Process for extracting rare-earth metals and yttrium from coals and ash-slag waste material of coal burning

FIELD: hydraulic metallurgy, namely processes for extracting rare and rare-earth metals from natural organic raw materials such as coals and their burning products such as ash-slag waste materials.

SUBSTANCE: method is realized by using nitric acid for leaching. Selective extraction of rare-earth metal nitrates is realized due to extraction with use of organic solutions of tributyl phosphate and due to using part of heat of coal burning for regeneration of nitric acid by thermal decomposition of refined products and contained in them nitrates of potassium, aluminum, iron and other metals.

EFFECT: lowered consumption of reagents (acids) for leaching rare-earth elements from coals or ash-slag waste materials, simplified process for extracting and purifying rare-earth metals after processing leaching solutions.

4 ex

 

The invention relates to the field of hydrometallurgy, in particular to chemical extraction technology of rare elements from natural organic raw materials (coal) and products of its combustion - ash waste.

There are various ways to extract valuable elements (including rare earth) from the mineral matter of coal, which are in chemical processing of ash waste after combustion of coals of various chemical reagents. The main method of processing the wastes waste is opening their acidic reagents, which can be used as mineral acids and organic cationogenic N+form.

When sulfuric acid showdown ash waste from the combustion of coal from Ekibastuz is achieved by extraction into solution up to 98-99% of rare earth metals /Vfront, Linedev and other Research opportunities for rare-earth elements from fly ash CHP plants// proc. Dokl. Int th int. "Rare earth metals: processing of raw materials, production of compounds and materials based on them". Krasnoyarsk. 1995. S-1091. The basis of Ekibastuz evils are: SiO2- 61,5%, Al2O3- 27,4%, Fe2O3- 5,6%, CaO - 1,2%, MgO of 0.5%, other elements are less than 4%. The process is carried out at 50°and the concentration of sulfuric acid 100 g/l

Also known with the ladies of the sulfuric acid leaching of radioactive, REE processing ash sulfuric acid solution with the addition of a solution of sodium chloride in the amount of 0.5-25 g/l for process intensification /Pat. 213839. Russia, IPC603 In 9/06, From 04 To 7/28. Vfront, Linedev, Oaaea, Ulichila. The method of preparation of fly ash from the combustion of coal for use as a building material. The process is carried out at concentrations of sulfuric acid 50-300 g/l with a ratio of T:W=1:4-10 and temperatures from 18 to 90°C. the Extraction can be carried out in a reactor with stirrer during 1-6 hours or heap leaching. In the latter case, the process temperature is maintained within the range of 18-40°C.

The method for extracting scandium and yttrium from ash waste hydrochloric acid /Have Adinew, Glasgow, Lpimage. The extraction of scandium and yttrium from ash waste //. ZH. - 1995. - T, VIP. With 1075-1078/. Removing a hold of ash and slag waste from the combustion of brown coal, composition: SiO2- 40,1%, Al2O3- 10,6%, Fe2O3- 8,5%, CaO - 7,4%, MgO - 8.3 Percent. The proposed leaching scandium and yttrium in 2-3 stage by re-use of effluents for leaching. Leaching is carried out with 10% HCl solution when heated. When this is achieved the degree of extraction in a solution of: scandium - 84% and yttrium - 92%. It is determined that cross in selachian leads to substantial saturation of the solution of salts of calcium, magnesium, iron and aluminum. The high concentration of salts it difficult to separate the solid from the liquid phase.

Closest to the proposed method for extracting rare earth metals from the ash wastes is the process /Glascow, Rbiiyeaa and other Sorption leaching of scandium from ash-slag wastes from the combustion of brown coal Borodinsky/ proc. reports of the International conference "Rare earth metals: processing of raw materials, production of compounds and materials based on them". Krasnoyarsk. 1995. S-106/combining leaching and sorption (sorption leaching). According to this method, acidified slurry of ash and slag waste mixed with sulfonic cation exchanger KU-2 at a temperature of 40-60°C. It provides a transition in the ion exchanger of scandium and rare earth metals. Simultaneously videochelsea well as calcium, its residual concentration in the ash not more than 0.2% when the content in the original ash about 20%. A significant part of the iron in these conditions is not leached.

