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Method of processing copper-containing materials Method of processing copper-containing materials includes dissolution of copper with sulphuric acid solution and crystallisation of copper sulphate pentahydrate from solution. After that, correction of mother liquor composition is realised for its re-use. Copper dissolution is realised in presence of copper nitrate, which serves as catalyst of dissolution process and which is prepared by dissolution of copper in nitric acid solution, concentration of which does not exceed 200 g/l. Correction of mother liquor composition is realised by control of copper nitrate concentration by qualitative reaction with diphenylamine in mother liquor. |
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Method of treating natural phosphate for extraction of rare-earth elements Invention relates to extraction of rare-earth elements from a natural phosphate. The method includes sulphuric acid decomposition of the phosphate into mineral fertilisers to obtain phosphogypsum. Further, the method includes separating the phosphogypsum and sulphuric acid decomposition of the phosphogypsum by successively treating multiple portions of phosphogypsum with one sulphuric acid solution. Further, the method includes separating the acid-resistant part of phosphogypsum, crystallising the rare-earth element concentrate, separating the rare-earth element concentrate by filtering to obtain a sulphuric acid filtrate and processing the rare-earth element concentrate. The separated acid-resistant part of phosphogypsum is treated with ammonia and carbon dioxide gas to obtain a calcium carbonate precipitate and ammonium sulphate solution. The calcium carbonate is separated, the ammonium sulphate is heat treated to obtain ammonium bisulphate and ammonia, which is returned for treating the acid-resistant part of phosphogypsum, and sulphuric acid solution, which is fed for decomposing the phosphate, is obtained from the obtained ammonium bisulphate and sulphuric acid filtrate with addition of sulphuric acid. |
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Method includes washing material with water to obtain a solid residue, preparing a sulphate solution, from which iron, copper and zinc are extracted by transferring iron into the residue in the form of iron hydroxide Fe(OH)3, precipitating copper from the filtrate with scrap iron, precipitating zinc from the filtrate with hydrogen sulphide. The filtrate, which contains Na2SO4 and acid, is then mixed with Ca(OH)2 to recycle the sodium sulphate and sulphuric acid in the form of a gypsum residue to obtain a filtrate with sodium hydroxide and accumulation of the filtrate for recycling NaOH. The solid residue obtained from the sulphide raw material is re-pulped and the pulp is treated with electric pulses with energy of 3.5-5.5 J, under the action of which pyrrhotine, chalcopyrite, sphalerite and sulphides decompose into iron, copper and zinc oxides and hydrogen sulphide. The liquid phase is then filtered out from the formed pulp and used as recycled water. The oxides of said metals are dissolved in sulphuric acid. The sulphate solution is filtered. Products containing iron, copper, zinc and gypsum are selectively extracted from the filtrate and gold is obtained from the residue, which contains quartz, sericite, gold and low-solubility minerals, by cyanation. Hydrogen sulphide is used to precipitate zinc. |
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Method of heap leaching of oxidised silicate nickel ore Method of heap leaching of silicate nickel ores includes ore crushing, ore mix preparation with fluoride adding from the group: sodium silicofluoride, fluoric calcium, ammonium fluoride and/or ammonium hydrofluoride amounting 1.3-1.7 wt % (in terms of fluorine). Then the mix is balled by pelletizing using the concentrated sulphuric acid in the ratio S : L = (88.0-94.0): (6.0-12.0) as a binder. After pelletizing the pellets are laid in a heap and leached by sulphuric acid solution. |
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Method of phosphite processing Method includes water processing, leaching of phosphate with sulphuric acid solution with concentration 3-6 wt % with conversion of REE, calcium and thorium into leaching solution and with obtaining gypsum product, extraction of REE, calcium and thorium from leaching solution by sorption with sulphoxide cationite. Leaching is carried out with sulphuric acid solution and L:S not less than 1.4:1. Sorption of REE, calcium and thorium is realised in stages. At first stage leaching solution is passed through cationite before beginning of REE breakthrough into forming primary depleted sulphuric acid solution. After that, desorption of calcium and thorium from saturated cationite with primary depleted sulphuric acid solution is carried out with obtaining primary calcium-thorium-containing strippant. At second stage remaining leaching solution is passed through cationite before REE breakthrough into secondary depleted sulphuric acid solution, which is applied for calcium and thorium desorption with obtaining secondary calcium-thorium-containing strippant. Then desorption of REE with solution of ammonium nitrate and precipitation of REE from strippant at pH 7.35-7.5 are realised. |
