Method of copper electrorefining leaded slimes treatment (versions)
SUBSTANCE: invention relates to copper electrorefining leaded slimes treatment, with slimes containing lead, antimony, aurum, silver and rare chalcogens and may be used to produce collective concentrates of precious metals. The method considers two versions of slimes treatment. Both versions include consecutive leaching of slimes and floatation. According to the first version, slime is leached in sulfuric acid solution at a temperature of 104-106°C, partial oxygen pressure of 0.02-0.1 MPa and stirring with oxygen absorption speed no less than 0.001 mole O2/m3-hour-Pa, filtered and floated. According to the second version, slime is subjected to liquid-phase sulfurisation at a temperature 160-200°C, then to leaching by ferric (II) sulfate, filtration and floatation. Extraction of aurum and silver in concentrate reached 99.8%.
EFFECT: slime is reduced and content of precious metals therein increases 3,6-4,8 times.
2 cl, 2 ex, 4 dwg, 1 tbl
The invention relates to ferrous metallurgy, in particular to a method of processing sludge of the electrorefining of copper, lead, antimony, gold, silver and rare chalcogen.
Chemical and phase compositions medeelektrolitnyj sludge - middlings metallurgical production, concentrating precious metals and rare chalcogen depend on the type and composition of feedstock (see table 1).
In Nickel sludge grain phases of precious metals are present mainly in individual state free from the oxide film surface. Features of their structure and composition make the processing is relatively simple. The main present admixture of Nickel can be easily removed together with copper-known, relatively simple to implement methods (autoclave sulfuric acid leaching, liquid-phase sulfamidate).
For lead sludge is characterized by the presence of aggregates of oxide phases impurities with precious metal chalcogenides. The clusters represented by the two types of particles.
The first represent the grains of sulphate of lead sheath chalcogenides silver and/or copper-silver covered with a thin film of secondary oxides of antimony and arsenic. They repeat the form and structure of inclusions anode metal (figures 1-2), however, due to the transformations, petechiae the s during electrorefining, excellent composition. The fraction of particles with such a structure, the sludge is higher, the lower the oxygen content in the anode in the form of cuprous oxide and the higher the ratio Pb/(Se+Te).
The latter represent the bulk conglomerates variable composition of secondary origin and are composed of fine particles of oxides of Sb-As, Arsenates and antimonates non-ferrous metals with inclusions of silver chalcogenides (figure 2).
Effective beneficiation of lead sludge is possible only when the organization of special operations for the removal of lead and antimony, as their extraction from sludge in the process of abuseive in sulfate environments impossible.
The literature describes methods of chemical leaching solutions of amines, chlorides of sodium and/or calcium, sodium hydroxide, a mixture of acetic acid and ammonium acetate) and mechanical extraction of lead (flotation) of medeelektrolitnyj sludge.
Application methods of chemical obespylevaniya involves large amounts of unstable, aggressive solutions (in the leaching of chlorides), and the high cost of reagents (leaching solutions of ethylene diamine).
The mechanical methods of enrichment are characterized by lower capital and conversion costs at comparable indicators on removing lead, but because of the nature of the chemical and phase SOS is ava sludge is not always effective.
The known method of flotation-autoclave processing of sludge electrorefining of Nickel (Intellence. Flotation-autoclave processing circuit anode slimes Nickel electrolysis. Non-ferrous metals, 1959, No. 7, p.36-40), according to which the sludge repulping and platinuum in sulfuric acid solution under the scheme, involving the primary flotation and two cleaners concentrate. The obtained foamy product is subjected to magnetic separation to remove iron and then leached with water in an autoclave at a temperature of 115-120°C and a partial pressure of oxygen of 1.5-2.0 MPa to remove sulphides of non-ferrous metals. In current products - magnetic concentrate (0,86% Pd) and the flotation tailings (~700 g/t Pd), extracted with 3.7% palladium. These products are offered melting on sulfide alloy, which, after crushing and grinding to attach to the original sludge. Direct extraction of palladium in concentrate - 95.3 per cent.
This method is not applicable for the processing of lead - and surmuslerdi slimes, because even long sulphuric acid repulpable not possible to clean the surface of the chalcogenide precious metals from the films of oxides of antimony and arsenic, to destroy the matrix component, and to individualize grain phases of precious metals. The inefficiency of magnetic separation for copper sludge obvious due to the lack of any significant what about the number of magnetic phases.
