Technology for processing tungsten ores. Extraction of weakly magnetic minerals using a high-intensity magnetic separator from ores of non-ferrous, rare earth and noble metals using the example of JSC Irgiredmet, Kovdor Mining and Processing Plant. for final qualifying work
Tungsten minerals, ores and concentrates
Tungsten is a rare element, its average content in earth's crust 10-4% (by mass). About 15 tungsten minerals are known, but only minerals of the wolframite and scheelite group are of practical importance.
Wolframite (Fe, Mn)WO4 is an isomorphic mixture (solid solution) of iron and manganese tungstates. If the mineral contains more than 80% iron tungstate, the mineral is called ferberite; if manganese tungstate predominates (more than 80%), it is called hübnerite. Mixtures lying in composition between these limits are called wolframites. Minerals of the wolframite group are colored black or brown and have a high density (7D-7.9 g/cm3) and a hardness of 5-5.5 on the mineralogical scale. The mineral contains 76.3-76.8% W03. Wolframite is weakly magnetic.
Scheelite CaWOA is calcium tungstate. The color of the mineral is white, gray, yellow, brown. Density 5.9-6.1 g/cm3, hardness on the mineralogical scale 4.5-5. Scheelite often contains an isomorphic admixture of powellite - CaMoO4. When irradiated ultraviolet rays Sheelite fluoresces blue-blue light. When the molybdenum content is more than 1%, the fluorescence becomes yellow. Scheelite is non-magnetic.
Tungsten ores are usually low in tungsten. The minimum W03 content in ores at which their exploitation is profitable is currently 0.14-0.15% for large deposits and 0.4-0.5% for small deposits.
Along with tungsten minerals, molybdenite, cassiterite, pyrite, arsenopyrite, chalcopyrite, tantalite or columbite, etc. are found in ores.
Based on the mineralogical composition, two types of deposits are distinguished - wolframite and scheelite, and based on the shape of ore formations - vein and contact types.
Tungsten minerals in vein deposits for the most part lie in quartz veins small thickness (0.3-1 m). The contact type of deposits is associated with contact zones of granite rocks and limestones. They are characterized by deposits of sheelite-bearing skarn (skarns are silicified limestones). Skarn-type ores include the largest Tyrn-Auz deposit in the USSR in the North Caucasus. When vein deposits are weathered, wolframite and scheelite accumulate, forming placers. In the latter, wolframite is often combined with cassiterite.
Tungsten ores are enriched, producing standard concentrates containing 55-65% W03. A high degree of enrichment of wolframite ores is achieved using various methods: gravity, flotation, magnetic and electrostatic separation.
When enriching scheelite ores, gravity-flotation or pure flotation schemes are used.
The extraction of tungsten into standard concentrates during the enrichment of tungsten ores ranges from 65-70% to 85-90%.
When enriching ores of complex composition or difficult to enrich, it is sometimes economically advantageous to remove middling products containing 10-20% W03 from the enrichment cycle for chemical (hydrometallurgical) processing, which results in the production of “artificial scheelite” or technical tungsten trioxide. Such combined schemes ensure high extraction of tungsten from ores.
The state standard (GOST 213-73) provides for the content of W03 in tungsten concentrates of the 1st grade not lower than 65%, of the 2nd grade - not lower than 60%. The content of impurities P, S, As, Sn, Cu, Pb, Sb, Bi is limited in them, ranging from hundredths of a percent to 1.0%, depending on the grade and purpose of the concentrate.
The explored reserves of tungsten as of 1981 are estimated at 2903 thousand tons, of which 1360 thousand tons are in China. The USSR, Canada, Australia, the USA, South and North Korea, Bolivia, Brazil, Portugal. Production of tungsten concentrates in capitalist and developing countries in the period 1971 - 1985 fluctuated between 20 - 25 thousand tons (in terms of metal content).
Methods for processing tungsten concentrates
The main product of direct processing of tungsten concentrates (in addition to ferrotungsten smelted for the needs of ferrous metallurgy) is tungsten trioxide. It serves as the starting material for tungsten and tungsten carbide - the main component of hard alloys.
Production schemes for processing tungsten concentrates are divided into two groups depending on the adopted decomposition method:
Tungsten concentrates are sintered with soda or treated with aqueous soda solutions in autoclaves. Tungsten concentrates are sometimes decomposed with aqueous solutions of sodium hydroxide.
Concentrates are decomposed with acids.
In cases where alkaline reagents are used for decomposition, solutions of sodium tungstate are obtained, from which, after purification from impurities, the final products are produced - ammonium paratungstate (PVA) or tungstic acid. 24
When the concentrate is decomposed with acids, a precipitate of technical tungstic acid is obtained, which is purified from impurities in subsequent operations.
Decomposition of tungsten concentrates. alkaline reagents Sintering with Na2C03
Sintering of wolframite with Na2C03. The interaction of wolframite with soda in the presence of oxygen actively occurs at 800-900 C and is described by the following reactions: 2FeW04 + 2Na2C03 + l/202 = 2Na2W04 + Fe203 + 2C02; (l) 3MnW04 + 3Na2C03 + l/202 = 3Na2W04 + Mn304 + 3C02. (2)
These reactions occur with a large decrease in the Gibbs energy and are practically irreversible. With the ratio in wolframite FeO:MnO = i:i AG°1001C = -260 kJ/mol. With an excess of Na2C03 in the charge of 10-15% above the stoichiometric amount, complete decomposition of the concentrate is achieved. To accelerate the oxidation of iron and manganese, 1-4% nitrate is sometimes added to the mixture.
Sintering of wolframite with Na2C03 at domestic enterprises is carried out in tubular rotary kilns lined with fireclay bricks. In order to avoid melting of the charge and the formation of accretions (accumulations) in zones of the furnace with a lower temperature, tailings from the leaching of cakes (containing iron and manganese oxides) are added to the charge, reducing the W03 content in it to 20-22%.
A furnace with a length of 20 m and an outer diameter of 2.2 m at a rotation speed of 0.4 rpm and an inclination angle of 3 has a charge capacity of 25 tons/day.
The components of the charge (crushed concentrate, Na2C03, saltpeter) are fed from bins into a screw mixer using automatic scales. The charge enters the furnace hopper, from which it is fed into the furnace. Upon exiting the furnace, cake pieces pass through crushing rolls and a wet grinding mill, from which the pulp is directed to a higher laminator (Fig. 1).
Sintering of scheelite with Na2C03. At temperatures of 800-900 C, the interaction of scheelite with Na2C03 can proceed through two reactions:
CaW04 + Na2CQ3 Na2W04 + CaC03; (1.3)
CaW04 + Na2C03 *=*■ Na2W04 + CaO + C02. (1.4)
Both reactions proceed with a relatively small change in the Gibbs energy.
Reaction (1.4) occurs to a noticeable extent above 850 C, when decomposition of CaCO3 is observed. The presence of calcium oxide in the cake leads to the formation of poorly soluble calcium tungstate when leaching the cake with water, which reduces the extraction of tungsten into the solution:
Na2W04 + Ca(OH)2 = CaW04 + 2NaOH. (1.5)
With a large excess of Na2C03 in the charge, this reaction is significantly suppressed by the interaction of Na2C04 with Ca(OH)2 with the formation of CaCO3.
To reduce the consumption of Na2C03 and prevent the formation of free calcium oxide, quartz sand is added to the charge to bind calcium oxide into poorly soluble silicates:
2CaW04 + 2Na2C03 + Si02 = 2Na2W04 + Ca2Si04 + 2C02;(l.6) AG°100IC = -106.5 kJ.
Still, in this case, to ensure a high degree of tungsten extraction into the solution, it is necessary to introduce a significant excess of Na2C03 into the charge (50-100% of the stoichiometric amount).
Sintering of the scheelite concentrate charge with Na2C03 and quartz sand is carried out in drum furnaces, as described above for wolframite at 850-900 °C. To prevent melting, leaching dumps (containing mainly calcium silicate) are added to the charge to reduce the W03 content to 20-22%.
Leaching of soda speco. When cakes are leached with water, sodium tungstate and soluble impurity salts (Na2Si03, Na2HP04, Na2HAs04, Na2Mo04, Na2S04), as well as excess Na2C03, pass into the solution. Leaching is carried out at 80-90 °C in steel reactors with mechanical stirring, operating in hierarchical conditions.
Concentrates with soda:
Elevator feeding concentrate to the mill; 2 - ball mill operating in a closed cycle with an air separator; 3 - auger; 4 - air separator; 5 - bag filter; 6 - automatic weighing dispensers; 7 - transport screw; 8 - screw mixer; 9 - charge hopper; 10 - feeder;
Drum oven; 12 - roll crusher; 13 - rod mill - lixiviant; 14 - reactor with stirrer
Wild mode, or drum rotating leaches of continuous operation. The latter are filled with crushing rods to crush pieces of cake.
The recovery of tungsten from the sinter into the solution is 98-99%. Strong solutions contain 150-200 g/l W03.
Autoclave is the only way to decompose tungsten concentrates
The autoclave-soda method was proposed and developed in the USSR1 in relation to the processing of scheelite concentrates and industrial products. Currently, the method is used at a number of domestic factories and in foreign countries.
The decomposition of scheelite with Na2C03 solutions is based on the exchange reaction
CaW04CrB)+Na2C03(pacTB)^Na2W04(pacTB)+CaC03(TB). (1.7)
At 200-225 °C and a corresponding excess of Na2C03, depending on the composition of the concentrate, decomposition proceeds with sufficient speed and completeness. The concentration equilibrium constants of reaction (1.7) are small, increase with temperature and depend on the soda equivalent (i.e., the number of moles of Na2C03 per 1 mole of CaW04).