To remove the main mass of the calcium sorbent is then treated with 1M solution of sodium sulfate, acidified to 0,1M sulfuric acid. When the ion exchanger is transferred to the H-form, and calcium is separated in the form of plaster.

The disadvantages of the known processes for the extraction of rare earth metals from the ash wastes are great supplies the d acid to neutralize oxides of macroelements (calcium, magnesium, strontium, aluminum, iron) ash wastes and problems of separation of rare earth metals from complex solutions.

So, the richest known for rare earth metals ash contain 100-1000 g/kg (0.01 and 0.1%) of rare earth elements with the content of only one of calcium oxide 10-20% or more, the neutralizing capacity is significantly higher neutralizing capacity of the oxides of rare earth metals. For such ash waste overrun acids due to their reaction only with calcium compounds, respectively 200-5000 times or more. The use of sorption leaching does not eliminate this disadvantage, since the practical use of sorption leaching involves a complete regeneration of the cation exchanger, which is achieved at the stage of desorption of cations by treatment of the resin with the same mineral acids.

Another significant disadvantage of the known processes is the problem of the separation of rare earth metals from sulfuric and hydrochloric acid solutions given the complexity of the salt composition.

Known sulfuric acid schema processing of rare earth concentrates, as a rule, multi-stage and include the operations of deposition of poorly soluble double sulfates of rare earth metals with sodium, and subsequent hydrolysis of precipitation by sodium hydroxide, rest the dimension of the formed hydroxides of rare earth metals with nitric acid, and cleaning their extraction.

The closest analogue in essential characteristics is EN 93051055 (IPC From 22 In 59/00, publ. 27.09.1996), in which is disclosed a method of extracting rare earth metals and yttrium from the coals and ash waste from burning them, including acid leaching and extraction of rare earth metals and yttrium from solutions by alkylphosphates. The technical result is to reduce the consumption of reagents (acid) leaching and simplification of the process of extraction and purification processing of the leaching solutions.

The technical result is achieved that rare earth metals and yttrium from the coals and ash waste from the combustion of gas is also extracted by acid leaching and extraction from solutions tributylphosphate, but unlike a close analogue of rare earth metals and yttrium leached with nitric acid, which regenerate due to the utilization of heat from the combustion of coal by thermal decomposition of the nitrates of the raffinate obtained after extraction and absorption of water flue gas.

The use of nitric acid for leaching provides, first of all, the use of universal, selective to rare earth metals and most widely used in practice of the process of extraction with tributyl phosphate. This process is well to study and can be used for selective separation of rare earth metals from most elements. Moreover, some of the main micropremie from leaching of ash waste nitric acid, nitrates, calcium, aluminum and iron, are wycliffism when extracting rare earth metals tributylphosphate, facilitating their extraction into the organic phase. Reextracted metal salts and regeneration of the extractant is not very difficult and is water.

On the other hand, the use of nitric acid at the stage of leaching of rare earth metals allows the use of excess heat from the combustion of coal for the regeneration of the acid by thermal decomposition of nitrates. This process for calcium nitrate begins to flow at a temperature of 500-600°and intensively at 600-700°C. Nitrates of aluminum and iron decompose at lower temperatures. The decomposition of these salts leads to the formation of oxides of the respective metals, nitrogen dioxide and oxygen, as, for example, calcium nitrate (1):

The absorption of the formed gases water ensures the regeneration of the original nitric acid according to reaction (2):

theoretical consumption of heat in the regeneration of acid required for leaching of 1 ton of calcium oxide, calculated from the enthalpy of reaction (1), is about 1.6·106kcal/so When CPE is it calorie coal 4000 kcal/kg specific consumption of coal will be 400 (kg coal)/(t Cao). When the ash content of the coal 10% and the content soluble in mineral acids of not more than 50% (based on calcium oxide) consumption of coal on the regeneration of the acid is about 2% of the total burned mass.

