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Method of opening perovskite concentrates Method includes preliminary mechanical treatment of perovskite concentrates and subsequent treatment of activated concentrates with nitric acid solution HNO3. HNO3 treatment is carried out on activated perovskite concentrates with stored overall energy which corresponds to change in the surface of coherent scattering regions and microdeformations of not less than 16 kJ/mol perovskite. The activated concentrates are treated with 30% HNO3 solution at temperature of 90-99°C. |
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Method of processing eudialyte concentrate Method involves decomposing the concentrate with inorganic acid to obtain a gel, heat treatment of the gel, regenerating the acid, water leaching the gel with transfer of rare-earth elements into the solution and a zirconium compound into an insoluble residue. The solution is then separated from the residue and the zirconium compound is separated from the residue. The concentrate is decomposed with acid consumption of 90-110% stoichiometric amount, and heat treatment of the gel, water leaching of the gel and acid regeneration are carried out simultaneously in autoclave conditions at temperature of 175-250°C for 1-4 hours to obtain a solution of rare-earth elements containing a free acid. The zirconium compound is separated from the insoluble residue by wet gravity separation. The inorganic acid used is hydrochloric acid or nitric acid. |
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Method of processing magnesite dolomite raw material Invention relates to the mining and processing industry and can be used in utilisation of wastes of magnesite ore mining and benefication. A method of a magnesite dolomite raw material processing includes raw material crushing, classification and further leaching of magnesium by an acid. Particles with coarseness of 0.2-4.0 mm are separated from the magnesite dolomite raw material. Leaching is carried out at a temperature of 15-20°C by 10-50% water solution of nitric acid with stoichiometric composition: dolomite from 1.1:1 to 1.3:1 with dolomite dissolution. A magnesite sediment is separated from the obtained solution, containing magnesium and calcium ions, by filtration, sulfuric acid is added into a solution with precipitation of calcium in the form of gypsum and obtaining nitric acid as a result of reaction Ca(NO3)2+H2SO4+2H2O→CaSO4·2H2O↓+2HNO3. Gypsum is separated by filtration, the solution is blown with carbon dioxide, the sediment of magnesium hydrocarbonates is obtained and separated by filtration. The obtained solution of nitric acid is conditioned to the required concentration and supplied to leaching. |
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Extraction of rare-earth elements from hard materials containing rare-earth metals Proposed method comprises acid leaching of solid materials ground to at least 100 mcm by the mix of sulphuric and nitric acids at their ratio of 6:1 to 1:1 parts by weight and concentration of the mix of acids lower than 15 wt % and F:S ratio varying from 2:1 to 6:1 parts by weight. At acid leaching vacuum-pulse effects are applied during the entire process of rare-earth elements conversion into solution and precipitation of remained solid material. Obtained precipitate of solid fossil and man-made materials is separated from leaching solution. Separation of rare-earth elements from leaching solution is performed with the help of ion-exchange or membrane filter. |
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Method of processing silicon-containing chemical concentrate of natural uranium Method involves leaching a concentrate with aqueous nitric acid solution at high temperature to obtain a pulp which consists of a solid phase and an aqueous phase, filtering off the aqueous phase in form of uranyl nitrate solution, extraction refining uranium using tributyl phosphate in a hydrocarbon diluent. The filtered uranyl nitrate solution, which contains uranium in concentration of 200-400 g/l, dissolved silicon in concentration of 1.0-3.2 g/l and nitric acid in concentration of 1-2 mol/l, is held until stabilisation of viscosity before being fed for extraction. |
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Method of production of gold nanoparticles from raw material containing iron and non-ferrous metals Invention relates to the technology of production of gold nanoparticles. The method of production of gold nanoparticles from the raw material containing iron and non-ferrous metals comprises preparation of the chlorazotic acid solution of gold using chlorazotic acid. Then floatation extraction of gold precursors is carried out with cationic surfactants from the solution, separation and evaporation of the organic phase to concentrate the gold precursors. Then the concentrate reduction is carried out to obtain dispersion of gold nanoparticles. At that the starting material is first treated with hydrochloric acid to form the insoluble precipitate. Production of chlorazotic acid solution is carried out by dissolving in chlorazotic acid solution of insoluble precipitate. Before floatation extraction of precursors the nitric acid is removed from chlorazotic acid solution with methyl or ethyl alcohol or hydrochloric acid. |
