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Technology of enrichment of wolframite 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 Irgiredmet OJSC, Kovdorsky Mining and Processing Plant Business plan for the enrichment of tungsten ore

Tungsten minerals, ores and concentrates

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

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

Scheelite CaWOA is calcium tungstate. The color of the mineral is white, gray, yellow, brown. Density 5.9-6.1 g/cm3, hardness on the mineralogical scale 4.5-5. Scheelite often contains an isomorphic admixture of powellite - CaMoO4. When irradiated 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.

Extraction of tungsten into conditioned concentrates during enrichment tungsten ores ranges from 65-70% to 85-90%.

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

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

The explored reserves of tungsten as of 1981 are estimated at 2903 thousand tons, of which 1360 thousand tons are in China. The USSR, Canada, Australia, the USA, South and North Korea, Bolivia, Brazil, and Portugal have significant reserves. Production 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.

Solutions neutralize hydrochloric acid, added in a thin stream with stirring (to avoid local peroxidation) to the heated solution of sodium tungstate.

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

Preparation of tungstic acid from solutions of sodium tungstate

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

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

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

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

And sodium). IN last years continuous washing of tungstic acid in pulsation columns was mastered, which 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 a technical mixture of polyhydric alcohols (C7 - C9) is added to prevent the release of the solid phase (due to the low solubility of amine salts in kerosene). Approximate composition of the organic phase: amines 10%, alcohols 15%, kerosene - the rest.

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

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

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

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

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

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

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

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

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

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

Decomposition of scheelite concentrates by acids

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

Decomposition reaction:

CaW04 + 2HC1 = W03H20 + CoCl2 (1.24)

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

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

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

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

The acid method differs from alkaline methods of decomposition of scheelite in a smaller number of operations 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, large volume 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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Navoi Mining and Metallurgical Plant

Navoi State Mining Institute

"Chemical and Metallurgical" Faculty"

Department of Metallurgy

Explanatory note

for final qualifying work

on the topic of: “Selection, justification and calculation of tungsten-molybdenum ore processing technology”

Graduate: K. Sayfiddinov

Navoi-2014

Introduction
1. General information on methods of beneficiation of tungsten ores
2. Enrichment of molybdenum-tungsten ores
2. Technological section
2.1 Calculation of crushing scheme with equipment selection
2.2 Calculation of the grinding scheme
2.3 Selection and calculation of semi-autogenous grinding mills
List of used literature

Introduction

Minerals are the basis of the national economy, and there is not a single industry where minerals or their processed products are not used.

Significant mineral reserves in many deposits of Uzbekistan make it possible to build large, highly mechanized mining, processing and metallurgical enterprises that extract and process many hundreds of millions of tons of minerals with high technical and economic indicators.

The mining industry deals with solid minerals from which, with modern technology, it is advisable to extract metals or other minerals. The main conditions for the development of mineral deposits are increasing their extraction from the subsoil and complex use. This is due to:

- significant material and labor costs during exploration and industrial development of new deposits;

- the growing need of various sectors of the national economy for almost all mineral components that make up the ore;

- the need to create waste-free technology and thereby preventing pollution environment production waste.

For these reasons, the possibility of industrial use of a deposit is determined not only by the value and content of the mineral, its reserves, geographical location, mining and transportation conditions, other economic and political factors, but also the presence of effective technology for processing mined ores.

1. General information about methods of beneficiation of tungsten ores

Tungsten ores are enriched, as a rule, in two stages - primary gravity enrichment and finishing of rough concentrates using various methods, which is explained by the low tungsten content in the processed ores (0.2 - 0.8% WO3) and high requirements for the quality of standard concentrates (55 - 65% WO3), The total enrichment degree is approximately 300 - 600.

Wolframite (huebnerite and ferberite) bedrock ores and placers usually contain a number of other heavy minerals, therefore, during the primary gravity enrichment of ores, they strive to isolate collective concentrates, which can contain from 5 to 20% WO3, as well as cassiterite, tantalite-columbite, magnetite, sulfides, etc. When finishing collective concentrates, it is necessary to obtain conditioned monomineral concentrates, for which flotation or flotogravity of sulfides, magnetic separation of magnetite in a weak magnetic field, and wolframite in a stronger one can be used. It is possible to use electric separation, gravitational enrichment on tables, flotation of gangue minerals and other processes to separate minerals so that the finished concentrates meet the requirements of GOSTs and technical specifications not only for the content of the base metal, but also for the content of harmful impurities.