With a soda equivalent of 1 and 2 at 225 C, the equilibrium constant (Kc = C / C cq) is 1.56 and
0.99 respectively. It follows from this that at 225 C the minimum required soda equivalent is 2 (i.e., the excess of Na2C03 is 100%). The real excess of Na2C03 is higher, since as equilibrium is approached the rate of the process slows down. For scheelite concentrates containing 45-55% W03 at 225 C, a soda equivalent of 2.6-3 is required. For industrial products containing 15-20% W03, 4-4.5 moles of Na2C03 per 1 mole of CaW04 are required.
The CaCO3 films formed on scheelite particles are porous and up to a thickness of 0.1-0.13 mm, their influence on the rate of decomposition of scheelite by Na2C03 solutions was not detected. With intense stirring, the rate of the process is determined by the rate of the chemical stage, which is confirmed by the high value of the apparent activation energy E = 75+84 kJ/mol. However, if the mixing speed is insufficient (which
Occurs in horizontal rotating autoclaves), an intermediate regime is realized: the rate of the process is determined by both the rate of supply of the reagent to the surface and the rate of chemical interaction.
0.2 0.3 0, it 0.5 0.5 0.7 0.8 Ш gШШУШгС031
As can be seen from Fig. 2, the specific reaction rate decreases approximately inversely with the increase in the ratio of molar concentrations of Na2W04:Na2C03 in solution. This
Cassock. Fig. 2. Dependence of the specific rate of decomposition of scheelite by soda solution in autoclave j on the molar ratio of Na2W04/Na2C03 concentrations in the solution at
Determines the need for a significant excess of Na2C03 against the minimum required, determined by the value of the equilibrium constant. To reduce the consumption of Na2C03, two-stage countercurrent leaching is carried out. In this case, the tailings after the first leaching, which contain little tungsten (15-20% of the original), are treated with a fresh solution containing a large excess of Na2C03. The resulting solution, which is recycled, enters the first stage of leaching.
Decomposition with Na2C03 solutions in autoclaves is also used for wolframite concentrates, but the reaction in this case is more complicated, as it is accompanied by hydrolytic decomposition of iron carbonate (manganese carbonate is only partially hydrolyzed). The decomposition of wolframite at 200-225 °C can be represented by the following reactions:
MnW04(TB)+Na2C03(paCT)^MiiC03(TB)+Na2W04(paCTB); (1.8)
FeW04(TB)+NaC03(pacT)*=iFeC03(TB)+Na2W04(paCTB); (1.9)
FeC03 + HjO^FeO + H2C03; (1.10)
Na2C03 + H2C03 = 2NaHC03. (l.ll)
The resulting iron oxide FeO at 200-225 °C undergoes a transformation according to the reaction:
3FeO + H20 = Fe304 + H2.
The formation of sodium bicarbonate leads to a decrease in the concentration of Na2C03 in the solution and requires a large excess of the reagent.
To achieve satisfactory decomposition rates of wolframite concentrates, it is necessary to finely grind them and increase the consumption of Na2C03 to 3.5-4.5 g-eq, depending on the composition of the concentrate. High-manganese wolframites are more difficult to decompose.
Adding NaOH or CaO to the autoclave pulp (which leads to causticization of Na2C03) improves the degree of decomposition.
The rate of decomposition of wolframite can be increased by introducing oxygen (air) into the autoclave pulp, which oxidizes Fe (II) and Mil (II), which leads to the destruction of the crystal lattice of the mineral on the reacting surface.
Secondary steam
Cassock. 3. Autoclave installation with a horizontally rotating autoclave: 1 - autoclave; 2 - loading pipe for pulp (steam is also introduced through it); 3 - pulp pump; 4 - pressure gauge; 5 - reactor-pulp heater; 6 - self-evaporator; 7 - droplet separator; 8 - pulp input into the self-evaporator; 9 - bumper made of armored steel; 10 - pipe for pulp removal; 11 - pulp collection
Leaching is carried out in steel horizontal rotating autoclaves heated with live steam (Fig. 3) and continuous vertical autoclaves with pulp mixing using bubbling steam. Approximate process mode: temperature 225 pressure in the autoclave ~2.5 MPa, T:L ratio = 1:(3.5*4), duration at each stage 2-4 hours.
Figure 4 shows a diagram of a battery of autoclaves. The initial autoclave pulp, heated by steam to 80-100 °C, is pumped into autoclaves, in which it is heated to
Secondary steam
Rve. 4. Scheme of a continuous autoclave installation: 1 - reactor for heating the initial pulp; 2 - piston pump; 3 - autoclave; 4 - throttle; 5 - self-evaporator; 6 - pulp collector
200-225 °C with live steam. During continuous operation, the pressure in the autoclave is maintained by releasing the pulp through a choke (a calibrated carbide washer). The pulp enters a self-evaporator - a vessel under a pressure of 0.15-0.2 MPa, where the pulp is rapidly cooled due to intense evaporation. The advantages of autoclave-soda decomposition of scheelite concentrates before sintering are the elimination of the furnace process and a slightly lower content of impurities in tungsten solutions (especially phosphorus and arsenic).
The disadvantages of this method include the high consumption of Na2C03. A high concentration of excess Na2C03 (80-120 g/l) entails an increased consumption of acids to neutralize solutions and, accordingly, high costs for the disposal of waste solutions.
Decomposition of tungstate conce n irate solutions and sodium hydroxide
Sodium hydroxide solutions decompose wolframite according to the exchange reaction:
Me WC>4 + 2Na0Hi=tNa2W04 + Me(0 H)2, (1.13)
Where Me is iron, manganese.
The value of the concentration constant of this reaction Kc = 2 at temperatures of 90, 120 and 150 °C is 0.68, respectively; 2.23 and 2.27.
Complete decomposition (98-99%) is achieved by treating the finely ground concentrate with a 25-40% sodium hydroxide solution at 110-120 °C. The required excess of alkali is 50% or higher. The decomposition is carried out in sealed steel reactors equipped with stirrers. Passing air into the solution accelerates the process due to the oxidation of iron (II) hydroxide Fe(OH)2 into hydrated iron (III) oxide Fe2O3-NH20 and manganese (II) hydroxide Mn(OH)2 into hydrated manganese oxide (IV) Mn02-lH20 .
The use of decomposition with alkali solutions is advisable only for high-grade wolframite concentrates (65-70% W02) with a small content of silica and silicates. When processing low-grade concentrates, highly contaminated solutions and difficult-to-filter sediments are obtained.
Processing of sodium tungstate solutions
Solutions of sodium tungstate containing 80-150 g/l W03, in order to obtain tungsten trioxide of the required purity, have so far been predominantly processed according to the traditional scheme, which includes: purification from compounds of impurity elements (Si, P, As, F, Mo); deposition
Calcium tungsten (artificial scheelite) followed by its decomposition with acids and the production of technical tungstic acid; dissolving tungstic acid in ammonia water, followed by evaporation of the solution and crystallization of ammonium paratungstate (PVA); calcination of PVA to obtain pure tungsten trioxide.
The main disadvantage of the scheme is that it is multi-stage, most operations are carried out in a periodic manner, and the duration of a number of stages. An extraction and ion exchange technology for converting Na2W04 solutions into (NH4)2W04 solutions has been developed and is already used at some enterprises. Below we briefly review the main stages of the traditional scheme and new extraction and ion exchange technology options.
Cleaning from impurities
Silicon removal. When the Si02 content in solutions exceeds 0.1% of the W03 content, preliminary removal of silicon is necessary. Purification is based on the hydrolytic decomposition of Na2Si03 by boiling a solution neutralized to pH = 8*9 with the release of silicic acid.
The solutions are neutralized with hydrochloric acid, which is added in a thin stream with stirring (to avoid local peroxidation) to the heated solution of sodium tungstate.
Purification from phosphorus and arsenic. To remove phosphate and arsenate ions, the method of precipitation of ammonium-magnesium salts Mg(NH4)P04 6H20 and Mg(NH4)AsC)4 6H20 is used. The solubility of these salts in water at 20 C is 0.058 and 0.038%, respectively. In the presence of excess Mg2+ and NH4 ions, solubility is lower.
Precipitation of phosphorus and arsenic impurities is carried out in the cold:
Na2HP04 + MgCl2 + NH4OH = Mg(NH4)P04 + 2NaCl +
Na2HAsQ4 + MgCl2 + NH4OH = Mg(NH4)AsQ4 + 2NaCl +
After standing for a long time (48 hours), crystalline precipitates of ammonium-magnesium salts fall out of the solution.
Purification from fluoride ions. With a high fluorite content in the initial concentrate, the content of fluoride ions reaches 5 g/l. Solutions are purified from fluoride ions by precipitation with magnesium fluoride from a neutralized solution to which MgCl2 is added. Fluorine removal can be combined with hydrolytic separation of silicic acid.
Molybdenum removal. Solutions of sodium tungstate must be cleaned of molybdenum if its content exceeds 0.1% of the W03 content (i.e. 0.1-0.2 t/l). At a molybdenum concentration of 5-10 g/l ( for example, when processing scheelite-powellite Tyrny-Auz concentrates), the separation of molybdenum becomes of particular importance, since the goal is to obtain a molybdenum chemical concentrate.
A common method is the precipitation of poorly soluble molybdenum trisulfide MoS3 from solution.