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Method to extract metals from silicate slags Method includes grinding of slag and leaching. To prevent formation of non-filtered pulps, specified for gel formation of silicic acid, slag is previously mixed with concentrated acid (nitric or hydrochloric), taken in quantities necessary to neutralise slag, pulp is maintained for 1-2 hours. At the same time precious components are leached, and the produced silicic acid coagulates, forming large agglomerates. After this the mass is additionally ground and leached with water. At the stage all salts are washed into the solution, and gels are not produced. Further the solution is separated by filtration or whirling and processed by available hydrometallurgical methods, and the hard silicate product is sent to a dump. |
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Preparation of uranium-bearing stock for extraction Proposed process comprises leaching of uranium by nitric acid and separation of water phase from undissolved precipitate. Then, undissolved precipitate is mixed with fluorine-bearing agent, dissolution of produced charge and/or charge as a suspension in nitric acid solution. Produced solution is returned to production process for extraction of uranium. Nitric acid concentration in solution makes at least 2 mol/l. Dilution is carried out at fluorine-ion concentration at, at least, 15 g/l. Dilution is performed at 60-100°C. |
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Method of cleaning of silver-bearing material Invention relates to cleaning of silver-bearing materials by hydrometallurgy processes, for example, scrap and wastes of microelectronics. Proposed method comprises dilution of silver-bearing material in nitric acid, addition of sodium nitrate to nitrate solution at mixing, extraction of silver salt precipitate and pits treatment to get metal silver. Note here that after addition of sodium nitrate the reaction mix is held for 1 hour to add sodium carbonate or bicarbonate to pulp pH of 8-10. Free silver salt precipitate as silver carbonate is separated from the solution by filtration. Sodium nitrite and carbonate or bicarbonate is added in the dry form. Note here that sodium nitrite is taken with 25% excess of stoichiometry. |
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Extraction of bismuth and germanium from wastes of bismuth orthogermanate crystals production Invention relates to hydrometallurgy of scattered elements, particularly, to extraction of bismuth and germanium from secondary stock sources, in particularly, to extraction of bismuth and germanium from oil-abrasive wastes of bismuth orthogermanium crystals production. Proposed method comprises hydrochloride-acid leaching of bismuth and extraction of bismuth from the solution by electrolysis. Said hydrochloride-acid leaching is performed with addition of surfactants to the solution to produce abrasive-germanium-bearing precipitate. Germanium is extracted from said precipitate by distillation of tetrachloride germanium in vapours of hydrochloric acid. Said surfactant represents the commercial mix of oxyethylated alkyl-phenols of commercial grade "АФ" 9-6 at the concentration of 0.01-0.1 wt %. |
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Method of extracting rare-earth elements from phosphogypsum Invention can be used in chemical industry to extract rare-earth elements from phosphogypsum. The method involves carbonising phosphogypsum to obtain phosphochalk, dissolving said phosphochalk in nitric acid to obtain a product suspension and then separating the insoluble residue - a black concentrate of rare-earth elements - by filtering. The product suspension is divided into two parts, one of which is fed for filtration to separate the insoluble residue - black concentrate, and the second is fed for premixing with nitric acid. The mixing process is carried out for 3-5 minutes with the ratio CaO/HNO3 equal to 0.75-2.25. When dissolving, pH is kept at 2.2-4. |
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Method of processing natural-origin uranium-bearing ore Proposed method comprises leaching the stock nitric acid solution to obtain suspension, introducing coprecipitator therein at 30-50°C and mixing. Then, clarified solution is separated from insoluble residue and directed for extraction. Said coprecipitator represents fresh solution of copolymer of acrylamide and chloride trimethyl ammonium ethyl acrylate of molecular weight of 3-15 millions with low charge density. Copolymer is introduced to concentration of 5.95-11.9 mg/l of insoluble residue. Prior to separation of clarified solution from insoluble residue settling is performed for 30-40 minutes. |
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Method of extracting rare-earth metals from phosphogypsum 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. |
