Method of producing scandium oxide

FIELD: metallurgy.

SUBSTANCE: 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 50-70C and held for one hour at mixing. Produced precipitate is flushed by ethanol at volume ratio of 1-10-11.

EFFECT: simplified process, higher purity scandium oxide.

3 ex

 

The invention relates to ferrous metallurgy, namely the recovery of scandium oxide from poor scandium concentrate obtained after processing by the carbonate leaching of red mud waste production in the processing of bauxite to alumina.

A known method of producing scandium oxide (patent RU 2069181, IPC C01F 17/00, 1996), including the dissolution of scandium-containing concentrate in mineral acid to bring the acid concentration in the solution to 260-400 g/DM3the sludge separation of scandium sulfate from the solution, washing and dissolution in water and precipitation from a solution of soluble compounds of scandium by processing, for example, oxalic acid, washing, drying and calcination to obtain commercial scandium oxide (99%).

The disadvantage of this method is a significant percentage of the total losses of scandium (up 11%), which are caused in particular by the fact that when washing the precipitate of sulphate of scandium sulfuric acid concentration used for deposition of scandium-containing concentrate, the solubility of scandium in it quite high, which leads to its effects.

Closest to the proposed technical solution is the method of producing oxide of scandium (patent RU 2257348, IPC C01F 17/00, 2005)(prototype), the essence of which is expressed by the following set of essential to recognize the s: the dissolution of the scandium-containing solution in mineral (salt, sulfuric, nitric) acid; clear scandium solution from impurities by processing solutions solfataras inorganic compound and then with barium chloride; processing the purified scandium solution of alkaline reagents, in particular NH4OH, obtaining low-solubility compounds of scandium: oxyhydrate or hydrooxicarbonate scandium; filtering the slurry to separate scandium precipitate from solution; sludge treatment with formic acid; separating the precipitate formate scandium from the mother liquor; washing the precipitate with formic acid, drying and calcining the precipitate to obtain commercial scandium oxide 99.99%purity.

The disadvantages of this method include its multi-stage, in particular to remove impurities at the first stage in scandium solution is injected sulfadimidine inorganic compounds and barium chloride, and then spend additional sludge treatment oxyhydrate scandium formic acid.

Thus, the authors goal was to develop a simple method to produce scandium oxide of high purity with a minimum total loss of scandium.

The problem is solved in the proposed method of producing oxide of scandium, including the dissolution of scandium-containing concentrate sulfuric acid, removing the acid-insoluble precipitate, the transfer of scandium in about the of Adak in the presence of ammonium compounds, filtering, washing, drying and calcination of the resulting sludge, characterized in that after removal of the acid-insoluble precipitate the sulfuric acid concentration in the filtrate is brought to 540-600 g/DM3as ammonium compounds using ammonium chloride, is introduced into the solution in the amount 26,7-53,5 g/DM3at a temperature of 50-70C, followed by exposure for 1 to 2 hours under stirring, and washing the obtained precipitate is carried out with ethyl alcohol based volume ratio 1 - 1011.

At the present time of patent and technical literature is not a method of obtaining oxide of scandium by dissolving scandium-containing concentrate sulfuric acid in the presence of ammonium chloride, is introduced into the solution in a certain amount and under certain conditions, obtained by washing the precipitate with ethyl alcohol in a certain volume ratios.

