Processing method of black-shale ores with rare metals extracting. RU patent 2493272.
| Method of processing chemical concentrate of natural uranium / 2490348 Invention relates to the technology of processing chemical concentrates of natural uranium, involving leaching (dissolving) the concentrate and extracting uranium using tributyl phosphate in a hydrocarbon diluent. The method involves dissolving the concentrate using aqueous nitric acid solution, feeding the obtained aqueous uranyl nitrate solution to the extract outputting step of a stepped extraction unit and extracting uranium with tributyl phosphate in a hydrocarbon diluent. Extraction is carried out by counterflow interaction of the aqueous and organic phases. Concentrate containing thorium impurities in ratio of 1 wt % to uranium is used. During extraction at the extract outputting step, the step for saturating the extractant with uranium is kept at least 87% of the maximum saturation of the extractant with uranium, and a portion of the aqueous phase, which is not more than 60 vol. % of the uranyl nitrate solution fed to the extract outputting step, after one of the extraction steps is removed from the extraction process and fed for dissolving the uranium concentrate. |
 Extraction method of natural uranium concentrate from sulphuric acid solutions of underground leaching, and plant for its implementation / 2489510Method involves use of an unbalanced solution consisting of a solution from the washing process of anionite from the acid and filtrate from the filter press, and their removal from the process together with a mother solution from deposition of natural uranium concentrate through an additional saturation operation together with a marketable reclaimed product. For that purpose, the plant includes a local solution recirculation circuit in the form of a collector for solutions of unbalanced and mother concentrate from deposition, which is connected to pipelines of the above solutions and equipped with solution supply pipelines attaching the collector through a gravity tank to an additional saturation column from the marketable reclaimed product and to a solution return pipeline attaching the gravity tank to the solution collector of the local solution recirculation circuit. |
 Method for extraction of uranium from ores / 2485193Method involves leaching of uranium and iron using sulphuric acid solution and ferric iron contained in the ore as an oxidiser. After leaching is completed, uranium is extracted from the solution so that mother solution containing ferrous iron is obtained. Then, acidification of the mother solution is performed using sulphuric acid and recovery of ferric iron is performed by oxidation of ferrous iron so that a reusable solution is obtained, and recirculation of that solution for leaching of uranium is performed. Recovery of ferric iron is performed by action on the mother solution of high-voltage pulse electric discharges at high voltage pulse amplitude of not less than 10 kV and at pulse repetition cycle at the interval of 400÷1400 pulse/sec. At that, prior to action on mother solution with high-voltage pulse electric discharges, it is subject to dispersion. |
 Method of uranium ore processing / 2481411Proposed process comprises crushing and grinding the ore, sulfuric acid leaching with addition of nitrogen acid as an oxidiser. Then, uranium is extracted and cleaned of impurities with the help of extractive agents mix to wash saturated extractive agent with the solution of sulfuric acid. After extraction, uranium is re-extracted to obtain uranium concentrate by means of 8-10%-solution of sodium carbonate. Uranium is deposited from re-extracted product by hydrogen peroxide with 50-100%-surplus from stoichiometry at equilibrium pH 3.6-4.2, mixing interval of 1- 1.5 h and sedimentation time of, at least, 1 h. |
| Method of extracting americium / 2477758 Invention relates to methods of extracting americium in form of americium dioxide from solutions. The invention can be used in the technology of extracting americium from production and radioactive wastes. The method involves concentrating nitric acid solution containing americium and impurities to americium content of not less than 100 mg/l by multi-step deposition of a precipitate containing americium, followed by dissolution thereof each time in a new portion of the starting solution. The precipitate containing americium is obtained from each portion of the solution by adding to 3.8-6.0 M nitric acid solution, which contains americium and impurities, ammonium hydroxide or an alkali metal hydroxide until achieving residual acidity of 0.1-0.2 M, oxalic acid to concentration of 10-50 g/l and adjusting acidity of the obtained reaction mixture to pH 0.6-2.3 if there are hydrolysable impurities in the starting solution and to pH 0.6-3.5 if not. The precipitate obtained by deposition from the americium-concentrated solution is then calcined and the calcined precipitate is then dissolved in nitric acid solution. Americium is then extracted from the obtained solution by a tributyl phosphate-based solid extractant, re-extracted, americium oxalate is deposited from the re-extract and then calcined to americium dioxide. |