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Method of treating rare-earth phosphate concentrate separated from apatite Invention 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. |
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Processing method of scraps of anodes of tantalum oxide-semiconductor capacitors Method involves cleaning of scraps by acid treatment with removal of manganese dioxide. Then, deoxidation of cleaned scraps, its hydration, grinding, dehydration at increased temperature is performed so that tantalum hydride powder is agglomerated and tantalum capacitor powder is obtained. At that, acid treatment is performed at room temperature using the solution containing 100-300 g/l of sulphuric acid and 110-300 g/l of hydrogen dioxide, or the solution containing 30-150 g/l of hydrochloric acid and 75-225 g/l of hydrogen dioxide. Scrap hydration is performed by treatment using the solution of hydrofluoric acid with concentration of 1-5%. The obtained capacitor tantalum powder provides specific charge of up to 7300 mcC/g, breakdown voltage of more than 200 V and leakage current of 0.0001-0.0003 mcA/mcC when being used in anodes of tantalum oxide-semiconductor capacitors. |
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Method of producing scandium oxide Proposed method comprises dissolving scandium-bearing concentrate in sulfuric acid, removing acid-insoluble residue, and precipitating scandium in the presence of ammonium compounds. Then, precipitate is filtered, flushed, dried and calcined to obtain scandium oxide precipitate. With acid-insoluble residue removed, sulfuric acid concentration in filtrate is increased to 540-600 g/dm3, ammonium chloride is added to solution in amount of 26.7-53.5 g/dm3 at 50-70°C and held for one hour at mixing. Produced precipitate is flushed by ethanol at volume ratio of 1-10-11. |
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Processing method of low-grade molybdenite concentrates Method involves two-stage processing of concentrates with water solution of nitric acid, pulp filtration so that cake and molybdenum-containing solution is obtained. Then, calcium molybdate suitable for making of ferromolybdenum is deposited from the solution. Decomposition of concentrates is performed at addition to water solution of nitric acid, sulphuric acid in the quantity sufficient for retention of the whole molybdenum in the solution in composition of water-soluble sulphate compounds of molybdenyl, and namely with anionic complex [MoO2(SO4)2]2-. |
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Processing method of micro production wastes of constant magnets Method involves oxidation of micro production wastes at temperature of 550-650°C in air atmosphere for destruction of crystal latitude Nd2Fe14B so that Fe2O3, Nd2O3, Fe2B is formed and moisture and oil is removed. Then, anhydrous fluorides of rare-earth metals are obtained and their metallothermic reduction is performed for production of constant magnets. After oxidation from oxidated microwastes is completed, rare-earth metals are leached with nitric acid with concentration of 1-2 mol/l at temperature of 20-80°C. Obtained nitrate solutions containing rare-earth metals and impurity elements are processed with solution of formic acid with extraction of formiates of rare-earth metals in the form of the deposit cleaned from impurity elements, which includes iron, aluminium, nickel, cobalt, copper and other transition metals. |
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Treatment method of sulphide concentrates containing precious metals Method involves leaching with further separation of non-soluble residue from the solution, its drying and further melting when it is mixed with sodium carbonate, silicon-containing flux, borax so that alloy of precious metals and slag is obtained. At that, original concentrate is subject to leaching by means of nitric acid solution. Melting is performed using the addition of sodium chloride to mixture. Concentrate is leached using nitric acid solution with mass concentration of 350-550 g/l. Sodium chloride is added to mixture for melting purpose in quantity which is more by 10-20% than stoichiometric quantity as per lead chloride obtaining reaction. |
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Method of processing natural uranium chemical concentrate Invention 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. |
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Method for obtaining dephosphorised concentrate of oolitic iron ores Invention refers to preparation of iron-ore raw material for metallurgical treatment by cleaning the latter from harmful impurities deteriorating the quality of obtained metals and alloys. Method for obtaining dephosphorised concentrate of oolitic iron ores involves high temperature treatment, cooling and leaching of concentrate with mineral acid. High temperature treatment of iron-bearing material is performed in the range of 1350-1450°C in reducing medium with participation of clinker minerals till molten metal and sinters are formed. They are cooled to magnetising roasting temperature of 750-860°C, crushed and separated with magnetic separation into clinker and concentrate. Then, concentrate is cooled to 50-90°C and supplied at that temperature for leaching with mineral acid for dilution of phosphorus. |