It is known that when sodium sulphide is added to solutions of sodium tungstate or molybdate, sulfosalts Na23S4 or oxosulfosalts Na23Sx04_x (where E is Mo or W) are formed:
Na2304 + 4NaHS = Na23S4 + 4NaOH. (1.16)
The equilibrium constant of this reaction for Na2Mo04 is significantly greater than for Na2W04(^^0 » Kzg). Therefore, if an amount of Na2S is added to the solution only sufficient to react with Na2Mo04 (with a slight excess), then molybdenum sulfosalt is predominantly formed. Upon subsequent acidification of the solution to pH = 2.5 * 3.0, the sulfosalt is destroyed with the release of molybdenum trisulfide:
Na2MoS4 + 2HC1 = MoS3 j + 2NaCl + H2S. (1.17)
Oxosulfosalts decompose with the release of oxosulfides (for example, MoSjO, etc.). Together with molybdenum trisulfide, a certain amount of tungsten trisulfide is coprecipitated. By dissolving the sulfide precipitate in a soda solution and re-precipitating molybdenum trisulfide, a molybdenum concentrate is obtained with a W03 content of no more than 2% with a loss of tungsten of 0.3-0.5% of the original amount.
After partial oxidative roasting of the precipitate - molybdenum trisulfide (at 450-500 °C) a molybdenum chemical concentrate containing 50-52% molybdenum is obtained.
The disadvantage of the method of deposition of molybdenum in the composition of trisulfide is the release of hydrogen sulfide by reaction (1.17), which requires costs for gas neutralization (they use the absorption of H2S in a scrubber irrigated with a solution of sodium hydroxide). Isolation of molybdenum trisulfide is carried out from a solution heated to 75-80 C. The operation is carried out in sealed steel reactors, rubberized or coated with acid-resistant enamel. Trisulfide precipitates are separated from the solution by filtration on a filter press.
Preparation of tungstic acid from solutions of sodium tungstate
Tungstic acid can be directly isolated from a solution of sodium tungstate with hydrochloric or nitric acids. However, this method is rarely used due to the difficulties of washing sediments from sodium ions, the content of which in tungsten trioxide is limited.
For the most part, calcium tungstate is initially precipitated from solution, which is then decomposed by acids. Calcium tungstate is precipitated by adding a CaC12 solution to a sodium tungstate solution heated to 80-90 C at a residual alkalinity of the solution of 0.3-0.7%. In this case, a white, finely crystalline, easily settled precipitate falls out; sodium ions remain in the mother solution, which ensures their low content in tungstic acid. 99-99.5% W is precipitated from the solution; mother liquors contain 0.05-0.07 g/l W03. The CaW04 precipitate washed with water in the form of a paste or pulp is sent for decomposition with hydrochloric acid when heated to 90°:
CaW04 + 2HC1 = H2W04i + CaCl2. (1.18)
During decomposition, the final acidity of the pulp is maintained high (90-100 g/l HCI), which ensures the separation of tungstic acid from impurities of phosphorus, arsenic and partly molybdenum compounds (molybdic acid dissolves in hydrochloric acid). Tungstic acid deposits require careful washing to remove impurities (especially calcium salts).
And sodium). In recent years, continuous washing of tungstic acid in pulsation columns has been developed, which has significantly simplified the operation.
At one of the enterprises in the USSR, when processing solutions of sodium tungstate, instead of hydrochloric acid, nitric acid is used to neutralize solutions and decompose CaW04 precipitates, and the precipitation of the latter is carried out by introducing Ca(N03)2 into solutions. In this case, nitrate mother liquors are utilized to obtain nitrate salts used as fertilizer.
Purification of technical tungstic acid and production of W03
Technical tungstic acid obtained by the method described above contains 0.2-0.3% impurities. As a result of acid calcination at 500-600 C, tungsten trioxide is obtained, suitable for the production of hard alloys based on tungsten carbide. However, for the production of tungsten, trioxide of higher purity is required with a total impurity content of no more than 0.05%.
The ammonia method for purifying tungstic acid is generally accepted. It easily dissolves in ammonia water, while most of the impurities remain in the sediment: silica, iron and manganese hydroxides and calcium (in the form of CaW04). However, ammonia solutions may contain an admixture of molybdenum and alkali metal salts.
A crystalline precipitate of PVA is isolated from the ammonia solution as a result of evaporation and subsequent cooling:
Evaporation
12(NH4)2W04 * (NH4)10H2W12O42 4H20 + 14NH3 +
In industrial practice, the composition of PVA is often written in oxide form: 5(NH4)20-12W03-5H20, which does not reflect its chemical nature as an isopolyacid salt.
Evaporation is carried out in periodic or continuous devices made of stainless steel. Typically, 75-80% tungsten is separated into crystals. It is undesirable to carry out deeper crystallization to avoid contamination of the crystals with impurities. It is significant that most of the molybdenum impurity (70-80%) remains in the mother solution. From the mother liquor, enriched with impurities, tungsten is precipitated in the form of CaW04 or H2W04, which is returned to the appropriate stages of the production scheme.
PVA crystals are squeezed out on a filter, then in a centrifuge, washed with cold water and dried.
Tungsten trioxide is obtained by thermal decomposition of tungstic acid or PVA:
H2W04 = "W03 + H20;
(NH4)10H2W12O42 4H20 = 12W03 + 10NH3 + 10H20. (1.20)
Calcination is carried out in rotating electric furnaces with a pipe made of heat-resistant steel 20Х23Н18. The calcination mode depends on the purpose of tungsten trioxide and the required size of its particles. So, to obtain VA tungsten wire (see below), PVA is calcined at 500-550 °C, HF and VT grade wires (tungsten without additives) - at 800-850 °C.
Tungstic acid is calcined at 750-850 °C. Tungsten trioxide made from PVA has larger particles than trioxide made from tungstic acid. In tungsten trioxide intended for the production of tungsten, the W03 content must be at least 99.95%; for the production of hard alloys - at least 99.9%.
Extraction and ion exchange methods for processing sodium tungstate solutions
The processing of sodium tungstate solutions is significantly simplified by extracting tungsten from solutions by extraction with an organic extractant, followed by re-extraction from the organic phase with an ammonia solution with the separation of PVA from the ammonia solution.
Since tungsten is found in solutions in the form of polymer anions in a wide range of pH = 7.5 + 2.0, anion-exchange extractants are used for extraction: salts of amines or quaternary ammonium bases. In particular, in industrial practice, trioctylamine sulfate salt (i?3NH)HS04 (where R is C8H17) is used. The highest rates of tungsten extraction are observed at pH=2*4.
Extraction is described by the equation:
4(i?3NH)HS04(opr) + Н2\У120*"(aq) + 2Н+(aq)ї=ї
Ї=ї(Д3ГШ)4Н4\У12О40(org) + 4Н80;(aq). (l.2l)
The amine is dissolved in kerosene, to which the technical mixture is added polyhydric alcohols(C7 - C9) to prevent the release of the solid phase (due to the low solubility of amine salts in kerosene). Approximate composition of the organic phase: amines 10%, alcohols 15%, kerosene - the rest.
Solutions purified from m-libdenum, as well as impurities of phosphorus, arsenic, silicon and fluorine are sent for extraction.
Tungsten is re-extracted from the organic phase with ammonia water (3-4% NH3), obtaining solutions of ammonium tungstate, from which PVA is isolated by evaporation and crystallization. Extraction is carried out in mixer-settler type devices or in pulsation columns with packing.
The advantages of extraction processing of sodium tungstate solutions are obvious: the number of operations in the technological scheme is reduced, the possibility of carrying out a continuous process for obtaining ammonium tungstate solutions from sodium tungstate solutions is created, and production space is reduced.
Wastewater extraction stage may contain an admixture of 80-100 mg/l of amines, as well as admixtures of higher alcohols and kerosene. To remove these environmentally harmful impurities, foam flotation and adsorption on activated carbon are used.
Extraction technology is used at foreign enterprises and is also implemented at domestic factories.
The use of ion exchange resins is a competing direction with extraction in the scheme for processing sodium tungstate solutions. For this purpose, low-basic anion exchangers containing amine groups (usually tertiary amines) or amphoteric resins (ampholytes) containing carboxyl and amine groups are used. At pH = 2.5 + 3.5, tungsten polyanions are sorbed on resins, and for some resins 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 respectively by the equations:
2tf2S04 + H4W12044; 5^«4H4W12O40 + 2SOf; (1.22)
I?4H4WI2O40 + 24NH4OH = 12(NH4)2W04 + 4DON + 12H20. (l.23)
The ion exchange method was developed and applied at one of the USSR enterprises. The required contact time of the resin with the solution is 8-12 hours. The process is carried out in a cascade of ion exchange columns with a suspended layer of resin in continuous mode. A difficult circumstance is the partial separation of PVA crystals at the elution stage, which requires their separation from the resin particles. As a result of elution, solutions containing 150-170 g/l W03 are obtained, which are sent to the evaporation and crystallization of PVA.
The disadvantage of ion exchange technology compared to extraction is unfavorable kinetics (contact duration 8-12 hours versus 5-10 minutes for extraction). At the same time, the advantages of ion exchangers include the absence of waste solutions containing organic impurities, as well as the fire safety and non-toxicity of resins.
Decomposition of scheelite concentrates by acids
In industrial practice, mainly when processing high-grade scheelite concentrates (70-75% W03), direct decomposition of scheelite with hydrochloric acid is used.
Decomposition reaction:
CaW04 + 2HC1 = W03H20 + CoCl2 (1.24)
Almost irreversible. However, the acid consumption is significantly higher than the stoichiometrically required one (250-300%) due to the inhibition of the process by films of tungstic acid on scheelite particles.
The decomposition is carried out in sealed reactors with stirrers, lined with acid-resistant enamel and heated through a steam jacket. The process is carried out at 100-110 C. The duration of decomposition varies from 4-6 to 12 hours, which depends on the degree of grinding, as well as the origin of the concentrate (scheelites from different deposits differ in reactivity).
A single treatment does not always lead to complete opening. In this case, after dissolving tungstic acid in ammonia water, the residue is re-treated with hydrochloric acid.