Widely known used in industry is a method of low-temperature liquid-phase sulfatization sludge, whereby the wet sludge pour concentrated sulfuric acid and maintained at a temperature of 120-180°C for 3-8 hours. After cooling the slurry of sulfatization leached with water and filtered. The method provides high recovery in the solution of non-ferrous metals and partial - arsenic in almost complete concentration of precious metals, of rare halogeno, as well as antimony and lead in the insoluble residue.
The disadvantage of this method is the low degree of enrichment of the sludge.
There's also a way of enriching slurry electrolysis Nickel and other products containing platinum metals, gold and silver (RF patent No. 2276195). According to the method, the slurry is subjected to pressure oxidation leaching in sulfuric acid at a temperature 108-110°C and a partial pressure of oxygen of 0.4-1.0 MPa, and then platinuum.
Prior to flotation, oxidation of low-temperature pressure leaching allows you to selectively translate into a solution of non-ferrous metals (except refractory oxides and ferrites Nickel), and a significant portion of sulfur, as well as to prepare slurry to flotation.
The disadvantage of this method is that it cannot be applied to PE is erbody lead medeelektrolitnyj slimes enriched in antimony and arsenic, as it does not take into account the peculiarities of their phase composition and, accordingly, does not provide for the individualization of grains phases of precious metals and activation of their surface the depth of the processes of the autoclave leaching. Conditions of autoclave leaching are necessary, but not sufficient to obtain floatated slurries.
The closest technical solution to the claimed method adopted in the closest prototype of the invention is a method of processing medeelektrolitnyj sludge implemented on two steel mills in China (meretukov M.A., Orlov A. M. metallurgy of noble metals. Foreign experience, M, metallurgy, 1991, s-243).
The method consists in the fact that the sludge is leached with diluted sulfuric acid at a temperature of 80°C in the presence of sodium chlorate, then treated copper scrap and activated carbon with the aim of identifying gold, passed into the solution during leaching, and filtered. To translate the silver in metallic form obtained cake is treated with iron powder. Restored silver platinuum in an acidic medium with the addition of iron powder, butylamino Aero-wlota and butyl xanthate as reagents.
The obtained concentrate coderit,9% silver, of 0.182% gold, 1.92% of selenium, 0,87% copper and 2,78% lead. The extraction of gold and silver in overall operations leaching of flotation is 99,04 and 99.2 percent, respectively.
Disadvantages closest analogue (prototype) is the possibility of formation of difficult recoverable selenate ion in the leaching process, and the need for clear dosing copper scrap to prevent secondary deposition at KEK significant amounts of selenium. In addition, sodium chlorate is an expensive reagent, special requirements for storage conditions and usage because of his explosiveness.
The present invention is to develop a method for processing sludge of the electrorefining of copper, providing a high degree of enrichment.
The technical result achieved by the invention is to obtain a rich collective concentrate, suitable for the production of precious metals and selenium by any known method, and product containing lead, antimony and arsenic, suitable for processing in lead production. The application of this method to lead copper sludge reduces the mass of material and to improve the content of precious metals 3.6-4.8 times when direct extraction sums of gold and silver in concentrate to 99.8%.
On the first Varian is the technical result is achieved by a method for enrichment of lead sludge of the electrorefining of copper, lead, antimony, arsenic, gold and silver, sequential operations autoclave oxidative leaching and flotation according to the invention, the leaching is carried out in an autoclave at a temperature 104-106°C using as oxidant oxygen partial pressure of 0.02-0.1 MPa, with stirring, until a homogeneous suspension and intensity to ensure the absorption of oxygen is not less than 0.001 mol O2/m3·h·PA.
And the other main variant technical result is achieved in that in the method of enrichment of lead sludge of the electrorefining of copper, lead, antimony, arsenic, gold and silver, including sequential operations liquid-phase sulfatization, leaching and flotation, according to the invention before the leaching is carried out liquid-phase sulfatization at a temperature of 160-200°C and leaching of lead sulphate solution iron (II)containing iron in an amount necessary to restore present in the liquid portion of the pulp sulfatization antimony, arsenic and selenium.