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Method of processing phosphogypsum Method of processing phosphogypsum includes preliminary aqueous treatment thereof; leaching the phosphogypsum by passing a sulphuric acid solution with concentration of 3-6 wt % through a layer thereof with displacement and separation of the aqueous solution and transfer of rare-earth elements and impurity components, including thorium, into the leaching solution; neutralising the washed phosphogypsum to obtain a gypsum product; the rare-earth elements and thorium are extracted from the leaching solution by sorption using a sulphoxide cationite and forming a sulphuric acid solution with rare-earth elements and thorium which is recycled; desorbing the rare-earth elements and thorium from the saturated cationite to obtain a strippant, wherein desorption of rare-earth elements is carried out by treating the cationite with an ammonium salt solution followed by precipitation of rare-earth elements from the strippant with an ammonium-containing precipitant and separating the rare-earth element precipitate. Leaching of the phosphogypsum with sulphuric acid solution is carried out with liquid to solid ratio of not less than 1.4:1. Desorption of rare-earth elements and thorium from the saturated cationite is carried out in the following sequence: first, thorium by treating the cationite with sulphuric acid solution with concentration of 3-6 wt % to obtain a thorium-containing strippant, then rare-earth elements to obtain a strippant containing rare-earth elements. |
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Method of processing beryllium fluorite-containing concentrates Concentrate with coarseness 45-85 mcm is subjected to sulphatisation with 93% sulphuric acid. First, low temperature sulphatisation is carried out at temperature 130-170°C with mixing with sulphuric acid, taken in quantity 50-70% of the required quantity. Then, high-temperature sulphatisation is carried out at 250-300°C with mixing with residual part of sulphuric acid. |
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Method of combined processing of beryllium concentrates Beryllium concentrate is activated by its milling until X-ray amorphous product with size of particles less than 5 mcm is obtained, and bertrandite-phenacite-fluorite concentrate is activated by addition in it of fluorine-containing compounds in quantity which ensures fluorine content 10÷25 wt %. Then, activated beryllium concentrate is processed with sulphuric acid and obtained acid pulp is added to activated bertrandite-phenacite-fluorite concentrate. After that, formed reaction mass is sulphatised for 1.5÷2 h at temperature 115÷125°C with continuous mechanical removal of reaction products from the surface of particles of activated concentrates with the following exposure at temperature 250÷300°C for not less than 1 h. |
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Method of processing of feed stock containing precious metals and sulphides Invention relates to hydraulic metallurgy of the precious metals in particular to processing of the aurum sulphide feed stock not containing organic carbon substance. Method of processing of the feed stock containing the precious metals and sulphides includes feed stock missing with water or sulphuric acid solution, and mixture treatment in autoclave with oxygen supply. Then sulphuric acid is separated from the oxidated pulp by washing with water. At that halogenide-ion is additionally injected to mixture supplied for autoclave treatment. Ions of chlorine, iodine and bromine are used as halogenide-ions, they are injected in form of the soluble salts or containing them natural minerals of carnallite (MgCl2·KCl·6H2O), or treated electrolytes of alkali and alkali-earth metals electrolysis. |
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Method of lithium concentrate processing Method includes concentrate sulphating by sulphuric acid, leaching of the sulphated concentrate, leached pulp separation to sulphate solution and insoluble cake. After cake washing its drying is performed. Then charge out of dry cake with lithium, potassium, sodium carbonates, magnesium, calcium, titanium, zinc oxides, trivalent chrome and cryolite is prepared. The charge is melted, produced melt is poured in the mould, cooled, moulding is removed and heat treated with ceramised glass creation. As the initial feed the lepidolite concentrate is used, it is sulphated at 95÷100°C for 4-6 minutes. During charge making out of dry cake with carbonates, oxides and cryolite the lithium carbonate and potassium carbonate consumption is 11.1÷11.3 wt % and 4.2÷4.3 wt %, respectively. |
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Method of working of lithium concentrates mixture Method includes preparation of charge out of lepidolite and spodumene concentrates, active charge preparation, making of sulphuric solution by sulphuric leaching with pulp separation by leaching to lithium sulphate solution and cake. At that prior to sulphuric leaching the activated charge is sulphated by sulphuric acid with consumption 1.2÷1.6 ml per 1 g of mixture for 4÷6 minutes, and sulphuric leaching is applied to sulphated charge for 40÷50 minutes. |
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Method of processing beryllium concentrate mixture Method includes activation of mixture, sulphatisation of activated mixture with sulphuric acid, leaching of sulphatised mixture, separation of leaching pulp into solution of beryllium sulphate and cake, sedimentation of beryllium hydroxide from solution. Activation of mixture is carried out by its milling until X-ray amorphous product with size of particles smaller than 5 mcm is obtained. Activated mixture is sulphatised for 45 min at temperature 100÷110°C with constant mechanical removal of reaction products from the surface of mixture particles by grinding and further carrying out exposure for not less than 2 h at temperature 280÷300°C. |