 Complex processing method of carbon-silicic black-shale ores / 2477327Invention refers to complex processing method of carbon-silicic black-shale ores, which contain vanadium, uranium, molybdenum and rare-earth elements. The above method involves ore crushing to the particle size of not more than 0.2 mm and two leaching stages. Oxidation sulphuric-acid leaching is performed at atmospheric pressure. Autoclave oxidation sulphuric-acid leaching is performed at the temperature of 130-150°C in presence of oxygen-containing gas and addition of a substance forming nitrogen oxide, as a catalyst of oxygen oxidation. Ion-exchange sorption of uranium, molybdenum, vanadium and rare-earth elements is performed from the obtained product solution. |
 Method of ion-exchange uranium extraction from sulfuric solutions and pulps / 2458164Method includes uranium sorption by anion exchange resin, uranium de-sorption from saturated anion exchange resin by sulphuric acid and obtaining finished product from strippant. Note that uranium de-sorption from saturated anion exchange resin is done by sulphuric acid solution with concentration 70-100 g/l with the presence of 1-2 mole/l of ammonia sulphate. |
| Method of producing uranium tetrafluoride / 2456243 Invention relates to chemical engineering of inorganic substances and can be used to produce uranium tetrafluoride. The method of producing uranium tetrafluoride involves reduction and fluorination of triuranium octoxide with vapour from decomposition of ammonium fluoride taken in excess of 100-130 mol. % of the stoichiometric amount at temperature in the range of 260-700°C. |
 Procedure for processing uranium hexafluoride and device of implementing same / 2453620Procedure for processing uranium hexafluoride involves supply of the main stream of gaseous uranium hexafluoride into uranium-fluorine plasma generator, supply of an additional flow of gaseous uranium hexafluoride into an additional circuit to the uranium-fluorine plasma generator, forming of a cluster of uranium-fluorine plasma out of the primary and secondary streams of uranium hexafluoride at the entrance to the uranium-fluorine plasma generator. Then uranium-fluorine plasma flow is formed in the separation chamber of the magnetic separator, removal of the neutral atomic fluorine from the uranium-fluorine plasma flow, condensation of uranium, collecting of molten metallic uranium, formation of a bar of metallic uranium and output of the formed uranium bar. The precession of a cluster of uranium-fluorine plasma is performed along a conical surface in the skin layer by means of magnetic and/or gas-dynamic scanning of additional flow of uranium hexafluoride. A device for implementation of the said procedure is also suggested. |
 Method of processing natural uranium chemical concentrate / 2451761Invention relates to processing natural uranium chemical concentrate. Proposed method comprises concentrate leaching by nitric acid solution to obtain suspension, adding coagulant into suspension and suspension separation. Clarified solution is separated from residue and directed to extraction. Note here that polyacrylamide-based anion coagulant is used and suspension with said coagulant is subjected to permanent magnetic field effects. Coagulant concentration and duration of magnetic field effects are selected to ensure concentration of insoluble residue now exceeding 100 mg/l in clarified solution. In extraction from clarified solution, no antifloating emulsions are observed. |
 Method of extracting rare-earth metals (rem) from phosphogypsum / 2492255Proposed method comprises REM and phosphorus leeching by sulfuric acid solution to obtain leaching solution and insoluble residue. Said insoluble solution is processed by calcium compound to pH over 5. PEM concentrate is extracted from said solution by crystallisation and fed to REM and phosphorus leaching stage. Prior to leaching phosphogypsum is subjected to flushing with water to obtain flushing solution containing REM and phosphorus. Said insoluble residue is flushed before processing by calcium compound. Obtained flushing solution is processed by calcium compound to produce pulp with pH not over that of REM phosphate precipitation beginning and combied with said flushing solution. REM is sorbed by cation exchangers and separated to desorb REM therefrom to produce desorbent and recovered cation exchanger. Said recovered cation exchanger is sent to REM sorption while desorbent is sent to REM concentrate production stage. Phosphorus and associated impurities are deposited from sorption mother pulp. Obtained pulp is separated in residue to be recovered and water phase to be used as circulating water. |