During the decomposition of scheelite-powellite concentrates containing 4-5% molybdenum, most of the molybdenum passes into the hydrochloric acid solution, which is explained by the high solubility of molybdic acid in hydrochloric acid. Thus, at 20 C in 270 g/l HC1, the solubilities of H2Mo04 and H2W04 are 182 and 0.03 g/l, respectively. Despite this, complete separation of molybdenum is not achieved. Tungstic acid precipitates contain 0.2-0.3% molybdenum, which cannot be extracted by repeated treatment with hydrochloric acid.
The acid method differs from the alkaline methods of scheelite decomposition in a smaller number of operations in the technological scheme. However, when processing concentrates with a relatively low content of W03 (50-55%) with a significant content of impurities, to obtain standard paravol-ammonium framate, it is necessary to carry out two or three ammonia purifications of tungstic acid, which is uneconomical. Therefore, decomposition with hydrochloric acid is mostly used in the processing of rich and pure scheelite concentrates.
The disadvantages of the decomposition method with hydrochloric acid are the high consumption of acid, the large volume of waste solutions of calcium chloride and the complexity of their disposal.
In light of the challenges of creating waste-free technologies, the nitrate method of decomposition of scheelite concentrates is of interest. In this case, mother solutions can be easily disposed of to obtain nitrate salts.
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State budgetary professional
educational institution of the Republic of Karelia
"Kostomuksha Polytechnic College"
Deputy Director of OD ___________________ Kubar T.S.
"_____"_________________________________2019
GRADUATE QUALIFYING WORK
Subject: “Maintaining the main method of beneficiation of tungsten ores and the use of auxiliary dehydration processes in the technological scheme of Primorsky GOK”
Group student: Kuzich S.E.
4th year, group OPI-15 (41C)
Specialty 02/21/18
"Beneficiation of mineral resources"
Head of the research and development work: Volkovich O.V.
teacher special disciplines
Kostomuksha
2019
Introduction………………………………………………………………………………...…3
- Technological part………………………………………………………6
1.1 general characteristics tungsten ores………………………………….6
1.2 Economic assessment of tungsten ores……………………………………10
- Technological scheme for beneficiation of tungsten ores using the example of Primorsky Mining and Processing Plant………………………………………………………..……11
2. Dehydration of enrichment products…………………………………......17
2.1. The essence of dehydration processes…………………………………..….17
2.2. Centrifugation…………………………………………………..…….24
3. Organization of safe working conditions…………………………………….30
3.1. Requirements for creating safe working conditions in the workplace…………………………………………………………..…...30
3.2. Requirements for maintaining safety in the workplace…….…..32
3.3. Safety requirements for enterprise employees…………32
Conclusion………………………………………………………………………………….…..…..34
List of sources and literature used……………………....…...36
Introduction
Mineral beneficiation - This is an industry that processes solid minerals with the intention of obtaining concentrates, i.e. products whose quality is higher than the quality of the original raw materials and meets the requirements for their further use in the national economy.Minerals are the basis of the national economy, and there is not a single industry where minerals or their processed products are not used.
One of these minerals is tungsten, a metal with unique properties. It has the highest boiling and melting points among metals, while having the lowest coefficient of thermal expansion. In addition, it is one of the hardest, heaviest, most stable and dense metals: the density of tungsten is comparable to the density of gold and uranium and 1.7 times higher than that of lead.The main tungsten minerals are scheelite, hübnerite and wolframite. Based on the type of minerals, ores can be classified into two types; scheelite and wolframite. When processing tungsten-containing ores, gravitational, flotation, magnetic, and also electrostatic,hydrometallurgical and other methods.
In recent years, metal-ceramic hard alloys made on the basis of tungsten carbide have been widely used. Such alloys are used as cutters, for the manufacture of drill bits, dies for cold wire drawing, dies, springs, parts of pneumatic tools, valves of internal combustion engines, heat-resistant parts of mechanisms operating at high temperatures. Surfacing hard alloys (stellites), consisting of tungsten (3-15%), chromium (25-35%) and cobalt (45-65%) with a small amount of carbon, are used for coating quickly wearing parts of mechanisms (turbine blades, excavator equipment and etc.). Tungsten alloys with nickel and copper are used in the manufacture of protective screens against gamma rays in medicine.
Metal tungsten is used in electrical engineering, radio engineering, X-ray engineering: for the manufacture of incandescent filaments in electric lamps, high-temperature heaters electric ovens, anticathodes and cathodes of X-ray tubes, electric vacuum equipment and much more. Tungsten compounds are used as dyes, to impart fire and water resistance to fabrics, in chemistry - as a sensitive reagent for alkaloids, nicotine, protein, and as a catalyst in the production of high-octane gasoline.
Tungsten is also widely used in the production of military and space equipment (armor plates, tank turrets, rifle and gun barrels, rocket cores, etc.).
The structure of tungsten consumption in the world is constantly changing. It is being replaced by other materials in some industries, but new areas of its application are emerging. Thus, in the first half of the 20th century, up to 90% of tungsten was spent on alloying steels. Currently, the industry is dominated by the production of tungsten carbide, and the use of tungsten metal is becoming increasingly important. IN Lately New opportunities are opening up for the use of tungsten as an environmentally friendly material. Tungsten can replace lead in the production of various ammunition, and can also be used in the manufacture of sports equipment, in particular golf clubs and balls. Developments in these areas are being carried out in the USA. In the future, tungsten should replace depleted uranium in the production of large-caliber ammunition. In the 1970s, when tungsten prices were around $170. for 1% WO content 3 per 1 ton of product, the USA, and then some NATO countries, replaced tungsten with depleted uranium in heavy ammunition, which, with the same technical characteristics, was significantly cheaper.
Tungsten, as a chemical element, belongs to the group of heavy metals and, from an environmental point of view, is classified as moderately toxic (Class II-III). The current source of tungsten pollution is environment are the processes of exploration, mining and processing (concentration and metallurgy) of tungsten-containing mineral raw materials. As a result of processing, such sources are unused solid waste, wastewater, and dusty tungsten-containing fine particles. Solid waste in the form of dumps and various tailings is generated during the enrichment of tungsten ores. Wastewater from processing plants is represented by tailings discharges, which are used as recycled water in grinding and flotation processes.
The purpose of the final qualifying work: to justify the technological scheme for the enrichment of tungsten ores using the example of Primorsky GOK and the essence of dehydration processes in this technological scheme.
IRKUTSK STATE TECHNICAL UNIVERSITY
As a manuscript
Artemova Olesya Stanislavovna
DEVELOPMENT OF TECHNOLOGY FOR EXTRACTING TUNGSTEN FROM STANDING TAILS OF THE DZHIDINSK VMK
Specialty 25.00.13- Mineral processing
dissertation for the degree of candidate of technical sciences
Irkutsk 2004
The work was carried out at Irkutsk State Technical University.
Scientific supervisor: Doctor of Technical Sciences,
Professor K.V. Fedotov
Official opponents: Doctor of Technical Sciences,
Professor Yu.P. Morozov
Candidate of Technical Sciences A.Ya. Mashovich
Leading organization: St. Petersburg State
Mining Institute (Technical University)
The defense will take place on December 22, 2004 at /O* hours at a meeting of the dissertation council D 212.073.02 of the Irkutsk State Technical University at the address: 664074, Irkutsk, st. Lermontova, 83, room. K-301
Scientific secretary of the dissertation council, professor
GENERAL DESCRIPTION OF WORK
Relevance of the work. Tungsten alloys are widely used in mechanical engineering, mining, metalworking industry, and in the production of electric lighting equipment. The main consumer of tungsten is metallurgy.
An increase in tungsten production is possible due to the involvement in processing of ores that are complex in composition, difficult to enrich, poor in the content of valuable components and off-balance ores, through the widespread use of gravity enrichment methods.
Involvement in recycling stale tails ore dressing at the Dzhida VMC will solve the current problem of the raw material base, increase the production of in-demand tungsten concentrate and improve the environmental situation in the Trans-Baikal region.
Purpose of the work: to scientifically substantiate, develop and test rational technological methods and modes of enrichment of stale tungsten-containing tailings from the Dzhida VMC.
The idea of the work: to study the relationship between the structural, material and phase compositions of the stale tailings of the Dzhida VMC with their technological properties, which makes it possible to create a technology for processing technogenic raw materials.
The following tasks were solved in the work: to assess the distribution of tungsten throughout the entire space of the main technogenic formation of the Dzhida VMC; study the material composition of the stale tailings of the Dzhizhinsky VMC; study the contrast of stale tailings in the original size in terms of W and 8 (II) content; to study the gravitational enrichment of stale tailings of the Dzhida VMC in various sizes; determine the feasibility of using magnetic enrichment to improve the quality of crude tungsten-containing concentrates; to optimize the technological scheme for the enrichment of technogenic raw materials of the general waste treatment plant of the Dzhida VMC; conduct pilot tests of the developed scheme for extracting W from the stale tailings of the DVMK.
Research methods: spectral, optical, optical-geometric, chemical, mineralogical, phase, gravitational and magnetic methods for analyzing the material composition and technological properties of initial mineral raw materials and enrichment products.
The reliability and validity of scientific statements and conclusions are ensured by a representative volume of laboratory research; confirmed by satisfactory convergence of calculated and experimentally obtained enrichment results, compliance with the results of laboratory and pilot tests.
NATIONAL I LIBRARY I SPEC gLYL!
Scientific novelty:
1. It has been established that technogenic tungsten-containing raw materials of the Dzhida VMC in any size are effectively enriched by the gravitational method.
2. Using generalized gravity concentration curves, the limiting technological indicators for processing stale tailings from the Dzhida VMC of various sizes using the gravity method were determined and the conditions for obtaining waste tailings with minimal tungsten losses were identified.