Example 1. Ash from burning coal basin rdaidah-Yuryakh (Republic of Sakha-Yakutia), containing: 0.11% yttrium, from 0.25% lanthanum, 0.52% of cerium, 0.022% of dysprosium, 0.008% of ytterbium and other rare earth elements is heated under stirring with 3M (˜17%) nitric acid solution at T:W=1:8, a temperature of 90°C for 1 hour. The output of the cake is 61%. The solution is extracted: 91% yttrium, 75% of lanthanum, 72% of cerium, 90% of dysprosium and 93% of ytterbium. The solution is evaporated in 1,5-2 times and emuleret with 80% tributyl phosphate in kerosene at A:B=1:1. The distribution coefficients in the organic phase, respectively: 40 for yttrium, more than 5 for lanthanum and cerium, for more than 30 dysprosium and ytterbium. Upon contact of the extract with water in the ratio A:B=5:1 is achieved fully reextracted rare earth metals.

The raffinate then evaporated, and the residue is calcined at a temperature of 700°C for 1 hour. The degree of decomposition of salts is more than 98%.

Example 2. Ash from the combustion of brown coal Borodino basin (Krasnoyarsk region), containing 0.008% yttrium, heated under stirring with a 4M (˜17%) nitric acid solution at T:W=1:5, temperature 90°C for 1hour. The output of the cake is 49%. The solution is extracted 89% yttrium. The solution and rinse water evaporated up to the original volume and mixed with 80% tributyl phosphate in kerosene at A:B=1:1. The distribution coefficients of yttrium in the organic phase are more than 20. Upon contact of the extract with water in the ratio A:B=5:1 is achieved fully reextracted salts of rare earth metals.

The raffinate evaporated, the residue is calcined at a temperature of 700°C for 1 hour. The degree of decomposition of salts is more than 98%.

Example 3. Ash from burning coal basin rdaidah-Yuryakh (Republic of Sakha-Yakutia), containing: 0.15% yttrium, 0.27% of lanthanum, 0.085% of cerium, 0.025% of dysprosium, 0.0079% of ytterbium and other rare earth elements, is heated under stirring with 4,5M (˜25%) nitric acid solution at T:W=1:3, a temperature of 90°C for 1 hour. The output of the cake is 67%. The solution is extracted: 90.8% yttrium, 72% of lanthanum, 70% of cerium, 97.5% of dysprosium and 91.3% of ytterbium. The extractors of the type "mixer-settler" hold countercurrent extraction REM 80% tributyl phosphate in kerosene at A:B=1:1 (3-speed), washing of the extract with 1M solution of nitric acid at A:B=30:1 (2 stage with recirculation of aqueous solution) and re-extraction of water A:B=10:1 (2 stages). Extracting rare earth metals from solution after leaching in the reextract is more than 95%. Get rare earth is a product with a content of impurities of less than 10% (mass) of the oxides of rare-earth metals.

The raffinate evaporated, and the residue is calcined at a temperature of 700°in a rotary kiln for 2 hours. Pair of nitric acid trap in the absorption column of water. The result is a 25% solution of nitric acid with a yield of 92% of its initial amount, taken on the leaching of ash.

Example 4. Coal basin rdaidah-Yuryakh (Republic of Sakha-Yakutia), containing: 149 g/t - yttrium, 160 g/t lanthanum, 380 g/t - cerium, 28 g/t dysprosium, 16 g/t - ytterbium and other rare earth elements, is heated under stirring with a 1M solution of nitric acid at T:W=1:5 at 90°C for 1 hour. The extract in solution, respectively, is: 74,3% for yttrium, 80.1% for lanthanum, 85% for cerium, 80% of dysprosium, 69% for ytterbium. After evaporation of the solution of the nitrates of rare-earth metals extracted by tributyl phosphate and extravert water. Removing the REM for the 1st stage of extraction is more than 75%.

