Method of extracting uranium from phosphate solutions
SUBSTANCE: invention relates to hydraulic metallurgy, particularly to extraction of uranium from used phosphate solutions. This process consists in adding the solvent to initial solution, said solvent being selected from the series: KMnO4, K2Cr2O7, HNO3, H2O2, KClO3. Then, uranium-bearing sediment is precipitated by acidity correction by ammonia to pH 2.8-4.0 at 20-35°C. Filtered precipitate is treated by 20-35% solution of NaOH at 80-85°C for 1.5-2.0 hours.
EFFECT: higher yield of uranium, return of high-enriched uranium to fuel cycle, lower costs of higher safety at long-term storage, accounting and control.
2 cl, 1 tbl, 4 ex
The invention relates to the field of hydrometallurgy, in particular to a method of extracting uranium from phosphate waste solutions resulting from the chemical analysis conducted by a standardized titrimetric method Davis-gray, on the uranium content in its various derivatives (oxides, nitrides, carbides and other).
The method for extracting uranium from solutions using ion-exchange resins that selectively sorbing uranium with subsequent elution (leaching) in desorbed (RF Patent No. 2159741, IPC C01G 43/00, B01D 15/04, publ. 27.11.2000).
The drawback of this method is low specific capacity of the resin and therefore no need for a bulky nuclear-safe equipment.
There is also known a method of extracting uranium from phosphoric acid solutions by multistage countercurrent extraction using as the organic phase of the mixture dialkylphosphorous acid and trialkylphosphine in inert diluent (Patent USSR No. 858572, IPC C22B 60/02, publ. 23.08.81).
A large number of equipment, and high capital costs required to create a closed extraction cycle, limit the application of this method.
The closest in technical essence and the tasks of the claimed invention prototype one is by the method of extraction of uranium from phosphate solutions, the resulting acidic opening monazite concentrates. The method is almost complete neutralization of the free acid with ammonia (pH~6,0) and subsequent separation of the uranium-containing precipitate by filtering (I. N. Beckman, Lectures MSU "Thorium", 2010,, 136 S.).
The disadvantage is the low rate of deposition of uranium, due to the fact that man-made solutions, resulting from the application of techniques of Davis-gray, at least part of the uranium has a valence U4+and does not form a deposited ammonia chemical complex. Furthermore, formed in the more alkaline environment of the ammonium phosphate has a substantially lower solubility compared to its di - and hydro-phosphate, which really complicates further processing due to the inclination of the system to spontaneous crystallization.
The present invention is directed to the achievement of the technical result consists in increasing the degree of extraction of uranium from concentrated on the phosphate ion solutions, obtaining the processed uranium concentrate and low uterine solutions.
The technical result is achieved in that in the method of extracting uranium from phosphoric acid solutions, including the introduction in the original precipitant solution, adjusting pH, filtering and processing of the resulting sludge, coz the ACLs to the invention in the original phosphate solution pre-enter the oxidizer, carry out the adjustment of the pH with ammonia to pH (2.8÷4,0) at a temperature of (20÷35°C, as obtained after filtration the precipitate is treated (20÷35)% solution of NaOH at a temperature of (80÷85)°C for (1,5÷2,0) hours.
This oxidizing agent may be selected from the range: KMnO4, K2Cr2O7, HNO3H2O2, KClO3.
The inventive method differs from the known fact that in the original solution is injected oxidant, for example a saturated aqueous solution of potassium permanganate to get a stable purple color, indicating the completion of the transition U4+=>U6+. Next, the free acid is neutralized with an aqueous solution of ammonia to a pH of 2.8÷4,0, while maintaining a temperature in the range 20÷35°C.
The need to maintain the indicated temperature range due to the fact that at temperatures over 35°C is thermal decomposition of potassium permanganate and, accordingly, its wasteful consumption, while lowering the temperature below 20°C significantly slows down the flow of liquid-phase oxidation-reduction reactions. The choice of the interval of pH values is determined by the fact that at a pH of less than 2.8 increasing uranium concentration in the mother solution and at pH 4.0, the formation of ammonium phosphate, considerably less soluble than hydrophosphate form.
