CN110295285B - Method for recovering zinc from oxygen-enriched solid sulfur reduction smelting slag - Google Patents

Method for recovering zinc from oxygen-enriched solid sulfur reduction smelting slag Download PDF

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CN110295285B
CN110295285B CN201910713362.8A CN201910713362A CN110295285B CN 110295285 B CN110295285 B CN 110295285B CN 201910713362 A CN201910713362 A CN 201910713362A CN 110295285 B CN110295285 B CN 110295285B
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zinc
leaching
slag
acid
indium
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CN110295285A (en
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巫文嵩
陈巍
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Sichuan Zhengxiang Environmental Protection Technology Co ltd
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    • CCHEMISTRY; METALLURGY
    • C22METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22BPRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B17/00Obtaining cadmium
    • C22B17/04Obtaining cadmium by wet processes
    • CCHEMISTRY; METALLURGY
    • C22METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22BPRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B19/00Obtaining zinc or zinc oxide
    • C22B19/20Obtaining zinc otherwise than by distilling
    • C22B19/22Obtaining zinc otherwise than by distilling with leaching with acids
    • CCHEMISTRY; METALLURGY
    • C22METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22BPRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B19/00Obtaining zinc or zinc oxide
    • C22B19/20Obtaining zinc otherwise than by distilling
    • C22B19/26Refining solutions containing zinc values, e.g. obtained by leaching zinc ores
    • CCHEMISTRY; METALLURGY
    • C22METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22BPRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B19/00Obtaining zinc or zinc oxide
    • C22B19/30Obtaining zinc or zinc oxide from metallic residues or scraps
    • CCHEMISTRY; METALLURGY
    • C22METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22BPRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B41/00Obtaining germanium
    • CCHEMISTRY; METALLURGY
    • C22METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22BPRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B58/00Obtaining gallium or indium
    • CCHEMISTRY; METALLURGY
    • C22METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22BPRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B7/00Working up raw materials other than ores, e.g. scrap, to produce non-ferrous metals and compounds thereof; Methods of a general interest or applied to the winning of more than two metals
    • CCHEMISTRY; METALLURGY
    • C22METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22BPRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B7/00Working up raw materials other than ores, e.g. scrap, to produce non-ferrous metals and compounds thereof; Methods of a general interest or applied to the winning of more than two metals
    • C22B7/001Dry processes
    • CCHEMISTRY; METALLURGY
    • C22METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22BPRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B7/00Working up raw materials other than ores, e.g. scrap, to produce non-ferrous metals and compounds thereof; Methods of a general interest or applied to the winning of more than two metals
    • C22B7/04Working-up slag
    • CCHEMISTRY; METALLURGY
    • C25ELECTROLYTIC OR ELECTROPHORETIC PROCESSES; APPARATUS THEREFOR
    • C25CPROCESSES FOR THE ELECTROLYTIC PRODUCTION, RECOVERY OR REFINING OF METALS; APPARATUS THEREFOR
    • C25C1/00Electrolytic production, recovery or refining of metals by electrolysis of solutions
    • C25C1/16Electrolytic production, recovery or refining of metals by electrolysis of solutions of zinc, cadmium or mercury
    • YGENERAL TAGGING OF NEW TECHNOLOGICAL DEVELOPMENTS; GENERAL TAGGING OF CROSS-SECTIONAL TECHNOLOGIES SPANNING OVER SEVERAL SECTIONS OF THE IPC; TECHNICAL SUBJECTS COVERED BY FORMER USPC CROSS-REFERENCE ART COLLECTIONS [XRACs] AND DIGESTS
    • Y02TECHNOLOGIES OR APPLICATIONS FOR MITIGATION OR ADAPTATION AGAINST CLIMATE CHANGE
    • Y02PCLIMATE CHANGE MITIGATION TECHNOLOGIES IN THE PRODUCTION OR PROCESSING OF GOODS
    • Y02P10/00Technologies related to metal processing
    • Y02P10/20Recycling

Abstract

The invention discloses a method for recovering zinc from oxygen-enriched solid sulfur reduction smelting slag, and belongs to the technical field of solid hazardous waste treatment. According to the method, zinc is extracted from the oxygen-enriched sulfur-fixing reduction smelting slag containing lead, zinc waste residues or lead plaster by a pyrogenic-wet method, the slag obtained after the lead, zinc waste residues or lead plaster is subjected to blank making and oxygen-enriched sulfur-fixing reduction smelting is fully recycled, the existing resources are further recovered, the pollution of valuable metals to the environment is avoided, and the method is safer and more environment-friendly; meanwhile, the method has the advantages of simple principle, reasonable flow, high zinc recovery rate and low cost.

Description

Method for recovering zinc from oxygen-enriched solid sulfur reduction smelting slag
Technical Field
The invention belongs to the technical field of solid hazardous waste treatment, and particularly relates to a method for recovering zinc from oxygen-enriched solid sulfur reduction smelting furnace slag.
Background
Due to the development of the existing metallurgical industry, a large amount of lead and zinc-containing waste residues are generated, and the waste residues contain a large amount of valuable metals, such as lead, zinc and the like, but are often ignored by enterprises. If the valuable metals can be recycled for the second time, the harm to the environment is avoided, and the resources are effectively utilized without waste.
In the process of recycling and disposing the waste lead-acid storage batteries, the waste lead-acid storage batteries are mainly divided into lead sulfate paste, lead grids, plastic particles, waste acid and the like after being disassembled, wherein the lead sulfate paste contains valuable metals such as lead, zinc, tin, antimony and the like and also contains a large amount of harmful sulfate, and the lead sulfate paste is dangerous metal solid waste. Lead-acid storage batteries are used as main consumer products all over the world, so that the recovery of lead and zinc containing waste residues and zinc in lead paste of waste lead-acid storage batteries is of great significance.
The method for smelting lead sulfate paste and lead and zinc-containing waste slag by briquetting and sulfur fixation reduction smelting is a method applied by many enterprises at present, and various products can be produced by the method, wherein the zinc content in the slag product is outstanding, but the operation of extracting zinc from slag has large difference in yield and high price due to different processes, and low-content valuable metals cannot be effectively recovered.
Disclosure of Invention
Aiming at the existing problems, the invention provides a method for recovering zinc from oxygen-enriched sulfur-fixing reduction smelting slag, which fully recycles the slag after the lead-containing, zinc-containing waste slag or lead plaster is made into briquettes and sulfur-fixing reduction smelting, not only further recovers the existing resources, but also avoids the pollution of valuable metals to the environment, and is safer and more environment-friendly; meanwhile, the method has the advantages of simple principle, reasonable flow, high zinc recovery rate and low cost.