Considering the high density of tungsten minerals (6 - 7.5 g/cm 3), during enrichment, gravitational enrichment methods can be successfully used on jigging machines, concentration tables, sluices, jet and screw separators, etc. For fine dissemination of valuable minerals, flotation or a combination is used gravitational processes with flotation. Considering the possibility of wolframite sludge during gravitational enrichment, flotation is used as auxiliary process even when enriching coarsely disseminated wolframite ores for more complete extraction of tungsten from sludge.

If there are large tungsten-rich ore pieces or large pieces of waste rock in the ore, sorting of ore with a particle size of 150 + 50 mm on belt conveyors can be used to separate the rich large-lump concentrate or pieces of rock that dilute the ore supplied for enrichment.

When beneficiating scheelite ores, gravity is also used, but most often a combination of gravity methods with flotation and flotation gravity, or flotation alone.

When sorting scheelite ores, luminescent installations are used. Scheelite, when irradiated with ultraviolet rays, glows with a bright blue light, which makes it possible to separate pieces of scheelite or pieces of waste rock.

Scheelite is an easily floated mineral characterized by high sludge properties. The extraction of scheelite increases significantly with flotation enrichment compared to gravity, therefore, in the enrichment of scheelite ores in the CIS countries, flotation has now begun to be used in all factories.

During the flotation of tungsten ores, a number of difficult technological problems arise that require the right decision depending on the material composition and association of individual minerals. In the process of flotation of wolframite, hübnerite and ferberite, it is difficult to separate from them iron oxides and hydroxides, tourmaline and other minerals containing neutralize their flotation properties with tungsten minerals.

Flotation of scheelite from ores with calcium-containing minerals (calcite, fluorite, apatite, etc.) is carried out by anionic fatty acid collectors, ensuring their good flotation with calcium cations of scheelite and other calcium-containing minerals. Separation of scheelite from calcium-containing minerals is possible only with the use of such regulators as liquid glass, sodium fluorosilicone, soda, etc.

2. Enrichment of molybdenum-tungsten ores

On Tyrnyauzskaya The factory enriches the molybdenum-tungsten ores of the Tyrnyauz deposit, which are complex in the material composition of not only valuable minerals with very fine dissemination, but also associated gangue minerals. Ore minerals - scheelite (tenths of a percent), molybdenite (hundredths of a percent), powellite, partially ferrimolybdite, chalcopyrite, bismuthite, pyrrhotite, pyrite, arsenopyrite. Nonmetallic minerals - skarns (50-70%), hornfels (21-48%), granite (1 - 12%), marble (0.4-2%), quartz, fluorite, calcite, apatite (3-10%) and etc.

In the upper part of the deposit, 50-60% of molybdenum is represented by powellite and ferrimolybdite, in the lower part their content decreases to 10-20%. Molybdenum is present in scheelite as an isomorphic impurity. Part of the molybdenite, oxidized from the surface, is covered with a film of powellite. Part of the molybdenum grows very finely with molybdoscheelite.

More than 50% of oxidized molybdenum is associated with scheelite in the form of powellite inclusions - a decomposition product of the Ca(W, Mo)O 4 solid solution. Such forms of tungsten and molybdenum can only be isolated into a collective concentrate with subsequent separation by hydrometallurgical methods.

Since 1978, the ore preparation scheme at the factory has been completely reconstructed. Previously, ore, after large crushing at the mine, was transported to the factory in trolleys via an overhead cableway. In the crushing department of the factory, the ore was crushed to - 12 mm, unloaded into bunkers and then crushed in one stage in ball mills operating in a closed cycle with double-spiral classifiers, up to 60% of the class - 0.074 mm.

A new ore preparation technology was developed jointly by the Mekhanobr Institute and the plant and put into operation in August 1978.

