rough concentrates using various methods, which is explained by the low tungsten content in the processed ores (0.2 - 0.8% WO3) and high requirements for the quality of standard concentrates (55 - 65% WO3), The total enrichment degree is approximately 300 - 600.
Wolframite (huebnerite and ferberite) bedrock ores and placers usually contain a number of other heavy minerals, therefore, during the primary gravity enrichment of ores, they strive to isolate collective concentrates, which can contain from 5 to 20% WO3, as well as cassiterite, tantalite-columbite, magnetite, sulfides, etc. When finishing collective concentrates, it is necessary to obtain conditioned monomineral concentrates, for which flotation or flotogravity of sulfides, magnetic separation of magnetite in a weak magnetic field, and wolframite in a stronger one can be used. It is possible to use electric separation, gravitational enrichment on tables, flotation of gangue minerals and other processes to separate minerals so that the finished concentrates meet the requirements of GOSTs and technical specifications not only for the content of the base metal, but also for the content of harmful impurities.
Considering the high density of tungsten minerals (6 - 7.5 g/cm 3), during enrichment gravitational enrichment methods can be successfully used on jigging machines, concentration tables, sluices, jet and screw separators, etc. For fine dissemination of valuable minerals, flotation or a combination is used gravitational processes with flotation. Considering the possibility of wolframite sludge during gravitational enrichment, flotation is used as an auxiliary process even when enriching coarsely disseminated wolframite ores for more complete extraction of tungsten from sludge.
If there are large tungsten-rich ore pieces or large pieces of waste rock in the ore, sorting of ore with a particle size of 150 + 50 mm on belt conveyors can be used to separate the rich large-lump concentrate or pieces of rock that dilute the ore supplied for enrichment.
When beneficiating scheelite ores, gravity is also used, but most often a combination of gravity methods with flotation and flotation gravity, or flotation alone.
When sorting scheelite ores, luminescent installations are used. Scheelite, when irradiated with ultraviolet rays, glows with a bright blue light, which makes it possible to separate pieces of scheelite or pieces of waste rock.
Scheelite is an easily floated mineral characterized by high sludge properties. The extraction of scheelite increases significantly with flotation enrichment compared to gravity, therefore, in the enrichment of scheelite ores in the CIS countries, flotation has now begun to be used in all factories.
During the flotation of tungsten ores, a number of difficult technological problems arise that require the right decision depending on the material composition and association of individual minerals. In the process of flotation of wolframite, hübnerite and ferberite, it is difficult to separate from them iron oxides and hydroxides, tourmaline and other minerals containing neutralize their flotation properties with tungsten minerals.
Flotation of scheelite from ores with calcium-containing minerals (calcite, fluorite, apatite, etc.) is carried out by anionic fatty acid collectors, ensuring their good flotation with calcium cations of scheelite and other calcium-containing minerals. Separation of scheelite from calcium-containing minerals is possible only with the use of such regulators as liquid glass, sodium fluorosilicone, soda, etc.
2. Enrichment of molybdenum-tungsten ores
On Tyrnyauzskaya The factory enriches the molybdenum-tungsten ores of the Tyrnyauz deposit, which are complex in the material composition of not only valuable minerals with very fine dissemination, but also associated gangue minerals. Ore minerals - scheelite (tenths of a percent), molybdenite (hundredths of a percent), powellite, partially ferrimolybdite, chalcopyrite, bismuthite, pyrrhotite, pyrite, arsenopyrite. Nonmetallic minerals - skarns (50-70%), hornfels (21-48%), granite (1 - 12%), marble (0.4-2%), quartz, fluorite, calcite, apatite (3-10%) and etc.
In the upper part of the deposit, 50-60% of molybdenum is represented by powellite and ferrimolybdite, in the lower part their content decreases to 10-20%. Molybdenum is present in scheelite as an isomorphic impurity. Part of the molybdenite, oxidized from the surface, is covered with a film of powellite. Part of the molybdenum grows very finely with molybdoscheelite.