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Method of processing manganese-containing material Method involves transfer of manganese and accompanying impurities into a solution through two-step treatment of the starting material with hydrochloric acid and absorption of chlorine with an alkaline solution. Further, impurities are separated to obtain a manganese salt solution which is then treated. The first step uses waste hydrochloric acid with concentration 1-10% with solid to liquid ratio equal to 1:(3-5). A manganese-containing residue is separated from the obtained pulp, where said residue is then treated at the second step with waste inhibited hydrochloric acid with concentration 20-24% and content of inhibitor of not less than 5 wt %, reaction with iron of which results in insoluble complex compounds, where said inhibitor is in form of quaternary ammonium salts, with molar ratio manganese:HCl=1.0:1.1. The insoluble residue of aluminosilicates is then separated and the manganese salt solution is then processed using existing methods. |
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Extraction of metals from sulphide minerals Method has been elaborated for two-stage dilution of nickel in leaching acid. Suspension of mineral and acid is activated by oxidation. It is performed during T1 time by means of electrolysis or alternatively chemically, by adding for example of oxidating acid to mineral. After activation the suspension is exposed in oxygen-free conditions during T2 time. During T2 time much quicker dilution of sulphide begins; quick decomposition of sulphide gives the possibility to nickel to be diluted and thus leached from mineral. Diluted nickel is extracted from leaching acid for example by electrochemical extraction. |
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Method of processing chemical concentrate of natural uranium Method involves leaching the concentrate with aqueous nitric acid solution at high temperature to obtain a pulp consisting a solid and an aqueous phase. The aqueous phase is then separated by filtration from the solid phase in form of uranium nitrate solution. Uranium is then extracted from the nitrate solution using tributyl phosphate in a hydrocarbon solvent. The extract is washed and uranium is re-extracted. Leaching is carried out by adding nitric acid and water in an amount which enables to obtain a nitrate solution in the aqueous phase of the pulp, said nitrate solution containing dissolved silicon in concentration of 2.5-3.7 g/l. The solid phase, which consists of insoluble concentrate residues, is separated by filtration from the solution which contains dissolved silicon, uranium in concentration of 170-250 g/l and nitric acid in concentration of 80-120 g/l. Filtration is carried out not more than 24 hours after leaching, preferably not more than 5 hours after leaching. |
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Method of processing chemical concentrate of natural uranium Method involves leaching in order to dissolve uranium when the concentrate reacts with nitric acid solution to obtain pulp from the concentrate. Uranium is then extracted from the pulp using tributyl phosphate in a hydrocarbon solvent. The extract is washed and uranium is re-extracted. Extraction is carried out from freshly prepared pulp which is obtained through direct-flow reaction at temperature 20-65°C of a stream of a suspension of the concentrate in water which is prepared beforehand and a stream of nitric acid solution with flow rate ratio which ensures nitric acid concentration in the pulp of 25-120 g/l. The period from the beginning of leaching to the beginning of extraction is not more than 10 minutes. |
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Method for obtaining cobalt and its compounds Method for obtaining cobalt and its compounds involves conversion of cobalt from cobalt-bearing raw material to the solution; deposition of cobaltic hydroxide (III) using oxidiser and neutraliser, dilution of cobaltic hydroxide (III) with conversion of cobalt (III) to cobalt (II). Then, the obtained solution is cleaned from impurities and metallic cobalt or its compounds are obtained. Cobaltic hydroxide (III) is diluted with the participation of recovered reduced forms of iron and/or copper salts in the range of oxidation-reduction potential of 300-700 mV relative to silver-chloride comparison electrode and pH 1-3, thus adjusting the temperature of the process by means of evaporation cooling. Recovery of reduced forms of iron and/or copper is performed in a separate unit using the reducer. At insufficient amount of copper and iron in raw material supplied to the process there performed is cleaning of leaching solution from copper by cementation and recirculation of cleaning product to the recovery stage of reduced forms of iron and/or copper. |