Thus, the proposed method allows to extract rare earth metals from coal or ash waste allocating them to concentrate with high rates of extraction (over 80%) and cleaning. Quality rare earth product content of impurities when using the extraction of rare earth metals from nitrate solutions by tributylphosphate can be brought to almost any level depending on the amounts of the steps of washing and Stripping. The use of the available heat from burning coal and thermal instability of nitrate helps to regenerate the basic chemical reagent nitric acid and repeatedly to reduce reagent costs.

The method of extraction of rare earth metals and yttrium from the coals and ash waste from burning them, including acid leaching and extraction of rare earth metals and yttrium from solutions by tributyl phosphate, characterized in that the rare earth metals and yttrium leached with nitric acid, which regenerate due to the utilization of heat from the combustion of coal by thermal decomposition of the nitrates of the raffinate obtained after extraction, and absorption of water flue gas.



 

Same patents:

FIELD: mining industry.

SUBSTANCE: invention relates to rare-earth element recovery technology in integrated processing of mineral stock, especially to hydrogen chloride technology of eudialyte concentrate. Method according to invention comprises introduction of nitrate ion into initial chloride solution containing rare-earth and impurity elements. Chloride-nitrate solution thus obtained is treated with tributyl phosphate extractant to transfer nitrates of rare-earth elements and a part of impurity elements into first organic phase, which is then separated from chloride-nitrate solution to form first extract and first raffinate containing major part of impurity elements. First extract is washed with nitrate solution and wash solution is then returned into initial chloride solution. Rare-earth element nitrates are reextracted with water and first reextract is separated. Concentrated 35-37% hydrochloric acid is added to the first raffinate to provide 5-10% excess of the acid relative to nitrate ion content. Raffinate is treated with tributyl phosphate to transfer nitric acid into second organic phase and to form chloride solution of the major part of impurity elements. Second organic phase is separated to form extract and second raffinate, after which nitric acid is reextracted from the second extract with alkali solution to form second reextract in the form of nitrate solution, which is used to wash the first extract.

EFFECT: increased rare-earth element recovery efficiency under better environmental conditions.

5 cl, 1 dwg, 3 ex

FIELD: hydrometallurgy; ore concentrates processing.

SUBSTANCE: the invention is pertaining to the field of hydrometallurgy, in particular, to processing of the loparite concentrate. The method includes a decomposing of the loparite concentrate at the temperatures of 103-105°C and the concentration of hydrofluoric acid of 38-42 mass % with production of the pulp containing fluorides of titanium, rare earth elements (REE), niobium, tantalum and sodium. The pulp is filtered at the temperature of 90-95°C with extraction into the fluorotitanium solution of fluorides of niobium and tantalum and no less than 58 % of sodium in terms ofNa2O and separation of the sediment containing fluorides of rare earth elements (REE) and a residual sodium. The produced solution is cooled down to 18-24°C with separation of the second sediment of sodium fluorotitanate. After that they extract niobium and a tantalum from the solution by octanol-1 extraction at a ratio of the organic and water phases as 1.1 : 1. The sediment of REE fluorides is washed from fluorotitanate by sodium water in a single phase at the temperature of 90-95°C and at the solid :liquid ratio = 1:2-2.5. The cleansing solution is separated and evaporated with extraction of the additional sediment of sodium fluorotitanate. After extraction of niobium and tantalum the fluorotitanium solution is evaporated and filtered with separation of the first sediment of sodium fluorotitanate from the concentrated solution of fluorotitanium acid, which is directed to extraction of titanium. The gained first, second and additional sediments of sodium fluorotitanate are combined and subjected to conversion with production of sodium fluorosilicate and the conversional fluorotitanium acid added to fluorotitanium solution before its evaporation. The technical result of the invention is a decrease in 2.0-2.5 times of the volume of the cleansing solutions at provision of a high degree of extraction of compounds of titanium and other target products. The produced sodium fluorotitanate contains the decreased amount of the impurity ingredients of calcium and strontium.