More than 50% of oxidized molybdenum is associated with scheelite in the form of powellite inclusions - a decomposition product of the Ca(W, Mo)O 4 solid solution. Such forms of tungsten and molybdenum can only be isolated into a collective concentrate with subsequent separation by hydrometallurgical methods.
Since 1978, the ore preparation scheme at the factory has been completely reconstructed. Previously, ore, after large crushing at the mine, was transported to the factory in trolleys via an overhead cableway. In the crushing department of the factory, the ore was crushed to - 12 mm, unloaded into bunkers and then crushed in one stage in ball mills operating in a closed cycle with double-spiral classifiers, up to 60% of the class - 0.074 mm.
A new ore preparation technology was developed jointly by the Mekhanobr Institute and the plant and put into operation in August 1978.
The ore preparation scheme provides for coarse crushing of the original ore up to -350 mm, screening according to the 74 mm class, separate storage of each class in bunkers in order to more accurately regulate the supply of large and small classes of ore to the autogenous grinding mill.
Self-grinding of coarse ore (-350 mm) is carried out in Cascade type mills with a diameter of 7 m (MMC-70X X23) with additional grinding of the coarse-grained fraction to 62% class -0.074 mm in MSHR-3600X5000 mills operating in a closed cycle with single-spiral classifiers 1KSN-3 and located in a new building on the mountainside at an elevation of about 2000 m above sea level between the mine and the operating factory.
Innings finished product from the autogenous vessel to flotation is carried out by hydraulic transport. The hydraulic transport route is a unique engineering structure, ensuring the transportation of pulp with a height difference of more than 600 m. It consists of two pipelines with a diameter of 630 mm, a length of 1750 m, equipped with stilling wells with a diameter of 1620 mm and a height of 5 m (126 wells for each pipeline).
The use of a hydraulic transport system made it possible to eliminate the cargo workshop cable cars, medium and fine crushing building, MSHR-3200X2100 mills at the processing plant. In the main building of the factory, two main flotation sections, new scheelite and molybdenum finishing departments, a liquid glass melting shop, and recycling water supply systems were built and put into operation. The thickening front for rough flotation concentrates and middlings has been significantly expanded due to the installation of thickeners with a diameter of 30 m, which reduces losses from thickening discharges.
The newly commissioned facilities are equipped with modern automated process control systems and local automation systems. Thus, in the autogenous building the automatic control system operates in direct control mode based on M-6000 computers. In the main building, a system for centralized control of the material composition of the pulp was introduced using X-ray spectral analyzers KRF-17 and KRF-18 in combination with an M-6000 computer. An automated system for sampling and delivery of samples (by pneumatic mail) to the express laboratory, controlled by the KM-2101 computer complex and issuing analyzes by teletype, has been mastered.
One of the most complex processing processes - finishing rough scheelite concentrates according to the method of N. S. Petrov - is equipped with an automatic monitoring and control system, which can work either in the “advisor” mode to the flotation operator, or in the mode of direct control of the process, regulating the flow rate of the suppressor (liquid glass), pulp level in cleaning operations and other process parameters.
The sulfide minerals flotation cycle is equipped with automatic control and dosing systems for collector (butyl xanthate) and suppressor (sodium sulfide) in the copper-molybdenum flotation cycle. The systems operate using ion-selective electrodes as sensors.
Due to the increase in production volume, the factory switched to processing new varieties of ores, characterized by a lower content of certain metals and a higher degree of oxidation. This required improvement of the reagent regime for flotation of sulfide-oxidized ores. In particular, a progressive technological solution was used in the sulfide cycle - a combination of two foaming agents of active and selective types. Reagents containing terpene alcohols are used as an active foaming agent, and a new reagent LV, developed for the enrichment of multicomponent ores, primarily Tyrnyauz ores, is used as a selective agent.