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Method of sulphuric acid decomposition of rem-containing phosphate raw material Invention relates to method of obtaining compounds of rare earth metals (REM) in complex processing of phosphate raw material, in particular apatites. Claimed is method of sulphuric acid decomposition of REM-containing phosphate raw material with concentration of REM in phosphogypsum. Method includes addition of sodium salt, potassium salt or their mixture. Sodium salts are added in amount 0.25-5.0 kg in terms of Na2O, potassium salts - in amount 0.25-5.0 kg in terms of K2O, and their mixture - in amount 0.25-5.0 kg in terms of Na2O and K2O per 1 kg of REM in terms of REM oxides in composition of phosphate raw material. |
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Method of processing brannerite-containing refractory uranium ores Milled to the coarseness of minus 0.3 mm ore is processed with a 1-40% solution of ammonium bifluoride with a ratio S:L=1:(1-5) and a temperature of 50-80°C for 1-4 hours. The obtained pulp is filtered, after which the filtrate is passed through waterless ammonia at a temperature of 25-60°C to pH 10. Then, the ammonium bifluoride solution is separated from the formed pulp, its concentration is increased to 1-40%, and supplied to re-processing of an initial raw material. The separated sediment is supplied to uranium extraction. Cakes, obtained as a result of the ore activation by solutions of ammonium bifluoride, are processed for 2-12 hours with a leaching solution with a ratio S:L=1:1, temperature of 60-80°C and residual acidity not lower than 20 g/l. The leaching solution represents a solution of sulphuric acid with the concentration 150-300 g/l, in which nitric acid is introduced as an oxidant, as well as NaCl as a complex forming reagent for gold. |
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Method of obtaining potassium alum In order to obtain potassium alum prepared is raw material represented by remnants of domanic formations, which contain aluminium, silicon dioxide, organic substance and include rare and rare earth elements. Leaching of acid-soluble components of raw material is carried out in autoclave with solution of sulphuric acid to its residual concentration 45-75 g/l. Obtained suspension is separated into liquid phase, which contains aluminium, potassium, sodium, rare metals, and solid phase, which contains silicon dioxide and organic substance. Potassium sulphate is added into hot liquid phase, obtained solution is cooled and crystallisation of potassium alum is carried out. Potassium sulphate is added from estimation of binding 80-90% of free aluminium sulphate in potassium alum with holding in solution of rare and rare-earth elements. |
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Method of extracting rare-earth metals (rem) from phosphogypsum Proposed method comprises REM sulphuric acid leaching from gypsum pulp with application of ultrasound oscillations, separation of said pulp to REM productive solution and cake, precipitation of REM collective concentrate from productive solution with production of water phase. Pulp is prepared on the basis of sulphuric acid solutions processed by electrochemical activation. Note here REM leaching is conducted under conditions of pulp circulation at combined effects of ultrasound oscillations at cavitation and magnetisation. Leaching pulp is divided into REM productive solution and first cake. REM are precipitated from productive solution as REM oxalates with production of REM collective concentrate. Water phase after precipitation of oxalates is divided into to parts. One part is re-restored by sulphuric acid and subjected to electrochemical activation for use in circulation while another part is neutralised to get the second cake to be flushed combined with first cake and directed for gypsum production. |
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Method of processing phosphogypsum Invention relates to the technology of processing phosphogypsum - wastes of enterprises, producing phosphoric fertilisers. The method includes opening phosphogypsum with sulphuric acid, further extraction of rare-earth elements (REE) and processing purified phosphogypsum with calcium oxide. In the course of opening with one solution of sulphuric acid successively processed are 1-3 lots of phosphogypsum with heating, the water phase is separated by filtration, the sediment is washed with water, apatite is added to the filtrate in a ratio of S:L=1:10-20, with the second heating at a temperature of 50-70°C and mixing for 1-2 hours with neutralisation with sulphuric acid to a concentration not lower than 0.1 mol/l. After that the sediment of the secondary phosphogypsum is separated by filtration and supplied to the beginning of the process. Calcium oxide or hydroxide and then ammonium hydroxide or carbonate are successively introduced into the filtrate until pH=2-3.5, the REE sediment is separated by filtration, and calcium hydroxide or oxide is introduced into the filtrate until pH=7-8, the sediment of feed tricalcium phosphate is separated by filtration, washed with water and discharged from the process. |