EFFECT: the invention ensures a decrease in two-two and a half times of the volume of the used cleansing solutions at provision of a high degree of extraction of compounds of titanium and other target products and a decreased amount of impurities of calcium and strontium in the sodium fluorotitanate.

7 cl, 1 dwg, 1 tbl, 3 ex

FIELD: non-ferrous metallurgy; methods of production of scandium-containing ligatures.

SUBSTANCE: the invention is pertaining to the field of non-ferrous metallurgy. The method of production of scandium-containing addition alloys includes a metallothermic restoration in halogenide melts. According to the invention the halogenide melt containing 1.0-1.4 mass % of scandium oxide is added with 1.4-1.7 mass % of zirconium oxide and conduct restoration by an alloy of aluminum with magnesium at the ratio of the halogenide melt to the aluminum-magnesium alloy from 1.2 up to 1.6. The technical result of the invention is production of a synthesized addition alloy containing scandium and zirconium with the maximal strengthening effect, decreased value of the produced addition alloy (by 30-40 %) due to decrease of consumption of the cost intensive scandium oxide by 50 %.

EFFECT: the invention ensures production of a synthesized scandium and zirconium ligature with maximal strength, allows to decrease significantly its production cost and consumption of expensive scandium oxide.

1 tbl, 1 ex

FIELD: metallurgy; hydrochemical methods of a complex processing of a multicomponent, polymetallic scrap.

SUBSTANCE: the invention is pertaining to the field of metallurgy, in particular, to the hydrochemical methods of a complex processing of a multicomponent, polymetallic scrap used in nonferrous metallurgy with extraction of valuable components and production of various commercial products. The technical result at reprocessing and neutralization of wastes of production of titanium tetrachloride consists in concentration of radioactive metals in the "head" of the process, transfer of the secondary wastes of production in an ecologically secure form suitable for a long-term entombment and-or storing, as well as in production of an additional commercial products - deficient and expensive black thermo- resistant inorganic pigments based on iron oxides, manganese and copper oxides. The method provides for a discharge of the spent melt of titanium chlorates into water; concentrating of a pulp by circulation; the pulp thickening; settling of metals oxyhydrates from the clarified solutions in succession in three stages: on the first stage - conduct a settling at pH = 3.-5.0 with separation of the formed settling of hydroxides of chrome, aluminum and scandium from the solution; on the second stage - conduct settling at presence of an oxidizing agent at pH = 2.5-3.5 within 20-50 hours with separation of the settling; on the third stage - conduct settling at pH = 9.5-11.0. The pulp at its circulation and concentration is added with sodium sulfite in amount of 5 - 15 g/dm3, then after circulation the pulp is treated with a solution of barium chloride in amount of 10-20 g/dm3 for cosettling of ions of thorium and radium, in the formed pulp of the first stage of settling introduce a high-molecular flocculant, and before settling process on the third stage of the process the solution is previously mixed with copper(II)-containing solution formed after lixiviation of a fusion cake of the process of cleanout of the industrial titanium tetrachloride from vanadium oxychloride by copper powder, then the produced settling of iron, manganese and copper oxyhydrates is filtered off, cleansed, dried and calcined at the temperature of 400-700°C.

EFFECT: the invention allows to concentrate radioactive metals in the "head" of the process, to transfer the process secondary wastes in the ecologically secure deficient and expensive black thermo-resistant inorganic pigments.

5 cl, 1 ex

FIELD: non-iron metallurgy, in particular scandium oxide recovery from industrial waste.