In the flotation cycle of oxidized minerals by fatty acid collectors, intensifying additives of a modifier reagent based on low molecular weight carboxylic acids are used. To improve the flotation properties of circulating industrial products pulp, regulation of their ionic composition has been introduced. Methods of chemical finishing of concentrates have found wider application.
From the autogenous grinding mill, the ore is sent to screening. Class +4 mm is further ground in a ball mill. Mill overflow and under-screen product (--4 mm) are subject to I and II classifications.
690 g/t soda and 5 g/t transformer oil are fed into the ball mill. The classifier discharge goes to the main molybdenum flotation, where 0.5 g/t xanthate and 46 g/t terpineol are fed. After I and II cleaning flotations, the molybdenum concentrate (1.2-1.5% Mo) is subjected to steaming with liquid glass (12 g/t) at 50-70°C, III cleaning flotation and further grinding to 95-98% class --0.074 mm with a supply of 3 g/t sodium cyanide and 6 g/t liquid glass.
The finished molybdenum concentrate contains about 48% Mo, 0.1% Cu and 0.5% WO 3 with a Mo extraction of 50%. The control flotation tailings of the III and IV cleaning operations are thickened and sent to copper-molybdenum flotation with a supply of 0.2 g/t xanthate and 2 g/t kerosene. The twice purified copper-molybdenum concentrate, after steaming with sodium sulfide, is sent to selective flotation, where a copper concentrate containing 8-10% Cu (with an extraction of about 45%), 0.2% Mo, 0.8% Bi is isolated.
The tailings of the control molybdenum flotation, containing up to 0 2% WO 3, are sent to scheelite flotation, which is carried out according to a very branched and complex scheme. After mixing with liquid glass (350 g/t), basic scheelite flotation is carried out with sodium oleate (40 g/t). After the first cleaning flotation and thickening to 60% solid, the scheelite concentrate is steamed with liquid glass (1600 g/t) at 80--90 °C. Next, the concentrate is cleaned twice more and again goes to steaming at 90--95 ° C with liquid glass (280 g/t) and is cleaned again three times.
2. Technological section
2.1 Calculation of crushing scheme with equipment selection
The designed concentration plant is intended for processing molybdenum-containing tungsten ores.
Medium-sized ore (f = 12 ± 14 units on Professor Protodyakonov’s scale) is characterized by a density c = 2.7 t/m 3 and is supplied to the factory with a moisture content of 1.5%. Maximum piece d=1000 mm.
In terms of productivity, the enrichment plant belongs to the category of medium productivity (Table 4/2/), according to the international classification - to group C.
To the factory ore D max. =1000 mm is supplied from open-pit mining.
1. Let's determine the productivity of the coarse crushing shop. We calculate productivity according to Razumov K.A. 1, pp. 39-40. The project adopted the delivery of ore 259 days a year, in 2 shifts of 7 hours, 5 days a week.
Ore strength factor /2/
where: Q c. etc. - daily productivity of the crushing shop, t/day
Coefficient taking into account the uneven properties of raw materials /2/
where: Q h..t. dr - hourly productivity of the crushing shop, t/h
k n - coefficient taking into account the uneven properties of raw materials,
n days - estimated number of working days per year,
n cm - number of shifts per day,
t cm - shift duration,
k" - coefficient for accounting for ore strength,
Calculation of annual working hours:
C = (n day n cm t cm) = 259 2 5 = 2590 (3)
Time utilization rate:
k in = 2590/8760 = 0.29 units = 29%
2. Calculation of crushing scheme. We carry out the calculation according to pp. 68-78 2.
According to the instructions, the moisture content of the initial ore is 1.5%, i.e. e.
Calculation procedure:
1. Determine the degree of fragmentation
2. Let us accept the degree of fragmentation.
3. Let’s determine the maximum size of products after crushing:
4. Let's determine the width of the crusher's discharge slots, taking the typical characteristics Z - coarsening of the crushed product relative to the size of the discharge slot.