The ore preparation scheme provides for coarse crushing of the original ore up to -350 mm, screening according to the 74 mm class, separate storage of each class in bunkers for the purpose of more precise regulation feeding large and small classes of ore into the autogenous grinding mill.

Self-grinding of coarse ore (-350 mm) is carried out in Cascade type mills with a diameter of 7 m (MMC-70X X23) with additional grinding of the coarse-grained fraction to 62% class -0.074 mm in MSHR-3600X5000 mills operating in a closed cycle with single-spiral classifiers 1KSN-3 and located in a new building on the mountainside at an elevation of about 2000 m above sea level between the mine and the operating factory.

Innings finished product from the autogenous vessel to flotation is carried out by hydraulic transport. The hydraulic transport route is a unique engineering structure that ensures the transportation of pulp with a height difference of more than 600 m. It consists of two pipelines with a diameter of 630 mm, a length of 1750 m, equipped with stilling wells with a diameter of 1620 mm and a height of 5 m (126 wells for each pipeline).

The use of a hydraulic transport system made it possible to eliminate the cargo ropeway workshop, the medium and fine crushing building, and the MShR-3200X2100 mills at the processing plant. In the main building of the factory, two main flotation sections, new scheelite and molybdenum finishing departments, a liquid glass melting shop, and recycling water supply systems were built and put into operation. The thickening front for rough flotation concentrates and middlings has been significantly expanded due to the installation of thickeners with a diameter of 30 m, which reduces losses from thickening discharges.

The newly commissioned facilities are equipped with modern automated process control systems and local automation systems. Thus, in the autogenous building the automatic control system operates in direct control mode based on M-6000 computers. In the main building, a system for centralized control of the material composition of the pulp was introduced using X-ray spectral analyzers KRF-17 and KRF-18 in combination with an M-6000 computer. An automated system for sampling and delivery of samples (by pneumatic mail) to the express laboratory, controlled by the KM-2101 computer complex and issuing analyzes by teletype, has been mastered.

One of the most complex processing processes - finishing rough scheelite concentrates according to the method of N. S. Petrov - is equipped with an automatic monitoring and control system, which can work either in the “advisor” mode to the flotation operator, or in the mode of direct control of the process, regulating the flow rate of the suppressor (liquid glass), pulp level in cleaning operations and other process parameters.

The sulfide minerals flotation cycle is equipped with automatic control and dosing systems for collector (butyl xanthate) and suppressor (sodium sulfide) in the copper-molybdenum flotation cycle. The systems operate using ion-selective electrodes as sensors.

Due to the increase in production volume, the factory switched to processing new varieties of ores, characterized by a lower content of certain metals and a higher degree of oxidation. This required improvement of the reagent regime for flotation of sulfide-oxidized ores. In particular, a progressive technological solution was used in the sulfide cycle - a combination of two foaming agents of active and selective types. Reagents containing terpene alcohols are used as an active foaming agent, and a new reagent LV, developed for the enrichment of multicomponent ores, primarily Tyrnyauz ores, is used as a selective agent.

In the flotation cycle of oxidized minerals by fatty acid collectors, intensifying additives of a modifier reagent based on low molecular weight carboxylic acids are used. To improve the flotation properties of circulating industrial products pulp, regulation of their ionic composition has been introduced. Methods of chemical finishing of concentrates have found wider application.

From the autogenous grinding mill, the ore is sent to screening. Class +4 mm is further ground in a ball mill. Mill overflow and under-screen product (--4 mm) are subject to I and II classifications.

690 g/t soda and 5 g/t transformer oil are fed into the ball mill. The classifier discharge goes to the main molybdenum flotation, where 0.5 g/t xanthate and 46 g/t terpineol are fed. After I and II cleaning flotations, the molybdenum concentrate (1.2-1.5% Mo) is subjected to steaming with liquid glass (12 g/t) at 50-70°C, III cleaning flotation and further grinding to 95-98% class --0.074 mm with a supply of 3 g/t sodium cyanide and 6 g/t liquid glass.