SUBSTANCE: method for preparation of scandium oxide from red mud being waste of alumina production includes: multiple subsequent leaching of red mud with mixture of sodium carbonate and hydrocarbonate solutions; washing and precipitate separation; addition into obtained solution zinc oxide, dissolved in sodium hydroxide; solution holding at elevated temperature under agitation; precipitate separation and treatment with sodium hydroxide solution at boiling temperature; separation, washing, and drying of obtained product followed by scandium oxide recovery using known methods. Leaching is carried out by passing through mixture of sodium carbonate and hydrocarbonate solutions gas-air mixture containing 10-17 vol.% of carbon dioxide, and repeated up to scandium oxide concentration not less than 50 g/m3; solid sodium hydroxide is introduced into solution to adjust concentration up to 2-3.5 g/m3 as calculated to Na2O (caustic); and mixture is hold at >=800C followed by flocculating agent addition, holding, and separation of precipitate being a titanium concentrate. Obtained mixture is electrolyzed with solid electrode, cathode current density of 2-4 A/dm3, at 50-750C for 1-2 h to purify from impurities. Zinc oxide solution in sodium hydroxide is added into purified after electrolysis solution up to ratio ZnO/Sc2O3 = (10-25):1, and flocculating agent is introduced. Solution is hold at 100-1020C for 4-8 h. Separated precipitate is treated with 5-12 % sodium hydroxide solution, flocculating agent is introduced again in amount of 2-3 g/m3, mixture is hold, and precipitate is separated. Method of present invention is useful in bauxite reprocessing to obtain alumina.

EFFECT: improved recovery ratio of finished product into concentrate; decreased impurity concentration in concentrate, reduced sodium hydrocarbonate consumption, as well as reduced process time due to decreased time of fine-dispersed precipitate.

2 cl, 2 ex

FIELD: chemical technology; deactivation and decontamination of radioactive industrial products and/or wastes.

SUBSTANCE: proposed method designed for deactivation and decontamination of radioactive industrial products and/or production wastes incorporating Th-232 and its daughter decay products (Ra-228, Ra-224), as well as rare-earth elements, Fe, Cr, Mn, Al, Ti, Zr, Nb, Ta, Ca, Mg, Na, K, and the like and that ensures high degree of coprecipitation of natural radionuclides of filtrates, confining of radioactive metals, and their conversion to environmentally safe form (non-dusting water-insoluble solid state) includes dissolution of wastes, their treatment with barium chloride, sulfuric acid, and lime milk, and separation of sediment from solution. Lime milk treatment is conducted to pH = 9-10 in the amount of 120-150% of that stoichiometrically required for precipitation of total content of metal oxyhydrate; then pulp is filtered and barium chloride is injected in filtrate in the amount of 0.4 - 1.8 kg of BaCl2 per 1 kg of CaCl2 contained in source solution or in pulp and pre-dissolved in sulfuric acid of chlorine compressors spent 5-20 times in the amount of 0.5 - 2.5 kg of H2SO4 per 1 kg of BaCl2. Then lime milk is added up to pH = 11 - 12 and acid chloride wash effluents of equipment and production floors are alternately introduced in sulfate pulp formed in the process at pulp-to-effluents ratio of 1 : (2-3) to pH = 6.5 - 8.5. Filtrate pulp produced in this way is filtered, decontaminated solution is discharged to sewerage system, sediment of barium and calcium sulfates and iron oxysulfate are mixed up with oxyhydrate sediment formed in source pulp neutralization, inert filler and 0.5 - 2 parts by weight of calcium sulfate are introduced in pasty mixture while continuously stirring them. Compound obtained in the process is placed in molds, held therein at temperature of 20 - 50 oC for 12 - 36 h, and compacted in blocks whose surfaces are treated with water-repelling material.

EFFECT: reduced radioactivity of filtrates upon separation of radioactive cakes.

8 cl, 1 dwg, 1 ex

FIELD: metallurgy; reworking wastes of alumina production process.