Experimentally, the authors found that when the pH of the solution 540-600 g/DM3H2SO4achieved saturation solubility of scandium and the introduction of a reagent containing both ions Cl-andcreate the conditions causing almost complete precipitation of scandium with simultaneous separation from many metals impurities. At a low content of scandium is in the original concentrate (1-2,5 g/DM 3) when transforming a scandium in the sediment reached a low residual content of scandium in solution, which reduces the loss of scandium and increases the yield of the final product. This eliminates the need for the preliminary stages of concentration and getting richer concentrate on scandium (for example, by the method prototype content of scandium in the original concentrate is 38 wt.%). It was established experimentally that significant values are terms of the introduction of chloride of ammonium sulfate in the solution obtained by dissolution of the scandium-containing concentrate in sulfuric acid and then bringing the pH of the solution to 540-600 g/DM3. So, with the introduction of ammonium chloride less than 26.7 g/DM3an increase in losses of scandium with a solution, with the introduction of more of 53.5 g/DM3not observed a noticeable increase in recovery of scandium in the sediment. Lowering the temperature below 50C leads to incomplete precipitation of the compounds of scandium in the sediment; the increase in temperature above 70C is impractical as it does not affect the output of scandium. Reducing the exposure time of the solution with the reagent less than 1 hour also does not provide sufficient recovery of scandium from the solution. The increase in time more than 2 hours is also impractical, as it does not affect the output of scandium. Significant t what is the use for washing the obtained precipitate of ethyl alcohol. Thus, the magnification ratio is more than 1:11 unreasonably increases the flow rate of the reagent, does not improve the results in the output of scandium; when the reduction ratio less than 1:10 insufficient sediment washed from impurities and salts of scandium contaminated by impurities. Thus, the whole set of essential features of the proposed solution allows their use to obtain the final product - the scandium oxide of high purity and with minimal losses.

The proposed method can be implemented as follows. As the original take poor scandium concentrate obtained after processing by the carbonate leaching of red mud waste production in the processing of bauxite to alumina. Scandium-containing concentrate is dissolved in sulfuric acid with a concentration of 150-300 g/DM3for 4 hours at a temperature of 955C. After removal of the acid-insoluble precipitate adjusting the sulfuric acid concentration in the filtrate to 540-600 g/DM3and at a temperature of 50-70C. is introduced into a solution of ammonium chloride NH4Cl in the number 26,7-53,5 g/DM3followed by exposure with stirring for 1-2 hours. Then incubated for another 24 hours without stirring and filtered. According to the chemical analysis of the obtained salt of scandium containing scandium to 20 wt.%. The precipitate is dried at a temperature of 11-120C and calcined at a temperature of 800-850C for at least 2 hours. According to x-ray phase and chemical analyses are trademark oxide of scandium Sc2O3purity 99,0%. The extraction of scandium is 97-98%. Loss of scandium does not exceed 2-4%.

The method allows for the separation of scandium from impurities in the solutions at the following ratios of concentrations as follows: Sc: (1-6) Ti: (0,5-2,5) Fe: 0,1-3,5) Zn: (5-13,5) Zr: (0,5-0,1) Ca: (0.1 to 0.5) Al: (0.01 to 0.3) Mg: (0,002-0,1) Th: (is 0.0002-0,008) U.

The proposed method is illustrated by the following examples.

Example 1. Take 50 g of scandium-containing concentrate composition, wt.%: Sc - 1,6; Ti - 5,3; Fe - 3,7; Zn and 0.3, Zr - 10,7; Na - 17,5; Ca - 4,2; Si - 1,4; Th - 3,5; U - 0,13; dissolved in 500 DM3sulfuric acid concentration of 300 g/DM3remove acid-insoluble precipitate and adjusting the sulfuric acid concentration in the filtrate to 540 g/DM3. The solution is injected 26.7 g NH4Cl (53,5 g/DM3) at 50C and maintained under stirring for 1 hour. Then incubated for 24 hours without stirring. The obtained precipitate was separated by filtration on a filter SCHOTT and washed with ethyl alcohol in a quantity of 50 ml (volume ratio of sediment: ethanol = 1:10). Then the precipitate is dried at 120C for 10 hours to obtain a constant weight and calcined at 850C for 2 hours. Get the scandium oxide purity 99,0% with a yield of 97.3%. Loss of scandium - 2,69%.