 Method of extracting rare-earth metals from phosphogypsum / 2491362Method includes leaching of rare-earth metals (REM) from phosphogypsum with 1-5% solution of sulphuric acid, REM sorption from leaching solution with cationite, REM desorption, precipitation of REM concentrate from desorbate, obtaining REM concentrate and mother liquor, which is used for REM desorption. Cationite after desorption is returned at sorption stage. Phosphor and fluorine are precipitated from mother liquor, phosphor -and fluorine-containing sediment are filtered and filtrate is used as return water in leaching. REM leaching and sorption are carried out simultaneously. Obtained pulp is filtered through mesh filter with separation of saturated REM cationite. After that, pulp is filtered with obtaining non-dissoluble residue and mother liquor of sorption. Before desorption cationite is treated with part of desorbate. |
| Method for quantitative determination of cerium in steels and alloys / 2491361 Method includes dissolution of a sample of analysed alloy and separation of cerium from the base of the alloy and macrocomponents. At the same time the base and macrocomponents are separated from cerium by serial deposition and extraction of the alloy base and macrocomponents of the alloy from the solution. Deposition is carried out with sodium diethyldithiocarbamate, extraction - with dithizone in chloroform. After separation of the organic phase, the cerium content is detected in water phase with the spectrometric method. |
| Method of extracting rare-earth metals from phosphogypsum / 2487185 Invention is meant for extracting rare-earth metals from phosphogypsum obtained in production of phosphorus fertiliser during sulphuric acid treatment of apatite. The method of extracting rare-earth metals from phosphogypsum involves converting phosphogypsum, dissolving the converted chalk to obtain an insoluble residue containing rare-earth metals. The obtained insoluble residue containing rare-earth metals is dissolved in nitric acid solution at solid-to-liquid ratio of 1:1.5 to obtain a solution and an insoluble residue. The insoluble residue is then washed with water; the obtained solution is mixed with the washing solution; the mixed solution is neutralised to acidity of 0.5-0.25 N with concentrated aqueous ammonia solution and taken for precipitation of rare-earth metal oxalates. The oxalates are precipitated with saturated oxalic acid solution; the residue is washed with 1.5-2.5% oxalic acid solution at solid-to-liquid ratio of 1:2-3. The oxalates are then dried and calcined until rare-earth metal oxides are obtained. |
| Solid extractant for extraction of scandium and method of its production / 2487184 Solid extractant is proposed (SEX) for extraction of scandium from scandium-containing solutions, containing a styrene divinyl benzene matrix with di-(2-ethyl hexyl)phosphoric acid. At the same time it additionally contains dibenzo-18-crown-6 at the following ratio of components, wt %: di-(2-ethyl hexyl)phosphoric acid 28-30, dibenzo-18-crown-6 28-30, styrene divinyl benzene - balance, besides, the ratio of styrene and divinyl benzene in the matrix is equal to 65÷70:30÷35. There is a method also suggested for production of the above extractant. |
 Method of extracting scandium / 2485049Invention relates to hydrometallurgical processing of mineral material, particularly scandium-containing "tailings" obtained during beneficiation of titanium-magnetite ore by wet magnetic separation. The method of extracting scandium is three-step sulphuric acid leaching of scandium, wherein at the first step, leaching is carried out with recycled solution after extraction of scandium at temperature of 30-50°C and solid to liquid ratio of 1:6-7 for 3-4 hours; the pulp is then divided into a solid phase and a liquid phase; at the second step, a portion of the solution obtained from the first step is returned to the solid phase and sulphuric acid is added to concentration of 340-360 g/l and sodium fluoride is added in amount of 20-25 kg fluorine/t solid; leaching is carried out at temperature of 95-98°C and solid to liquid ratio of 1:2.5-3 for 3-4 hours; further, at the third step, the pulp is diluted in solid to liquid ratio of 1:6.5-7.5; treatment is carried out at temperature of 95-98°C for 3-4 hours. |
| Method of producing scandium-bearing concentrate from red mud / 2484164 Proposed method comprises sulfuric acid leaching of scandium from red mud, pulp filtration, scandium sorption from sulfuric acid solutions, desorption from organic phase by carbonate solution to obtain column effluent. Then, scandium poorly soluble compounds are precipitated from column effluent, precipitate is filtered out, flushed, dried and annealed to get scandium-bearing concentrate. Note here that said leaching is performed by 10.0-13.5%-sulfuric acid at pulp initial vibration cavitation at rotary velocity of 35-60 m/s for 15-35 min. Scandium is precipitated from column effluent by potassium caprinate in amount of 75-100 g/t of scandium at pH 3.5-4.5 and exposure for 15-25 min. |