D is stijene technical result contributes to the fact, that is obtained after filtration, the residue treated with 20÷35% sodium hydroxide solution at a temperature of 80÷85°C for 1,5÷2,0 hours. This is done in order to reduce the phosphorus content in selected from the original solution of uranium concentrate, thereby ensuring its absolute suitability for further refinery processing.
The essence of the proposed technical solution is illustrated by examples of specific implementation.
The table shows the typical chemical composition of spent uranium phosphate solutions as a result of chemical analysis by standardized titrimetric method Davis-gray, on the uranium content in its various derivatives (oxides, nitrides, carbides and other). (Branch instruction. Procedure for potentiometric determination of uranium, 2000, 33 S. AOI 001.493-00).
|Uranium total||(a 2.0 to 4.6) g/l|
|Phosphoric acid||384 g/l; 4M; 12N|
|Sulfamic acid||5.7 g/l|
|The bichromate is Aliya||1.2 g/l|
|Nitric acid||43,4 g/l; 0,7 N|
|Molybdate ammonium||0.36 g/l|
|Ferric sulfate||6,72 g/l|
|Sulfuric acid||2.16 g/l|
|Ammonium Vanadate||0,044 g/l|
Example 1. The original solution of the above composition, taken in an amount of 1.0 l, was treated with 25% ammonia to obtain a pH=6,8. After filtering received 1,66 l stock solution containing 0.6 g/l of uranium. Thus, its degree of deposition did not exceed 80%.
Example 2. In 1.0 l of the starting solution introduced ≈20 ml of a saturated solution of KMnO4to get a stable weakly-violet color. Then, avoiding overheating of the solution, introduced a 25% ammonia to pH=3,5. After filtration, the residual uranium content in the mother solution was 0.03 g/l Specific activity such solutions at a ratio of isotopes 235/238≈9 does not exceed 3.7·105Bq/kg of This can be attributed to the category of low and greatly simplifies and reduces the cost of disposal.
Example 3. In 1.0 l of the starting solution was added ≈25 ml of potassium permanganate solution until obtaining the SLA is on-violet color, and then, gradually adding ammonia, brought the pH to a value of 4.5. The precipitation was filtered, and the solution was analyzed for uranium content, which was 0.07 g/L. Therefore, further reduction of the acidity of the system does not contribute to the completeness of precipitation.
Example 4. 30 l of the starting solution was treated with ammonia according to the proposed procedure, the precipitate was filtered and treated with 25% sodium hydroxide solution at a temperature of 85°C for 1.5 hours. The weight of the resulting chemical concentrate amounted to 422 g when the uranium content of 25.0 wt.%, which corresponds to the degree of extraction of >96%.
Thus, as seen from the above examples, the claimed technical solution in the form of a set of proposed operations and parameters, novel, technically feasible and is cost effective.
Economic efficiency from the use of the invention due to the high degree of extraction of uranium, the return of highly enriched uranium in the fuel cycle, the elimination of a significant amount of intermediate level liquid waste by transferring them into the category of low, reduce costs associated with maintaining their safe long-term storage, accounting and control.
1. The method of extraction of uranium from phosphoric acid solutions, including the store the precipitation of the uranium-containing precipitate from the original solution by adjusting the pH, filtering and processing of the resulting sludge, characterized in that in the original phosphate solution pre-enter the oxidizer, the adjustment of the acidity spend ammonia up to pH values of 2.8÷4,0 at a temperature of 20÷35°C, and obtained after filtration, the residue treated with 20÷35% NaOH solution at a temperature of 80÷85°C for 1,5÷2,0 hours.
2. The method according to p. 1, characterized in that the oxidizing agent is chosen from the series: KMnO4, K2Cr2O7, HNO3H2O2, KClO3.
SUBSTANCE: method involves leaching a concentrate with aqueous nitric acid solution at high temperature to obtain a pulp which consists of a solid phase and an aqueous phase, filtering off the aqueous phase in form of uranyl nitrate solution, extraction refining uranium using tributyl phosphate in a hydrocarbon diluent. The filtered uranyl nitrate solution, which contains uranium in concentration of 200-400 g/l, dissolved silicon in concentration of 1.0-3.2 g/l and nitric acid in concentration of 1-2 mol/l, is held until stabilisation of viscosity before being fed for extraction.