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Manufacturing method of concentrate of precious metals from sulphide copper-nickel raw material Method involves separation of Bessemer matte into metallised and sulphide fractions and oxidation leaching of metallised fraction with solution at the controlled chlorine supply. At that, ratio of sulphur and copper contained in metallised fraction is 0.3-0.7. Leaching is performed in the value range of oxidation-reduction potential of 430-470 mV, preferably 440-450 mV. Sludge of baths of electrolytic nickel refining of processing technology of sulphide Bessemer matte fraction is supplied together with metal faction of Bessemer matte for leaching. |
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Method for extracting rare metals from ash-slag masses of used underground gas generator Intensified extraction of underground water is performed from used underground gas generator through the main water drain wells. Then, there created is cone of depression in the section of used underground gas generator for minimisation of underground water level and maintenance of maximum depression in the section of used underground gas generator. Treatment of ash-slag masses is performed using the solvent injected to the used underground gas generator, and leaching and collection of rare metal dissolved in the solvent extracted from underground gas generator is performed in surface chemical complex. |
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Procedure for extraction of aluminium and iron from ash-and-slag waste Procedure for extraction of aluminium and iron from ash-and-slag waste consists in treatment with solution of sulphuric acid and in extraction of aluminium containing components into solution. Before extraction of aluminium containing components into solution waste is subjected to classification and multi-stage magnetic separation at periodic increase of field of magnetic induction for complete extraction of magnetic fraction containing iron. |
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Method of extracting lanthanides from apatite concentrate Apatite concentrate is decomposed with nitric acid to obtain nitrate-phosphate solution. Calcium nitrate and sodium fluosilicate are successively extracted from the solution. The solution is neutralised with ammonia to form a precipitate of phosphates of the main part of lanthanides and a mother solution with the residual part of lanthanides. The mother solution is treated by bringing it into contact with an amorphous titanium phosphate-based sorbent at pH of the solution equal to 1.4-1.7 with concentration of the residual part of lanthanides in the sorbent. The lanthanide-rich sorbent is separated, washed with water and treated with 0.5-3.0 M mineral acid with desorption of lanthanides. The amorphous sorbent used is one of the sorbents with the following composition: TiOHPO4·nH2O, (Ti1-xZrx)OHPO4·nH2O, TiONH4PO4·nH2O, (Ti1-xZrx)ONH4PO4·nH2O, where n≥1, x=0.05-0.5. |
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Heap leaching method of antimony ores Method involves ore pre-benefication with X-ray radiometric separation and further heap leaching of antimony from ore. At that, leaching is performed with solution containing sulphuric acid and sodium chloride in presence of manganese compounds of oxidation degree of not less than 3 or with solution containing chlorhydric acid, active chlorine and sodium chloride, or with solution containing sulphuric acid and sodium chloride in presence of chlorinated lime. Productive antimonous chloride solution obtained after leaching is supplied for removal of antimony with cementation, hydrolysis or electric deposition. After removal of antimony the solutions are regenerated with release of manganese compounds and returned for leaching. |
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Procedure consists in leaching mineral raw material containing gold in water cyanic solution with active oxygen at hydro-dynamic effect facilitating cavitation mode at swirling flow of suspension of mineral stock containing gas and solid phase. Also, leaching is performed at increased rate of swirling of suspension flow till generation of the mode of super-cavitation in flow for destruction of crystal lattice of minerals and simultaneous after-crumbling solid phase with participation of free radicals of guaranteed electron of atomic oxygen and hydrogen formed in the process of super-cavitation. The transition reactor consists of an inlet branch, of a disperser with a connecting pipe for supply of oxygen into a united cavity of cylinder channels, of a cavitation chamber, of a cascade of hydro-dynamic emitters; of damping nozzle and of an outlet branch. Also, the disperser is made of porous material with nano structured pores structurally facilitating division of suspension into small flows. The cavitation chamber is made in form of an elastic collar controlling flow area into mutually perpendicular directions by means of clamping mechanisms forming a cascade of controlled hydro-dynamic emitters. |