The technical scheme adopted by the invention is as follows:
a method for recovering zinc from oxygen-enriched solid sulfur reduction smelting slag comprises the following steps:
s1, material preparation and blank making: mixing lead-zinc-containing waste residues or lead plaster, a sulfur-fixing agent and a flux to obtain material powder, wherein the water content of the material powder is 12-15%; briquetting and blank making are carried out on the material powder under the pressure of 30-50 MPa, and then the material powder is dried until the water content is 5% -6% and the Pb content is 18% -28%, so as to obtain a blank block;
s2, sulfur fixation reduction smelting: adding the briquettes and coke to 7.8m2Carrying out slagging reaction in the oxygen-enriched sulfur-fixing reduction smelting furnace to obtain furnace slag;
s3, converting in a fuming furnace: mixing the slag with low-grade waste slag containing lead and zinc to obtain a mixture, adding the mixture into a fuming furnace for converting, and obtaining secondary zinc oxide smoke dust at 1250-1300 ℃; feeding the secondary zinc oxide smoke dust into a rotary kiln for roasting to obtain secondary zinc oxide roasted sand;
s4, neutral leaching: adding zinc hypoxide calcine into mixed waste acid liquor to carry out neutral leaching with a liquid-solid ratio of 7-8: 1, adding an oxidant to carry out solid-liquid separation when the pH value reaches 5-5.2, and obtaining neutral leaching residue and neutral leaching liquor;
s5, a three-stage purification process of neutral leachate, wherein the first stage comprises the following steps: adding zinc powder into the neutral leaching solution at 50-55 ℃, and filtering by a filter press to obtain copper-cadmium-containing slag and filtrate 1; and a second stage: heating the filtrate 1 to 80-90 ℃ by using steam, adding antimonate and zinc powder to remove impurities, and filtering by using a filter press to obtain cobalt-nickel-containing slag and filtrate 2; a third stage: cooling the filtrate 2 to below 70 ℃, adding zinc powder to remove Cd, and filtering by a filter press to obtain dregs and a filtrate 3; mixing the filtrate 3 and the waste electrolyte according to the volume ratio of 1: 15-20, adding the mixture into an electrodeposition tank for electrodeposition to obtain cathode zinc and waste liquid after electrodeposition is finished, and stripping the cathode zinc to obtain a zinc precipitation sheet; melting and separating out zinc sheets by adopting a power frequency induction furnace, adding a clarifying agent, maintaining the temperature of zinc liquid in the furnace at 470-490 ℃, and casting into zinc ingots;
s6, indium and germanium enrichment and recovery: preparing a mixed acid solution, adding the mixed acid solution and neutral leaching residues into a leaching tank, performing solid-liquid separation after leaching to obtain acidic leaching residues and an acidic leaching solution, delivering the acidic leaching residues to an oxygen-enriched sulfur-fixation reduction smelting furnace for smelting, and delivering the acidic leaching solution into an indium-germanium enrichment tank; when the temperature of the acid leaching solution is 75-85 ℃, Zn powder is used as a displacer for displacement, then filtration is carried out to obtain indium-germanium-enriched slag, and the indium-germanium-enriched slag is subjected to acid leaching and separation to obtain acid leaching slag and acid leaching solution; extracting the pickle liquor to obtain an indium-rich organic phase and a germanium-containing raffinate, performing hydrochloric acid back extraction on the indium-rich organic phase, and adding a zinc plate into the indium-containing liquor for replacement to obtain sponge indium; pressing the sponge indium into a briquette, and then casting a crude indium ingot; washing the germanium-containing raffinate with water, concentrating, and adding oxalic acid to precipitate germanium to obtain germanium-containing concentrate;
s7, the recovery operation of the copper-containing cadmium slag comprises the following steps: and recovering cadmium by adopting sulfuric acid leaching, zinc powder replacement and sponge cadmium rectification processes.
Specifically, the low-grade waste residue containing lead and zinc mainly comprises the following components: pb 20-28% and Zn 5-8%.
Specifically, the mixed waste acid solution in the step S4 is a mixed solution of battery regenerated sulfuric acid, zinc electrolyte and washing water of the waste lead acid storage battery, the initial acid concentration of the mixed waste acid solution is 70-120 g/L, the oxidant is manganese dioxide, and the addition amount of the mixed waste acid solution is 1.2 times of that of bivalent iron in the solution; the washing water is slag washing water of each procedure.
Further, the mass ratio of the lead-zinc-containing waste residue or lead paste, the sulfur-fixing agent and the flux in the step S1 is as follows: 100: 5-8: 5 to 10.
Further, the briquette-to-coke mass ratio in S2 is 100: 9 to 12.
Further, the sulfur-fixing reduction smelting conditions are as follows: the coke rate is 9-12%, and the blast intensity is 35-45 m3/min·m215-18 kpa of wind pressure, slag type Fe/SiO2CaO: 20-26: 23-30: 16-20, and the oxygen-enriched concentration is 23-25%.
Further, the blowing conditions of the fuming furnace in the zinc extraction are as follows: the zinc content of the mixture is 12-18%, and the total blast volume is 19.5-23.6 Km 3/h; total wind pressure: 55-58 kPa, primary air pressure: 45-51 kPa; secondary air pressure: 55-58 kPa; negative pressure of a tertiary air port: -30 to-80 Pa.
Further, the acid concentration of the mixed acid liquid in the S6 is 150 g/L; the solid-to-solid ratio of the leaching solution is 5:1, the leaching temperature is 80-90 ℃, the leaching time is 8 hours, and the final residual acid is 15-20 g/L.
Further, the replacement time in the S6 is 4 hours, and the pH value of the replacement end point is 4.8-5.0.
Further, the specific operation of S7 is:
a sulfuric acid leaching stage: the leaching base solution is waste liquid of washing slag water and zinc electrolysis, the solid-to-solid ratio of copper-cadmium-containing slag to the leaching base solution is 5:1, the initial leaching acid is 145g/L, the leaching temperature is 70-80 ℃, the oxidant of iron is manganese dioxide ore powder, the leaching time is 4-5 h, the end point is when 3-5 g/L of residual acid is leached, the leached copper slag and the leaching liquid are obtained, and the leached copper slag is returned to the fuming furnace for volatilization and zinc extraction;
putting the leachate containing acid more than or equal to 5g/L into an iron removing tank to measure the content of ferrous iron, when the ferrous iron does not reach the standard, putting hydrogen peroxide into the iron removing tank until the pH value is 5-5.2, putting a Co removing agent to remove Co, adding the agent for 1h, filtering, returning filter residues into a rotary kiln to volatilize, and performing harmless treatment; filtrateCd content up to 18kg/m3Then flows into the replacement groove;
replacing Cd with zinc powder at the replacement temperature of less than 60 ℃ for 1.5-2 h, and performing filter pressing and hot water washing when the Cd content of the solution is less than 50mg/L to obtain sponge cadmium; the filtrate returns to the step of neutral leaching of the secondary zinc oxide calcine to be used as a supplementary solution; the sponge cadmium is rectified by continuous distillation and purification by a tray type electric heating vacuum rectifying furnace, so that the content of the rectified cadmium reaches 99.995 percent, the total amount of impurities is less than 0.005 percent, the rectified slag contains 25 to 30 percent of zinc, and the rectified slag is returned to a fuming furnace to volatilize and extract the zinc.
The invention has the beneficial effects that: according to the method for recovering zinc from the oxygen-enriched sulfur-fixing reduction smelting slag, zinc is extracted from the oxygen-enriched sulfur-fixing reduction smelting slag containing lead, zinc waste residues or lead plaster by a pyrogenic-wet method, the method fully recycles the slag obtained after the lead, zinc waste residues or lead plaster is subjected to blank block making and sulfur-fixing reduction smelting, not only is the existing resource further recovered, but also the pollution of valuable metals to the environment is avoided, and the method is safer and more environment-friendly; meanwhile, the method has the advantages of simple principle, reasonable flow, high zinc recovery rate and low cost.
Detailed Description
The embodiments of the present invention can be obtained by different substitutions in specific ranges based on the above technical solutions, and therefore, the following embodiments are only preferred embodiments of the embodiments, and any technical substitutions made by the above technical solutions are within the protection scope of the present invention.