5. Let’s check the compliance of the selected crushing scheme with the manufactured equipment.
The requirements that crushers must satisfy are listed in Table 1.
Table 1
In terms of the width of the receiving opening and the range of adjustment of the discharge slot, crushers of the ShchDP 12X15 brand are suitable.
Let's calculate the productivity of the crusher using the formula (109/2/):
Q cat. = m 3 / h
Q fraction. = Q cat. · with n · k f · k cr. · k ow. · k c, m 3 / h (7)
where c n is the bulk density of ore = 1.6 t/m 3,
Q cat. - passport capacity of the crusher, m 3 / h
k f . , k ow. , kcr, kc - correction factors for strength (crushingability), bulk density, ore size and moisture content.
The value of the coefficients is found from the table k f =1.6; k cr =1.05; k ow. =1%;
Q cat. = S pr. / S n · Q n = 125 / 155 · 310 ? 250 m 3 /h
Let's find the actual productivity of the crusher for the conditions defined by the project:
Q fraction. = 250 · 1.6 · 1.00 · 1.05 · 1 · 1 = 420 t/h
Based on the calculation results, we determine the number of crushers:
We accept 12 x 15 boards for installation - 1 pc.
2.2 Calculation of the grinding scheme
The grinding scheme chosen in the project is a type of VA Razumov K.A. page 86.
Calculation procedure:
1. Determining the hourly productivity of the grinding shop , which is actually the hourly productivity of the entire factory, since the grinding shop is the main ore preparation building:
where 343 is the number of working days in a year
24 - continuous work week 3 shifts of 8 hours (3x8=24 hours)
Kv - equipment utilization factor
Kn - coefficient taking into account the uneven properties of raw materials
We accept: K in =0.9 K n =1.0
The coarse ore warehouse provides a two-day supply of ore:
V= 48,127.89 / 2.7 = 2398.22
We accept the initial data
Let's ask ourselves about liquefaction in plums and sands classification:
R 10 =3 R 11 =0.28
(R 13 is based on row 2 p. 262 depending on the size of the drain)
in 1 -0.074 =10% - class content - 0.074 mm in crushed ore
in 10 -0.074 =80% - class content - 0.074 mm in the classification plum.
We accept the optimal circulation load With opt = 200%.
Calculation procedure:
Grinding stages I and II are represented by a type VA scheme, page 86 fig. 23.
Calculation of scheme B comes down to determining the weights of products 2 and 5 (product yields are found according to general formula g n = Q n: Q 1)
Q 7 = Q 1 C opt = 134.9 · 2 = 269.8 t/h;
Q 4 = Q 5 = Q 3 + Q 7 = 404.7 t/h;
g 4 = g 5 = 300%;
g 3 = g 6 = 100%
The calculation is carried out according to Razumov K.A. 1 pp. 107-108.
1. Calculation of scheme A
Q 8 = Q 10 ; Q 11 = Q 12 ;
Q 9 = Q 8 + Q 12 = 134.88 + 89.26 = 224.14 t/h
g 1 = 100%; g 8 = g 10 = 99.987%;
g 11 = g 12 =Q 12: Q 1 = 89.26: 134.88 = 66.2%;
g 9 = Q 9: Q 1 = 224.14: 134.88 = 166.17%
Process flow diagramschleniyamolybdenum-tungsten ores.
CalculationByqualitative-quantitative scheme.
Initial data for calculating qualitative-quantitative schemess.