SUBSTANCE: proposed method includes preparation of batch of charge containing red mud and carbon reductant, heating the charge in melting unit to solid-phase iron reduction temperature, three-phase reduction of ferric oxides in charge by carbon reductant and saturation of iron with carbon in charge thus prepared, melting the reduced charge for obtaining metal phase in form of cast iron and slag phase in form of primary slag, separation of cast iron from primary slag in melt heated to temperature of 40 C, reduction of silicon and titanium from oxides contained in primary slag by aluminum and removal of cast iron and primary slag from melting unit; during preparation of charge, concentrate of titanomagnetite ore containing titanium oxide in the amount from 1 to 15% is added to red mud; besides that, additional amount of carbon reductant and additives are introduced; after separation of primary slag from cast iron in melting unit, cast iron is heated to 1500-1550 C and product containing ferric oxide is added to it; iron is reduced by carbon of cast iron for converting the cast iron into steel at obtaining secondary slag; main portion of steel is removed from melting unit, secondary slag is added to primary slag and silicon and titanium are converted into steel residue in melting unit by reduction with aluminum, thus obtaining final slag-saturated slag and master alloy containing iron, titanium and silicon; main portion of master alloy is removed from melting unit; after removal of final slag for converting the master alloy residue to steel in melting unit, titanium and silicon are converted into slag phase by oxidation and next portion of charge is fed to slag phase formed after converting the master alloy residue to steel. Proposed method ensures high efficiency due to obtaining iron-titanium silicon master alloy in form of independent product and production of alumina from high-alumina final slag or high-alumina cement and concentrate of rare-earth metals.

EFFECT: enhanced efficiency due to avoidance of intermediate remelting of steel.

10 cl, 2 dwg

The invention relates to a technology developing concentrates of rare earth elements from natural phosphate concentrates

The invention relates to the technology of extraction of rare earth elements (REE) from phosphogypsum obtained by sulphuric acid processing of Apatite concentrate on mineral fertilizers

The invention relates to the metallurgy of rare metals, more specifically to the technology of aluminum alloys with rare-earth elements, scandium, yttrium and lanthanides

FIELD: mining industry.

SUBSTANCE: invention relates to rare-earth element recovery technology in integrated processing of mineral stock, especially to hydrogen chloride technology of eudialyte concentrate. Method according to invention comprises introduction of nitrate ion into initial chloride solution containing rare-earth and impurity elements. Chloride-nitrate solution thus obtained is treated with tributyl phosphate extractant to transfer nitrates of rare-earth elements and a part of impurity elements into first organic phase, which is then separated from chloride-nitrate solution to form first extract and first raffinate containing major part of impurity elements. First extract is washed with nitrate solution and wash solution is then returned into initial chloride solution. Rare-earth element nitrates are reextracted with water and first reextract is separated. Concentrated 35-37% hydrochloric acid is added to the first raffinate to provide 5-10% excess of the acid relative to nitrate ion content. Raffinate is treated with tributyl phosphate to transfer nitric acid into second organic phase and to form chloride solution of the major part of impurity elements. Second organic phase is separated to form extract and second raffinate, after which nitric acid is reextracted from the second extract with alkali solution to form second reextract in the form of nitrate solution, which is used to wash the first extract.

EFFECT: increased rare-earth element recovery efficiency under better environmental conditions.

5 cl, 1 dwg, 3 ex

FIELD: methods of determination of content of palladium and platinum in ores containing considerable amount of iron, copper, zinc and other metals.

SUBSTANCE: proposed method includes decomposition of ore by hydrofluoric and nitric acids followed by further decomposition by aqua regia, boiling-off to moist salts, dissolving of them in hydrochloric acid and extraction. Determination of content of palladium is carried out in organic phase thus obtained and that of platinum is carried out in hydrochloric acid phase. Extractants used for such determination are s-alkylisothiouronium halides and alcohols of C5-C8 fractions, as well as kerosene, benzene, toluene and xylols used as diluents. Used as s-alkylisothiouronium halides are chlorides, bromides and iodides from C7 to C14 and their fractions.

EFFECT: extended range of reagents and inert diluents.

2 cl, 2 ex

FIELD: hydrometallurgy; ore concentrates processing.