Neutral leaching of secondary zinc oxide calcine in the embodiment:
ZnO·SiO2+H2SO4=ZnSO4+SiO2·H2O FeO·SiO2+H2SO4=FeSO4+SiO2·H2O
PbO·SiO2+H2SO4=PbSO4+SiO2·H2O ZnO+H2SO4=ZnSO4+H2O
PbO+H2SO4=PbSO4+H2O In2O3+3H2SO4=In2(SO4)3+3H2O
GeO2+H2SO4=GeSO4+H2O GaO+H2SO4=GaSO4+H2O
In2(SO4)3+3Zn=3ZnSO4+2In↓ Ge(SO4)2+2Zn=2ZnSO4+Ge↓
Figure BDA0002154513130000051
Figure BDA0002154513130000052
three-stage purification of the secondary zinc oxide calcine neutral leaching solution:
GaSO4+Zn=ZnSO4+Ga↓ Zn+Cu2+=Zn2++Cu↓
Zn+Cd2+=Zn2++Cd↓ Zn+Co2+=Zn2++Co↓
8C2H5OCS2Na+2CuSO4+CoSO4=Cu2(C2H5OCS2)2↓+2Co(C2H5OCS2)3↓+NaSO4Cu+2Cl-+Cu2+=Cu2Cl2
example 1
A method for recovering zinc from oxygen-enriched solid sulfur reduction smelting slag through a firework furnace comprises the following steps:
s1, material preparation and blank making: taking lead-zinc-containing waste residues of certain electrolytic zinc enterprises, and mixing the lead-zinc-containing waste residues with a sulfur-fixing agent and a flux in a mass ratio of 100: 5: 5, mixing to obtain material powder, wherein the water content of the material powder is 12%; briquetting and blank-making the material powder under 30MPa, and drying until the water content is 5% and the Pb content is 20% -25% to obtain a briquette;
s2, oxygen-enriched sulfur-fixing reduction smelting: mixing the briquettes with coke in a mass ratio of 100: 9 add 7.8m2Carrying out slagging reaction in the oxygen-enriched sulfur-fixing reduction smelting furnace, wherein the sulfur-fixing reduction smelting conditions are as follows: coke rate 9% and blast intensity 35m3/min·m215kpa of wind pressure, slag type Fe/SiO2CaO is 20: 23: 16, obtaining slag with the oxygen-enriched concentration of 23 percent;
s3, converting in a fuming furnace: mixing the furnace slag with low-grade waste slag containing lead and zinc to obtain a mixture, adding the mixture into a fuming furnace for converting, wherein the converting conditions of the fuming furnace are as follows: the zinc content of the mixture is 15 to 18 percent, and the total blast volume is 19.5Km3H; total wind pressure: 55kPa, primary air pressure: 45 kPa; secondary air pressure: 55 kPa; negative pressure of a tertiary air port: -30 Pa; at 1250 ℃, generating metal oxide by heavy metal gas in the material and oxygen in furnace gas, entering the furnace gas, cooling by waste heat, and collecting dust by a cloth bag to recover secondary zinc oxide smoke dust; granulating the secondary zinc oxide smoke dust and the high-indium secondary zinc oxide, and feeding the granulated secondary zinc oxide smoke dust and furnace gas into a rotary kiln in a countercurrent manner for roasting at the roasting temperature of 1200 ℃ to obtain secondary zinc oxide calcine;
s4, neutral leaching: adding roasted secondary zinc oxide into a mixed waste acid solution with the initial acid concentration of 90-100 g/L to perform neutral leaching at the temperature of 60 ℃, wherein the liquid-solid ratio is 8:1, adding manganese dioxide which is 1.2 times of the ferrous iron in solution, and when the pH value reaches 5-5.2, adding Fe (OH) into the leaching solution3Hydrolyzing, coagulating and settling with impurity ions, and performing solid-liquid separation to obtain neutral leaching residue and neutral leaching solution;
s5, neutral leaching and purifying: a first stage: adding zinc powder when the temperature of the neutral leaching solution is 50 ℃, wherein the adding amount of the zinc powder is 2g/L, reacting for 1h, and filtering by a filter press to obtain copper-cadmium-containing slag and filtrate 1; and a second stage: heating the filtrate 1 to 80 ℃ by using steam, adding antimonate and zinc powder to remove impurities, wherein the adding amount of the zinc powder is 3g/L, the adding amount of the antimonate is 0.6 times of the mass of cobalt in the solution, reacting for 3 hours, filtering by using a filter press to obtain cobalt-containing nickel slag and filtrate 2, and sending the cobalt-containing nickel slag to a cadmium recovery workshop; a third stage: cooling filtrate 2 to below 70 deg.C, adding zinc powder to remove Cd, reacting for 1 hr, filtering with filter press to obtain residue and filtrate 3,returning the dregs to the first stage purifying tank; mixing the filtrate 3 and the waste electrolyte at the outlet of the cooling tower according to the volume ratio of 1:15, adding the mixture into an electrodeposition tank which takes quaternary alloy plates (lead, silver, calcium and strontium) as an anode and an aluminum plate as a cathode through a chute to perform electrodeposition, wherein the electrodeposition conditions are as follows: the main component of the electrolyte is sulfuric acid 150g/L, Zn of 50g/L, and the current density is 180A/m2Period 24h, cell voltage 3.4V. After electrodeposition is finished, cathode zinc and waste liquid are obtained, and the cathode zinc is stripped to obtain separated zinc sheets; melting and separating out zinc sheets by adopting a power frequency induction furnace, adding a clarifying agent (caustic soda with the addition of 30-40 kg/t of zinc), maintaining the temperature of zinc liquid in the furnace at 470 ℃, and casting into zinc ingots;
s6, indium and germanium enrichment and recovery: mixing the waste liquid, concentrated sulfuric acid and washing water to prepare mixed acid liquid, wherein the acid concentration is 150 g/L; adding the mixed acid solution and the neutral leaching residues into a leaching tank, wherein the liquid-solid ratio is 5:1, the leaching temperature is 80 ℃, the leaching time is 8 hours, the end-point residual acid is 20g/L, then carrying out solid-liquid separation to obtain acidic leaching residues and acidic leaching solution, sending the acidic leaching residues to an oxygen-enriched sulfur-fixing reduction smelting furnace for smelting, and sending the acidic leaching solution into an indium-germanium enrichment tank; adding 50Kg/g of Zn powder into the acid leachate at the temperature of 75 ℃ for replacement, wherein the replacement time is 4 hours, the pH value of the replacement end point is 4.8-5.0, then filtering, and taking the filtrate as neutral leaching base solution of zinc hypoxide calcine to obtain indium and germanium enriched slag; acid leaching and separating the indium-germanium enriched slag to obtain acid leaching slag and acid leaching liquid; returning the acid leaching residue to a fuming furnace for roasting; extracting the pickle liquor to obtain an indium-rich organic phase and a germanium-containing raffinate, performing back extraction on the indium-rich organic phase by hydrochloric acid, and adding a zinc plate to the indium-containing liquid for replacement to obtain sponge indium; pressing into balls, and casting crude indium ingots at 300 ℃. Washing the germanium-containing raffinate with water, concentrating, adding oxalic acid with the mass concentration of 7% to precipitate germanium, and obtaining germanium-containing concentrate;
the solvent for extracting the pickle liquor is 30 percent of P204And 70% of 260#Solvent oil, extracting with 150g/L H after 3 stages2SO4Washing at the 2 level; the concentration of the hydrochloric acid for back extraction is as follows: 6mol/L, the back extraction temperature is less than 40 ℃, and H with the concentration of 150g/L is used after 3 grades of back extraction2SO4Washing for 2 grades to obtain indium-containing liquid; the indium-containing liquid is sent into a replacement box, a zinc plate is hung for indium replacement for a week at room temperature, and the method is startedThe acidity pH is 1.0, and the In content of the solution after replacement is less than or equal to 50 mg/L;
s7, the recovery operation of the copper-containing cadmium slag comprises the following steps: and recovering cadmium by adopting sulfuric acid leaching, zinc powder replacement and sponge cadmium rectification processes. The method specifically comprises the following steps: a sulfuric acid leaching stage: the leaching base solution is washing slag water and the waste liquid of S5, the solid-to-solid ratio of the copper-cadmium-containing slag to the leaching base solution is 5:1, the initial leaching acid is 145g/L, the leaching temperature is 70 ℃, the addition amount of the oxidant manganese dioxide ore powder is 1.2 times of the divalent iron amount, the leaching time is 4 hours, the end point is when the residual acid is 3-5 g/L, the leached copper slag and the leaching liquid are obtained, and the leached copper slag is returned to the fuming furnace for volatilization and zinc extraction;
putting the leachate containing acid more than or equal to 5g/L into an iron removing tank to measure the content of ferrous iron, and when the ferrous iron does not reach the standard, adding hydrogen peroxide to the leachate until the pH value is 5-5.2, and adding C with the mass of 1.2 times that of cobalt2H5OCS2Removing Co from Na, adding the agent for 1h, filtering, returning filter residue as iron slag, and volatilizing in a rotary kiln for harmless treatment; the Cd content of the filtrate reaches 18kg/m3Then flows into the replacement groove;
replacing Cd with zinc powder at the replacement temperature of less than 60 ℃ for 1.5h, and performing filter pressing and hot water washing when the Cd content of the solution is less than 50mg/L to obtain sponge cadmium; the filtrate returns to the step of neutral leaching of the secondary zinc oxide calcine to be used as a supplementary solution; the sponge cadmium is rectified by continuous distillation and purification by a tray type electric heating vacuum rectifying furnace, so that the content of the rectified cadmium reaches 99.995 percent, the total amount of impurities is less than 0.005 percent, the rectified slag contains 25 to 27 percent of zinc, and the rectified slag is returned to a fuming furnace to volatilize and extract the zinc.