Prior to flotation autoclave leaching sludge provides training material to flotation - surface activation chalcogenides, the individual is the realizations of the phases of precious metals and impurity elements through a series-parallel processes 1-5, moreover, the contribution of each of them is determined by the modes leaching:
1) dissolving a matrix component and oxide films from the surface of chalcogenides with translation in the liquid part of the slurry of antimony, arsenic, silver, and other items that are part of them;
2) leaching of copper and tellurium from selenaselena copper-silver;
3) exchange interaction of silver with a non-stoichiometric solenoceridae copper-silver;
4) the ordering of the crystal lattice of selenaselena copper-silver and compounds forming matrix component and a film on the surface of the nuclei under the action of high temperatures;
5) oxidation passed into the liquid part of the slurry of antimony and arsenic to the highest oxidation States and their secondary deposition in the cake in the form of mixed oxides of antimony-arsenic-and antimony-lead.
In conditions where oxygen transport is not limiting stage, the oxidation of components of the sludge (processes 2, 5) ahead of the processes of ordering, lattices (process 4). This achieves an extremely deep obezbedjivanje sludge (copper content in the insoluble residue of less than 0.05%) and high rates of subsequent flotation separation, determined by the depth of oxidation of antimony and arsenic phases.
When implementing autoclave leaching under conditions of scarcity, color is Yes, the role of the processes of ordering, lattices becomes very significant. Pestiviruses film on the surface of the chalcogenide and the matrix component become more resistant with respect to the liquid portion of the pulp, which leads to a decrease (down to zero) flotation activity slurries.
Reducing the temperature of the leaching slows down all processes during pressure oxidation leaching. However, the kinetics of chemical reactions (processes 2, 5) is more influenced by this factor than the rate of ordering of crystal lattices (process 4).
Thus, obtaining floatated slurries in autoclave leaching is possible only with a substantial excess of the rate of oxidation (processes 2, 5) over speed the ordering of crystal lattices (process 4) components of the sludge, which is achieved by maintaining the operating parameters in the specified range. The output values of oxygen partial pressure, the rate of absorption of oxygen and temperature at the lower boundary of the range (0.02 MPa, 0.001 mol O2/m3·h·PA and 104°C, respectively) leads to the change of the ratio of the above speeds and obtaining relatirely slurries. Increasing the partial pressure of oxygen and temperature in excess of 0.15 MPa and 106°C, respectively, leads to increased losses of selenium with leaching solutions, but no significant effect on pokazatel the flotation.
If sulfatization at temperatures above 160°C occur oxidation and dissolution of a large number of components of the sludge: antimony and arsenic into solution in higher oxidation States; chalcogenides of copper, copper-silver and silver oxidized with the formation of sulphate of copper, sulphate of silver and articlesbysubject (Se8(HSO4)2).
When the leaching pulp sulfatization water octacarbonyl dis-proportionum on se acid and elemental selenium, which interacts with silver and forms on the surface of the lead sulfate precipitate secondary silver selenide (figure 3, reaction 1-2). When the hydrolysis of the sulfate of antimony in solution and on the surface of solid particles (including chalcogenides), is formed precipitate bulk flotation activity of particles
When the leaching pulp sulfatization containing octacarbonyl and bisulfate silver sulphate solution iron (II) is immediate formation of silver selenide in the form of an individual right crystals cubic shape (figure 4, reaction 3). Antimony and arsenic are restored and quickly and completely precipitate in the form of spheroidal grains, and the formation of sediment is not on the surface of solid particles, as nab is udaetsya in the leaching water, and in solution
Lowering the temperature of sulfatization below 160°C does not give a positive effect, because the oxidation process of selenaselena copper-silver sludge to ekstsentrisiteta is suppressed, the temperature rise more than 200°C is not expected to have a positive impact on process performance, however, will lead to increased energy consumption.
Next, the results of laboratory and ukrupnennom-laboratory experiments.
The method according to paragraph 1
Example 1. A portion of the copper slurry weighing 150 g of composition, %: 11,5 Ag, 4,9 Se, 15,3 Pb, 7,3 Sb, 15,9 C videlacele in the autoclave 1 DM3with the mixing device, providing the speed of absorption of oxygen 0,21 molló2/m3·h·PA, at a temperature of 104-106°C and a partial pressure of oxygen of 0.1 MPa. The obtained cake was fluoroware in a closed loop with two cleanings concentrate and control the flotation tails.