3. New patterns of separation processes have been established that determine the gravitational enrichment of tungsten-containing technogenic raw materials in a particle size of +0.1 mm.
4. For the stale tailings of the Dzhida VMC, a reliable and significant correlation between the contents of WO3 and S(II) was revealed.
Practical significance: a technology has been developed for the enrichment of stale tailings from the Dzhidinsky VMC, which ensures the effective extraction of tungsten and makes it possible to obtain a standard tungsten concentrate.
Approbation of the work: the main content of the dissertation work and its individual provisions were presented at the annual scientific and technical conferences of the Irkutsk State Technical University (Irkutsk, 2001-2004), the All-Russian school-seminar of young scientists “Leonov Readings - 2004” (Irkutsk , 2004), scientific symposium “Miner’s Week - 2001” (Moscow, 2001), All-Russian scientific and practical conference “New technologies in metallurgy, chemistry, enrichment and ecology” (St. Petersburg, 2004 .), Plaksinsky readings - 2004. The dissertation work was presented in full at the Department of Mineral Processing and Environmental Engineering at ISTU, 2004 and at the Department of Mineral Processing at SPGGI (TU), 2004.
Publications. 8 printed publications have been published on the topic of the dissertation work.
Structure and scope of work. The dissertation consists of an introduction, 3 chapters, a conclusion, 104 bibliographic sources and contains 139 pages, including 14 figures, 27 tables and 3 appendices.
The author expresses deep gratitude to the scientific supervisor, Doctor of Technical Sciences, Prof. K.V. Fedotov for professional and friendly leadership; prof. HE. Belkova - for valuable advice and useful critical comments expressed during the discussion of the dissertation work; G.A. Badenikova - for consulting on the calculation of the technological scheme. The author sincerely thanks the department staff for their comprehensive assistance and support provided during the preparation of the dissertation.
The objective prerequisites for the involvement of man-made formations in production turnover are:
The inevitability of preserving natural resource potential. This is achieved by reducing the extraction of primary mineral resources and reducing the amount of damage caused to the environment;
The need to replace primary resources with secondary ones. Determined by the needs of production for material and raw materials, including those industries whose natural resource base is practically exhausted;
The possibility of using technogenic waste is ensured by the introduction of scientific and technological progress.
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.
Ore processing waste storage facilities are objects of increased environmental hazard due to their negative impact on the air basin, ground and surface water, and soil cover over vast areas.
Payments for pollution are a form of compensation for economic damage from emissions and discharges of pollutants into the environment, as well as for the disposal of waste on the territory of the Russian Federation.
The Dzhida ore field belongs to the high-temperature deep hydrothermal quartz-wolframite (or quartz-gübnerite) type of deposits, playing vital role in tungsten mining. The main ore mineral is wolframite, the composition of which ranges from ferberite to pobnerite with all intermediate members of the series. Scheelite is a less common tungstate.
Wolframite ores are enriched mainly by gravity; Gravity methods of wet enrichment are usually used on jigging machines, hydrocyclones and concentration tables. To obtain quality concentrates, magnetic separation is used.
Until 1976, ores at the Dzhida VMC factory were processed according to a two-stage gravity scheme, including heavy-medium concentration in hydrocyclones, two-stage concentration of narrowly classified ore materials on three-deck tables of the SK-22 type, additional grinding and enrichment of industrial products in a separate cycle. The sludge was enriched according to a separate gravitational scheme using domestic and foreign sludge concentration tables.
From 1974 to 1996 Only tungsten ore enrichment tailings were stored. In 1985-86, ores were processed using a gravity-flotation technological scheme. Therefore, gravity enrichment tailings and sulfide flotogravity product were dumped into the main tailings pond. Since the mid-80s, due to the increased flow of ore supplied from the Inkursky mine, the share of large waste has increased
classes, up to 1-3 mm. After the Dzhidinsky GOK was shut down in 1996, the settling pond self-destructed due to evaporation and filtration.
In 2000, the “emergency discharge tailings storage facility” (EDT) was identified as an independent object due to its rather significant difference from the main tailings storage facility in terms of the conditions of occurrence, the scale of reserves, the quality and degree of safety of technogenic sands. Another secondary tailings storage facility is alluvial technogenic sediments (ATS), which include redeposited molybdenum ore flotation tailings in the area of the river valley. Modoncul.
The basic standards for payment for waste disposal within the established limits for the Dzhida VMC are 90,620,000 rubles. Annual environmental damage from land degradation due to the disposal of stale ore processing tailings is estimated at 20,990,200 rubles.
Thus, the involvement of stale ore dressing tailings of the Dzhida VMC in the processing will allow: 1) to solve the problem of the enterprise’s raw material base; 2) increase the production of the sought-after "-concentrate" and 3) improve the environmental situation in the Trans-Baikal region.
Material composition and technological properties of technogenic mineral formation of the Dzhida VMC
Geological sampling of the stale tailings of the Dzhida VMC was carried out. During the inspection of the secondary tailings dump (emergency discharge tailings dump (EDT)), 13 samples were taken. 5 samples were taken from the ATO deposit area. The sampling area of the main tailings dump (MTD) was 1015 thousand m2 (101.5 hectares), 385 private samples were taken. The weight of the selected samples is 5 tons. All selected samples were analyzed for the content of "03 and 8 (I).
OTO, CHAT and ATO were statistically compared in terms of "03" content using the Student's t test. With a confidence level of 95%, it was established: 1) the absence of a significant statistical difference in "03" content between private samples of side tailings; 2) the average results of testing the general waste dumps in terms of content "03 in 1999 and 2000 refer to the same general population; 3) the average results of testing the main and side tailings dumps in terms of content "03 significantly differ from each other and the mineral raw materials of all tailings dumps cannot be processed according to the same technology.
The subject of our research is general relativity.
The material composition of the mineral raw materials of the OTO of the Dzhida VMC was established based on the analysis of ordinary and group technological samples, as well as the products of their processing. Random samples were analyzed for the content of "03 and 8(11). Group samples were used for mineralogical, chemical, phase and sieve analyses.
According to the spectral semi-quantitative analysis of a representative analytical sample, the main useful component- " and minor ones - Pb, /u, Cu, Au and Contents "03 in the form of scheelite
quite stable in all size classes of various sand varieties and averages 0.042-0.044%. The content of WO3 in the form of hübnerite varies in various classes size. High contents of WO3 in the form of hübnerite were observed in particles of +1 mm size (from 0.067 to 0.145%) and especially in the -0.08+0 mm class (from 0.210 to 0.273%). This feature is typical for light and dark sands and is preserved for the average sample.
The results of spectral, chemical, mineralogical and phase analyzes confirm that the properties of hübnerite, as the main mineral form of \UOz, will determine the technology of enrichment of mineral raw materials of the OTO of the Dzhida VMC.
The granulometric characteristics of OTO raw materials with the distribution of tungsten by size class are shown in Fig. 1.2.
It can be seen that the bulk of the OTO sample material (~58%) has a particle size of -1+0.25 mm, 17% each falls on the large (-3+1 mm) and small (-0.25+0.1 mm) classes . The share of material with a particle size of -0.1 mm is about 8%, of which half (4.13%) is of the slurry class -0.044+0 mm.
Tungsten is characterized by a slight fluctuation (0.04-0.05%) in the content in size classes from -3 +1 mm to -0.25+0.1 mm and a sharp increase (up to 0.38%) in the size class -0 .1+0.044 mm. In the slurry class -0.044+0 mm, the tungsten content is reduced to 0.19%. That is, 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%.
As a result of the analysis of data on the dissemination of hübnerite and scheelite in the OTO mineral raw material of the original size and crushed to - 0.5 mm (see Table 1).
Table 1 - Distribution of grains and intergrowths of pobnerite and scheelite by size class of initial and crushed mineral raw materials _
Size classes, mm Distribution, %
Huebnerite Scheelite
Free grains | Splices Free grains | Splices
OTO material in original size (- 5 +0 mm)
3+1 36,1 63,9 37,2 62,8
1+0,5 53,6 46,4 56,8 43,2
0,5+0,25 79,2 20,8 79,2 20,8
0,25+0,125 88,1 11,9 90,1 9,9
0,125+0,063 93,6 6,4 93,0 7,0
0,063+0 96,0 4,0 97,0 3,0
Amount 62.8 37.2 64.5 35.5
OTO material, crushed to - 0.5 +0 mm
0,5+0,25 71,5 28,5 67,1 32,9
0,25+0,125 75,3 24,7 77,9 22,1
0,125+0,063 89,8 10,2 86,1 13,9
0,063+0 90,4 9,6 99,3 6,7
Amount 80.1 19.9 78.5 21.5
It was concluded that it is necessary to classify deslimed mineral raw materials OTO according to a particle size of 0.1 mm and separate enrichment of the resulting classes. From large class should: 1) separate the free grains into a rough concentrate, 2) tailings containing intergrowths should be subjected to additional grinding, desliming, combining with the desliming class -0.1+0 mm of the initial mineral raw material and gravity enrichment to extract fine grains of scheelite and pobnerite into the middling product.
To assess the contrast of OTO mineral raw materials, a technological sample was used, which is a combination of 385 individual samples. The results of fractionation of individual samples according to the content of WO3 and sulfide sulfur are shown in Fig. 3, 4.
0 Y OS 0.2 "l M o l O 2 SS * _ " 8
S(kk|Yupytetr "oknsmm" fr**m.% Contained gulfkshoy
Rice. 3 Conditional contrast curves of the original Fig. 4 Conditional contrast curves of the original
mineral raw materials OTO by content Ch/O) mineral raw materials OTO by content 8 (II)
It was found that the contrast indices for the content of WO3 and S (II) are equal to 0.44 and 0.48, respectively. Taking into account the classification of ores by contrast, the studied mineral raw materials in terms of WO3 and S (II) content belong to the category of non-contrast ores. Radiometric enrichment is not
suitable for extracting tungsten from small-sized stale tailings of the Dzhida VMC.