EFFECT: preventing escape of the aqueous phase with the uranium extract, which improves efficiency of the extraction stage, lowers content of impurities in the uranium extract and enables to obtain a product which meets ASTM C 788-03 requirements.
1 dwg, 1 tbl
SUBSTANCE: invention relates to method of uranium extraction from mother liquors. Method includes obtaining resin, modified by aminophosphonic groups, and obtaining mother liquor, which contains from 25 to 278 g/l of sulphate and uranium. After that, mother liquor is passed through resin, modified by aminophosphonic groups, in acid form to separate uranium from mother liquor. Then, elution of uranium from resin is realised.
EFFECT: possibility of sorption extraction of uranium from solutions, which contain high concentrations of sulfate.
7 cl, 1 tbl, 3 ex
SUBSTANCE: proposed process comprises leaching of uranium by nitric acid and separation of water phase from undissolved precipitate. Then, undissolved precipitate is mixed with fluorine-bearing agent, dissolution of produced charge and/or charge as a suspension in nitric acid solution. Produced solution is returned to production process for extraction of uranium. Nitric acid concentration in solution makes at least 2 mol/l. Dilution is carried out at fluorine-ion concentration at, at least, 15 g/l. Dilution is performed at 60-100°C.
EFFECT: decreased losses of uranium, minimised wastes.
4 cl, 1 dwg, 1 tbl
SUBSTANCE: method involves dissolving wastes in concentrated nitric acid, oxalate precipitation from the solution, drying and calcining the americium oxalate to americium dioxide. The solution obtained by dissolving wastes with high concentration of impurity cations, one of which is ferric iron, is mixed with a reducing agent for reducing ferric iron to ferrous iron. After reduction, the solution with acidity by nitric acid of 1-2.5 mol/l is taken for extraction of americium with a solid extractant based on different-radical phosphine oxide, followed by washing and re-extraction of americium. Oxalate precipitation is carried out from the re-extract with americium concentration of not less than 3 g/l and nitric acid concentration of not less than 3 mol/l, said precipitation being carried out in two steps: adding an oxalate ion to the americium-containing solution in weight ratio to americium of (2-7):1 and then adding water to the separated precipitate in volume ratio to the precipitate of (3-8):1 and the oxalate ion in weight ratio to americium of (1-4):1. The obtained reaction mixture is boiled and taken for separation of americium oxalate from the solution.
EFFECT: high output of the product and degree of purity thereof.
SUBSTANCE: metallic uranium obtaining method involves electrolysis of uranium dioxide in the melt of lithium and potassium chlorides in an electrolysis unit with a graphite anode and a metal cathode and release of metallic uranium on the cathode and carbon dioxide on the anode. First, mixtures of uranium dioxide and carbon are prepared in molar ratio of 6:1 and 1:1 by crushing the corresponding powders; the obtained powders are briquetted into pellets. To the anode space of the electrolysis unit, which is formed with a vessel with porous walls, which is arranged in a ceramic melting pot, there loaded are pellets obtained from mixture of uranium dioxide and carbon, and melt of lithium and potassium chlorides. To the cathode space of the electrolysis unit, which is formed with the vessel walls with porous walls and the ceramic melting pot, there loaded is melt of lithium and potassium chlorides and uranium tetrachloride in the quantity of 5-15 wt % of lithium and potassium chlorides. Electrolysis is performed at the electrolyte temperature of 500-600°C, cathode density of current of 0.5-1.5 A/cm2, anode density of current of 0.05-1.5 A/cm2, in argon atmosphere with periodic loading to anode space of pellets of mixture of uranium dioxide and carbon.
EFFECT: current yield of metallic uranium is 80-90% of theoretical.