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Method of purifying phosphate-fluoride concentrate of ree Invention relates to purification of a phosphate-fluoride concentrate of rare earth elements (REE), obtained in the complex apatite processing. A method of purification of the phosphate-fluoride concentrate of REE, which contains admixtures of calcium and thorium, includes processing of the concentrate with a solution of sulphuric acid with a concentration of 4-6 wt % in the presence of sulphoxide cationite. REE, admixtures of thorium and calcium are absorbed by sulphoxide cationite, transfer of fluorine together with phosphorus into the sulphuric acid solution, separation of the sulphuric acid solution from sulphoxide cationite, desorption from cationite of REE and admixtures of thorium and calcium with an ammonium salt solution with obtaining desorbate and its neutralisation with an ammonium compound in three stages. At the first stage neutralisation is carried out until pH 4.2-5.0 is achieved with formation and separation of a thorium-containing residue, at the second stage - until pH 7.0-7.5 is achieved with formation and separation of a concentrate of REE, and at the third stage - until pH not less than 8.5 is achieved with formation and separation of a calcium-containing residue. |
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Method of processing phosphogypsum for production of concentrate of rare earth metals and gypsum Method includes preparation of phosphogypsum pulp, leaching rare earth metals (REM) and phosphorus with sulphuric acid. After that, pulp is separated into REM and a phosphorus-containing solution and gypsum in the form of insoluble sediment, its neutralisation and REM sorption with cationite from the solution with obtaining mother liquor. After that, REM desorption is performed with obtaining a strippant and separation of REM concentrate from the strippant. Stage leaching is performed in the method, with supply of phosphogypsum to each stage and sulphuric acid to the first stage. Before neutralisation gypsum is subjected to water washing with obtaining a washing solution, supplied to REM sorption with cationite. Sorption mother liquor is divided into two parts, one of which is used to prepare phosphogypsum pulp, with precipitation of phosphorus and fluorine with basic calcium compound from the second one. The obtained sediment is separated from water phase and supplied to utilisation, and water phase is used in circulation. |
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Proposed method comprises mixing of initial titanium-bearing slag with soda ash, mix sintering and leaching of produced cake, first, in water for production of iron-titanium-bearing precipitate and, then, in muriate for production of titanium-bearing precipitate. Then, pulp is filtered to separate the precipitate to get titanium dioxide concentrate. note here that initial slag sintering with soda ash occurs at 900°C and Na2CO3:slag ratio equal to (0.98-1.15):1. Cake sintering in water is conducted to transfer sodium silicate to solution while titanium dioxide concentrate is made by calcination of precipitate resulted from hydrochloride-acid treatment. Note here that said initial titanium slag represents that of reducing fusion of ilmenite. |
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Phosphosemihydrate processing method Invention refers to processing of freshly obtained phosphosemihydrate and can be used to obtain concentrate of rare-earth elements (REE) and gypsum product for construction materials. Phosphosemihydrate is processed with water solution containing fluorine-ion. Sulphuric acid leaching is performed with displacement and separation of the water solution containing fluorine-ion, as well as with conversion of REE and impurity components to a leaching solution and production of a phosphosemihydrate layer saturated with a sulphuric-acid solution. Then, water displacement of the remaining amount of sulphuric acid solution is performed so that washed phosphosemihydrate and a leaching solution is obtained Phosphosemihydrate is neutralised with a calcium-containing reagent so that a gypsum product is obtained. Rare-earth elements and impurity components are extracted from the leaching solution by sorption using sulphoxy cationite so that a lean sulphuric-acid solution is formed; REE and impurity components are desorbed from saturated cationite by its processing with an ammonium sulphate solution so that a strippant is obtained; REE and impurity components are deposited from the strippant with an ammonium-containing precipitator in two stages and REE deposit is separated. |
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Method of processing phosphogypsum Method of processing phosphogypsum involves step-by-step agitation sulphuric-acid leaching of rare-earth metals and phosphorus while feeding sulphuric acid to the head step, using the obtained leaching solution of the head step at subsequent leaching steps, separating the undissolved residue from pulp of a tail step and washing with water, treating the leaching solution of the tail step to obtain a mother solution, using the mother solution and the washing solution for leaching. Leaching of the rare-earth metals and phosphorus at the second and subsequent steps is carried out from a mixture of phosphogypsum and the leached pulp from the previous step. Sulphuric acid is fed to the head leaching step in an amount which enables to extract rare-earth metals and phosphorus into the solution at the head step and subsequent steps at pH values at the tail leaching step not higher than pH at the onset of precipitation of rare-earth metal phosphates. The tail step for leaching rare-earth metals and phosphorus is carried out while simultaneously treating the leaching solution by extracting rare-earth metals by sorption with a cationite. The rare-earth metal-saturated cationite is separated from the mother pulp and taken for producing a rare-earth metal concentrate. A portion of the mother solution is pre-purified from phosphorus by precipitation thereof with a basic calcium compound. The obtained phosphorus-containing precipitate is fed for recycling. |