Example 2. Take 50 g of scandium-containing conc is the composition, wt.%: Sc - 1,6; Ti - 5,3; Fe - 3,7; Zn and 0.3, Zr - 10,7; Na - 17,5; Ca - 4,2; Si - 1,4; Th - 3,5; U - 0,13; dissolved in 500 DM3sulfuric acid concentration of 300 g/DM3remove acid-insoluble precipitate and adjusting the sulfuric acid concentration in the filtrate to 600 g/DM3. In the enter solution of 13.4 g of NH4Cl (26.7 g/ DM3) at 70C and maintained for 2 hours under stirring. Then incubated for 24 hours without stirring. The obtained precipitate was separated by filtration on a filter SCHOTT and washed with ethyl alcohol 55 ml (volume ratio of sediment: ethanol = 1:11). Then the precipitate is dried at 120C for 12 hours and calcined at 800C for 2 hours. Get the scandium oxide purity 99,0% with a yield of 97.8%. Loss of scandium amount of 2.16%.

Example 3. Take 50 g of scandium-containing concentrate composition, wt.%: Sc - 1,96; Ti - 2,85; Fe - 1,42; Zn - of 6.71; Zr - 18,13; Na - 17,7; Ca - 0,4; Si - 1,1; Th - 0,17; U - 0,015; dissolved in 500 DM3sulfuric acid concentration of 300 g/DM3remove acid-insoluble precipitate and adjusting the sulfuric acid concentration in the filtrate to 600 g/DM3. The solution is injected 26.7 g NH4Cl (53,5 g/DM3) at 50C and maintained for 2 hours under stirring. Then incubated for 24 hours without stirring. The obtained precipitate was separated by filtration on a filter SCHOTT and washed with ethyl alcohol is m 55 ml (volume ratio of sediment: ethanol = 1:10). Then the precipitate is dried at 120C for 12 hours and calcined at 800C for 2 hours. Get the scandium oxide purity 99,0% with a yield of 97.8%. Loss of scandium amount to 2.2%.

Thus, the authors propose a technically simple method of obtaining a trademark of scandium oxide of high purity 99,0% with the yield up to 97-98% of the poor scandium concentrate obtained after processing by the carbonate leaching of red mud waste production in the processing of bauxite to alumina.

The method of producing oxide of scandium, including the dissolution of scandium-containing concentrate sulfuric acid, removing the acid-insoluble precipitate, the translation of scandium in the sediment in the presence of ammonium compounds, filtering, washing, drying and calcination to obtain a precipitate of oxide of scandium, characterized in that after removal of the acid-insoluble precipitate the sulfuric acid concentration in the filtrate is brought to 540-600 g/DM3as ammonium compounds using ammonium chloride, is introduced into the solution in the amount 26,7-53,5 g/DM3at a temperature of 50-70C, followed by exposure for 1-2 h under stirring, and washing the obtained precipitate is carried out with ethyl alcohol in a volume ratio of 1 - 1011.



 

Same patents:

FIELD: chemistry.

SUBSTANCE: method of extracting yttrium (III) from salt solutions involves floatation extraction using an organic phase and a collector. The organic phase used is isooctyl alcohol. The collector used is an anionic surfactant - sodiium dodecyl suphate in a concentration which corresponds to the stoichiometry: Y+3+SDS-=Y[DS]3, where Y+3 is a yttrium cation, DS- is a dodecyl sulphate ion. Floatation extraction is carried out at pH=7.0-7.8 and ratio of the organic phase to the aqueous phase of 1/20-1/40.

EFFECT: high degree of extraction of yttrium.

2 dwg, 1 ex

FIELD: metallurgy.

SUBSTANCE: 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 130-150C 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.

EFFECT: increasing extraction degree of vanadium, uranium, molybdenum; improving the complexity of ore use owing to associated extraction of rare-earth elements.

18 cl, 1 dwg

FIELD: chemistry.

SUBSTANCE: 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.

EFFECT: high efficiency of the method owing to higher extraction of rare-earth elements without a filtration step.

6 tbl, 6 ex

FIELD: metallurgy.