The results of the correlation analysis, with the help of which a mathematical relationship was revealed between the concentrations of \\Sos and 8 (II) (Stoz = 0»0232 + 0.038C5(u)And r = 0.827; the correlation is valid and reliable), confirm the conclusions about the inappropriateness of using radiometric separation.
The results of the analysis of the separation of OTO mineral grains in heavy liquids prepared on the basis of selenium bromide were used to calculate and construct gravity enrichment curves (Fig. 5), from the form of which, especially the curve, it follows that the OTO of the Dzhida VMC in any size is suitable for mineral raw materials gravity enrichment method.
Taking into account the shortcomings in the use of gravity concentration curves, especially the curve for determining the metal content in floating fractions with a given yield or recovery, generalized gravity concentration curves were constructed (Figure 6), the results of the analysis of which are given in Table. 2.
Table 2 - Forecast technological indicators of enrichment of different size classes of stale tailings from the Dzhida VMC using the gravity method_
g Size class, mm Maximum losses \U with tailings, % Tailings yield, % XV content, %
in the tails at the end
3+1 0,0400 25 82,5 0,207 0,1
3+0,5 0,0400 25 84 0,19 0,18
3+0,25 0,0440 25 90 0,15 0,28
3+0,1 0,0416 25 84,5 0,07 0,175
3+0,044 0,0483 25 87 0,064 0,27
1+0,5 0,04 25 84,5 0,16 0,2
1+0,044 0,0500 25 87 0,038 0,29
0,5+0,25 0,05 25 92,5 0,04 0,45
0,5+0,044 0,0552 25 88 0,025 0,365
0,25+0,1 0,03 25 79 0,0108 0,1
0,25+0,044 0,0633 15 78 0,02 0,3
0,1+0,044 0,193 7 82,5 0,018 1,017
In terms of gravity washability, the classes -0.25+0.044 and -0.1+0.044 mm are significantly different from materials of other sizes. The best technological indicators of gravitational enrichment of mineral raw materials are predicted for the size class -0.1+0.044 mm: ^ |*0M4=82.5%, =0.018% and e* =7%.
The results of electromagnetic fractionation of heavy fractions (HF), gravitational analysis using the Sochnev S-5 universal magnet and magnetic separation of HF showed that the total yield of highly magnetic and non-magnetic fractions is 21.47% and the losses in them are 4.5%. Minimum losses "with a non-magnetic fraction and the maximum content" in the combined weakly magnetic product are predicted provided that the separation power in a strong magnetic field has a particle size of -0.1+0 mm.
Rice. 5 Gravity enrichment curves for stale tailings of the Dzhida VMC
e) class -0.1+0.044 mm
Rice. 6 Generalized gravity concentration curves for various size classes of mineral raw materials GTO
Development of a technological scheme for the enrichment of stale ore dressing tailings of the Dzhidinsky VM K
The results of technological testing of various methods of gravitational enrichment of stale tailings of the Dzhidinsky VMC are presented in Table. 3.
Table 3 - Results of testing gravity devices
Comparable technological indicators were obtained for the extraction of WO3 into rough concentrate during the enrichment of unclassified stale tailings using both screw separation and centrifugal separation. Minimal losses of WO3 with tailings were detected during enrichment in a centrifugal concentrator of class -0.1+0 mm.
In table Figure 4 shows the granulometric composition of the rough W-concentrate with a particle size of -0.1+0 mm.
Table 4 - Granulometric composition of rough W-concentrate
Size class, mm Yield of classes, % Content Distribution of AUOz
Absolute Relative, %
1+0,071 13,97 0,11 1,5345 2,046
0,071+0,044 33,64 0,13 4,332 5,831
0,044+0,020 29,26 2,14 62,6164 83,488
0,020+0 23,13 0,28 6,4764 8,635
Total 100.00 0.75 75.0005 100.0
In the concentrate, the main amount of WO3 is in the class -0.044+0.020 mm.
According to mineralogical analysis, compared to the source material, the concentrate contains a higher mass fraction of pobnerite (1.7%) and ore sulfide minerals, especially pyrite (16.33%). The content of rock-forming materials is 76.9%. The quality of rough W-concentrate can be increased by the sequential use of magnetic and centrifugal separation.
The results of testing gravitational devices for extracting >V03 from the tailings of the primary gravitational enrichment of mineral raw materials OTO in a particle size of +0.1 mm (Table 5) have proven that the most effective device is the KKEL80No concentrator
Table 5 - Results of testing gravity devices
Product G,% ßwo>, % rßwo> st">, %
screw separator
Concentrate 19.25 0.12 2.3345 29.55
Tails 80.75 0.07 5.5656 70.45
Initial sample 100.00 0.079 7.9001 100.00
wing gateway
Concentrate 15.75 0.17 2.6750 33.90
Tails 84.25 0.06 5.2880 66.10
Initial sample 100.00 0.08 7.9630 100.00
concentration table
Concentrate 23.73 0.15 3.56 44.50
Tails 76.27 0.06 4.44 55.50
Initial sample 100.00 0.08 8.00 100.00
centrifugal concentrator KC-MD3
Concentrate 39.25 0.175 6.885 85.00
Tails 60.75 0.020 1.215 15.00
Initial sample 100.00 0.081 8.100 100.00
When optimizing the technological scheme for the beneficiation of mineral raw materials of the OTO of the Dzhida VMC, the following were taken into account: 1) technological schemes for processing finely disseminated wolframite ores from domestic and foreign enrichment plants; 2) technical characteristics of the modern equipment used and its dimensions; 3) the possibility of using the same equipment for simultaneous implementation of two operations, for example, separation of minerals by size and dehydration; 4) economic costs for the hardware design of the technological scheme; 5) the results presented in Chapter 2; 6) GOST requirements for the quality of tungsten concentrates.
During semi-industrial testing of the developed technology (Figure 7-8 and Table 6), 15 tons of initial mineral raw materials were processed in 24 hours.
The results of spectral analysis of a representative sample of the obtained concentrate confirm that the W-concentrate III of magnetic separation is standard and corresponds to the KVG (T) grade of GOST 213-73.
Fig. 8 Results of technological testing of the scheme for finishing rough concentrates and middling products from the stale tailings of the Dzhida VMC
Table 6 - Results of testing the technological scheme
Product
Conditioned concentrate 0.14 62.700 8.778 49.875
Dump tailings 99.86 0.088 8.822 50.125
Initial ore 100.00 0.176 17.600 100.000
CONCLUSION
The work provides a solution to a pressing scientific and production problem: scientifically substantiated, developed and, to a certain extent, implemented effective technological methods for extracting tungsten from the stale ore dressing tailings of the Dzhida VMC.
The main results of the research, development and their practical implementation are as follows:
The main useful component is tungsten, the content of which stale tailings are a non-contrasting ore, represented mainly by hübnerite, which determines the technological properties of technogenic raw materials. Tungsten is unevenly distributed among size classes and its main amount is concentrated in the size
It has been proven that the only effective method for enriching W-containing stale tailings of the Dzhida VMC is gravity. Based on the analysis of generalized gravity enrichment curves of stale W-containing tailings, it was established that dump tailings with minimal tungsten losses are a distinctive feature of the enrichment of technogenic raw materials in a size of -0.1+Ohm. New patterns of separation processes have been established that determine the technological indicators of gravitational enrichment of stale tailings from the Dzhida VMC in a size of +0.1 mm.
It has been proven that among the gravitational devices used in the mining industry for the beneficiation of W-containing ores, the screw separator and the centrifugal concentrator KKEL80N are suitable for maximum extraction of tungsten from the technogenic raw materials of the Dzhida VMC into rough W-concentrates. The effectiveness of using the KKEL80K concentrator has also been confirmed for additional extraction of tungsten from tailings of primary enrichment of technogenic W-containing raw materials in size - 0.1 mm.
3. Optimized technology system extraction of tungsten from the stale ore dressing tailings of the Dzhidinsky VMC made it possible to obtain a standard W-concentrate, solve the problem of depletion of mineral resources of the Dzhidinsky VMC and reduce the negative impact production activities enterprises on the environment.
Preferred use of gravity equipment. During semi-industrial testing of the developed technology for extracting tungsten from the stale tailings of the Dzhida VMC, a standard “-concentrate” was obtained with a “03 content of 62.7% with an extraction of 49.9%. The payback period for the processing plant for processing stale tailings from the Dzhida VMC in order to extract tungsten was 0.55 years.
The main provisions of the dissertation work were published in the following works:
1. Fedotov K.V., Artemova O.S., Polinskina I.V. Assessment of the possibility of processing stale tailings of the Dzhida VMC, Ore dressing: Sat. scientific works - Irkutsk: ISTU Publishing House, 2002. - 204 pp., pp. 74-78.
2. Fedotov K.V., Senchenko A.E., Artemova O.S., Polinkina I.V. Application of a centrifugal separator with continuous discharge of concentrate for the extraction of tungsten and gold from the tailings of the Dzhidinsky VMC, Environmental problems and new technologies complex processing mineral raw materials: Materials of the International Meeting “Plaksin Readings - 2002”. - M.: P99, Publishing House PKTs "Altex", 2002 - 130 p., P.96-97.
3. Zelinskaya E.V., Artemova O.S. The possibility of regulating the selectivity of the action of the collector during the flotation of tungsten-containing ores from stale tailings, Directed changes in the physico-chemical properties of minerals in mineral processing processes (Plaksin Readings), materials of the international meeting. - M.: Altex, 2003. -145 p., pp. 67-68.