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Processing method of black-shale ores Processing method of black-shale ores includes crushing, counterflow two-stage leaching by sulfuric acid solution upon heating, separation of pulps formed after leaching at both stages by filtration. Then valuable soluble materials are washed from deposit at the second stage with strengthened and washing solutions being produced, marketable filtrate is clarified at the first stage for its further processing. Ore is crushed till the size of 0.2 mm, leaching at the first stage is performed by cycling acid solution with vanadium under atmospheric pressure, temperature of 65-95°C during 2-3 hours, till residual content of free sulphuric acid is equal to 5-15 g/l. Leaching at the second stage is performed at sulphuric acid rate of 9-12% from the quantity of initial hard material under pressure of 10-15 atm and temperature of 140-160°C during 2-3 hours. Cake filtered after the first stage is unpulped by part of strengthened solution which content is specified within 35-45% of total quantity. |
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Processing method of black-shale ores with rare metals extracting 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. |
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Method of extracting rare-earth metals (rem) from phosphogypsum Proposed 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. |
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Method of extracting rare-earth metals from phosphogypsum Method 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. |
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Method of producing titanium dioxide Invention can be used in chemical industry. In order to obtain titanium dioxide, a mixture is prepared, leaching is carried out in sulphuric acid solution to form titanyl sulphate TiOSO4 and iron sulphates FeSO4 and Fe2(SO4)3, followed by precipitation of iron sulphate FeSO4 and hydrolysis of titanyl sulphate TiOSO4 to obtain hydrated titanium dioxide TiO(OH)2 and firing. The mixture is formed by adding potassium hydrogen sulphate KHSO4. Before leaching, the obtained mixture is melted at temperature of 300-400°C to obtain a melt containing potassium titanate K2TiO3. The melt is then leached using sulphuric acid solution with concentration of 5-10%. |
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Method for extraction of uranium from ores Method 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. |
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Nickel matte processing method Method involves leaching of matte using a nickel solution at increased pressure and temperature so that nickel sulphate solution is obtained. Nickel sulphate solution is cleaned from impurities of iron, copper and cobalt, and nickel is extracted by means of electrowinning from clean solution of nickel sulphate so that cathode nickel and sulphuric anolyte containing nickel sulphate and sulphuric acid is obtained. Sulphuric anolyte is subject to extraction treatment using a mixture of tertiary amines and aliphatic alcohols as an extracting agent with extraction to the extract of sulphuric acid and obtainment of sulphuric raffinate and extract. Sulphuric acid is re-extracted with water from the extract, and sulphuric raffinate is supplied to leaching of the matte. |
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Method of producing scandium-bearing concentrate from red mud 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. |
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Method of extracting manganese from manganese-bearing raw stock Method of processing manganese oxide materials containing heterovalent manganese oxides comprises leaching crushed raw stock by sulfuric acid aqueous solution in the presence of bivalent iron sulphate, iron precipitation and manganese extraction from productional solution. Note here that said leaching is performed on adding reducing agent in the form of metal iron or iron sulphate (Fe2+) at 60-95°C for 60-300 min. Leaching is carried out at initial concentration of H2SO4 in leaching solution of up to 100 g/dm3 and final acidity in productional solution relative to hydrogen ion exponent pH<2. |
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Method to process sludge of neutralisation of acid mine waters Invention relates to the field of hydrometallurgy of heavy non-ferrous metals. The method for processing of sludges of neutralisation of acid mine waters includes its preliminary grinding, afterwards sulfuric leaching is carried out during mixing by means of treatment of the sludge with acid mine waters and sulfuric acid and addition of iminodiacetatic ampholyte for simultaneous sorption of copper and zinc. Ampholyte is separated from the produced pulp, and its desorption is carried out with sulfuric acid with formation of desorbed iminodiacetic ampholyte and sulfate solution. The desorbed ampholyte is returned to the stage of leaching and simultaneous sorption. From the sulfate solution by means of electrolysis copper is serially extracted, and then - zinc. The treated sulfate solution is returned to the stage of desorption. The produced pulp after separation of ampholyte from it is neutralised with lime, afterwards it is separated with solid residue and liquid part. The remaining solid residue is dried and ground with production of a gypsum-containing end product. |