On set operations autoclave oxidative leaching of flotation sludge enriched 4.2 times. Flotation concentrate contains 51,0% silver, 18% selenium, 9.1% of lead, 6.1% of antimony. Tails containing 0.2% silver. Direct extraction of silver in concentrate 99.5%, tails removed 85,1% lead and 79.3 percent of antimony; removing selenium in the solution of ~ 1% of the content in the initial sludge.
Example 2. A portion of the copper sludge mass 4.2 kg SOS is ava %: 12,4 Ag, 5,1 Se, 15,3 Pb, Sb 7,5, 17,6 C videlacele in an autoclave with a capacity of 25 DM3with the mixing device, providing the speed of absorption of oxygen 0,017 molló2/m3·h·PA, at a temperature of 104-106°C, partial pressure of oxygen of 0.1 MPa. The obtained cake was fluoroware in a closed loop with three cleanings concentrate and control the flotation tails.
On set operations autoclave oxidative leaching of flotation sludge enriched 3.6 times. Flotation concentrate contains 45,0% silver, 16.2% of selenium, 13.7 percent lead, 6.9% of antimony. The tails contain 0.5% silver. Direct extraction of silver in concentrate is 98.9 per cent; in the tails removed 73,5% lead and 73,8% antimony; removing selenium in the solution of ~ 1% of the content in the initial sludge.
The method according to paragraph 2
Example 3. A portion of the copper slurry weighing 150 g of the composition of 11.5 Ag, 4,9 Se, 15,3 Pb, 7,3 Sb, 15,9 C was subjected to liquid-phase sulfatization at a temperature of 180°C, then videlacele in a solution of sulphate of iron (II)containing 27,2 g of iron at a temperature of ~100°C. the resulting cake was fluoroware in a closed loop with three cleanings concentrate and control the flotation tails.
On the set of operations of the liquid-phase sulfatization - leaching in a solution of sulphate of iron (II) - flotation sludge enriched 4.8 times. Flotation concentrate contains 55% silver, 8% lead, 2.7% antimony. The tails contain 0.5% with the ribs. Direct extraction of silver in concentrate is 98,7%, tails removed 89% of lead and 78.8% antimony; removing selenium in solution <0.1% of the content in the initial sludge.
1. A method of enrichment of lead sludge of the electrorefining of copper, lead, antimony, arsenic, gold and silver, including sequential operations leaching and flotation, wherein the leaching is carried out in an autoclave at a temperature 104-106°C, using as oxidant oxygen partial pressure of 0.02-0.1 MPa, with stirring, until a homogeneous suspension in intensity to ensure the absorption of oxygen is not less than 0.001 mol O3/m3·h·PA.
2. A method of enrichment of lead sludge of the electrorefining of copper, lead, antimony, arsenic, gold and silver, including sequential operations leaching and flotation, characterized in that before the leaching is carried out liquid-phase sulfatization at a temperature of 160-200°C, and the leaching of lead sulphate solution iron (II)containing iron in an amount necessary to restore present in the liquid portion of the pulp sulfatization antimony, arsenic and selenium.
SUBSTANCE: method includes preparation of a nepheline-lime-soda charge, its sintering in a tubular rotary furnace by heat released when burning fossil coal. After sintering, leaching, desiliconisation and carbonisation of an aluminate solution is carried out to produce alumina and soda products. The fossil coal to burn is a brown coal, the solid residue of which contains calcium oxide CaO of at least 30 wt %, and silicon oxide SiO2 of not more than 40 wt %. Brown coal from the Kansko-Achinskiy field is burnt.
EFFECT: reduced consumption of lime in charge preparation and lower content of silicon oxide in an aluminate solution, using a less scarce fossil coal as fuel.
2 cl, 2 tbl, 1 ex
SUBSTANCE: method of sulphide stock containing noble metals comprises mixing stock with water solution of reagents and autoclave oxidising treatment by water solution of reagents on feeding oxygen and adding component with halogenide-ion to produce pulp. Then, pulp is divided into solution and solid residue. Note here that autoclave oxidising treatment is carried out by water solution containing component with halogenide-ion at 160-250°C and oxygen partial pressure of 0.5-5.0 MPa. Extraction of noble metals is carried out by leaching from solid residue by sulfite-sulfate solutions.