SUBSTANCE: method involves selective extraction of salts in volumes of nanopores of nanoporous conducting materials due to effect of electrostatic interaction of dipole moments of solvated ionic complexes of transition, rare-earth and actinoid elements with electric field of double electric layer of "nanopore wall - solution" boundary line. The method is implemented by subsequent filling of nanopore of nanoporous conducting material with the solution containing ionic complexes of transition, and/or rare-earth and/or actinoid elements, displacement from nanopore of ionic complexes of transition, rare-earth and actinoid elements weakly localised in nanopores by means of pressure of gases or liquids, by filling of nanopore with solution of inorganic acid of high concentration, and by extracting from nanopores of residual ionic complexes of transition, rare-earth and actinoid elements by means of pressure of gases or liquids. The above method can be implemented in an electrochemical cell.

EFFECT: obtaining cheap and competitive compounds of the above elements of high technical purity.

45 cl, 18 dwg, 4 ex

FIELD: metallurgy.

SUBSTANCE: invention can be used in the technology of obtaining the compounds of rare-earth metals at complex processing of apatites, and namely for obtaining of concentrate of rare-earth metals (REM) from phosphogypsum. Method involves sorption of rare-earth metals. At that, prior to sorption, phosphogypsum is crushed in water so that pulp is obtained in the ratio Solid : Liquid=1:(5-10). Sorption is performed by introducing to the obtained pulp of sorbent containing sulphate and phosphate functional groups, at the ratio of Solid : Sorbent=1:(5-10) and mixing during 3-6 h.

EFFECT: increasing REM extraction degree to finished product.

5 tbl, 5 ex

FIELD: metallurgy.

SUBSTANCE: method involves oxidation of micro production wastes at temperature of 550-650C 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 20-80C. 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.

EFFECT: regeneration of rare-earth metals from production wastes of magnets and obtaining raw material containing rare-earth metals for reutilisation in production of rare-earth constant magnets.

2 cl, 2 tbl, 7 ex

FIELD: chemistry.

SUBSTANCE: invention relates to methods of extracting a concentrate of rare-earth elements from wet-process phosphoric acid, which is obtained in a dihydrate process of processing an apatite concentrate, and can be used in chemical and related industries. The method involves sorption of rare-earth elements and thorium contained in wet-process phosphoric acid at temperature 20-85C, wherein the sorbent used is a sulphoxide cationite, washing the saturated sorbent with water, desorption of rare-earth elements and thorium with concentrated ammonium sulphate solution to form a desorbate, and treating the desorbate with an ammonia-containing precipitant in form of ammonium carbonate or ammonia gas, which is fed in two steps, wherein at the first step the precipitant is fed until achieving pH 4.5-5.0 with precipitation and separation of a thorium-containing precipitate, and at the second step - until achieving pH of not less than 7 with precipitation and separation of a concentrate of rare-earth elements.

EFFECT: invention increases extraction of rare-earth elements while obtaining a non-radioactive concentrate of rare-earth elements.

4 cl, 4 ex

FIELD: metallurgy.

SUBSTANCE: method to extract holmium (III) cations from nitrate solutions includes ion floatation using an anion-type surfactant as a collector. Besides, the collector is dodecyl sodium sulfate in a concentration corresponding to stoichiometry of the following reaction: Ho+3+3C12H25OSO3Na=Ho[C12H25OSO3]3+3Na+, where Ho+3 - holmium cation, C12H25OSO3Na - sodium dodecyl sulfate. Moreover, ion floatation is carried out at pH=6.6-7.4, which makes it possible to achieve 90% extraction of holmium from aqueous solutions of its salts.

EFFECT: higher extent of holmium extraction.

1 dwg, 1 tbl, 1 ex

FIELD: metallurgy.

SUBSTANCE: invention relates to the method for production of pure lanthanum or its oxides from lean or industrial raw materials by method of ion floatation. The method to extract lanthanum La+3 cations from aqueous solutions of salts includes ion floatation using an anion-type surfactant as a collector. Besides, the collector is dodecyl sodium sulfate in a concentration corresponding to the stoichiometric reaction: La+3+3NaDS=La[DS]3+3Na+, where La+3 - lanthanum cation, NaDS - dodecyl sodium sulfate. Moreover, ion floatation is carried out at pH=7.8-8.1, which makes it possible to achieve 98% extraction of lanthanum from aqueous solutions of its salts.