Extraction of tungsten into the final concentrate - e tungsten 17 = 68%
Extraction of tungsten into collective concentrate - e tungsten 15 =86%
Extraction of tungsten into molybdenum concentrate - e tungsten 21 = 4%
Extraction of molybdenum into the final concentrate - e Mo 21 = 77%
Extraction of molybdenum into tungsten flotation tailings - e Mo 18 =98%
Extraction of molybdenum into control flotation concentrate - eMo 19 =18%
Extraction of molybdenum into collective concentrate - e Mo 15 = 104%
Yield of collective concentrate - g 15 = 36%
Yield of tungsten concentrate - g 17 = 14%
Yield of molybdenum concentrate - g 21 = 15%
Yield of control flotation concentrate - g 19 =28%
Determining the yield of enrichment products
G 18 = g 15 - G 17 =36-14=22%
G 22 = g 18 - G 21 =22-15=7%
G 14 = g 13 + g 19 + g 22 =100+28+7=135%
G 16 = g 14 - G 15 =135-36=99%
G 20 = g 16 - G 19 =99-28=71%
Determining the masses of enrichment products
Q 13 = 127.89t/h.
Q 1 4 = Q 13 XG 14 = 127.89x1.35=172.6 t/h
Q 1 5 = Q 13 XG 15 = 127.89x0.36=46.0 t/h
Q 1 6 = Q 13 XG 16 = 127.89x0.99=126.6t/h
Q 1 7 = Q 13 XG 17 = 127.89x0.14=17.9 t/h
Q 1 8 = Q 13 XG 18 = 127.89x0.22=28.1 t/h
Q 1 9 = Q 13 XG 19 = 127.89x0.28=35.8 t/h
Q 20 = Q 13 XG 20 = 127.89x0.71=90.8 t/h
Q 21 = Q 13 XG 21 = 127.89x0.15=19.1 t/h
Q 22 = Q 13 XG 22 = 127.89x0.07=8.9 t/h
Determining the recovery of enrichment products
For tungsten
e tungsten 13 =100 %
e tungsten 18 = e tungsten 15 - e tungsten 17 =86-68=28 %
e tungsten 22 = e tungsten 18 - e tungsten 21 =28-14=14 %
e tungsten 14 = e tungsten 13 + e tungsten 22 + e tungsten 19 =100+14+10=124 %
e tungsten 16 = e tungsten 14 - e tungsten 15 =124-86=38%
e tungsten 20 = e tungsten 13 - e tungsten 17 + e tungsten 21 =100 - 68+4=28%
e tungsten 19 = e tungsten 16 - e tungsten 20 =38-28=10 %
for molybdenum
e Mo 13 =100%
e Mo 22 = e Mo 18 - e Mo 21 =98-77=11 %
e Mo 14 = e Mo 13 + e Mo 22 + e Mo 19 =100+11+18=129 %
e Mo 16 = e Mo 14 - e Mo 15 =129-94=35 %
e Mo 17 = e Mo 15 - e Mo 18 =104-98=6%
e Mo 20 = e Mo 13 - e Mo 17 + e Mo 21 =100 - 6+77=17%
e Mo 19 = e Mo 16 - e Mo 20 =35-17=18%
Determining the amount of metals in the product Oh enrichment
For tungsten
14 =124 x0.5 / 135=0.46%
15 =86x0.5 / 36=1.19%
16 =38 x0.5 / 99=0.19%
17 =68 x0.5 / 14=2.43%
18 =28 x0.5 / 22=0.64%
19 =10 x0.5 / 28=0.18%
20 =28 x0.5 / 71=0.2%
21 =14 x0.5 / 15=0.46%
22 =14 x0.5 / 7=1%
For molybdenum
14 =129 x0.04/ 135=0.04%
15 =94x0.04/ 36=0.1%
16 =35 x0.04 / 99=0.01%
17 =6 x0.04 / 14=0.017%
18 =98 x0.04 / 22=0.18%
19 =18 x0.04 / 28=0.025%
20 =17 x0.04 / 71=0.009%
21 =77 x0.04 / 15=0.2%
22 =11 x0.04 / 7=0.06%
Table 3. Table of qualitative-quantitative enrichment scheme
|
Operation no. cont. |
|
Q, t/h |
, % |
copper , % |
copper , % |
zinc , % |
zinc , % |
|
|
I |
Grinding stage I |
|||||||
|
arrives |
||||||||
|
crushed ore |
||||||||
|
comes out |
||||||||
|
crushed ore |
||||||||
|
II |
Classification |
|||||||
|
arrives |
||||||||
|
CrushedbChennsth product IArt. grinding |
||||||||
|
CrushedbChennsth product II st .grinding |
||||||||
|
comes out |
||||||||
|
drain |
||||||||
|
sands |
||||||||
|
III |
Grinding I I stage |
|||||||
|
arrives |
||||||||
|
Sands classification |