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Method of uranium ore processing Proposed 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. |
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Method of metal extraction from silicate nickel ores Proposed method comprises ore pretreatment by crushing, classification and grading, biological degradation of ore silicate minerals by multiple ore interaction with silicate bacteria cultural medium without mixing with replacement of said cultural medium at pH, at least, 0.4. Then, metals are leached from biological degradation cakes by cultural solutions after extraction of silicon therefrom and additions of sulfuric acid to concentration of 50-450 g/l. After leaching, metals are extracted form cake leaching solution. Note here that cultural medium is replaced on reaching redox potential in solution of minus 250 mV. After biological degradation and before leaching, cakes are flushed with water. |
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Complex processing method of carbon-silicic black-shale ores Invention 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. |
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Method of extracting rare-earth elements from phosphogypsum Invention relates to the technology of producing compounds of rare-earth elements during complex processing of apatites, particularly extraction of rare-earth elements from phosphogypsum. The method involves preparation of pulp from phosphogypsum and sorption of rare-earth elements on a sorbent. The pulp is prepared from ground phosphogypsum and sulphuric acid solution with pH=0.5-2.5 until achieving liquid:solid ratio of 4-7. Sorption is carried out directly from the phosphogypsum pulp on a sorbent with sulphuric acid functional groups for 5-7 hours with solid:sorbent ratio of 4-7. |
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Method to extract metals from sulphide mineral raw materials Method includes leaching of ground raw materials in a solution of sulphuric acid with concentration of more than 2.0 g/l, containing ions of trivalent iron of more than 10-12 g/l, while mixing, at the temperature up to 100°C, solid phase content to 60%, at least in two serially connected reservoirs. The pulp discharged from the last reservoir is separated into solid and liquid phases. At the same time the solid phase is returned for leaching into the first reservoir. Iron oxidation in the liquid phase is carried out with iron-oxidising bacteria adsorbed on a neutral carrier at the pH 1.4-2.2 and 90°C with aeration by gas containing oxygen and carbonic acid. Then the liquid phase is returned after iron oxidation into leaching reservoirs, and metals are extracted from the produced phases. Besides, leaching is carried out with aeration by oxygen-containing gas. The pulp discharged from each reservoir is separated into solid and liquid phases. The solid phase is sent for leaching to the next reservoir, and the liquid phase is prepared prior to oxidation with bacteria. Duration of leaching is increased in each subsequent reservoir. |
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Method to process metal-containing sulphide mineral raw materials with extraction of metals Invention relates to the method for extraction of metals from metal-containing sulphide mineral raw materials. The method includes leaching with mixing of a sulphuric acid solution in presence of trivalent iron ions in at least two serially connected tubs, separation of leaching products into liquid and solid phases, iron oxidation in a liquid phase, return of the liquid phase after iron oxidation into leaching reservoirs, intermediate extraction of metals from liquid phases. At the same time the initial raw materials prior to leaching are exposed to preliminary acid treatment at PH=0.8-1.4, S:L=1:1. Leaching is carried out in two stages at the temperature of 75-95°, PH=1.0-1.2 and S:L=1:(3-6), concentration of trivalent iron ions of 30-45 g/l at each stage with separation of leaching products after each stage into liquid and solid phases and iron oxidation in a liquid phase after each stage and with extraction of metals at each stage from liquid phases after iron oxidation. To the first stage raw materials exposed to preliminary acid treatment are sent, and to the second stage - a solid phase is sent, produced after separation of leaching products at the first stage. The liquid phase, which was produced after oxidation of iron at the second stage, is returned to the last leaching reservoir of the same stage. |
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Method of processing titanium-containing concentrate Invention can be used in chemical industry. The method of processing titanium-containing concentrate involves decomposition of the titanium-containing concentrate with sulphuric acid solution while heating with transfer of titanium into the sulphuric acid solution and separating the solid residue. Ammonium sulphate is added to the titanium-containing sulphuric acid solution in an amount which ensures concentration thereof in the solution of 300-450 g/l, with crystallisation of the ammonium-titanium-containing solid phase which is separated and dissolved in water to obtain sulphuric acid solution with pH=1-2. A silicon-sodium reagent is added to the obtained solution and sodium hydroxide is also added to ensure molar ratio TiO2:SiO2:Na2O=1:(0.75-5.5):(0.5-5) in the suspension. The silicon-sodium reagent used is crystalline sodium silicate or liquid sodium glass. The suspension is held in sealed conditions at temperature 150-250°C for 20-40 hours to form a titanium-silicon sodium-containing residue which is separated, washed with water and dried at 70-150°C to obtain the end product. |