SUBSTANCE: method involves dissolving a chemical concentrate of natural uranium in nitric acid solution, extracting and re-extracting uranium. The dissolved concentrate contains 1.2-3.7 wt % iron to uranium, 1.4-4.0 wt % sulphur to uranuim and 0-0.7 wt % phosphorus to uranium in nitric acid solution. Nitric acid and water are taken in an amount which provides the following concentration in the solution fed for extraction: uranium 450-480 g/l, iron (III) ions 0.1-0.3 mol/l, sulphate ions 0.2-0.6 mol/l, phosphate ions 0-0.10 mol/l, and free nitric acid 0.8-2.4 mol/l, and saturation of extractant with uranium during extraction is maintained in accordance with the ratio: Y ≤90.691-34.316·[SO4]+7.611·([Fe]-[PO4])+5.887·[HNO3]-9.921·[SO4]·[HNO3]+19.841·[SO4]2+7.481·([Fe]-[PO4])·[HNO3]-64.728·([Fe]-[PO4])·[SO4]+92.701·[SO4]·[HNO3]·([Fe]-[PO4])-185.402·[SO4]2·([Fe]-[PO4]), where Y is saturation of the extractant with uranium, %, and concentration in the solution fed for extraction, mol/l: [SO4] - sulphate ions, [PO4] - phosphate ions, [HNO3] - nitric acid, [Fe] - iron (III) ions.
EFFECT: obtaining raffinates with low uranium content.
SUBSTANCE: method includes sorption of rich components from production solutions by ion-exchange material counterflow under controlled pH of environment and oxidation-reduction potential Eh. Sorption is performed by ion-exchange materials in stages from production solutions containing uranium, molybdenum, vanadium and rare earth elements. At the first stage uranium and molybdenum are extracted by anion-exchange material sorption. At the second stage vanadium is extracted by anion-exchange material sorption with hydrogen dioxide available at Eh of 750-800 mV, pH of 1.8-2.0 and temperature of 60°C, at that vanadium sorption is performed till complete destruction of hydrogen dioxide and till Eh is below 400 mV. Then barren solutions are transferred to cationite at pH of 2.0-2.5 and Eh of 300-350 mV for extraction of rare earth elements.
EFFECT: sorption concentration and selective separation of uranium and molybdenum from vanadium, and vanadium from rare earth elements, and rare earth elements from iron and aluminium, intensification of sorption process, reduction of flow diagram and possibility of environmentally sound oxidants use.
1 dwg, 4 tbl, 1 ex
SUBSTANCE: 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.
EFFECT: high-efficiency extraction of rich components, possibility of pulps separation by filtration after leaching with high properties thus reducing costs for separation processes.
3 cl, 1 dwg, 1 tbl
SUBSTANCE: processing method of black-shale ores with rare metals extracting includes leaching of ore by sulphuric acid solution with dilution of rare metals. Leaching is performed in autoclave by sulphuric acid solution consisting of free and combined sulphuric acid with ratio of H2SO4(free):H2SO4(comb)=2:1, and containing 25-45 g/l of iron sulphate, 70-90 g/l of aluminium sulphate and 0.5 g/l of nitric acid. At that the process is performed under pressure in autoclave equal to 10-15 atm with mixing at temperature of 140-160°C in concentration range of general H2SO4(gen) equal to 350-450 g/l under pulp density S: L=1:0.7-0.9, preferably 1:0.8, under constant oxidation-reduction potential Eh in the system equal to 350-450 mV during 2-3 hours till residual concentration of free H2SO4(free) is within 45-75 g/l.
EFFECT: increasing break-down of ore and extraction of rare metals: vanadium, uranium, molybdenum and rare-earth elements, reducing consumption of acid and improving efficiency of autoclave volume usage.
1 tbl, 1 ex
SUBSTANCE: invention relates to the technology of processing chemical concentrates of natural uranium, involving leaching (dissolving) the concentrate and extracting uranium using tributyl phosphate in a hydrocarbon diluent. The method involves dissolving the concentrate using aqueous nitric acid solution, feeding the obtained aqueous uranyl nitrate solution to the extract outputting step of a stepped extraction unit and extracting uranium with tributyl phosphate in a hydrocarbon diluent. Extraction is carried out by counterflow interaction of the aqueous and organic phases. Concentrate containing thorium impurities in ratio of 1 wt % to uranium is used. During extraction at the extract outputting step, the step for saturating the extractant with uranium is kept at least 87% of the maximum saturation of the extractant with uranium, and a portion of the aqueous phase, which is not more than 60 vol. % of the uranyl nitrate solution fed to the extract outputting step, after one of the extraction steps is removed from the extraction process and fed for dissolving the uranium concentrate.