SUBSTANCE: the invention is pertaining to the field of hydrometallurgy, in particular, to processing of the loparite concentrate. The method includes a decomposing of the loparite concentrate at the temperatures of 103-105°C and the concentration of hydrofluoric acid of 38-42 mass % with production of the pulp containing fluorides of titanium, rare earth elements (REE), niobium, tantalum and sodium. The pulp is filtered at the temperature of 90-95°C with extraction into the fluorotitanium solution of fluorides of niobium and tantalum and no less than 58 % of sodium in terms ofNa2O and separation of the sediment containing fluorides of rare earth elements (REE) and a residual sodium. The produced solution is cooled down to 18-24°C with separation of the second sediment of sodium fluorotitanate. After that they extract niobium and a tantalum from the solution by octanol-1 extraction at a ratio of the organic and water phases as 1.1 : 1. The sediment of REE fluorides is washed from fluorotitanate by sodium water in a single phase at the temperature of 90-95°C and at the solid :liquid ratio = 1:2-2.5. The cleansing solution is separated and evaporated with extraction of the additional sediment of sodium fluorotitanate. After extraction of niobium and tantalum the fluorotitanium solution is evaporated and filtered with separation of the first sediment of sodium fluorotitanate from the concentrated solution of fluorotitanium acid, which is directed to extraction of titanium. The gained first, second and additional sediments of sodium fluorotitanate are combined and subjected to conversion with production of sodium fluorosilicate and the conversional fluorotitanium acid added to fluorotitanium solution before its evaporation. The technical result of the invention is a decrease in 2.0-2.5 times of the volume of the cleansing solutions at provision of a high degree of extraction of compounds of titanium and other target products. The produced sodium fluorotitanate contains the decreased amount of the impurity ingredients of calcium and strontium.

EFFECT: the invention ensures a decrease in two-two and a half times of the volume of the used cleansing solutions at provision of a high degree of extraction of compounds of titanium and other target products and a decreased amount of impurities of calcium and strontium in the sodium fluorotitanate.

7 cl, 1 dwg, 1 tbl, 3 ex

FIELD: metallurgy of rare and dispersed metals, chemical technology.

SUBSTANCE: invention relates to a method for extraction separation of tantalum and niobium. Method involves extraction separation of tantalum from niobium with organic solvent. As an organic solvent method involves using a mixture of methyl isobutyl ketone taken in the amount 40-80 vol.% with aliphatic (C7-C9)-alcohol taken in the amount 20-60 vol.%. At the extraction process tantalum transfers into organic phase and niobium - into aqueous phase. Then organic and aqueous phases are separated. Invention provides enhancing the extraction degree of tantalum into organic phase and to enhance the separation degree of tantalum and niobium in extraction.

EFFECT: improved separating method.

5 tbl, 5 ex

The invention relates to the processing orangutango raw materials

The invention relates to the field of hydrometallurgical processing of tantalum raw materials and are aimed at achieving its complex use

The invention relates to the extraction of substances organic extractants from aqueous solutions and can be used in ferrous and nonferrous metallurgy, as well as for the treatment of industrial and domestic wastewater

The invention relates to methods for determination of palladium and platinum ores containing large amounts of iron, copper, zinc and other metals

The invention relates to a method of extracting metal from ore or concentrate containing Nickel and/or cobalt and other metals

The invention relates to the field of non-ferrous metallurgy, in particular to the technology of rare and scattered elements, and can be used for the recovery of indium from zinc sulfate solutions with a high content of silica

FIELD: mineral dressing.

SUBSTANCE: method comprises charging, chemically enriching concentrate, cleaning, and discharging desired product. Chemical enrichment is carried out by way of single or multiple processing in acid or in acids and then in alkali or alkali mixture, while heating material to 900-1000°C and holding it at this temperature in inert gas medium at stirring.

EFFECT: enhanced diamond cleaning efficiency.

6 cl, 1 tbl

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