The finished molybdenum concentrate contains about 48% Mo, 0.1% Cu and 0.5% WO 3 with a Mo recovery of 50%. The control flotation tailings of the III and IV cleaning operations are thickened and sent to copper-molybdenum flotation with a supply of 0.2 g/t xanthate and 2 g/t kerosene. The twice purified copper-molybdenum concentrate, after steaming with sodium sulfide, is sent to selective flotation, where a copper concentrate containing 8-10% Cu (with an extraction of about 45%), 0.2% Mo, 0.8% Bi is isolated.

The tailings of the control molybdenum flotation, containing up to 0 2% WO 3, are sent to scheelite flotation, carried out through a very branched and complex scheme. After mixing with liquid glass (350 g/t), basic scheelite flotation is carried out with sodium oleate (40 g/t). After the first cleaning flotation and thickening to 60% solid, the scheelite concentrate is steamed with liquid glass (1600 g/t) at 80--90 °C. Next, the concentrate is cleaned twice more and again goes to steaming at 90--95 ° C with liquid glass (280 g/t) and is cleaned again three times.

2. Technological section

2.1 Calculation of crushing scheme with equipment selection

The designed concentration plant is intended for processing molybdenum-containing tungsten ores.

Medium-sized ore (f = 12 ± 14 units on Professor Protodyakonov’s scale) is characterized by a density c = 2.7 t/m 3 and is supplied to the factory with a moisture content of 1.5%. Maximum piece d=1000 mm.

In terms of productivity, the enrichment plant belongs to the category of medium productivity (Table 4/2/), according to the international classification - to group C.

To the factory ore D max. =1000 mm is supplied from open-pit mining.

1. Let's determine the productivity of the coarse crushing shop. We calculate productivity according to Razumov K.A. 1, pp. 39-40. The project adopted the delivery of ore 259 days a year, in 2 shifts of 7 hours, 5 days a week.

Ore strength factor /2/

where: Q c. etc. - daily productivity of the crushing shop, t/day

Coefficient taking into account the uneven properties of raw materials /2/

where: Q h..t. dr - hourly productivity of the crushing shop, t/h

k n - coefficient taking into account the uneven properties of raw materials,

n days - estimated number of working days per year,

n cm - number of shifts per day,

t cm - shift duration,

k" - coefficient for accounting for ore strength,

Calculation of annual working hours:

C = (n day n cm t cm) = 259 2 5 = 2590 (3)

Time utilization rate:

k in = 2590/8760 = 0.29 units = 29%

2. Calculation of crushing scheme. We carry out the calculation according to pp. 68-78 2.

According to the instructions, the moisture content of the initial ore is 1.5%, i.e. e.

Calculation procedure:

1. Determine the degree of fragmentation

2. Let us accept the degree of fragmentation.

3. Let’s determine the maximum size of products after crushing:

4. Let's determine the width of the crusher's discharge slots, taking the typical characteristics Z - coarsening of the crushed product relative to the size of the discharge slot.

5. Let’s check the compliance of the selected crushing scheme with the manufactured equipment.

The requirements that crushers must satisfy are listed in Table 1.

Table 1

In terms of the width of the receiving opening and the range of adjustment of the discharge slot, crushers of the ShchDP 12X15 brand are suitable.

Let's calculate the productivity of the crusher using the formula (109/2/):

Q cat. = m 3 / h

Q fraction. = Q cat. · with n · k f · k cr. · k ow. · k c, m 3 / h (7)

where c n is the bulk density of ore = 1.6 t/m 3,

Q cat. - passport capacity of the crusher, m 3 / h

k f . , k ow. , kcr, kc - correction factors for strength (crushingability), bulk density, ore size and moisture content.

The value of the coefficients is found from the table k f =1.6; k cr =1.05; k ow. =1%;

Q cat. = S pr. / S n · Q n = 125 / 155 · 310 ? 250 m 3 /h

Let's find the actual productivity of the crusher for the conditions defined by the project:

Q fraction. = 250 · 1.6 · 1.00 · 1.05 · 1 · 1 = 420 t/h

Based on the calculation results, we determine the number of crushers:

We accept 12 x 15 boards for installation - 1 pc.

2.2 Calculation of the grinding scheme

The grinding scheme chosen in the project is a type of VA Razumov K.A. page 86.