EFFECT: high extraction of uranium and nitric acid from the raffinate.
SUBSTANCE: invention relates to a method of processing silver-containing concentrates. Oxidation-chlorination roasting is carried out with the application of chlorides of alkali metals with obtaining a chloride cinder, further leaching of the chloride cinder and separation of a cake from the solution. The cake of autoclave leaching of silver-containing concentrates, with the application as such of copper sulphide silver-containing concentrates, is subjected to oxidation-chlorination roasting. Leaching of the chloride cinder is performed in water with obtaining a chloride solution and a cake of water leaching. The obtained chloride solution is processed with sodium sulphide and calcium chloride with the separation of residues of copper sulphide and calcium sulphate dehydrate. A water solution of alkali metal chloride is evaporated to saturation and directed to oxidation-chlorination roasting. The cake of water leaching is leached in a sodium thiosuphfate solution. The obtained productive solution is processed with sodium sulphide, the silver-containing sediment of sulphides is separated from the regenerated sodium thiosulphate solution, and the regenerated solution is directed to a stage of leaching the cake of water leaching, with the sediment of sulphides being processed with obtaining silver.
EFFECT: reduction of expenditures and increase of silver extraction.
4 tbl, 1 ex
SUBSTANCE: 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.
EFFECT: higher oxidative potential of leaching solution, lower consumption of reagents and their concentration.
9 cl, 3 dwg, 1 tbl, 1 ex
SUBSTANCE: 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.
EFFECT: technical result consists in the reduction of expenditures due to the creation of economically efficient technology.
4 cl, 2 tbl, 2 ex
SUBSTANCE: method includes precipitation of rhenium sulfides by processing with sulfide-containing precipitator in presence of reducing reagent in form of hydrazine-containing compound and heating reaction mixture. Precipitation of rhenium sulfides in carried out from hydrochloric acid solution, containing iron chloride. First, hydrazine-containing compound in form of hydrazine chloride, hydrazine sulfate or hydrazine hydrate is introduced into solution in amount 2-4 g/l counted for pure hydrazine. After that, reaction mixture is heated at temperature 60-90°C for 20-40 minutes. Then, counted mass of sulfide-containing reagent is introduced and pulp is heated at specified temperature.
EFFECT: reduction of consumption of reducing reagent and high extraction of rhenium into sulfide sediment from hydrochloric solutions, containing iron salts.
1 tbl, 1 ex
SUBSTANCE: invention relates to the technology of production of gold nanoparticles. The method of production of gold nanoparticles from the raw material containing iron and non-ferrous metals comprises preparation of the chlorazotic acid solution of gold using chlorazotic acid. Then floatation extraction of gold precursors is carried out with cationic surfactants from the solution, separation and evaporation of the organic phase to concentrate the gold precursors. Then the concentrate reduction is carried out to obtain dispersion of gold nanoparticles. At that the starting material is first treated with hydrochloric acid to form the insoluble precipitate. Production of chlorazotic acid solution is carried out by dissolving in chlorazotic acid solution of insoluble precipitate. Before floatation extraction of precursors the nitric acid is removed from chlorazotic acid solution with methyl or ethyl alcohol or hydrochloric acid.
EFFECT: improvement of efficiency of the method of production of nanoparticles, namely the increase in the number of gold nanoparticles obtained or its hybrids with noble metals.
FIELD: process engineering.
SUBSTANCE: invention relates to cleaning of silver-bearing materials by hydrometallurgy processes, for example, scrap and wastes of microelectronics. Proposed method comprises dilution of silver-bearing material in nitric acid, addition of sodium nitrate to nitrate solution at mixing, extraction of silver salt precipitate and pits treatment to get metal silver. Note here that after addition of sodium nitrate the reaction mix is held for 1 hour to add sodium carbonate or bicarbonate to pulp pH of 8-10. Free silver salt precipitate as silver carbonate is separated from the solution by filtration. Sodium nitrite and carbonate or bicarbonate is added in the dry form. Note here that sodium nitrite is taken with 25% excess of stoichiometry.