EFFECT: reduced number of processes, lower costs.
11 cl, 3 tbl, 1 ex
SUBSTANCE: proposed method consists in valuable metals are decomposed in salt melt containing 60-95 wt % of NaOH and 5-40 wt % of Na2SO4. Then, melt decomposition product is converted into solid phase by cooling to room temperature. After cooling, minced melt decomposition product is converted in water at temperature lower than 80°C to produce water suspension and water fraction is separated by filtration for components to be extracted therefrom.
EFFECT: higher efficiency of extraction.
22 cl, 1 dwg, 3 tbl
SUBSTANCE: manganese dioxide obtaining method involves dilution of manganese-containing raw material in nitric acid so that solution of manganese nitrates and nitrates of calcium, potassium, magnesium and sodium impurities contained in the ore is obtained. Then, thermal decomposition of nitrates in autoclave is performed. Thermal decomposition is performed at constant pressure drop in autoclave, starting from pressure of 0.6 MPa and reducing it to the end of the process to 0.15 MPa. At that, pulp is constantly mixed at thermal decomposition with the mixer rotating at speed of 1-15 rpm and with superimposition of vibration on it with frequency of 20-50 Hz. Method can be implemented at chemical plants provided with pressure autoclaves.
EFFECT: obtaining manganese dioxide of improved quality.
2 tbl, 2 ex
SUBSTANCE: invention relates to a method for treatment of low-grade oxidized zinc ores and concentrates with zinc, manganese, iron, lead, silver, calcium and silicon dioxide recovery. The method comprises crushing, fine crushing, lixiviating, settling-down of the above components from solutions, wherein the lixiviating shall be performed in stages: during the first and the second stage - lixiviation by sulphuric acid solution over the reducing substance, the oxidation-reduction potential φ (ORP): at the first stage φ=380-420 mV, at the second stage φ=420-460 mV. At the third stage lixiviation is performed by an ammonium carbonate solution with the composition (g/dm3): 60-110 NH3 general 30-65 C02 general while at the fourth stage it is performed by nitric acid NHO3. At the fifth stage sulphating roasting with H2SO4 term with further water leaching is performed, the sixth stage being lixiviation with ammonium fluoride aqueous solution of acid NH4HF2. The obtained soutions shall be puriied and used for zinc, manganese, iron, lead, silver, calcium, silicon dioxide and sodium sulfate, ammonium nitrate recovery. EFFECT: enhanced zinc recovery and rational utilization of the stock.
16 cl, 1 dwg, 3 ex
SUBSTANCE: method involves grinding ore, mixing the ground ore with sodium bisulphate taken in stoichiometric amount required for binding manganese and impurities into sulphates. The mixture is calcined in three steps to obtain coal tar: at the first step at temperature 200-300°C for 1-2 hours, at the second step at temperature 400-500°C for 0.5-1.5 hours, at the third step at temperature 600-700°C for 2-4 hours. The coal tar is leached with water at temperature 40-80°C for 0.5-1 hour and weight ratio coal tar: water equal to 1:(3-4). After filtering the obtained pulp, sludge is separated and the filtrate is treated with sodium carbonate solution taken in stoichiometric amount required for binding and depositing manganese (II) and iron (II) compounds. After filtering the obtained suspension, the precipitate of manganese (II) and iron (II) carbonates is dried and washed to obtain a manganese concentrate. Calcination exhaust gases are absorbed with the filtrate from the manganese carbonate extraction step to obtain sodium bisulphate solution. Through evaporation of the obtained solution, crystallisation and drying, sodium bisulphate is obtained, which is taken for mixing with the ground ore to obtain the mixture.
EFFECT: simple process and zero-discharge scheme of the process of processing manganese ore.