Example 2
A method for recovering zinc from oxygen-enriched solid sulfur reduction smelting slag comprises the following steps:
s1, material preparation and blank making: taking lead-zinc-containing waste residues of certain electrolytic zinc enterprises, and mixing the lead-zinc-containing waste residues with a sulfur-fixing agent and a flux in a mass ratio of 100: 7: 8, mixing to obtain material powder, wherein the water content of the material powder is 14%; briquetting and blank-making the material powder under 40MPa, and drying until the water content is 5% and the Pb content is 25% -28%, so as to obtain a briquette;
s2, oxygen-enriched sulfur-fixing reduction smelting: mixing the briquettes with coke in a mass ratio of 100: 11 add 7.8m2Slagging reaction in oxygen-enriched sulfur-fixing reduction smelting furnaceOxygen-enriched sulfur-fixing reduction smelting conditions are as follows: coke rate 10%, blast intensity 39m3/min·m216kpa wind pressure, slag type Fe/SiO2CaO is 24: 27: 18, obtaining slag with the oxygen-enriched concentration of 24 percent;
s3, converting in a fuming furnace: mixing the furnace slag with low-grade waste slag containing lead and zinc to obtain a mixture, adding the mixture into a fuming furnace for converting, wherein the converting conditions of the fuming furnace are as follows: the zinc content of the mixture is 17 to 18 percent, and the total blast volume is 21.5Km3H; total wind pressure: 56kPa, primary air pressure: 48 kPa; secondary air pressure: 57 kPa; negative pressure of a tertiary air port: -50 Pa; at 1280 ℃, generating metal oxide from heavy metal gas in the material and oxygen in the furnace gas, entering the furnace gas, cooling by waste heat, and collecting dust by a cloth bag to recover secondary zinc oxide smoke dust; granulating the secondary zinc oxide smoke dust and the high-indium secondary zinc oxide, and feeding the granulated secondary zinc oxide smoke dust and furnace gas into a rotary kiln in a countercurrent manner for roasting at the roasting temperature of 1250 ℃ to obtain secondary zinc oxide calcine;
s4, neutral leaching: adding roasted secondary zinc oxide into a mixed waste acid solution with the initial acid concentration of 100-120 g/L to perform neutral leaching at 60 ℃, wherein the liquid-solid ratio is 7:1, adding manganese dioxide which is 1.2 times of the ferrous iron in solution, and when the pH value reaches 5-5.2, Fe (OH) in the leaching solution3Hydrolyzing, coagulating and settling with impurity ions, and performing solid-liquid separation to obtain neutral leaching residue and neutral leaching solution;
s5, neutral leaching and purifying: a first stage: adding zinc powder when the temperature of the neutral leachate is 52 ℃, wherein the adding amount of the zinc powder is 3g/L, reacting for 1h, and filtering by a filter press to obtain copper-cadmium-containing slag and filtrate 1; and a second stage: heating the filtrate 1 to 85 ℃ by using steam, adding antimony salt and zinc powder to remove impurities, wherein the adding amount of the zinc powder is 4.5g/L, the adding amount of the antimony salt is 0.8 times of the mass of cobalt in the solution, reacting for 3 hours, filtering by using a filter press to obtain cobalt-containing nickel slag and filtrate 2, and sending the cobalt-containing nickel slag to a cadmium recovery workshop; a third stage: cooling the filtrate 2 to below 70 ℃, adding zinc powder to remove Cd, wherein the adding amount of the zinc powder is 1.5g/L, reacting for 1h, filtering by a filter press to obtain dregs and a filtrate 3, and returning the dregs to the first section of purification tank; mixing the filtrate 3 and the waste electrolyte at the outlet of a cooling tower according to the volume ratio of 1:18, adding the mixture into an electrodeposition tank which takes quaternary alloy plates (lead, silver, calcium and strontium) as an anode and aluminum plates as a cathode through a chuteElectrodeposition, technical conditions electrolyte main component sulfuric acid: 150g/L, Zn50g/L, other elements strictly controlled, and current density of 200A/m2Period 24h, cell voltage 3.4V. After electrodeposition is finished, cathode zinc and waste liquid are obtained, and the cathode zinc is stripped to obtain separated zinc sheets; melting and separating out zinc sheets by adopting a power frequency induction furnace, adding a clarifying agent (caustic soda with the addition of 30-40 kg/t of zinc), maintaining the temperature of zinc liquid in the furnace at 480 ℃, and casting into zinc ingots;
s6, indium and germanium enrichment and recovery: mixing the waste liquid, concentrated sulfuric acid and washing water to prepare mixed acid liquid, wherein the acid concentration is 150 g/L; adding the mixed acid solution and the neutral leaching residues into a leaching tank, wherein the liquid-solid ratio is 5:1, the leaching temperature is 80 ℃, the leaching time is 8 hours, the end-point residual acid is 15g/L, then carrying out solid-liquid separation to obtain acidic leaching residues and acidic leaching solution, sending the acidic leaching residues to an oxygen-enriched sulfur-fixing reduction smelting furnace for smelting, and sending the acidic leaching solution into an indium-germanium enrichment tank; when the temperature of the acid leachate is 75 ℃, adding 55Kg/g of indium into Zn powder, carrying out replacement for 4 hours, wherein the pH value of the replacement end point is 4.8-5.0, then filtering, and taking the filtrate as neutral leaching base solution of zinc hypoxide calcine to obtain indium and germanium enriched slag; acid leaching and separating the indium-germanium enriched slag to obtain acid leaching slag and acid leaching liquid; returning the acid leaching residue to a fuming furnace for roasting; extracting the pickle liquor to obtain an indium-rich organic phase and a germanium-containing raffinate, performing back extraction on the indium-rich organic phase by hydrochloric acid, and adding a zinc plate to the indium-containing liquid for replacement to obtain sponge indium; pressing into balls, and casting crude indium ingots at 400 ℃. Washing the germanium-containing raffinate with water, concentrating, adding oxalic acid with the mass concentration of 7% to precipitate germanium, and obtaining germanium-containing concentrate;
the solvent for extracting the pickle liquor is 30 percent of P204And 70% of 260#Solvent oil, extracting with 100g/L H after 3 stages2SO4Washing at the 2 level; the concentration of the hydrochloric acid for back extraction is as follows: 6mol/L, the back extraction temperature is less than 40 ℃, and H with the concentration of 100g/L is used after 3 grades of back extraction2SO4Washing for 2 grades to obtain indium-containing liquid; the indium-containing solution is sent into a replacement box and a zinc-coated plate for indium replacement for a week at room temperature, the initial acidity pH is 2.0, and the In content of the replaced solution is less than or equal to 50 mg/L;