Calculation procedure:

1. Determining the hourly productivity of the grinding shop , which is actually the hourly productivity of the entire factory, since the grinding shop is the main ore preparation building:

where 343 is the number of working days in a year

24 - continuous work week 3 shifts of 8 hours (3x8=24 hours)

Kv - equipment utilization factor

Kn - coefficient taking into account the uneven properties of raw materials

We accept: K in =0.9 K n =1.0

The coarse ore warehouse provides a two-day supply of ore:

V= 48,127.89 / 2.7 = 2398.22

We accept the initial data

Let's ask ourselves about liquefaction in plums and sands classification:

R 10 =3 R 11 =0.28

(R 13 is based on row 2 p. 262 depending on the size of the drain)

in 1 -0.074 =10% - class content - 0.074 mm in crushed ore

in 10 -0.074 =80% - class content - 0.074 mm in the classification plum.

We accept the optimal circulation load With opt = 200%.

Calculation procedure:

Grinding stages I and II are represented by a type VA scheme, page 86 fig. 23.

The calculation of scheme B comes down to determining the weights of products 2 and 5 (the yields of products are found according to the general formula r n = Q n: Q 1)

Q 7 = Q 1 C opt = 134.9 · 2 = 269.8 t/h;

Q 4 = Q 5 = Q 3 + Q 7 = 404.7 t/h;

g 4 = g 5 = 300%;

g 3 = g 6 = 100%

The calculation is carried out according to Razumov K.A. 1 pp. 107-108.

1. Calculation of scheme A

Q 8 = Q 10 ; Q 11 = Q 12 ;

Q 9 = Q 8 + Q 12 = 134.88 + 89.26 = 224.14 t/h

g 1 = 100%; g 8 = g 10 = 99.987%;

g 11 = g 12 =Q 12: Q 1 = 89.26: 134.88 = 66.2%;

g 9 = Q 9: Q 1 = 224.14: 134.88 = 166.17%

Process flow diagramschleniyamolybdenum-tungsten ores.

CalculationByqualitative-quantitative scheme.

Initial data for calculating qualitative-quantitative schemess.

Extraction of tungsten into the final concentrate - e tungsten 17 = 68%

Extraction of tungsten into collective concentrate - e tungsten 15 =86%

Extraction of tungsten into molybdenum concentrate - e tungsten 21 = 4%

Extraction of molybdenum into the final concentrate - e Mo 21 = 77%

Extraction of molybdenum into tungsten flotation tailings - e Mo 18 =98%

Extraction of molybdenum into control flotation concentrate - eMo 19 =18%

Extraction of molybdenum into collective concentrate - e Mo 15 = 104%

Yield of collective concentrate - g 15 = 36%

Yield of tungsten concentrate - g 17 = 14%

Yield of molybdenum concentrate - g 21 = 15%

Yield of control flotation concentrate - g 19 =28%

Determining the yield of enrichment products

G 18 = g 15 - G 17 =36-14=22%

G 22 = g 18 - G 21 =22-15=7%

G 14 = g 13 + g 19 + g 22 =100+28+7=135%

G 16 = g 14 - G 15 =135-36=99%

G 20 = g 16 - G 19 =99-28=71%

Determining the masses of enrichment products

Q 13 = 127.89t/h.