||||||||
|
comes out |
||||||||
|
Shreddedsth product |
||||||||
|
IV |
Collective Wo 3 -Mo flotation |
|||||||
|
arrives |
||||||||
|
Drain classification |
||||||||
|
|
||||||||
|
TailsMo flotationAnd |
||||||||
|
comes out |
||||||||
|
concentrate |
||||||||
|
tails |
||||||||
|
V |
Control flotation |
|||||||
|
arrives |
||||||||
|
Tailcollective flotation |
||||||||
|
comes out |
||||||||
|
concentrate |
||||||||
|
tails |
||||||||
|
VI |
Tungsten flotation |
|||||||
|
arrives |
||||||||
|
Concentratecollective flotation |
||||||||
|
comes out |
||||||||
|
concentrate |
||||||||
|
tails |
||||||||
|
Mo flotation |
||||||||
|
arrives |
||||||||
|
Tails Wo 3 flotation |
||||||||
|
comes out |
||||||||
|
concentrate |
||||||||
|
tails |
Calculation of water-sludge scheme .
The purpose of calculating the water-sludge scheme is to: ensure optimal liquid: solid ratios in the operations of the scheme; determining the amount of water added to operations or, conversely, released from products during dehydration operations; determination of L:T ratios in the products of the scheme; determination of the total water requirement and specific water consumption per ton of processed ore.
To obtain high technological indicators of ore processing, each operation of the technological scheme must be carried out at optimal values of the L:T ratio. These values are established based on data from ore dressing tests and the operating practices of existing processing plants.
The relatively low specific water consumption per ton of processed ore is explained by the presence of intra-factory water circulation at the designed plant, since the thickener drains are fed into the grinding - classification cycle. Water consumption for flushing floors, washing equipment and for other purposes is 10-15% of the total consumption.
Table 3. Table of qualitative-quantitative enrichment scheme.
|
Opera no.walkie-talkies cont. |
Name of operations and products |
Q, t/h |
, % |
R |
W |
|
|
I |
Grinding stage I |
|||||
|
arrives |
||||||
|
crushed ore |
0 , 0 25 |
|||||
|
comes out |
||||||
|
crushed ore |
||||||
|
II |
Classification |
|||||
|
arrives |
||||||
|
CrushedbChennsth product IArt. grinding |
||||||
|
CrushedbChennsth product II st .grinding |
||||||
|
comes out |
||||||
|
drain |
||||||
|
sands |
||||||
|
III |
Grinding I I stage |
|||||
|
arrives |
||||||
|
Sands classification |
||||||
|
comes out |
||||||
|
Shreddedsth product |
||||||
|
IV |
Collective Wo 3 -Mo flotation |
|||||
|
arrives |
||||||
|
Drain classification |
||||||
|
Control flotation concentrate |
||||||
|
Tails Mo flotationAnd |
||||||
|
comes out |
||||||
|
concentrate |
||||||
|
Tails |
||||||
|
V |
Control flotation |
|||||
|
arrives |
||||||
|
Tailcollective flotation |
||||||
|
comes out |
||||||
|
concentrate |
||||||
|
Tails |
||||||
|
VI |
Tungsten flotation |
|||||
|
Incoming |
||||||
|
Concentratecollective flotation |
||||||
|
It turns out |
||||||
|
Concentrate |
||||||
|
Tails |
||||||
|
Mo flotation |
||||||
|
Incoming |
||||||
|
Tails tungstenflotation |
||||||
|
It turns out |
||||||
|
concentrate |
||||||
|
tails |
Crusher selection and calculation.