EFFECT: higher purity and yield, simplified process.
2 cl, 2 ex
SUBSTANCE: proposed method comprises leaching the stock nitric acid solution to obtain suspension, introducing coprecipitator therein at 30-50°C and mixing. Then, clarified solution is separated from insoluble residue and directed for extraction. Said coprecipitator represents fresh solution of copolymer of acrylamide and chloride trimethyl ammonium ethyl acrylate of molecular weight of 3-15 millions with low charge density. Copolymer is introduced to concentration of 5.95-11.9 mg/l of insoluble residue. Prior to separation of clarified solution from insoluble residue settling is performed for 30-40 minutes.
EFFECT: lower processing costs.
3 cl, 1 dwg, 1 tbl
SUBSTANCE: method includes grinding of initial material, cyanide leaching with production of a product solution of gold with mercury admixtures, introduction of a sulfide-containing reagent for mercury deposition, sorption of gold onto activated coal with return of the reuse cyanide solution for leaching, desorption of gold and electrolysis of gold from a strippant. The sulfide-containing reagent is introduced in the form of an aqueous solution of a mixture of sodium sulfide and calcium oxide at their mass ratio of 4.3-4.4 per 900-1100 wt parts of the reuse cyanide solution. After separation of mercury in the form of a sparingly soluble residue, the suspension is separated to produce a clarified solution, from which the gold is adsorbed onto activated coal.
EFFECT: practically complete separation of mercury without negative impact at gold sorption.
4 cl, 1 dwg, 1 ex
SUBSTANCE: invention relates to a method of producing bismuth potassium citrate. Bismuth potassium citrate is obtained by treating bismuth citrate with aqueous potassium hydroxide solution. The method is carried out with molar ratio of potassium hydroxide to bismuth citrate of 1.0-1.5, and with weight ratio of potassium hydroxide solution to bismuth citrate of 0.5-2.0. The product is obtained in form a paste.
EFFECT: simple process and reduced consumption of reactants.
1 tbl, 3 ex
SUBSTANCE: method of noble metal extraction from solid stock comprises dissolving of noble metal and base metals in acid. Noble metal is extracted with the help of substituted quaternary ammonium salts (SQAS). Noble metal can be oxidised and reduced. Said substituted quaternary ammonium salts represent the following form H0-3R4-1NX, where H= hydrogen, R= organic group, N= nitrogen and X= halogen. This method uses, for example, tetramethyl ammonium chloride. Au-SQAS is separated by flushing with solvent. Rh-SQAS is dissolved in acid and oxidised to precipitate the salts, and separated. SQAS is added to filtrate and cooled to precipitate Rh-AQAS to be separated. Rh-SQAS is cleaned before formation of final product. Other metals are separated by boiling the initial acid solution with precipitation of metal salts, cooling and separation. The pulp is separated by dissolution and separation.
EFFECT: simplified extraction.
20 cl, 4 dwg, 4 tbl, 2 ex
FIELD: powder metallurgy, namely processes for producing silver powder used in electrical engineering industry branches, possibly for making electrodes of chemical electric current sources, electric contacts and so on.
SUBSTANCE: method comprises steps of depositing silver chloride from solution of silver nitrate with use of water soluble chloride at temperature 20 - 50°C and pH 1 - 5; decanting mother liquor; treating suspension with solution of alkali metal hydroxide at concentration in reaction medium 12 - 200 g/l; reducing silver from suspension by means of Formalin or formate at temperature 40 - 90 c for 10 -60 min; washing out successively in hot deionized water, in ammonium solution and in cold deionized water; filtering and drying deposit of silver powder at 70 - 120°C.; sifting dried powder through sieve with mesh 250 micrometers.
EFFECT: improved electrochemical, chemical and physical properties of silver powders.
2 cl, 1 tbl, 1 ex