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Method of phosphogypsum processing for manufacture of concentrate of rare-earth elements and gypsum Method of phosphogypsum processing involves leaching of phosphogypsum with sulphuric acid solution with change-over of phosphorus and rare-earth elements to the solution, and gypsum residues is obtained, rare-earth elements are extracted from the solution and the gypsum residue is neutralised with the main calcium compound. In addition, leaching is performed with sulphuric acid solution with concentration of 1-5 wt %. After that, rare-earth elements are extracted from the solution by sorption using sulfocationite in hydrogen or ammonia form with further desorption of rare-earth elements with ammonia sulphate solution. After desorption to the obtained strippant there added is ammonia or ammonium carbonate with deposition and separation of hydroxide or carbon-bearing concentrate of rare-earth elements. Extraction of rare-earth elements of medium and yttrium groups to concentrates is 41-67% and 28-51.4% respectively. Specific consumption of neutralising calcium compound per 1 kg of phosphogypsum has been reduced at least by 1.6 times. |
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Method of producing vanadium oxide Invention relates to vanadium oxide extraction. The method of producing vanadium oxide involves preparation of starting vanadium-containing material for burning, burning with lime treatment, leaching with sulphuric acid, separating the solid substance and liquid, precipitation of ammonium polyvanadate with an ammonium salt and removing ammonia by calcination or reduction to obtain vanadium oxide. The solid starting material used at said steps has total amount of alkali metal of not more than 0.3 wt % and total amount of Cl- and NO3 - ions of not more than 0.1 wt %, and the liquid starting material has total amount of alkali metal of not more than 0.1 g/l and total amount of of Cl- and NO3 - ions of not more than 0.1 g/l. |
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Method involves stages of (a) material interaction with acid leaching solution in presence at least of one iron compound and acidophilic microorganisms at least capable of oxidating ferrous iron, and (b) leaching. Leaching stage (b) is performed at control of molar ratio of dissolved ferric iron to dissolved molybdenum and it is assumed equal at least to 6:1, preferably at least to 7:1 and after leaching there performed is stage of (c) molybdenum extraction at least from one solid and liquid residue of the leaching process. Finally, molybdenum is extracted from leaching residue of leaching process. Final degree of Mo extraction from sulphide material containing molybdenum is 89%. |
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Processing method of sulphide-oxidated copper ores with copper and silver extraction Processing method of sulphide-oxidated copper ores with copper and silver extraction involves collective flotation of sulphide and oxidated copper minerals from crushed ore with extraction of collective concentrate. Then, leaching of collective concentrate is performed at mixing with water solution of sulphuric acid with participation of ozone, hydrogen peroxide and ions of ferric iron. Then, dehydration and washing of cake of concentrate leaching and copper extraction from copper-bearing solutions is performed. Copper and silver is extracted from cakes by flotation without using any foaming agent and using reagent of isobutyl dithiophosphate at pH 6-8 or reagent DSP017 containing isobutyl dithiophosphate and thionic carbamate. |
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Procedure for processing vanadium containing raw material Procedure consists in preparation of water pulp from raw material, in introduction of sulphuric acid and anionite into it for leaching and in extraction of vanadium from pulp by sorption. Upon sorption saturated anionite is withdrawn and washed; vanadium is de-sorbed from anionite and regenerated anionite is introduced to the stage of leaching and sorption. Also, water vanadium pulp is prepared from vanadium containing raw material. As such there is used oxidised slag or slime at pH 11.5-7.5. Sulphuric acid is introduced into prepared water pulp at S:L=1:2 to pH 4.5-4.0. Vanadium is extracted from pulp by counter-flow sorption at pH 4.5-1.8 with following saturated ionite washing at drainage. Value of pH in pulp is maintained at 4.5-4.0 at withdrawal of saturated anionite, while at introduction of regenerated anionite - 2.0-1.8. |
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Procedure for extraction of metals from silicate nickel ore Procedure for extraction of metals from silicate nickel ore consists in preparation of silicate nickel ore by crushing and classification, in silicon leaching from ore with cultural medium of silicate bacteria and in successive extraction nickel from cake. Silicate minerals of ore are bio-degraded at leaching silicon with cultural medium of silicate bacteria; bio-degradation is performed at pH as high, as 4, without mixing and with replacement of cultural medium. Nickel is extracted from cake of bio-degradation by leaching with utilisation of solution of bio-degradation upon silicon has been extracted from it and by adding sulphuric acid to concentration 50÷450 g/l. Further, metal is extracted from leaching solution of cake bio-degradation. |