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Procedure for processing final tailings of galvanic production Procedure for processing final tailings of galvanic production consists in crumbling, leaching, separation of solution from sedimentation and in extracting heavy non-ferrous metals from produced solution. Also, final tailings are crumbled with mechanic-chemical activation by wet crumbling in form of pulp suspension at pH≤3 and ratio s (solid): l (liquid) = 1:(0.4-1) and temperature 60-90°C. |
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Procedure for production of refined silver Procedure for production of refined silver consists in leaching industrial products containing silver in nitric acid, in separation of non-soluble residue from nitrate solution of silver, in silver reduction, in separation of produced silver powder from solution and its re-melting. Upon separation of non-soluble residue nitrate solution of silver is treated with hydroxide of alkali metal to pH (1-3) and is heated to temperature 80-100°C. Silver is reduced by treatment of nitrate solution of silver with sulphite of alkali metal or ammonia, and reduction is finished, when value of redox potential (relative to a chlorine-silver reference electrode) reaches 200-250 mV. Before re-melting produced silver powder is flushed in diluted solution of hydrofluoric acid. |
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Procedure for extraction of rare earth elements from phospho-gypsum Procedure consists in acid extraction of compounds of rare earth elements from phospho-gypsum by mixing extract suspension, in separation of insoluble sedimentation of crystal gypsum from extraction solution and in extracting compounds of rare earth elements from extraction solution. Also, acid extraction is performed with solution of mixture of sulphuric and nitric acids at ratio from 3.2 to 1.2 and concentration from 1 to 3 wt % at ratio L:S (liquid : solid) from 4 to 5 during 8-12 min with simultaneous hydro-acoustic effect onto mixed extraction suspension. Rare earth elements are extracted from extraction solution by means of cation-exchanging sorption with infiltration of extraction solution through cation-exchanging filter. |
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Procedure for processing gold-antimonial concentrates Procedure consists in concentrate stage leaching with conversion of antimony into solution at first stage and conversion of gold into solution at second stage. At the first stage leaching is performed with solution containing sulphuric acid and sodium chloride at presence of compounds of manganese at oxidation degree not below 3 or with solution of active chlorine in hydrochloric acid solution; further, obtained productive solution of antimony chloride is directed for antimony extraction by cementation, hydrolysis and electric settling. At the second stage cake of the first stage is leached with solution containing HCl and compounds of manganese at degree of oxidation not below 3 in form of salts of manganic acid or manganese minerals, or solution of HCl with active chlorine, which is followed with extraction of gold from solution. Upon extraction of gold and antimony solutions are re-generated with extracting compounds of manganese by means of chemical compounds and are returned to leaching. |
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Method of processing gold-containing quartz ore for gold extraction Proposed method comprises crushing, fine-crushing and vatting of gold from crushed ore. Note here that prior to fin-crushing, crushed ore is subjected to bacterial vatting using silicate bacteria. Vatting gold from crushed ore is carried out by means of nitrohydrochloric acid. |
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Processing method of slurries of galvanic productions Method involves mixing of slurries with additions, crushing, heat treatment, leaching, filtration and extraction of heavy metals from solution. Mixing is performed at crushing with method of mechanical-chemical activation with additions in the form of compounds containing chloride- or sulphate ions in ratio of chloride- or sulphate ions to sum of metals included in slurry of not less than 1:1. Heat treatment is performed at 550-600°C. Further leaching of obtained sinter is performed with acid waste water of galvanic production with similar composition as to content of ions of heavy metals at pH ≤ 3. Extraction of heavy metals from obtained solution is performed with floatation at pH=8-12. |
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Method for processing mixed copper ores containing precious metals Method involves crushing of initial ore to fraction size of 30÷80 mm. Then there performed is heap leaching, extraction, re-extraction and electric extraction of cathode copper from liquid phase of leachant. After leaching the final crushing of solid leachant phase and floatation of finally crushed solid phase is performed at pH=7-8 created with caustic soda so that copper concentrate is obtained and gold is extracted to it. |
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Method involves dissolution of Tc-Ru alloy, extraction of ruthenium fraction in the form of hydroxide, conversion of ruthenium hydroxide by means of annealing to RuO4 ruthenium oxide and its hydrogen reduction to metal. Alloy is dissolved in acid solution in presence of catalyst at bubbling through solution of ozone-oxygen or ozone-air mixture. Reduction of ruthenium oxide to metal is performed in hydrogen-inert gas mixture. As catalyst there used are argentum salts Ag(I) or cobalt salts Co(II). |