 Method of extracting rare-earth metals from aqueous solutions / 2484163Proposed method comprises extraction of rare-earth metal cations by organic phase including extragent solution in inert diluter. Naphthenic acid is used as said extragent. Kerosene is used as inert diluter. Extraction is conducted in three stages at relationship between organic and aqueous phases O:A=1·(9-11) at every stage. Note here that, at first stage, europium cations (III) are extracted at content of naphthenic acid in kerosene of 10-13 vol. % and aqueous solution pH 5.0-5.1. At second stage, samarium cations (III) are extracted at content of naphthenic acid in kerosene of 13-16 vol. % and aqueous solution pH 4.6-4.7. At third stage cerium and lanthanum cations (III) are extracted at the same content of extragent and pH 5.0-5.1. |
| Method of extracting rare-earth metals from technological and productive solutions and pulps / 2484162 Method of extracting rare-earth metals from solutions containing iron (III) and aluminium comprises sorption of rare-earth metals on sorbent. Ampholyte with iminodiacetic functional groups is used as said sorbent. Sorption is carried out after preliminary neutralisation or acidification of solution to pH 4-5 by whatever alkaline or acid agent to add ampholyte in obtained pulp with separation of solid fraction. Sorption is conducted at ampholyte:pulp ratio of 1:50-1:150, phase contact time of 3-6 h and in the presence of reducing agent. |
 Method of treating rare-earth phosphate concentrate separated from apatite / 2484018Invention relates to methods of separating deactivated rare-earth elements during nitric acid treatment of apatite concentrate from nitrate-phosphate solutions. The method of treating a rare-earth phosphate concentrate isolated from apatite involves decomposition of the rare-earth phosphate concentrate with nitric acid, treating the obtained solution with oxalic acid with precipitation of rare-earth oxalates in two steps, at the first step of precipitation of oxalates of thorium and rare-earth elements, 5-10% oxalic acid in stoichiometric amount is added to rare-earth elements present in the solution, and at the second step of precipitation of rare-earth oxalates, 110-115% oxalic acid in stoichiometric amount is added to rare-earth elements present in the initial solution, and the rare-earth oxalates are then calcined to rare-earth oxides. |
 Extraction method of molybdenum from diluted acid solutions of complex composition / 2477329Invention can be used for extraction, concentration and cleaning of molybdenum from companion elements (Fe3+, Cu2+, Zn2+, Ni2+, Co2+, Al3+, Sn4+, Sb3+, rare-earth elements3+, etc.) at processing of different liquid and solid molybdenum-containing wastes and middling products. Extraction method of molybdenum from diluted acid solutions containing companion elements of molybdenum involves molybdenum deposition in the form of its salt. Besides, deposition with molybdenum purification is performed in the form of its cesium salt of 12-molybdophosphoric acid containing the following: Cs3-xHxPMo12O40·nH2O (x=0-1, n=9-12). Deposition is performed at heating up to 40-80°C by subsequent addition of orthophosphate-ion in the form of soluble phosphate or orthophosphoric acid, strong acid, for example sulphuric acid or sodium hydroxide up to pH 1-3 and soluble cesium salt, including mother solutions obtained during leaching of pollucite. |
FIELD: metallurgy.
SUBSTANCE: processing method of black-shale ores with rare metals extracting includes leaching of ore by sulphuric acid solution with dilution of rare metals. Leaching is performed in autoclave by sulphuric acid solution consisting of free and combined sulphuric acid with ratio of H2SO4(free):H2SO4(comb)=2:1, and containing 25-45 g/l of iron sulphate, 70-90 g/l of aluminium sulphate and 0.5 g/l of nitric acid. At that the process is performed under pressure in autoclave equal to 10-15 atm with mixing at temperature of 140-160°C in concentration range of general H2SO4(gen) equal to 350-450 g/l under pulp density S: L=1:0.7-0.9, preferably 1:0.8, under constant oxidation-reduction potential Eh in the system equal to 350-450 mV during 2-3 hours till residual concentration of free H2SO4(free) is within 45-75 g/l.
EFFECT: increasing break-down of ore and extraction of rare metals: vanadium, uranium, molybdenum and rare-earth elements, reducing consumption of acid and improving efficiency of autoclave volume usage.
1 tbl, 1 ex
The invention relates to the autoclave hydrometallurgy and can be used for extraction of rare metals from poor, resistant, ultra-dispersed ores.
Attempts fundamental approach to the development of theoretical bases and creation of a method of complex extraction of rare metals (vanadium, uranium, molybdenum, rare-earth elements) from the black schist ores were numerous. In the basis of the suggested ways laid the pyrometallurgical oxidative firing at temperatures 750-850°C. the Disadvantage of pyrometallurgical methods is to burn carbon formation of harmful gases and dust containing radioactive elements.