s7, the recovery operation of the copper-containing cadmium slag comprises the following steps: and recovering cadmium by adopting sulfuric acid leaching, zinc powder replacement and sponge cadmium rectification processes. The method specifically comprises the following steps: a sulfuric acid leaching stage: the leaching base solution is washing slag water and the waste liquid of S5, the solid-to-solid ratio of the copper-containing cadmium slag to the leaching base solution is 5.5:1, the initial leaching acid is 155g/L, the leaching temperature is 75 ℃, the addition amount of the oxidant manganese dioxide ore powder is 1.3 times of the divalent iron amount, the leaching time is 4.5 hours, the end point is when the residual acid is 3-5 g/L, the leached copper slag and the leaching liquid are obtained, and the leached copper slag is returned to the fuming furnace for volatilization and zinc extraction;
putting the leachate containing acid more than or equal to 5g/L into an iron removing tank to measure the content of ferrous iron, and when the ferrous iron does not reach the standard, adding hydrogen peroxide to the leachate until the pH value is 5-5.2, and adding C with the mass of 1.4 times that of cobalt2H5OCS2Removing Co from Na, adding the agent for 1h, filtering, returning filter residue as iron slag, and volatilizing in a rotary kiln for harmless treatment; the Cd content of the filtrate reaches 18kg/m3Then flows into the replacement groove;
replacing Cd with zinc powder at the replacement temperature of less than 60 ℃ for 1.8h, and performing filter pressing and hot water washing when the Cd content of the solution is less than 50mg/L to obtain sponge cadmium; the filtrate returns to the step of neutral leaching of the secondary zinc oxide calcine to be used as a supplementary solution; the sponge cadmium is rectified by continuous distillation and purification by a tray type electric heating vacuum rectifying furnace, so that the content of the rectified cadmium reaches 99.995 percent, the total amount of impurities is less than 0.005 percent, the rectified slag contains 26 to 28 percent of zinc, and the rectified slag is returned to a fuming furnace to volatilize and extract the zinc.
Example 3
A method for recovering zinc from oxygen-enriched solid sulfur reduction smelting slag comprises the following steps:
s1, material preparation and blank making: taking lead-zinc-containing waste residues of certain electrolytic zinc enterprises, and mixing the lead-zinc-containing waste residues with a sulfur-fixing agent and a flux in a mass ratio of 100: 8: 10, mixing to obtain material powder, wherein the water content of the material powder is 15%; briquetting and blank-making the material powder under 50MPa, and drying until the water content is 6% and the Pb content is 18% -22% to obtain a briquette;
s2, oxygen-enriched sulfur-fixing reduction smelting: mixing the briquettes with coke in a mass ratio of 100: 12 add 7.8m2Carrying out slagging reaction in the oxygen-enriched sulfur-fixing reduction smelting furnace, wherein the oxygen-enriched sulfur-fixing reduction smelting conditions are as follows: coke rate 12% and blast intensity 45m3/min·m218kpa of wind pressure, slag type Fe/SiO2CaO is 26: 30: 20, obtaining slag with the oxygen-enriched concentration of 25 percent;
s3, converting in a fuming furnace: mixing the furnace slag with low-grade waste slag containing lead and zinc to obtain a mixture, adding the mixture into a fuming furnace for converting, wherein the converting conditions of the fuming furnace are as follows: the zinc content of the mixture is 12 to 15 percent, and the total blast volume is 23.6Km3H; total wind pressure: 58kPa, primary air pressure: 51 kPa; secondary air pressure: 58 kPa; negative pressure of a tertiary air port: -80 Pa; at 1300 ℃, generating metal oxide from heavy metal gas in the material and oxygen in the furnace gas, entering the furnace gas, cooling by waste heat, and collecting dust by a cloth bag to recover secondary zinc oxide smoke dust; granulating the secondary zinc oxide smoke dust and the high-indium secondary zinc oxide, and feeding the granulated secondary zinc oxide smoke dust and furnace gas into a rotary kiln in a countercurrent manner for roasting at the roasting temperature of 1300 ℃ to obtain secondary zinc oxide calcine;
s4, neutral leaching: adding roasted secondary zinc oxide into mixed waste acid liquor with initial acid concentration of 70-100 g/L for neutral leaching at 60 ℃, adding manganese dioxide with the mass of 1.2 times of ferrous iron solution at a liquid-solid ratio of 8:1, and when the pH value reaches 5-5.2, adding Fe (OH) into the leachate3Hydrolyzing, coagulating and settling with impurity ions, and performing solid-liquid separation to obtain neutral leaching residue and neutral leaching solution;
s5, neutral leaching and purifying: a first stage: adding zinc powder when the temperature of the neutral leachate is 55 ℃, wherein the adding amount of the zinc powder is 4g/L, reacting for 1h, and filtering by a filter press to obtain copper-cadmium-containing slag and filtrate 1; and a second stage: heating the filtrate 1 to 90 ℃ by using steam, adding antimonate and zinc powder to remove impurities, wherein the adding amount of the zinc powder is 6g/L, the adding amount of the antimonate is 1 time of the mass of cobalt in the solution, reacting for 3 hours, filtering by using a filter press to obtain cobalt-containing nickel slag and filtrate 2, and sending the cobalt-containing nickel slag to a cadmium recovery workshop; a third stage: cooling the filtrate 2 to below 70 ℃, adding zinc powder to remove Cd, wherein the adding amount of the zinc powder is 2g/L, reacting for 1h, filtering by a filter press to obtain dregs and a filtrate 3, and returning the dregs to the first section of purification tank; mixing the filtrate 3 and the waste electrolyte at the outlet of a cooling tower according to the volume ratio of 1:20, adding the mixture into an electrodeposition tank which takes quaternary alloy plates (lead, silver, calcium and strontium) as an anode and an aluminum plate as a cathode through a chute for electrodeposition, wherein the electrolyte has the following main components of sulfuric acid: 150g/L, Zn50g/L, other elements strictly controlled, and current density 160A/m2Period 24h, cell voltage 3.4V. Obtaining cathode zinc and waste liquid after the electrodeposition is finished,stripping the cathode zinc to obtain a zinc precipitation sheet; melting and separating out zinc sheets by adopting a power frequency induction furnace, adding a clarifying agent (caustic soda with the addition of 30-40 kg/t of zinc), maintaining the temperature of zinc liquid in the furnace at 490 ℃, and casting into zinc ingots;