4. Fedotov K.V., Artemova O.S. Problems of processing stale tungsten-containing products Modern methods of processing mineral raw materials: Conference materials. Irkutsk: Irk. State Those. Univ., 2004 - 86 s.
5. Artemova O. S., Gaiduk A. A. Extraction of tungsten from stale tailings of the Dzhida tungsten-molybdenum plant. Prospects for the development of technology, ecology and automation of chemical, food and metallurgical industries: Materials of a scientific and practical conference. - Irkutsk: ISTU Publishing House. - 2004 - 100 p.
6. Artemova O.S. Assessment of the uneven distribution of tungsten in the Dzhida tailings dump. Modern methods for assessing the technological properties of mineral raw materials of precious metals and diamonds and advanced technologies for their processing (Plaksin Readings): Proceedings of the international meeting. Irkutsk, September 13-17, 2004 - M.: Altex, 2004. - 232 s.
7. Artemova O.S., Fedotov K.V., Belkova O.N. Prospects for the use of the technogenic deposit of the Dzhidinsky VMC. All-Russian scientific and practical conference “New technologies in metallurgy, chemistry, enrichment and ecology”, St. Petersburg, 2004.
Signed for publication on November 12, 2004. Format 60x84 1/16. Printing paper. Offset printing. Conditional oven l. Academician-ed.l. 125. Circulation 400 copies. Law 460.
ID No. 06506 dated December 26, 2001 Irkutsk State Technical University 664074, Irkutsk, st. Lermontova, 83
RNB Russian Fund
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. Need for use
1.3. Technogenic mineral formation of the Dzhida VMC
1.4. Goals and objectives of the study. Research methods. Provisions for defense
2. RESEARCH OF THE SUBSTANTIAL COMPOSITION AND TECHNOLOGICAL PROPERTIES OF STELLED TAILINGS OF THE DZHIDINSK VMK
2.1. Geological testing and evaluation of tungsten distribution
2.2. Material composition of mineral raw materials
2.3. Technological properties of mineral raw materials
2.3.1. Grading
2.3.2. Study of the possibility of radiometric separation of mineral raw materials in the original size
2.3.3. Gravity analysis
2.3.4. Magnetic analysis
3. DEVELOPMENT OF A TECHNOLOGICAL SCHEME FOR THE EXTRACTION OF TUNGSTEN FROM STANDING TAILS OF THE DZHIDINSK VMK
3.1. Technological testing of various gravity devices for the enrichment of stale tailings of various sizes
3.2. Optimization of the general waste processing scheme
3.3. Pilot testing of the developed technological scheme for the enrichment of general waste and an industrial plant
Introduction Dissertation on geosciences, on the topic "Development of technology for extracting tungsten from the stale tailings of the Dzhida VMC"
The sciences of mineral processing are, first of all, aimed at developing the theoretical foundations of mineral separation processes and the creation of processing apparatus, at revealing the relationship between the distribution patterns of components and separation conditions in processing 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. Poor use of resource-saving technologies contributes to large losses of minerals during the extraction and enrichment of raw materials.
An analysis of the development of equipment and technology for mineral processing over the past 10-15 years indicates significant achievements of domestic fundamental science in the field of knowledge of the basic phenomena and patterns in the separation of mineral complexes, which makes it possible to create highly efficient processes and technologies for the primary processing of ores of complex 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 intensity, energy intensity and wear resistance.
In addition, due to the departmental affiliation of mining and processing enterprises, complex raw materials were processed only taking into account the necessary industry needs for a specific metal, which led to the irrational use of natural mineral resources and increased costs for waste storage. Currently, more than 12 billion tons of waste have been accumulated, the content of valuable components in which in some cases exceeds their content in natural deposits.
In addition to the above negative trends, since the 90s, the environmental situation at mining and processing enterprises has sharply worsened (in a number of regions, threatening the existence of not only biota, but also humans), there has been a progressive decline in the production of non-ferrous and ferrous metal ores, mining and chemical raw materials, deterioration in the quality of processed ores and, as a consequence, the involvement in the processing of difficult-to-process ores of complex material composition, characterized by a low content of valuable components, fine dissemination and similar technological properties of minerals. Thus, over the past 20 years, the content of non-ferrous metals in ores has decreased by 1.3-1.5 times, iron by 1.25 times, gold by 1.2 times, the share of difficult ores and coal has increased from 15% to 40% of the total mass of raw materials supplied for enrichment.
Human impact on the natural environment in the process economic activity is now becoming global. In terms of the scale of extracted and transported rocks, transformation of relief, impact on the redistribution and dynamics of surface and groundwater, activation of geochemical transfer, etc. this activity is comparable to geological processes.
The unprecedented scale of extracted mineral resources leads to their rapid depletion, the accumulation of large amounts of waste on the Earth’s surface, in the atmosphere and hydrosphere, the gradual degradation of natural landscapes, a reduction in biodiversity, and a decrease in the natural potential of territories and their life-supporting functions.
Ore processing waste storage facilities are objects of increased environmental hazard due to their negative impact on the air basin, ground and surface water, and soil cover over vast areas. Along with this, tailings dumps are little-studied technogenic deposits, the use of which will make it possible to obtain additional sources of ore and mineral raw materials while significantly reducing the scale of disturbance of the geological environment in the region.
Production of products from technogenic deposits, as a rule, is several times cheaper than from raw materials specially mined for this purpose, and is characterized by a quick return on investment. However, the complex chemical, mineralogical and granulometric composition of tailings, as well as the wide range of minerals they contain (from main and associated components to the simplest building materials) make it difficult to calculate the total economic effect of their processing and determine an individual approach to the assessment of each tailings.
Consequently, at the moment a number of insoluble contradictions have emerged between the change in the nature of the mineral resource base, i.e. the need to involve difficult-to-process ores and technogenic deposits in the processing, the environmentally aggravated situation in mining regions and the state of technology, technology and organization of primary processing of mineral raw materials.
The issues of using waste from the enrichment of polymetallic, gold-containing and rare metals have both economic and environmental aspects.
In achieving the current level of development of the theory and practice of processing tailings from the enrichment of non-ferrous, rare and precious metal ores, V.A. made a great contribution. Chanturia, V.Z. Kozin, V.M. Avdokhin, S.B. Leonov, J.I.A. Barsky, A.A. Abramov, V.I. Karmazin, S.I. Mitrofanov and others.
An important component of the overall strategy of the ore industry, incl. tungsten, is the increased use of ore processing waste as additional sources of ore and mineral raw materials, with a significant reduction in the scale of disturbance of the geological environment in the region and the negative impact on all components of the environment.
In the field of using ore processing waste, the most important thing is a detailed mineralogical and technological study of each specific, individual technogenic deposit, the results of which will make it possible to develop 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 Irkutsk State Technical University on the topic “Fundamental and technological research in the field of processing of mineral and technogenic raw materials for the purpose of their integrated use, taking into account environmental problems in complex industrial systems" and paper topic No. 118 "Study of the beneficiation of stale tailings of the Dzhida VMC."
The purpose of the work is to scientifically substantiate, develop and test rational technological methods for the enrichment of stale tungsten-containing tailings from the Dzhida VMC.
The following tasks were solved in the work:
Assess the distribution of tungsten throughout the entire space of the main technogenic formation of the Dzhida VMC;
To study the material composition of the stale tailings of the Dzhizhinsky MMC;
Investigate the contrast of stale tailings in the original size according to the content of W and S (II); to study the gravitational enrichment of stale tailings of the Dzhida VMC in various sizes;
To determine the feasibility of using magnetic enrichment to improve the quality of rough tungsten-containing concentrates;
Optimize the technological scheme for the enrichment of technogenic raw materials of the OTO of the Dzhida VMC; conduct pilot tests of the developed scheme for extracting W from the stale tailings of DVMC;
To develop a circuit diagram of devices for the industrial processing of stale tailings from the Dzhida VMC.
To carry out the research, a representative technological sample of stale tailings from the Dzhida VMC was used.
When solving the formulated problems, the following research methods were used: 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 basic scientific provisions are submitted for defense: The patterns of distribution of initial technogenic mineral raw materials and tungsten by size classes have been established. The need for primary (preliminary) classification by size of 3 mm has been proven.
Installed quantitative characteristics stale tailings of ore processing of ores from the Dzhida VMC in terms of WO3 and sulfide sulfur content. It has been proven that the initial mineral raw materials belong to the category of non-contrasting ores. A reliable and reliable correlation between the contents of WO3 and S (II) was revealed.
Quantitative patterns of gravitational enrichment of stale tailings from the Dzhida VMC have been established. It has been proven that for source material of any size, an effective method for extracting W is gravitational enrichment. Forecast technological indicators of gravitational enrichment of initial mineral raw materials in various sizes have been determined.
Quantitative patterns of distribution of stale ore dressing tailings of the Dzhida VMC into fractions of different specific magnetic susceptibility have been established. The effectiveness of the sequential use of magnetic and centrifugal separation has been proven to improve the quality of rough W-containing products. The technological modes of magnetic separation have been optimized.
Conclusion Dissertation on the topic "Beneficiation of mineral resources", Artemova, Olesya Stanislavovna
The main results of the research, development and their practical implementation are as follows:
1. An analysis of the current situation in the Russian Federation with mineral resources of the ore industry, in particular tungsten, was carried out. Using the example of the Dzhidinsky VMC, it is shown that the problem of involving stale ore dressing tailings in the processing is relevant, having technological, economic and environmental significance.
2. The material composition and technological properties of the main W-containing technogenic formation of the Dzhida VMC have been established.
The main useful component is tungsten, the content of which stale tailings are a non-contrasting ore, represented mainly by hübnerite, which determines the technological properties of technogenic raw materials. Tungsten is unevenly distributed across size classes and its main amount is concentrated in sizes -0.5+0.1 and -0.1+0.02 mm.