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Procedure for hydro-metallurgical treatment of rhenium containing molybdenum concentrate Procedure for hydro-metallurgical treatment of rhenium containing molybdenite concentrate consists in rhenium and molybdenum autoclave leaching with solution of nitric acid and in producing solution containing nitric and sulphuric acids. Further, residue in form of molybdenum acid is filtered and washed; molybdenum acid is dissolved in ammonia water and molybdenum and rhenium are extracted. Upon autoclave leaching rhenium is extracted from solution by sorption in two stages. At each stage duration of phase contact is 22-24 hours. Summary concentration of sulphuric and nitric acids at the first stage is maintained at ≤120 g/l and pH value at the second stage is maintained at 2-4. Molybdenum is extracted from a merged solution produced from the solution after sorption extraction of rhenium and from ammonia solution of dissolved molybdenum acid. Molybdenum is extracted by sorption in two stages at duration of phase contact 22-24 hours and maintaining pH value=1.5-2.0 at the first stage and pH=2.5-4.0 at the second stage. |
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Method of selective extraction of palladium, platinum and rhodium out of concentrates Method consists in successive leaching and in dissolving each extracted metal: palladium - with solution of nitric acid, platinum - with mixture of nitric and hydrochloric acid and rhodium with solution containing hydrochloric acid; also each producing solution is removed. Notably, non-soluble residue with dissolved rhodium is subjected to leaching after dissolving of platinum; this leaching is carried out by dissolving sodium hypochlorite in solution of nitric acid diluted with water. All extracted metals are leached by bubbling reaction mass with air supplied via a porous partition from beneath. |
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Procedure for processing chemical concentrate of natural uranium Invention refers to procedures for processing chemical concentrate of natural uranium. The procedure consists in dissolving natural uranium in solution of nitric acid, in directing solution to uranium extraction with tri-butyl-phosphate and hydrocarbon thinner, in washing extract with re-extract and in uranium re-extracting. There is dissolved chemical concentrate of natural uranium with increased content of phosphorus and sulphur. Uranium is extracted from solutions with increased concentration of phosphorus and sulphur and with concentration of nitric acid 30÷80 g/l to 60÷75% saturation of extractant with uranium. Also this level of extractant saturation with uranium in the said range in % is maintained not exceeding value equal to (0.3×[HNO3]+51), where [HNO3] is concentration of nitric acid in solution directed to extraction, in g/l. |
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Procedure for complex processing nickel-cobalt raw material Procedure consists in autoclave nitrogen-acid nickel leaching from cobalt and in extracting non-soluble iron-containing residue. Residue upon washing is withdrawn from the process in form of iron containing raw material. Residue is extracted from producing solution by neutralisation to pH=3.5-4.5 with magnesium containing reagent, whereupon residue is recycled for autoclave raw material leaching. Further there is performed sorption from neutralised producing solution in two successive stages. Nickel is adsorbed on the first stage, while cobalt - on the second. Produced raffinate containing magnesium nitrate is divided in two parts; one part is recycled to nitrogen-acid autoclave raw-material leaching, while the other part is subject to thermo-decomposition producing magnesium oxide, water phase and nitrogen oxides, which are combined with nitrogen oxides extracted at stage of autoclave leaching and converted into nitric acid. Produced nitric acid and water phase formed at thermo-decomposition of raffinate are directed to return, while magnesium oxide is directed to neutralisation as magnesium containing reagent and/or as a commodity product. |
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Procedure for processing oxidised nickel-, cobalt-, iron-, and magnesium-containing ore Procedure consists in crushing and screening ore lumps with separation into small and big fractions. Further there is performed separate said fractions of ore pelleting with solutions of acid; thus there are formed independent objects out of pellets of the said fractions of ore. Nickel, cobalt and associated metals leaching out of an object of pellets of small fraction ore is performed with solution of acid of source concentration not less, than 0.25 moles/dm3; thus mother solution is produced. Nickel, cobalt and associated metals leaching out of an object of pellets of big fraction ore is performed with acidulous said mother solution till producing solution containing salts of nickel, cobalt, iron and magnesium is obtained. Further producing solution is neutralised to the value of pH=1-3, and nickel and cobalt are ion-change selectively sorbed out of producing solution. Produced raffinate contains water soluble salts of iron and magnesium. Upon sorption of nickel and cobalt one part of the said raffinate is recycled to ore leaching; while the other part of raffinate is withdrawn to utilisation. |