s6, indium and germanium enrichment and recovery: mixing the waste liquid, concentrated sulfuric acid and washing water to prepare mixed acid liquid, wherein the acid concentration is 150 g/L; adding the mixed acid solution and the neutral leaching residues into a leaching tank, wherein the liquid-solid ratio is 5:1, the leaching temperature is 80 ℃, the leaching time is 8 hours, the end-point residual acid is 18g/L, then carrying out solid-liquid separation to obtain acidic leaching residues and acidic leaching solution, sending the acidic leaching residues to an oxygen-enriched sulfur-fixing reduction smelting furnace for smelting, and sending the acidic leaching solution into an indium-germanium enrichment tank; when the temperature of the acid leachate is 75 ℃, adding 60Kg/g of indium into Zn powder, carrying out replacement for 4 hours, wherein the pH value of the replacement end point is 4.8-5.0, then filtering, and taking the filtrate as neutral leaching base solution of zinc hypoxide calcine to obtain indium and germanium enriched slag; acid leaching and separating the indium-germanium enriched slag to obtain acid leaching slag and acid leaching liquid; returning the acid leaching residue to a fuming furnace for roasting; extracting the pickle liquor to obtain an indium-rich organic phase and a germanium-containing raffinate, performing back extraction on the indium-rich organic phase by hydrochloric acid, and adding a zinc plate to the indium-containing liquid for replacement to obtain sponge indium; pressing into a mass, and casting crude indium ingots at 260 ℃. Washing the germanium-containing raffinate with water, concentrating, adding oxalic acid with the mass concentration of 7% to precipitate germanium, and obtaining germanium-containing concentrate;
the solvent for extracting the pickle liquor is 30 percent of P204And 70% of 260#Solvent oil, extracting with 120g/L H after 3 stages2SO4Washing at the 2 level; the concentration of the hydrochloric acid for back extraction is as follows: 6mol/L, the back extraction temperature is less than 40 ℃, and H with the concentration of 120g/L is used after 3 grades of back extraction2SO4Washing for 2 grades to obtain indium-containing liquid; the indium-containing solution is sent into a replacement box and a zinc-coated plate for indium replacement for a week at room temperature, the initial acidity pH is 1.5, and the In content of the replaced solution is less than or equal to 50 mg/L;
s7, the recovery operation of the copper-containing cadmium slag comprises the following steps: and recovering cadmium by adopting sulfuric acid leaching, zinc powder replacement and sponge cadmium rectification processes. The method specifically comprises the following steps: a sulfuric acid leaching stage: the leaching base solution is the washing slag water and the waste liquid obtained in the step S5, the solid-to-solid ratio of the copper-containing cadmium slag to the leaching base solution is 6:1, the initial leaching acid is 125g/L, the leaching temperature is 80 ℃, the addition amount of the oxidant manganese dioxide ore powder is 1.2 times of the divalent iron amount, the leaching time is 5 hours, the end point is when the residual acid is 3-5 g/L, the leached copper slag and the leaching liquid are obtained, and the leached copper slag is returned to the fuming furnace for volatilization and zinc extraction;
putting the leachate containing acid more than or equal to 5g/L into an iron removing tank to measure the content of ferrous iron, and when the ferrous iron does not reach the standard, adding hydrogen peroxide to the leachate until the pH value is 5-5.2, and adding C with the mass of 1.4 times that of cobalt2H5OCS2Removing Co from Na, adding the agent for 1h, filtering, returning filter residue as iron slag, and volatilizing in a rotary kiln for harmless treatment; the Cd content of the filtrate reaches 18kg/m3Then flows into the replacement groove;
replacing Cd with zinc powder at the replacement temperature of less than 60 ℃ for 2h, and performing filter pressing and hot water washing when the Cd content of the solution is less than 50mg/L to obtain sponge cadmium; the filtrate returns to the step of neutral leaching of the secondary zinc oxide calcine to be used as a supplementary solution; the sponge cadmium is rectified by continuous distillation and purification by a tray type electric heating vacuum rectifying furnace, so that the content of the rectified cadmium reaches 99.995 percent, the total amount of impurities is less than 0.005 percent, the rectified slag contains 27 to 30 percent of zinc, and the rectified slag is returned to a fuming furnace to volatilize and extract the zinc.
The recovery rates of the metals obtained in examples 1 to 3 are shown in Table 1 below.
TABLE 1 Metal recovery (%)
Valuable metal Zinc Indium (In) Cadmium (Cd)
Example 1 96~97 92~94 95~96
Example 2 96~98 92~95 95~96
Example 3 96~98 93~95 95~96
The invention is not limited to the foregoing embodiments. The invention extends to any novel feature or any novel combination of features disclosed in this specification and any novel method or process steps or any novel combination of features disclosed.

Claims (3)

1. A method for recovering zinc from oxygen-enriched solid sulfur reduction smelting slag is characterized by comprising the following steps:
s1, material preparation and blank making: taking lead-zinc-containing waste residues of certain electrolytic zinc enterprises, and mixing the lead-zinc-containing waste residues with a sulfur-fixing agent and a flux in a mass ratio of 100: 5: 5, mixing to obtain material powder, wherein the water content of the material powder is 12%; briquetting and blank-making the material powder under 30MPa, and drying until the water content is 5% and the Pb content is 20% -25% to obtain a briquette;
s2, oxygen-enriched sulfur-fixing reduction smelting: mixing the briquettes with coke in a mass ratio of 100: 9 add 7.8m2Carrying out slagging reaction in the oxygen-enriched sulfur-fixing reduction smelting furnace, wherein the sulfur-fixing reduction smelting conditions are as follows: coke rate 9% and blast intensity 35m3/min·m215kPa blast pressure, slag type Fe/SiO2CaO is 20: 23: 16, obtaining slag with the oxygen-enriched concentration of 23 percent;
s3, converting in a fuming furnace: mixing the furnace slag with low-grade waste slag containing lead and zinc to obtain a mixture, adding the mixture into a fuming furnace for converting, wherein the converting conditions of the fuming furnace are as follows: the zinc content of the mixture is 15 to 18 percent, and the total blast volume is 19.5Km3H; total wind pressure: 55kPaPrimary air pressure: 45 kPa; secondary air pressure: 55 kPa; negative pressure of a tertiary air port: -30 Pa; at 1250 ℃, generating metal oxide by heavy metal gas in the material and oxygen in furnace gas, entering the furnace gas, cooling by waste heat, and collecting dust by a cloth bag to recover secondary zinc oxide smoke dust; granulating the secondary zinc oxide smoke dust and the high-indium secondary zinc oxide, and feeding the granulated secondary zinc oxide smoke dust and furnace gas into a rotary kiln in a countercurrent manner for roasting at the roasting temperature of 1200 ℃ to obtain secondary zinc oxide calcine;