Q 1 4 = Q 13 XG 14 = 127.89x1.35=172.6 t/h

Q 1 5 = Q 13 XG 15 = 127.89x0.36=46.0 t/h

Q 1 6 = Q 13 XG 16 = 127.89x0.99=126.6t/h

Q 1 7 = Q 13 XG 17 = 127.89x0.14=17.9 t/h

Q 1 8 = Q 13 XG 18 = 127.89x0.22=28.1 t/h

Q 1 9 = Q 13 XG 19 = 127.89x0.28=35.8 t/h

Q 20 = Q 13 XG 20 = 127.89x0.71=90.8 t/h

Q 21 = Q 13 XG 21 = 127.89x0.15=19.1 t/h

Q 22 = Q 13 XG 22 = 127.89x0.07=8.9 t/h

Determining the extraction of enrichment products

For tungsten

e tungsten 13 =100 %

e tungsten 18 = e tungsten 15 - e tungsten 17 =86-68=28 %

e tungsten 22 = e tungsten 18 - e tungsten 21 =28-14=14 %

e tungsten 14 = e tungsten 13 + e tungsten 22 + e tungsten 19 =100+14+10=124 %

e tungsten 16 = e tungsten 14 - e tungsten 15 =124-86=38%

e tungsten 20 = e tungsten 13 - e tungsten 17 + e tungsten 21 =100 - 68+4=28%

e tungsten 19 = e tungsten 16 - e tungsten 20 =38-28=10 %

for molybdenum

e Mo 13 =100%

e Mo 22 = e Mo 18 - e Mo 21 =98-77=11 %

e Mo 14 = e Mo 13 + e Mo 22 + e Mo 19 =100+11+18=129 %

e Mo 16 = e Mo 14 - e Mo 15 =129-94=35 %

e Mo 17 = e Mo 15 - e Mo 18 =104-98=6%

e Mo 20 = e Mo 13 - e Mo 17 + e Mo 21 =100 - 6+77=17%

e Mo 19 = e Mo 16 - e Mo 20 =35-17=18%

Determining the amount of metals in the product Oh enrichment

For tungsten

14 =124 x0.5 / 135=0.46%

15 =86x0.5 / 36=1.19%

16 =38 x0.5 / 99=0.19%

17 =68 x0.5 / 14=2.43%

18 =28 x0.5 / 22=0.64%

19 =10 x0.5 / 28=0.18%

20 =28 x0.5 / 71=0.2%

21 =14 x0.5 / 15=0.46%

22 =14 x0.5 / 7=1%

For molybdenum

14 =129 x0.04/ 135=0.04%

15 =94x0.04/ 36=0.1%

16 =35 x0.04 / 99=0.01%

17 =6 x0.04 / 14=0.017%

18 =98 x0.04 / 22=0.18%

19 =18 x0.04 / 28=0.025%

20 =17 x0.04 / 71=0.009%

21 =77 x0.04 / 15=0.2%

22 =11 x0.04 / 7=0.06%

Table 3. Table of qualitative-quantitative enrichment scheme

Operation no. cont.		*Q, t/h*	*, %*	*copper* *, %*	*copper* *, %*	*zinc* *, %*	*zinc* *, %*

I	*Grinding stage I*
	arrives
	crushed ore
	comes out
	crushed ore
II	*Classification*
	arrives
	CrushedbChennsth product IArt. grinding
	CrushedbChennsth product II st .grinding
	comes out
	drain
	sands
*III*	*Grinding I* I *stage*
	arrives
	Sands classification
	comes out
	Shreddedsth product
IV	*Collective* Wo 3 *-Mo* *flotation*
	arrives
	Drain classification

	TailsMo flotationAnd
	comes out
	concentrate
	tails
V	*Control flotation*
	arrives
	Tailcollective flotation
	comes out
	concentrate
	tails
VI	*Tungsten* *flotation*
	arrives
	Concentratecollective flotation
	comes out
	concentrate
	tails
	Mo *flotation*
	arrives
	Tails Wo 3 flotation
	comes out
	concentrate
	tails

Calculation of water-sludge scheme .

The purpose of calculating the water-sludge scheme is to: ensure optimal liquid: solid ratios in the operations of the scheme; determining the amount of water added to operations or, conversely, released from products during dehydration operations; determination of L:T ratios in the products of the scheme; determination of the total water requirement and specific water consumption per ton of processed ore.

To get high technological indicators ore processing, each operation of the technological scheme must be carried out at optimal values of the L:T ratio. These values are established based on data from ore dressing tests and the operating practices of existing processing plants.

The relatively low specific water consumption per ton of processed ore is explained by the presence of intra-factory water circulation at the designed plant, since the thickener drains are fed into the grinding - classification cycle. Water consumption for flushing floors, washing equipment and for other purposes is 10-15% of the total consumption.

Table 3. Table of qualitative-quantitative enrichment scheme.