2 cl, 1 dwg, 2 ex
SUBSTANCE: method involves leaching alkali and alkali-earth metals with a solution of a chlorine-containing reagent and separating the insoluble residue containing manganese dioxide. The ore undergoes preliminary decarboxylation via thermal treatment at temperature 750-1000°C for 2-4 hours to obtain coal tar. The chlorine-containing reagent used when leaching the coal tar is 10-40% aqueous ammonium chloride solution, taken in weight ratio ore: ammonium chloride equal to 1:1-2. Leaching is carried out at temperature 20-100°C for 1-2 hours. After separating the insoluble residue, the filtrate is carbonised with exhaust gases from the ore decarboxylation step, followed by separation of the obtained calcium carbonate and return of the aqueous ammonium chloride solution to the coal tar leaching step.
EFFECT: obtaining quality end product - manganese dioxide concentrate and a by-product - calcium carbonate using a zero-discharge process scheme.
1 dwg, 1 tbl, 1 ex
SUBSTANCE: concentrate is subjected to two-stage oxidising roasting. Note here that, before first stage, said concentrate is mixed with sulfur-binding additive to perform first roasting stage at 550-650°C for 15-30 min. Prior to second roasting stage, molybdenum concentrate is added to calcine produced at first stage in amount of 10-30 wt % from concentrate used in first roasting step. Second roasting is performed at 600-670°C for 30-40 min with subsequent leaching of molybdenum and rhenium from calcine obtained in first step.
EFFECT: higher yield of molybdenum.
2 tbl, 2 ex
SUBSTANCE: proposed method comprises irradiating ores by SHF-field and processing them by acid and/or oxidiser solution to transfer noble metals into solution. Prior to irradiation by SHF-field initial material is subjected to fractionation in upflow with variable hydrodynamic conditions at linear speed of said upflow of 10-50 m/h to produce concentrated fraction. Concentrated fraction is subjected to said irradiation. Note here that irradiation is executed in microwave range at load that allows heating the materials to 180-280°C. Then, noble metals are leashed into solution.
EFFECT: higher yield of noble metals.
1 tbl, 2 ex
SUBSTANCE: method for industrial production of pure MgCO3 involves crushing olivine-containing rock and bringing the crushed rock into contact with water and CO2. At the first step, which is carried out under pressure, a dissolution reaction takes place according to the equation Mg2SiO4(s)+4H+=2Mg2++SiO2(aq)+2H2O. At the second step, deposition is carried out at higher pH. The following reactions take place: Mg2++HCO3 -=MgCO3(s)+H+ and Mg2++CO3 2-=MgCO3(s). The presence of HCO3 - and H+ ions is mainly a result of reaction of CO2 and water.
EFFECT: invention enables to produce pure magnesium carbonate from rock while binding free carbon dioxide gas.
19 cl, 4 dwg, 2 tbl, 2 ex
SUBSTANCE: electrolyser consists of electrolysis bath, of anode, cathode and of collector of cathode lead. Also, electrolysis bath is made out of heat resistant concrete. The anode is installed in a recess of the bath and is made in form of a graphite bottom with a steel current conductor, while cathode corresponds to two cylinders out of graphite. Chutes for drain of cathode lead into collectors are arranged directly under the graphite cylinders.
EFFECT: raised efficiency of electrolyser, reduced power expenditures and labour input, raised reliability of electrolyser operation.
6 cl, 1 tbl, 2 dwg
SUBSTANCE: procedure consists in preliminary crumbling slag, in its mechanical purification from compounds of iron and in skimming oxide films of brass off it. There is prepared solution containing solution of ammonia salt and surface active substance; solution is then heated to temperature close to boiling temperature. Further there is prepared water suspension by supplying slag into prepared solution with continuous mixing and maintaining said temperature; produced suspension is heated above 101°C at continuous mixing during 15-22 min to purify particles of brass from oxides. Fine dispersed brass is separated; left solution is separated from zinc oxide and copper oxide by mixing during 35-45 min and cooled; by filtering copper oxide and crystallisation of zinc oxide they are separated in cooled solution; the solution is washed with cold flushing water of temperature below 10-24°C to stop chemical reactions. Further, undesirable impurities are turned into solution by washing with hot flushing water at temperature close to boiling; and zinc oxide is separated by filtering.
EFFECT: continuous waste-less production of brass, zinc oxide and copper oxide at reduced expenditures.