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Procedure for processing grinding wastes from manufacture of permanent magnets Invention refers to procedure of extraction of rare earth elements (REE) from grinding wastes of manufacture of permanent magnets. The procedure consists in dissolving grinding wastes in sulphuric acid, in extraction of double salts and sodium REE-Na by sedimentation and in washing double salts. Further, double salts are conversed in hydroxide of REE and hydroxides of REE are washed. Upon washing hydroxides of REE are oxalate conversed, dried and burned producing sum of REE oxides. After sedimentation of double salts of rare earth elements and sodium REE-Na mother solutions are processed producing ferriferous and cobalt cakes. |
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Procedure for extraction of nickel from silicate ore by heap or underground leaching Procedure consists in leaching with solution of sulphuric acid and in receiving productive solution. Productive solutions are received by heap or underground leaching with solution of sulphuric acid. During process periods of ore leaching and concentration are alternated, while concentration of sulphuric acid in leaching solution is maintained under a differential mode with change of higher concentration of 50-250 g/l in a period of ore leaching to a lowered one - 1-10 g/l in periods of mode of ore buddling. Further, impurities of iron, aluminium, magnesium and silicon are removed from productive solutions; nickel is extracted into concentrate and re-circulated solutions are returned to leaching after strengthening with sulphuric acids. |
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Extraction method of nickel from oxidised nickel ores Method involves heap leaching of nickel with sulphuric acid solution, neutralisation of production solution, sorption of nickel on ionite from it, processing of strippant so that nickel is obtained, supply of raffinate solution for heap leaching of ore during its revolution and circulation. At that, some part of raffinate is subject to neutralisation with lime milk and its cleaning from impurities of iron, manganese and magnesium by directing in the form of neutralised pulp to waste heap for deposition of iron, manganese and magnesium on it with further mixing with the main volume of cycling solution of raffinate and final strengthening with sulphuric acid. Volume of some part of raffinate, which is subject to neutralisation and cleaning, as per the amount of iron, manganese and magnesium, corresponds to their transition to production solution during leaching of nickel from the ore during circulation of cycling solution. |
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Method of enriching anatase mechanical concentrates involves burning (1) an anatase concentrate in a fluidised bed furnace or a drum furnace; reducing (2) the burnt product in the fluidised bed furnace or drum furnace using a hydrogen or natural gas as a reducing agent; dry and wet separation (3) of the reduced product in a weak magnetic field in magnetic separators fitted with a permanent magnet and a drum, where the magnetic fraction formed during reduction is discarded; dry separation (4) in a strong, high-gradient magnetic field of the magnetic fraction obtained during separation in a weak magnetic field in roller or drum separators with a rare-earth permanent magnet, with extraction of silicates, secondary phosphates, monazite, calzirtite, zircolinite and uranium and thorium-containing minerals; leaching (5) of the magnetic fraction obtained from separation in strong magnetic field in mixing tanks or fluidised bed columns, with a hydrochloric acid solution; filtering the leached product on a belt filter; drying of the filtered product in a rotary drier or fluidised bed drier; oxidation (6) of the dried product in the drum furnace or fluidised bed furnace; fast cooling of the oxidised product in water or compressed air in a drum cooling device or by immersing in water; leaching (7) the fast-cooled product in mixing tanks or columns, or with hydrochloric acid or sulphuric acid; filtering the product from the second leaching (7) on a belt filter; and drying of the filtered product in a rotary or fluidised bed drier; and final dry separation (8) of the product of the second leaching in a strong, high-gradient magnetic field in roller or drum separators with a rare-earth permanent magnet, while discarding the magnetic fraction and extracting the non-magnetic fraction as the final product (P), i.e. the synthetic rutile. |
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Procedure for extaction of metals from gold containing sulphide-oxidised copper ores Procedure consists in crushing and crumbling gold containing ore to size 60 % of class minus 0.063 mm, in gravitation concentration in centrifugal concentrators with successive refinement of crude gold containing concentrate on tables concentrators to production of gold containing concentrate. Collective copper concentrate is extracted from gravitation rejects by collective flotation of sulphide and oxidised minerals of copper. Collective copper concentrate is subjected to leaching at mixing with water solution of sulphuric acid at concentration not less 2 g/l with ozone at concentration in ozone-oxygen gas mixture over 85 g/l, hydrogen peroxide and ions of oxidised iron with concentration not less 2 g/l. Further, solid phase of concentrate leaching is dehydrated and washed, copper is extracted from copper containing solutions and copper and silver are extracted by flotation from solid phase of concentrate leaching. |
Another patent 2551173.