It has been proven that the only effective method for enriching W-containing stale tailings of the Dzhida VMC is gravity. Based on the analysis of generalized gravity enrichment curves of stale W-containing tailings, it was established that dump tailings with minimal tungsten losses are a distinctive feature of the enrichment of technogenic raw materials in a size of -0.1+0 mm. New patterns of separation processes have been established that determine the technological indicators of gravitational enrichment of stale tailings from the Dzhida VMC in a size of +0.1 mm.
It has been proven that among the gravitational devices used in the mining industry for the enrichment of W-containing ores, a screw separator and a KNELSON centrifugal concentrator are suitable for maximum extraction of tungsten from the technogenic raw materials of the Dzhida VMC into rough W-concentrates. The effectiveness of using the KNELSON concentrator has also been confirmed for the additional extraction of tungsten from the tailings of the primary enrichment of technogenic W-containing raw materials in a particle size of 0.1 mm.
3. An optimized technological scheme for extracting tungsten from the stale ore dressing tailings of the Dzhidinsky VMC made it possible to obtain a standard W-concentrate, solve the problem of depletion of mineral resources of the Dzhidinsky VMC and reduce the negative impact of the enterprise’s production activities on the environment.
The essential features of the developed technology for extracting tungsten from the stale tailings of the Dzhida VMC are:
Narrow classification by feed size of primary enrichment operations;
Preferred use of gravity equipment.
During semi-industrial testing of the developed technology for extracting tungsten from the stale tailings of the Dzhida VMC, a standard W-concentrate was obtained with a WO3 content of 62.7% with an extraction of 49.9%. The payback period for the processing plant for processing stale tailings from the Dzhida VMC in order to extract tungsten was 0.55 years.
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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. Composition of tin concentrates obtained from processing 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. To enrich the under-sieve product of screens, jigging machines of various designs are used, 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%.
The enrichment of primary tin ores, characterized by the complexity of their material composition and uneven dissemination of cassiterite, is carried out according to more complex multi-stage schemes using not only gravitational methods, but also flotation gravity, 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, hübnerite 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 ones, in two stages - primary gravity 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 gravitational 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%.
Tungsten minerals and ores
Of the tungsten minerals, the minerals of the wolframite and scheelite group are of practical importance.
Wolframite (xFeWO4 yMnWO4) is an isomorphic mixture of iron and manganese tungstates. If a mineral contains more than 80% iron, the mineral is called ferberite. If the mineral contains more than 80% manganese, then the mineral is called hubernite.
Scheelite CaWO4 is almost pure calcium tungstate.
Tungsten ores contain small amounts of tungsten. The minimum WO3 content at which their processing is advisable. is 0.14-0.15% for large deposits and 0.4-0.5% for small deposits. In ores, tungsten is accompanied by tin in the form of cassiterite, as well as the minerals molybdenum, bismuth, arsenic and copper. The main gangue rock is silica.
Tungsten ores undergo beneficiation. Wolframite ores are enriched using the gravity method, and scheelite ores are enriched by flotation.
Tungsten ore enrichment schemes are varied and complex. They combine gravitational enrichment with magnetic separation, flotation gravity and flotation. By combining various enrichment methods, concentrates containing up to 55-72% WO3 are obtained from ores. The extraction of tungsten from ore into concentrate is 82-90%.
The composition of tungsten concentrates varies within the following limits,%: WO3-40-72; MnO-0.008-18; SiO2-5-10; Mo-0.008-0.25; S-0.5-4; Sn-0.03-1.5; As-0.01-0.05; P-0.01-0.11; Cu-0.1-0.22.
Technological schemes for processing tungsten concentrates are divided into two groups: alkaline and acidic.
Methods for processing tungsten concentrates
Regardless of the method of processing wolframite and scheelite concentrates, the first stage of their processing is opening, which is the transformation of tungsten minerals into easily soluble chemical compounds.
Wolframite concentrates are opened by sintering or fusion with soda at a temperature of 800-900°C, which is based on chemical reactions:
4FeWO4 + 4Na2CO3 + O2 = 4Na2WO4 + 2Fe2O3 +4CO2 (1)
6MnWO4 + 6Na2CO3 + O2 = 6Na2WO4 + 2Mn3O4 +6CO2 (2)
When sintering scheelite concentrates at a temperature of 800-900°C, the following reactions occur:
CaWO4 + Na2CO3 = Na2WO4+ CaCO3 (3)
CaWO4 + Na2CO3 = Na2WO4+ CaO + CO2 (4)
In order to reduce soda consumption and prevent the formation of free calcium oxide, silica is added to the charge to bind calcium oxide into a sparingly soluble silicate:
2CaWO4 + 2Na2CO3 + SiO2 = 2Na2WO4+ Ca2SiO4 + CO2 (5)
Sintering of scheelite concentrate with soda and silica is carried out in drum furnaces at a temperature of 850-900°C.
The resulting cake (alloy) is leached with water. During leaching, sodium tungstate Na2WO4 and soluble impurities (Na2SiO3, Na2HPO4, Na2AsO4, Na2MoO4, Na2SO4) and excess soda pass into the solution. Leaching is carried out at a temperature of 80-90°C in steel reactors with mechanical stirring, operating in batch mode, or in continuous drum rotary kilns. The recovery of tungsten into the solution is 98-99%. The solution after leaching contains 150-200 g/l WO3. The solution is filtered, and after separating the solid residue, it is sent for purification from silicon, arsenic, phosphorus and molybdenum.
Purification from silicon is based on the hydrolytic decomposition of Na2SiO3 by boiling a solution neutralized at pH = 8-9. Neutralization of excess soda in the solution is carried out with hydrochloric acid. As a result of hydrolysis, slightly soluble silicic acid is formed:
Na2SiO3 + 2H2O = 2NaOH + H2SiO3 (6)
To remove phosphorus and arsenic, the method of precipitation of phosphate and arsenate ions in the form of poorly soluble ammonium-magnesium salts is used:
Na2HPO4 + MgCl2+ NH4OH = Mg(NH4)PO4 + 2NaCl + H2O (7)
Na2HAsO4 + MgCl2+ NH4OH = Mg(NH4)AsO4 + 2NaCl + H2O (8)
Purification from molybdenum is based on the decomposition of molybdenum sulfosalt, which is formed when sodium sulfide is added to a solution of sodium tungstate:
Na2MoO4 + 4NaHS = Na2MoS4 + 4NaOH (9)
Upon subsequent acidification of the solution to pH = 2.5-3.0, the sulfosalt is destroyed with the release of slightly soluble molybdenum trisulfide:
Na2MoS4 + 2HCl = MoS3 + 2NaCl + H2S (10)
Calcium tungstate is first precipitated from a purified solution of sodium tungstate using CaCl2:
Na2WO4 + CaCl2 = CaWO4 + 2NaCl. (eleven)
The reaction is carried out in a boiling solution containing 0.3-0.5% alkali
while stirring with a mechanical stirrer. The washed sediment of calcium tungstate in the form of a pulp or paste is subjected to decomposition with hydrochloric acid:
CaWO4 + 2HCl = H2WO4 + CaCl2 (12)
During decomposition, the high acidity of the pulp is maintained at about 90-120 g/l HCl, which ensures the separation of impurities of phosphorus, arsenic and partly molybdenum, which are soluble in hydrochloric acid, from the tungstic acid sediment.
Tungstic acid can also be obtained from a purified solution of sodium tungstate by direct precipitation with hydrochloric acid. When the solution is acidified with hydrochloric acid, H2WO4 precipitates as a result of hydrolysis of sodium tungstate:
Na2WO4 + 2H2O = 2NaOH + H2WO4 (11)
The alkali formed as a result of the hydrolysis reaction reacts with hydrochloric acid:
2NaOH + 2HCl = 2NaCl + 2H2O (12)
The addition of reactions (8.11) and (8.12) gives the total reaction of precipitation of tungstic acid with hydrochloric acid:
Na2WO4 + 2HCl = 2NaCl + H2WO4 (13)
However, in this case, great difficulties arise in washing the sediment from sodium ions. Therefore, at present, the latter method of tungstic acid deposition is used very rarely.
The technical tungstic acid obtained by precipitation contains impurities and therefore needs to be purified.
The most widely used method is the ammonia method for purifying technical tungsten acid. It is based on the fact that tungstic acid is highly soluble in ammonia solutions, while a significant part of the impurities it contains are insoluble in ammonia solutions:
H2WO4 + 2NH4OH = (NH4)2WO4 + 2H2O (14)
Ammonia solutions of tungstic acid may contain impurities of molybdenum and alkali metal salts.
Deeper cleaning is achieved by isolating large crystals of ammonium paratungstate from the ammonia solution, which are obtained by evaporating the solution:
12(NH4)2WO4 = (NH4)10W12O41 5H2O + 14NH3 + 2H2O (15)
tungsten acid anhydride precipitation
Deeper crystallization is impractical to avoid contamination of the crystals with impurities. From the mother liquor, enriched with impurities, tungsten is precipitated in the form of CaWO4 or H2WO4 and returned to the previous stages.
Paratungstate crystals are squeezed out on filters, then in a centrifuge, washed with cold water and dried.
Tungsten oxide WO3 is obtained by calcining tungstic acid or paratungstate in a rotating tubular furnace with a stainless steel pipe and heated by electricity at a temperature of 500-850oC:
H2WO4 = WO3 + H2O (16)
(NH4)10W12O41 5H2O = 12WO3 + 10NH3 +10H2O (17)
In tungsten trioxide intended for the production of tungsten, the WO3 content must be no lower than 99.95%, and for the production of hard alloys - no lower than 99.9%