There is a method of direct leaching of rare metals from raw materials in solutions of acids, which use a variety of oxidants and complexation (Palant A.A. Direct extraction of vanadium concentrate. J. «Metals», №5, 1996). Introduction oxidants, with the redox potential of more than 330 mV, allows oxidize uranium, vanadium, molybdenum, and iron in the higher degrees of oxidation that well retrieved synthetic ionitami or mineral sorbents. The difference of the chemical properties of anionic forms of uranium, vanadium, molybdenum and phosphorus is low and does not allow to achieve a clear separation of elements one technique.
There is a method of extraction of vanadium in the solution of vanadium containing material (Sat. «Chemistry and technology of vanadium compounds», Perm, Materials of the first all-Union meeting on chemistry, technology and application of the vanadium compounds, 1974, .103-108). The method involves burning in the atmospheric conditions at temperatures 350-400°C tonnes of vanadium raw materials with solid additions of iron (+III) and sulphates ammonium and leaching cinder solutions of sulfuric acid.
There is a method (patent of Russian Federation №2148669, IPC 22 34/22, publ. 10.05.2000,), in which oxidative roasting raw materials lead in atmospheric conditions at a temperature of 150-350°C after wetting it with a solution of sulfuric acid and subsequent leaching acid. Consumption of a sulfuric acid on technological cycle support in stages: roasting: leaching=(60-80):(20-40)%.
A common shortcoming of these ways of processing containing raw materials is the increased consumption of reagents and low vanadium minerals.
A method of refining products containing metal sulfides (RF patent №2245380, IPC 22 3/08, publ. 27.01.2005,), consisting in the implementation of the leaching of processed products in the sulphamate solution concentration from 1.8 to 3.5 g/DM 3 at a temperature from 0 to 150 C in the presence of ferric ions the concentrations of more than 1 g/DM 3 and regeneration ferric undertaken compounds of elements potentials transition from the highest degree of valency in the lower higher than that of iron added to a solution by increasing the concentration of ferrous ions of iron.
Drawback of this method of processing is that when using it due to the relatively «soft» conditions leaching in atmospheric conditions not provided with high extraction of valuable components, and what happens to permanent loss of a sulfuric acid in the form of SO2 SO3) and, leaving from the reaction zone in the environment.
The closest to the technical nature of the claimed invention is a method for processing of quartzite Karatau, including ore leaching solution of sulfuric acid with the dissolution of rare metals (provisional patents KZ №12431 And, IPC 22 34/22, 22 60/02, publ. 17.12.2002, bul. №12).
The disadvantage of a low opening of refractory minerals of rare metals.
The technical result of the invention is to increase the opening of valuable components (vanadium, uranium, molybdenum and rare earth elements), increase of their extraction.
The technical result is achieved in processing technology black schist with the extraction of ores of rare metals, including ore leaching solution of sulfuric acid with the dissolution of rare metals, the leaching of lead in the autoclave acid, consisting of free and bound of sulphuric acid in the ratio of H 2 SO 4(FL) :H 2 SO 4(tie) =2:1, containing 25 to 45 g/l sulfate, 70-90 g/l of aluminum sulfate and 0.5 g/l nitric acid at a temperature of 140-160°C, in the range of the total concentration of H 2 SO 4 (Ls)equal to 350-450 g/l, with a density of pulp s:l=1:0.7-0.9, preferably 1:0,8, at constant redox potential in the system of equal Eh 350-450 mV, for 2-3 hours, to a residual concentration of free H 2 SO 4(FL) within 45 to 75 g/HP In addition, in the autoclave leaching of lead with pneumatic stirring under pressure equal to 13.5-15 ATM, and when the content of the mentioned sulphate salts 220-320 g/HP