EFFECT: higher extent of lanthanum extraction.

2 dwg, 1 ex

FIELD: metallurgy.

SUBSTANCE: method for extracting rare-earth elements from the technological and productive solutions containing iron (III) and aluminium, with a pH-0.52.5, includes the sorption of rare-earth elements with strong-acid cation resin. As the strong-acid cation resin the microporous strong-acid cation resin is used based on hypercrosslinked polystyrene having a size of micropores 1-2 nm.

EFFECT: higher efficiency of the process due to greater sorption capacity of the said strong-acid cation resin, high kinetics of sorption and selectivity, improvement of the subsequent quality of eluates and simplification of the process of their further processing.

5 tbl, 5 ex

FIELD: metallurgy.

SUBSTANCE: 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-.

EFFECT: creation of economic and environmentally safe technology allowing to process low-grade molybdenite concentrates as per a short scheme, which provides sufficient increase in throughout extraction of molybdenum and its associated metals from ores to commodity products, and thus, contributes to more rational use of mineral resources.

2 cl, 3 tbl, 2 ex

FIELD: metallurgy.

SUBSTANCE: method involves oxidation of micro production wastes at temperature of 550-650C 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 20-80C. 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.

EFFECT: regeneration of rare-earth metals from production wastes of magnets and obtaining raw material containing rare-earth metals for reutilisation in production of rare-earth constant magnets.

2 cl, 2 tbl, 7 ex

FIELD: metallurgy.

SUBSTANCE: 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.

EFFECT: improving the extraction of precious metals; obtaining purer alloy and reducing the sulphide concentrate treatment costs.

3 cl, 4 tbl, 1 ex

FIELD: process engineering.

SUBSTANCE: 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.

EFFECT: solution suitable for further extraction.

3 cl, 2 tbl

FIELD: metallurgy.

SUBSTANCE: 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 1350-1450C in reducing medium with participation of clinker minerals till molten metal and sinters are formed. They are cooled to magnetising roasting temperature of 750-860C, crushed and separated with magnetic separation into clinker and concentrate. Then, concentrate is cooled to 50-90C and supplied at that temperature for leaching with mineral acid for dilution of phosphorus.

EFFECT: improved process efficiency.

5 cl, 2 tbl, 1 ex

FIELD: chemistry.

SUBSTANCE: 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.

EFFECT: obtaining high-quality products.

5 cl, 2 tbl, 2 ex

FIELD: metallurgy.

SUBSTANCE: 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.

EFFECT: improved process economy.

23 cl, 2 dwg

FIELD: chemistry.

SUBSTANCE: 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.

EFFECT: obtaining clean nuclear materials, suitable for producing uranium hexafluoride for enrichment.

2 tbl, 2 ex

FIELD: chemistry.

SUBSTANCE: 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 20-65C 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.

EFFECT: protection of extraction from formation of non-demixing emulsions, providing given purity of uranium from ballast impurities and obtaining raffinates which can be removed into underground collector sand layers.

2 tbl, 1 ex

FIELD: metallurgy.

SUBSTANCE: 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.

EFFECT: improving the opening rate of initial raw material and simplifying the process.

12 cl, 2 dwg, 3 tbl, 6 ex

FIELD: chemistry.

SUBSTANCE: invention relates to complexes of lanthanides and organic ligands which are luminescent in the visible spectrum and are used in electroluminescent devices, means of protecting security paper and documents from falsification etc. Disclosed are novel luminescent coordination compounds of lanthanides of formula: where Ln is Eu3+, Tb3+, Dy3+, Sm3+, Gd3+.

EFFECT: said compounds have high luminescence intensity and considerable thermal tolerance of up to 400C, which enables use thereof in modern production of light-emitting diodes.

4 dwg, 2 tbl

Up!