5 cl, 1 dwg
SUBSTANCE: complex includes bay of slag preparation successively consisting of furnace, crusher, magnetic separator, vibration screen, and accumulator of processed slag. Also, the accumulator of processed slag is connected to a mixer-separator of the bay for preparation of zinc containing stock, an input of which is parallel connected with industrial water, chemicals, return water, return solution and a heat source, while an output is connected to the bay for preparation of fine dispersed brass to re-melt. The bay for preparation of fine dispersed brass consists of a drier of brass and of moulding device connected to the furnace. The bay for separation of zinc oxide and copper oxide consists of the mixer whereto chemicals are introduced and of filter for separation to copper oxide and to solution containing zinc oxide. The filter of separation to copper oxide and to solution containing zinc oxide is successively connected to a crystalliser, to the filter of separation of zinc oxide from solution returned to the mixer-separator of the bay for preparing zinc-containing stock, to a mixer-separator of washing, to the filter of separation of zinc oxide from water coming to rotation into the mixer of the bay for preparation of zinc containing stock.
EFFECT: non-waste processing industrial waste producing brass, zinc oxide and copper oxide at minimal number of process operations and minimal consumption of power.
SUBSTANCE: procedure consists in bromination of tetra-methyl-lead with solution of bromine in tetra-chloride of carbon at ratio 1:3 and in distilled purification of lead bromide (II) at temperature 750-850°C in flow of hydrogen or inert gas. Further, bromide of lead (II) is reduced to metal with 20 % water solution of tetra-hydro-borate of potassium and is melted in flow of hydrogen at temperature 600-650°C.
EFFECT: high output of metal lead.
SUBSTANCE: method involves processing by using ammonium chloride, obtaining of chlorides solution, extraction of zinc from the solution. At that, sulphide, oxidated or mixed zinc-lead containing ores are subject to processing by using solid ammonium chloride in quantity of 100-130% of stoichiometric ratio at temperature of 200-320°C. Solution of chlorides is obtained by water leaching of the obtained mixture of chlorides. Iron is separated in the form of hydroxide from the obtained solution at pH 4. After iron is separated, zinc is extracted in the form of zinc hydroxide by adding ammonium to the solution to pH 7. Obtained hydroxides are calcinated till oxides are obtained. Lead is extracted from residual water leaching in the form of chloride by leaching with sodium chloride solution with concentration of 300-320 g/l at temperature of 70-95°C.
EFFECT: possibility of developing effective and environmentally safe technology for processing of sulphide, oxidated or mixed zinc-lead containing ores.
SUBSTANCE: invention refers to procedures for extracting valuable metals from wastes including wastes of refining production. The procedure for processing silver containing lead wastes for extracting silver and lead in form of products consists in leaching wastes with solution containing hydrochloric acid to transit lead into solution and silver to sedimentation. Also leaching is carried out in three stages. At the first stage leaching is peformed with solution of lead chloride (PbCl2) at temperature 0°-25°C in concentrated hydrochloric acid (HCl) by placing wastes in this solution heated to temperature 40°-80°C. Major part of lead transits into solution. During the second and the third stages non-dissolved wastes are leached in 30-42% HCl and successively filtered till pure powder-like silver is produced in a non-soluble residue. Upon the first stage solution is cooled to 0°-25°C, crystals of PbCl2 are extracted out of it and combined with solutions of HCl after leaching at the first and the third stages; a new portion of wastes is directed to leaching. Implementation of this procedure at a lead plant of average output can collect yearly income of approximately 100 million dollars from sale of lead chloride.
EFFECT: raised efficiency of wastes processing, additional economic profit from implementation of this procedure at various metallurgical productions.
4 tbl, 2 ex
SUBSTANCE: method of extracting lead (II) ions (Pb2+) from acidic solutions involves sorption of Pb2+ ions by bringing the solution into contact with an anionite. Sorption is carried out at 70-80° C from solutions containing 80-120 g/l hydrochloric acid and chlorides of ammonia, alkali or alkali-earth metals on an AMP anionite which contains exchange groups or on an AM-2b anionite which contains exchange groups pre-treated with distilled water.
EFFECT: finding optimum conditions for efficient sorption of lead ions.
2 dwg, 4 tbl, 4 ex
SUBSTANCE: method includes sulphur removing in two stages. At the first stage lead sulfate from active mass is put into contact with Na2CO3 in solution, receiving dispersion, containing carbonised active mass on the basis of main lead carbonates. At the second stage this dispersion is put into interaction with CO2 with formation of dispersion, containing sweet active mass on the basis of PbCO3. Between these two stages it is implemented granulometric division with following sulphur removing of coarse grain.