The choice of crusher type and size depends on physical properties ore, required crusher capacity, crushed product size and ore hardness.
Tungsten-molybdenum ore by strength category is an ore of medium strength.
The maximum size of a piece of ore entering the crushing operation is 1000 mm.
To crush the ore coming from the mine, I install a jaw crusher with a simple swing jaw ShchDP 12x15. *
Crusher productivity, Q is equal to:
Q =q*L*i, t/h,
where q - specific productivity of the jaw crusher per 1 cm 2 of the discharge slot area, t/(cm 2 * h);
L is the length of the discharge slot of the neck crusher, cm;
i - width of the unloading slot, see /4/
According to the experience of the crushing department processing plant The specific productivity of the jaw crusher is 0.13 t/cm2 * hour.
The productivity of a jaw crusher will be determined by:
Q= 0.13*150*15.5 = 302.25 t/h.
The crusher accepted for installation provides the specified ore productivity.
The maximum size of a piece in the crusher feed will be:
120*0.8 = 96 cm.
Selection and calculation of grate screen
A grate screen with a hole size of 95 cm (950 mm) is installed in front of the crusher.
The required screening area is determined by the formula:
where Q* - productivity, t/h;
a is a coefficient equal to the width of the gap between the grates, mm. /5/ According to the layout conditions, the width of the grate screen is taken to be 2.7 m, length 4.5 m.
The practice of the crushing department of the factory shows that the ore delivered from the quarry contains about 4.5% of pieces with a particle size of more than 950 mm. Pieces of this size are delivered by a front-end loader to the ore yard, where they are crushed and again fed by the loader to the grate screen.
2.3 Selection and calculation of semi-autogenous grinding mills
IN Lately during processing gold ores In world and domestic practice, in the first stage of grinding, semi-autogenous grinding mills with subsequent cyanidation are becoming increasingly common. In this case, the loss of gold from iron scrap and crumbs is eliminated, the consumption of cyanide during cyanidation is reduced, and the sanitary conditions of working on quartz silicate ores are improved. Therefore, I accept a semi-autogenous grinding (SAG) mill for installation in the first stage of grinding.
1. Find the specific productivity for the newly formed class of the operating SSI mill, t/(m 3 * h):
where Q is the productivity of the operating mill, t/h;
- class content -0.074 mm in the mill discharge, %;
- class content -0.074 mm in the original product,%;
D is the diameter of the operating mill, m;
L is the length of the operating mill, m.
2. We determine the specific productivity of the designed mill according to the newly formed class:
where q 1 is the specific productivity of a working mill in the same class;
K and is a coefficient that takes into account differences in the grindability of the ore designed for processing and the ore being processed (Ki = 1);
K k - coefficient taking into account the difference in the size of the initial and final grinding products at the existing and designed factories (K k = 1);
K D is a coefficient that takes into account the difference in the diameters of the drums of the designed and operating mills:
K D = ,
where D and D 1 respectively, the nominal diameters of the drums of the mills being designed for installation and those in operation. (K D =1.1);
Kt is a coefficient that takes into account differences in the type of designed and operating mills (Kt=1).
q = 0.77*1*1*1.1*1 =0.85 t/(m 3 * h).
I accept for installation an autogenous grinding mill "Cascade" with a diameter of 7 m and a length of 2.3 m with a working volume of 81.05 m3
3. We determine the productivity of the mills for ore using the formula:
where V is the working volume of the mill. /4/
4. Determine the estimated number of mills:
n- 101/125.72 = 0.8;
then the accepted one will be equal to 1. The Cascade mill provides the specified productivity.
Screen selection and calculation II screening stage .
Draining of semi-autogenous mills using pumps...
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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 remoteness of the manufacturing plant and, as a consequence, for a long time waiting for spare parts.
In search of an alternative method of 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. top scores 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
|
Indicators |
Initial food |
Concentrate |
|||
|
Recovery, % |
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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%),