s4, neutral leaching: adding roasted secondary zinc oxide into a mixed waste acid solution with the initial acid concentration of 90-100 g/L to perform neutral leaching at the temperature of 60 ℃, wherein the liquid-solid ratio is 8:1, adding manganese dioxide which is 1.2 times of the ferrous iron in solution, and when the pH value reaches 5-5.2, adding Fe (OH) into the leaching solution3Hydrolyzing, coagulating and settling with impurity ions, and performing solid-liquid separation to obtain neutral leaching residue and neutral leaching solution;
s5, neutral leaching and purifying: a first stage: adding zinc powder when the temperature of the neutral leaching solution is 50 ℃, wherein the adding amount of the zinc powder is 2g/L, reacting for 1h, and filtering by a filter press to obtain copper-cadmium-containing slag and filtrate 1; and a second stage: heating the filtrate 1 to 80 ℃ by using steam, adding antimonate and zinc powder to remove impurities, wherein the adding amount of the zinc powder is 3g/L, the adding amount of the antimonate is 0.6 times of the mass of cobalt in the solution, reacting for 3 hours, filtering by using a filter press to obtain cobalt-containing nickel slag and filtrate 2, and sending the cobalt-containing nickel slag to a cadmium recovery workshop; a third stage: cooling the filtrate 2 to below 70 ℃, adding zinc powder to remove Cd, wherein the adding amount of the zinc powder is 1g/L, reacting for 1h, filtering by a filter press to obtain dregs and a filtrate 3, and returning the dregs to the first section of purification tank; mixing the filtrate 3 and the waste electrolyte at the outlet of the cooling tower according to the volume ratio of 1:15, adding the mixture into an electrodeposition tank which takes a lead, silver, calcium and strontium quaternary alloy plate as an anode and an aluminum plate as a cathode through a chute to perform electrodeposition, wherein the electrodeposition conditions are as follows: the main component of the electrolyte is sulfuric acid 150g/L, Zn of 50g/L, and the current density is 180A/m2The period is 24h, the bath voltage is 3.4V, cathode zinc and waste liquid are obtained after electrodeposition is finished, and the cathode zinc is stripped to obtain separated zinc sheets; melting and separating out zinc sheets by adopting a power frequency induction furnace, adding a clarifying agent, wherein the clarifying agent is caustic soda, the adding amount of the caustic soda is 30-40 kg/t of zinc, and casting into zinc ingots when the temperature of zinc liquid in the furnace is maintained at 470 ℃;
s6, indium and germanium enrichment and recovery: mixing the waste liquid, concentrated sulfuric acid and washing water to prepare mixed acid liquid, wherein the acid concentration is 150 g/L; adding the mixed acid solution and the neutral leaching residues into a leaching tank, wherein the liquid-solid ratio is 5:1, the leaching temperature is 80 ℃, the leaching time is 8 hours, the end-point residual acid is 20g/L, then carrying out solid-liquid separation to obtain acidic leaching residues and acidic leaching solution, sending the acidic leaching residues to an oxygen-enriched sulfur-fixing reduction smelting furnace for smelting, and sending the acidic leaching solution into an indium-germanium enrichment tank; adding 50Kg/g of Zn powder into the acid leachate at the temperature of 75 ℃ for replacement, wherein the replacement time is 4 hours, the pH value of the replacement end point is 4.8-5.0, then filtering, and taking the filtrate as neutral leaching base solution of zinc hypoxide calcine to obtain indium and germanium enriched slag; acid leaching and separating the indium-germanium enriched slag to obtain acid leaching slag and acid leaching liquid; returning the acid leaching residue to a fuming furnace for roasting; extracting the pickle liquor to obtain an indium-rich organic phase and a germanium-containing raffinate, performing back extraction on the indium-rich organic phase by hydrochloric acid, and adding a zinc plate to the indium-containing liquid for replacement to obtain sponge indium; pressing into briquettes, casting crude indium ingots at 300 ℃, washing germanium-containing raffinate with water, concentrating, adding oxalic acid with the mass concentration of 7% to precipitate germanium, and obtaining germanium-containing enriched substances;
the solvent for extracting the pickle liquor is 30 percent of P204And 70% of 260#Solvent oil, extracting with 150g/L H after 3 stages2SO4Washing at the 2 level; the concentration of the hydrochloric acid for back extraction is as follows: 6mol/L, the back extraction temperature is less than 40 ℃, and H with the concentration of 150g/L is used after 3 grades of back extraction2SO4Washing for 2 grades to obtain indium-containing liquid; the indium-containing solution is sent into a replacement box and a zinc-coated plate for indium replacement for a week at room temperature, the initial acidity pH is 1.0, and the In content of the replaced solution is less than or equal to 50 mg/L;
s7, the recovery operation of the copper-containing cadmium slag comprises the following steps: the cadmium is recovered by adopting sulfuric acid leaching, zinc powder replacement and sponge cadmium rectification processes, and the method specifically comprises the following steps: a sulfuric acid leaching stage: the leaching base solution is washing slag water and the waste liquid of S5, the solid-to-solid ratio of the copper-cadmium-containing slag to the leaching base solution is 5:1, the initial leaching acid is 145g/L, the leaching temperature is 70 ℃, the addition amount of the oxidant manganese dioxide ore powder is 1.2 times of the divalent iron amount, the leaching time is 4 hours, the end point is when the residual acid is 3-5 g/L, the leached copper slag and the leaching liquid are obtained, and the leached copper slag is returned to the fuming furnace for volatilization and zinc extraction;
placing the leachate containing acid more than or equal to 5g/L in an iron removing tank to determine the content of ferrous iron, and adding hydrogen peroxide when the ferrous iron does not reach the standardWhen the pH value of the water is 5-5.2, C with the mass of 1.2 times of that of the cobalt is added2H5OCS2Removing Co from Na, adding the agent for 1h, filtering, returning filter residue as iron slag, and volatilizing in a rotary kiln for harmless treatment; the Cd content of the filtrate reaches 18kg/m3Then flows into the replacement groove;
replacing Cd with zinc powder at the replacement temperature of less than 60 ℃ for 1.5h, and performing filter pressing and hot water washing when the Cd content of the solution is less than 50mg/L to obtain sponge cadmium; the filtrate returns to the step of neutral leaching of the secondary zinc oxide calcine to be used as a supplementary solution; the sponge cadmium is rectified by continuous distillation and purification by a tray type electric heating vacuum rectifying furnace, so that the content of the rectified cadmium reaches 99.995 percent, the total amount of impurities is less than 0.005 percent, the rectified slag contains 25 to 27 percent of zinc, and the rectified slag is returned to a fuming furnace to volatilize and extract the zinc.