EFFECT: effective sulphur removing with almost total removing of sodium during sulphur removing.
28 cl, 3 dwg, 6 ex
SUBSTANCE: concentrate is grinned up to size not more than 0.32 mm and treated by mixture 18-35% of nitrogenous and 40-45% hydrofluoric acids at mass correlation of acids 2.7-5.3:1, "Т:Ж"=1:3-5 and temeprature 90-120°C with formation of pulp. Nitro-acid lead (II), undecomposed minerals and radionuclides are transferred into sedimentation, and tantalum and niobium - into fluoride - nitro-acid solution. Pulp is cooled up to 10-20°C, lead-bearing deposit is separated and washed it by 35-50% nitric acid at "Т:Ж"=1:1.5-2.0. Scrub solution is affixed to fluoride - nitro-acid solution. From washed sediment it is leached nitro-acid lead (II) by water at correlation "Т:Ж"=1:5-10 and temperature 50-60°C with separation of undecomposed minerals and radionuclides and receiving of nitro-acid lead solution. From combined fluoride - nitro-acid solution it is implemented extraction of tantalum and niobium by neutral extragent. In the capacity of neutral extragent it is used triiso-amyl-phosphate or octanol-1. Extraction is implemented on 2-6 stages at O:B=1.5-2.0:1 with transferring of tantalum into organic phase and niobium - into aqueous. Extraction by triiso-amyl-phosphate is implemented at 2-4 stages, and extraction by octanol-1 - at 5-6 stages.
EFFECT: increasing of extraction ratio of tantalum and niobium into solution and reduction of operation amount at lead exudation in the form of well-soluted salt of nitro-acid lead.
8 cl, 5 ex
SUBSTANCE: invention relates to ferrous metallurgy, mainly to devices for reprocessing of powdered lead- and zinc-containing raw materials, in which there can be copper and precious metals. Aggregate for reprocessing of powdered lead- and zinc-containing raw materials contains rectangular upright smelting chamber with burner facility, gas cooler stack, partition with water-cooled copper elements, separating smelting chamber from gas cooler stack, electric furnace, separated from the smelting chamber by partition with water-cooled copper elements, coffer chord, facilities for discharge of smelting products, bottom, herewith correlation of difference of level of bottom edges to distance from the smelting chamber crown to bottom edge of partition, separating electric furnace from the smelting chamber, is 0.15-0.29, and relation of distance from bottom edge of this partition up to bottom to difference of level of bottom edges is 1.25-2.10. On walls of gas cooler stack of aggregate there are installed not more than two tuyers on the level of bottom edge of partition, separating gas cooler stack from smelting chamber, with inclination into the side of bottom on-the-mitre to horizontal plane, specified by formula α=arctg(k-ΔH/B), where α - angle of tuyers slope; k - coefficient of angle of tuyers slope, equal to 1.11-1.25; ΔH - difference of level of bottom edges of partitions; B - inside width of gas cooler stack. At mounting of two tuyers they are located by one on each of opposite side walls of gas cooler stack with reflector displacement relative to its cross-axial section. Additionally each of it is located at a distance of cross-axial section of gas cooler stack, relation of which to inside length of gas cooler stack is 0.25-0.30.
EFFECT: it is provided simultaneous increasing of direct lead extraction into crude metal and unit specific capacity.
3 cl, 4 dwg, 2 tbl, 16 ex
SUBSTANCE: method of determining noble metals involves drying a sample with grain size 1 mm to constant weight at temperature 105-110°C and using the dried sample for second and subsequent one-off determination of noble metals. An undried sample is used during the first one-off determination, wherein sample material is mixed with a charge mixture, the obtained mixture is molten and the amount of noble metals in the melt is determined. The sample is dried during the first determination and the weight ratio of moisture in the sample is determined. Content of noble metals in the sample is determined using the formula: where Cme is content of noble metals in the sample, g/t, Mme is mass of the noble metal detected in the melt, mg, m1 is the mass of the sample material used in the first determination, g, W is the weight ratio of moisture in the sample.
EFFECT: invention increases rapidness of determining noble metals and labour efficiency.