Opera no.walkie-talkies cont.	Name of operations and products	*Q, t/h*	*, %*	R	W


I	*Grinding stage I*
	arrives
	crushed ore			0 , 0 25
	comes out
	crushed ore
II	*Classification*
	arrives
	CrushedbChennsth product IArt. grinding
	CrushedbChennsth product II st .grinding
	comes out
	drain
	sands
*III*	*Grinding I* I *stage*
	arrives
	Sands classification
	comes out
	Shreddedsth product
IV	*Collective* Wo 3 *-Mo* *flotation*
	arrives
	Drain classification
	Control flotation concentrate
	Tails Mo flotationAnd
	comes out
	concentrate
	Tails
V	*Control flotation*
	arrives
	Tailcollective flotation
	comes out
	concentrate
	Tails
VI	*Tungsten* *flotation*
	Incoming
	Concentratecollective flotation
	It turns out
	Concentrate
	Tails
	Mo *flotation*
	Incoming
	Tails tungstenflotation
	It turns out
	concentrate
	tails

Crusher selection and calculation.

The choice of crusher type and size depends on physical properties ore, required crusher capacity, crushed product size and ore hardness.

Tungsten-molybdenum ore by strength category is an ore of medium strength.

The maximum size of a piece of ore entering the crushing operation is 1000 mm.

To crush the ore coming from the mine, I install a jaw crusher with a simple swing jaw ShchDP 12x15. *

Crusher productivity, Q is equal to:

Q =q*L*i, t/h,

where q - specific productivity of the jaw crusher per 1 cm 2 of the discharge slot area, t/(cm 2 * h);

L is the length of the discharge slot of the neck crusher, cm;

i - width of the unloading slot, see /4/

According to the practice of operating the crushing department of the processing plant, the specific productivity of the jaw crusher is 0.13 t/cm 2 * hour.

The productivity of a jaw crusher will be determined by:

Q= 0.13*150*15.5 = 302.25 t/h.

The crusher accepted for installation provides the specified ore productivity.

The maximum size of a piece in the crusher feed will be:

120*0.8 = 96 cm.

Selection and calculation of grate screen

A grate screen with a hole size of 95 cm (950 mm) is installed in front of the crusher.

The required screening area is determined by the formula:

where Q* - productivity, t/h;

a is a coefficient equal to the width of the gap between the grates, mm. /5/ According to the layout conditions, the width of the grate screen is taken to be 2.7 m, length 4.5 m.

The practice of the crushing department of the factory shows that the ore delivered from the quarry contains about 4.5% of pieces with a particle size of more than 950 mm. Pieces of this size are delivered by a front-end loader to the ore yard, where they are crushed and again fed by the loader to the grate screen.

2.3 Selection and calculation of semi-autogenous grinding mills

IN Lately during processing gold ores In world and domestic practice, in the first stage of grinding, semi-autogenous grinding mills with subsequent cyanidation are becoming increasingly common. In this case, the loss of gold from iron scrap and crumbs is eliminated, the consumption of cyanide during cyanidation is reduced, and the sanitary conditions of working on quartz silicate ores are improved. Therefore, I accept a semi-autogenous grinding (SAG) mill for installation in the first stage of grinding.

1. Find the specific productivity for the newly formed class of the operating SSI mill, t/(m 3 * h):

where Q is the productivity of the operating mill, t/h;

- class content -0.074 mm in the mill discharge, %;

- class content -0.074 mm in the original product,%;

D is the diameter of the operating mill, m;

L is the length of the operating mill, m.

2. We determine the specific productivity of the designed mill according to the newly formed class:

where q 1 is the specific productivity of a working mill in the same class;

K and is a coefficient that takes into account differences in the grindability of the ore designed for processing and the ore being processed (Ki = 1);

K k - coefficient taking into account the difference in the size of the initial and final grinding products at the existing and designed factories (K k = 1);

K D is a coefficient that takes into account the difference in the diameters of the drums of the designed and operating mills:

K D = ,

where D and D 1 respectively, the nominal diameters of the drums of the mills being designed for installation and those in operation. (K D =1.1);

Kt is a coefficient that takes into account differences in the type of designed and operating mills (Kt=1).

q = 0.7711*1.11 =0.85 t/(m 3 h).