It is established that the high-efficiency extraction of valuable components from the black schist ores can be realized in conditions of autoclave leaching those solutions of ions of iron (+III), aluminum (+III), vanadium (+IV) and nitric acid under pressure. Only the application of pressure up to 10-15 ATM. temperature 140-160°C, the total concentration of H 2 SO 4 in the range of 350-450 g/l at s:l=1:0.7-0.9, preferably 1:0,8, at constant redox potential of the system is equal Eh 350-450 mV for 2-3 hours, allows to increase the extraction of vanadium and other valuable components in the conditions of the direct-flow scheme of the material flow and the autoclave. These conditions result in the partial dissolution and solid phase with the formation of fine particulate. In the conduct of the leaching process in less than 2 hours you can not oxidize , and with a duration of process over 3 hours in an atmosphere of oxygen is the oxidation of vanadium to the highest degree of oxidation and his co-precipitation with . The free acid H 2 SO 4(FL) as an independently introduced in the process, and is formed by the hydrolysis of sulfates, iron, and destruction of sulphide minerals in the primary ore. To maintain the temperature of the process within 140-160°C allows you to get a free sulphuric acid by oxidation of sulphide sulphur ore and hydrolysis sulfate formation of insoluble hematite and . Sulfuric acid concentration equal to 350-450 g/l, is the sum of free H 2 SO 4(FL) and related H 2 SO 4(tie) at a ratio equal to 2:1. In conjunction with other distinctive features of this ratio acids translates vanadium in soluble form according to the following reactions:
A l 2 O 3 + 3 H 2 S O 4 → A l 2 ( S O 4 ) 3 + 3 H 2 O ( 1 )Maintenance of the leaching process at the expense of free and bound acids in the ratio of H 2 SO 4(FL) :H 2 SO 4(tie) =2:1 in the range of the total concentration of H 2 SO 4(Ls) equal to 350-450 g/l, allows achieve a sufficiently high extraction of precious metals at the residual concentration of free sulphuric acid 45-75 g/HP
At decrease in H 2 SO 4(FL) less than 45 g/l is observed coprecipitation of vanadium with sparingly soluble compounds of iron. At a concentration of more than 75 g/l obvious overrun acid and reagents for neutralization of productive solutions.
Recommended ratio s:l=1:0.7-0.9, preferably, 1:0,8, lets get pulp necessary mobility and high density up to 65% of solid compared with the prototype, suitable for the extraction of valuable components. When the ratio of the s:l less than 1:0.7 pulp badly mixed and when T:G more than 1:0.9 reduces the efficiency of the use of the autoclave.
The redox potential of the system is supported by elements whose potential is higher than that of iron, this capacity is equal Eh 350-450 mV and is a necessary and sufficient condition for the opening of valuable components.
Maintaining the concentration of the oxidant ions of iron (+III), aluminum (+III) and vanadium (+IV), nitric acid and sulfuric anhydride (SO3 ) under the pressure of 10-15 ATM, increases the speed and depth of opening of rare metals from spinel (oxides in lower oxidation States) and reduction of the time of the implementation process.
Salt content includes aluminium sulphates(+III), iron(+III), vanadium(+IV) and free sulphuric acid. The iron, vanadium and sulphurous anhydride provide oxidative leaching process, and sulfate participates in the process of formation of free H 2 SO 4(FL) . Upper limit salt 320 g/l is the limit of solubility, and the lower 220 g/l provides the required acidity in the process of leaching.
The proposed method of autoclave leaching provides effective extraction of valuable components to the content in the sludge, %: V 2 O 5-0.05, U-0.0004, Mo-0.001, REE-0.01), as well as reduce the consumption of a sulfuric acid in 2-2 .5 times in comparison with the prototype by reduction of losses in the atmosphere SO 3 and SO 2 , hydrolysis iron sulfates and oxidation of sulphide sulphur ore, which is involved in the process. Improving the efficiency of the volume of the autoclave is achieved by increasing the density of pulp, providing efficient leaching.
Example.Autoclave leaching process is conducted on ore containing 0.57% V 2 O 5 , 0.02% U, 0.03% Mo and 0.09% REE. Hanging of ore is 400 g, particle size - 0.2 mm (100%). kneaded with a solution containing 200 g/l H 2 SO 4
(free) and 0.5 g/l nitric acid (HNO 3 ) and 40 g/l sulphate and 90 g/l of aluminum sulfate with s:l=1:0.8. The pulp is loaded into the autoclave and lead the process at a temperature of 150 C, pressure of 10 ATM. with stirring, for 3 hours, and the ratio of H 2 SO 4(FL) :H 2 SO 4(tie) =2:1. At the end of the process autoclave leaching cool, the solution is filtered, and the residue is washed with a 3% solution of H 2 SO 4 and hot water in the 2 stage of the calculation of the s:l=1:1.
Results of experiments are given in table.
2. The method according to claim 1, characterized in that the autoclave leaching are pneumatically stirring under the pressure of 13.5-15 ATM.
3. The method according to claim 1, characterized in that the autoclave leaching is conducted at the content of the mentioned sulphate salts 220-320 g/HP