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Processing method of resistant copper raw material containing noble metals Invention relates to method of treatment of resistant sulphide-oxidised copper ores, as well as analogous concentrates, containing different forms of sulphide minerals and noble metals. Treatment method of resistant copper raw material containing noble metals includes leaching in water solution of sulphur acid at its concentration not less than 60 g/dm3. Leaching is implemented at atmospheric pressure at presence of hydrogen peroxide and ions of ferric iron with conversion of copper into liquid phase. After leaching it is implemented separation of liquid and solid phases and extraction of metals from it. Additionally from solid phase it is implemented extraction of noble metals and copper by means of flotation in sulphuric acid at pH not more than 2. |
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Dissolving method of alloy tc-ru Method includes dissolving of alloy with oxidation of ally components. Additionally dissolving is implemented by acid bathes, containing silver salts (I) with barbotage of acid bath by ozone-bearing gas. In the capacity of acid bathes there are used nitro-acid or chlorine-acid solutions. In the capacity of ozone-bearing gas it is used nitro-acid or ozone-air mixture. |
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Method of enrichment of persistent base ores and extraction from it of precious metals Invention relates to method of enrichment of persistent and base ores and extraction from its of noble metals, particularly gold, platinum, palladium. Method includes irradiation of ore by microwave frequency field, magnetic separation with formation of non-magnetic residue and magnetic and weak-magnetic fractions. Magnetic and/or weak-magnetic fractions are treated by solution of acid and/or oxidant at irradiation by microwave frequency field with receiving of suspension and passing of noble metals into solution. Additionally initial ore before irradiation by microwave frequency field subject to mechanical activation with usage of surface active agents with receiving of grinned ore with grain size 50-500 mcm. Received after treatment by solution of acid and/or oxidant at irradiation by microwave frequency field suspension is filtered with receiving of filtrate which is a concentrate of noble metals. |
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Preparation method of resistant sulphide ore and concentrates to leaching Method of ores preparation for leaching includes disintegration and/or ore sintering, bacterial oxidation of ore by means of treatment by sulfuric solution, containing components, activating bacteria growth, with introduction of colonies of bacteria strains, oxidising iron, copper and sulfur in ore. Before bacterial oxidation it is implemented preliminary treatment of ore oxidising sulfide minerals with usage of oxidising solution, received as a result of sulfuric solution barbotage by air, irradiated by ultraviolet rays in the wave range, providing ozone generation, and electrolysis of solution and/or pulp in electro-chemical reactor. Additionally colonies of bacteria strains during bacterial oxidation are introduced on sludges of sulfide minerals in sulfuric solution. |
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Processing method of plumb-microlitic concentrate Concentrate is grinned up to size not more than 0.32 mm and treated by mixture 18-35% of nitrogenous and 40-45% hydrofluoric acids at mass correlation of acids 2.7-5.3:1, "Т:Ж"=1:3-5 and temeprature 90-120°C with formation of pulp. Nitro-acid lead (II), undecomposed minerals and radionuclides are transferred into sedimentation, and tantalum and niobium - into fluoride - nitro-acid solution. Pulp is cooled up to 10-20°C, lead-bearing deposit is separated and washed it by 35-50% nitric acid at "Т:Ж"=1:1.5-2.0. Scrub solution is affixed to fluoride - nitro-acid solution. From washed sediment it is leached nitro-acid lead (II) by water at correlation "Т:Ж"=1:5-10 and temperature 50-60°C with separation of undecomposed minerals and radionuclides and receiving of nitro-acid lead solution. From combined fluoride - nitro-acid solution it is implemented extraction of tantalum and niobium by neutral extragent. In the capacity of neutral extragent it is used triiso-amyl-phosphate or octanol-1. Extraction is implemented on 2-6 stages at O:B=1.5-2.0:1 with transferring of tantalum into organic phase and niobium - into aqueous. Extraction by triiso-amyl-phosphate is implemented at 2-4 stages, and extraction by octanol-1 - at 5-6 stages. |
Another patent 2528327.