I accept for installation an autogenous grinding mill "Cascade" with a diameter of 7 m and a length of 2.3 m with a working volume of 81.05 m3

3. We determine the productivity of the mills for ore using the formula:

where V is the working volume of the mill. /4/

4. Determine the estimated number of mills:

n- 101/125.72 = 0.8;

then the accepted one will be equal to 1. The Cascade mill provides the specified productivity.

Screen selection and calculation II screening stage .

Draining of semi-autogenous mills using pumps...

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%

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

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

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

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

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

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

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

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

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

The separator operation diagram is shown in Fig. 1

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

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Explanatory note

for final qualifying work

on the topic of: “Selection, justification and calculation of tungsten-molybdenum ore processing technology”

Graduate: K. Sayfiddinov

Navoi-2014

Introduction

Minerals are the basis of the national economy, and there is not a single industry where minerals or their processed products are not used.

Significant mineral reserves in many deposits of Uzbekistan make it possible to build large, highly mechanized mining, processing and metallurgical enterprises that extract and process many hundreds of millions of tons of minerals with high technical and economic indicators.

The mining industry deals with solid minerals from which, with modern technology, it is advisable to extract metals or other minerals. The main conditions for the development of mineral deposits are increasing their extraction from the subsoil and complex use. This is due to:

- significant material and labor costs during exploration and industrial development of new deposits;

- the growing need of various sectors of the national economy for almost all mineral components that make up the ore;

- the need to create waste-free technology and thereby preventing pollution environment production waste.

1. General information about methods of beneficiation of tungsten ores

2. Enrichment of molybdenum-tungsten ores

A new ore preparation technology was developed jointly by the Mekhanobr Institute and the plant and put into operation in August 1978.

The ore preparation scheme provides for coarse crushing of the original ore up to -350 mm, screening according to the 74 mm class, separate storage of each class in bunkers for the purpose of more precise regulation feeding large and small classes of ore into the autogenous grinding mill.

The sulfide minerals flotation cycle is equipped with automatic control and dosing systems for collector (butyl xanthate) and suppressor (sodium sulfide) in the copper-molybdenum flotation cycle. The systems operate using ion-selective electrodes as sensors.

From the autogenous grinding mill, the ore is sent to screening. Class +4 mm is further ground in a ball mill. Mill overflow and under-screen product (--4 mm) are subject to I and II classifications.

2. Technological section

2.1 Calculation of crushing scheme with equipment selection

2.2 Calculation of the grinding scheme

Table 3. Table of qualitative-quantitative enrichment scheme.

2.3 Selection and calculation of semi-autogenous grinding mills

1. Find the specific productivity for the newly formed class of the operating SSI mill, t/(m 3 * h):

where Q is the productivity of the operating mill, t/h;

- class content -0.074 mm in the mill discharge, %;

- class content -0.074 mm in the original product,%;

D is the diameter of the operating mill, m;

L is the length of the operating mill, m.

2. We determine the specific productivity of the designed mill according to the newly formed class:

where q 1 is the specific productivity of a working mill in the same class;

K and is a coefficient that takes into account differences in the grindability of the ore designed for processing and the ore being processed (Ki = 1);

K k - coefficient taking into account the difference in the size of the initial and final grinding products at the existing and designed factories (K k = 1);

K D is a coefficient that takes into account the difference in the diameters of the drums of the designed and operating mills:

K D = ,

where D and D 1 respectively, the nominal diameters of the drums of the mills being designed for installation and those in operation. (K D =1.1);

Kt is a coefficient that takes into account differences in the type of designed and operating mills (Kt=1).

q = 0.77*1*1*1.1*1 =0.85 t/(m 3 * h).

I accept for installation an autogenous grinding mill "Cascade" with a diameter of 7 m and a length of 2.3 m with a working volume of 81.05 m3

3. We determine the productivity of the mills for ore using the formula:

where V is the working volume of the mill. /4/

4. Determine the estimated number of mills:

n- 101/125.72 = 0.8;

then the accepted one will be equal to 1. The Cascade mill provides the specified productivity.

Screen selection and calculation II screening stage .

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q = 0.7711*1.11 =0.85 t/(m 3 h).