2. A method for recovering zinc from oxygen-enriched solid sulfur reduction smelting slag is characterized by comprising the following steps:
s1, material preparation and blank making: taking lead-zinc-containing waste residues of certain electrolytic zinc enterprises, and mixing the lead-zinc-containing waste residues with a sulfur-fixing agent and a flux in a mass ratio of 100: 7: 8, mixing to obtain material powder, wherein the water content of the material powder is 14%; briquetting and blank-making the material powder under 40MPa, and drying until the water content is 5% and the Pb content is 25% -28%, so as to obtain a briquette;
s2, oxygen-enriched sulfur-fixing reduction smelting: mixing the briquettes with coke in a mass ratio of 100: 11 add 7.8m2Carrying out slagging reaction in the oxygen-enriched sulfur-fixing reduction smelting furnace, wherein the oxygen-enriched sulfur-fixing reduction smelting conditions are as follows: coke rate 10%, blast intensity 39m3/min·m216kPa blast pressure, slag type Fe/SiO2CaO is 24: 27: 18, obtaining slag with the oxygen-enriched concentration of 24 percent;
s3, converting in a fuming furnace: mixing the furnace slag with low-grade waste slag containing lead and zinc to obtain a mixture, adding the mixture into a fuming furnace for converting, wherein the converting conditions of the fuming furnace are as follows: the zinc content of the mixture is 17 to 18 percent, and the total blast volume is 21.5Km3H; total wind pressure: 56kPa, primary air pressure: 48 kPa; secondary air pressure: 57 kPa; negative pressure of a tertiary air port: -50 Pa; at 1280 ℃, the heavy metal gas in the material and the oxygen in the furnace gas generate metal oxide which enters the furnace gas, the waste heat is cooled, and the cloth bag dust is collected and recycledCollecting secondary zinc oxide smoke dust; granulating the secondary zinc oxide smoke dust and the high-indium secondary zinc oxide, and feeding the granulated secondary zinc oxide smoke dust and furnace gas into a rotary kiln in a countercurrent manner for roasting at the roasting temperature of 1250 ℃ to obtain secondary zinc oxide calcine;
s4, neutral leaching: adding roasted secondary zinc oxide into a mixed waste acid solution with the initial acid concentration of 100-120 g/L to perform neutral leaching at 60 ℃, wherein the liquid-solid ratio is 7:1, adding manganese dioxide which is 1.2 times of the ferrous iron in solution, and when the pH value reaches 5-5.2, Fe (OH) in the leaching solution3Hydrolyzing, coagulating and settling with impurity ions, and performing solid-liquid separation to obtain neutral leaching residue and neutral leaching solution;
s5, neutral leaching and purifying: a first stage: adding zinc powder when the temperature of the neutral leachate is 52 ℃, wherein the adding amount of the zinc powder is 3g/L, reacting for 1h, and filtering by a filter press to obtain copper-cadmium-containing slag and filtrate 1; and a second stage: heating the filtrate 1 to 85 ℃ by using steam, adding antimony salt and zinc powder to remove impurities, wherein the adding amount of the zinc powder is 4.5g/L, the adding amount of the antimony salt is 0.8 times of the mass of cobalt in the solution, reacting for 3 hours, filtering by using a filter press to obtain cobalt-containing nickel slag and filtrate 2, and sending the cobalt-containing nickel slag to a cadmium recovery workshop; a third stage: cooling the filtrate 2 to below 70 ℃, adding zinc powder to remove Cd, wherein the adding amount of the zinc powder is 1.5g/L, reacting for 1h, filtering by a filter press to obtain dregs and a filtrate 3, and returning the dregs to the first section of purification tank; mixing the filtrate 3 and the waste electrolyte at the outlet of a cooling tower according to the volume ratio of 1:18, adding the mixture into an electrodeposition tank which takes a lead, silver, calcium and strontium quaternary alloy plate as an anode and an aluminum plate as a cathode through a chute for electrodeposition, wherein the electrolyte has the following main components of sulfuric acid: 150g/L, Zn50g/L, other elements strictly controlled, and current density of 200A/m2The period is 24h, the bath voltage is 3.4V, cathode zinc and waste liquid are obtained after electrodeposition is finished, and the cathode zinc is stripped to obtain separated zinc sheets; melting and separating out zinc sheets by adopting a power frequency induction furnace, adding a clarifying agent, wherein the clarifying agent is caustic soda, the adding amount of the caustic soda is 30-40 kg/t of zinc, and casting into zinc ingots by maintaining the temperature of zinc liquid in the furnace at 480 ℃;
s6, indium and germanium enrichment and recovery: mixing the waste liquid, concentrated sulfuric acid and washing water to prepare mixed acid liquid, wherein the acid concentration is 150 g/L; adding the mixed acid solution and the neutral leaching residues into a leaching tank, wherein the liquid-solid ratio is 5:1, the leaching temperature is 80 ℃, the leaching time is 8 hours, the end-point residual acid is 15g/L, then carrying out solid-liquid separation to obtain acidic leaching residues and acidic leaching solution, sending the acidic leaching residues to an oxygen-enriched sulfur-fixing reduction smelting furnace for smelting, and sending the acidic leaching solution into an indium-germanium enrichment tank; when the temperature of the acid leachate is 75 ℃, adding 55Kg/g of indium into Zn powder, carrying out replacement for 4 hours, wherein the pH value of the replacement end point is 4.8-5.0, then filtering, and taking the filtrate as neutral leaching base solution of zinc hypoxide calcine to obtain indium and germanium enriched slag; acid leaching and separating the indium-germanium enriched slag to obtain acid leaching slag and acid leaching liquid; returning the acid leaching residue to a fuming furnace for roasting; extracting the pickle liquor to obtain an indium-rich organic phase and a germanium-containing raffinate, performing back extraction on the indium-rich organic phase by hydrochloric acid, and adding a zinc plate to the indium-containing liquid for replacement to obtain sponge indium; pressing into balls, casting crude indium ingots at 400 ℃, washing germanium-containing raffinate with water, concentrating, adding oxalic acid with the mass concentration of 7% to precipitate germanium, and obtaining germanium-containing enriched substances;
the solvent for extracting the pickle liquor is 30 percent of P204And 70% of 260#Solvent oil, extracting with 100g/L H after 3 stages2SO4Washing at the 2 level; the concentration of the hydrochloric acid for back extraction is as follows: 6mol/L, the back extraction temperature is less than 40 ℃, and H with the concentration of 100g/L is used after 3 grades of back extraction2SO4Washing for 2 grades to obtain indium-containing liquid; the indium-containing solution is sent into a replacement box and a zinc-coated plate for indium replacement for a week at room temperature, the initial acidity pH is 2.0, and the In content of the replaced solution is less than or equal to 50 mg/L;
s7, the recovery operation of the copper-containing cadmium slag comprises the following steps: the cadmium is recovered by adopting sulfuric acid leaching, zinc powder replacement and sponge cadmium rectification processes, and the method specifically comprises the following steps: a sulfuric acid leaching stage: the leaching base solution is washing slag water and the waste liquid of S5, the solid-to-solid ratio of the copper-containing cadmium slag to the leaching base solution is 5.5:1, the initial leaching acid is 155g/L, the leaching temperature is 75 ℃, the addition amount of the oxidant manganese dioxide ore powder is 1.3 times of the divalent iron amount, the leaching time is 4.5 hours, the end point is when the residual acid is 3-5 g/L, the leached copper slag and the leaching liquid are obtained, and the leached copper slag is returned to the fuming furnace for volatilization and zinc extraction;
putting the leachate containing acid more than or equal to 5g/L into an iron removing tank to measure the content of ferrous iron, and when the ferrous iron does not reach the standard, adding hydrogen peroxide to the leachate until the pH value is 5-5.2, and adding C with the mass of 1.4 times that of cobalt2H5OCS2Removing Co from Na, adding chemical for 1 hr, filtering to obtain iron residue, returning to rotary kiln, and volatilizing to make harmless treatmentProcessing; the Cd content of the filtrate reaches 18kg/m3Then flows into the replacement groove;
replacing Cd with zinc powder at the replacement temperature of less than 60 ℃ for 1.8h, and performing filter pressing and hot water washing when the Cd content of the solution is less than 50mg/L to obtain sponge cadmium; the filtrate returns to the step of neutral leaching of the secondary zinc oxide calcine to be used as a supplementary solution; the sponge cadmium is rectified by continuous distillation and purification by a tray type electric heating vacuum rectifying furnace, so that the content of the rectified cadmium reaches 99.995 percent, the total amount of impurities is less than 0.005 percent, the rectified slag contains 26 to 28 percent of zinc, and the rectified slag is returned to a fuming furnace to volatilize and extract the zinc.
3. A method for recovering zinc from oxygen-enriched solid sulfur reduction smelting slag is characterized by comprising the following steps:
s1, material preparation and blank making: taking lead-zinc-containing waste residues of certain electrolytic zinc enterprises, and mixing the lead-zinc-containing waste residues with a sulfur-fixing agent and a flux in a mass ratio of 100: 8: 10, mixing to obtain material powder, wherein the water content of the material powder is 15%; briquetting and blank-making the material powder under 50MPa, and drying until the water content is 6% and the Pb content is 18% -22% to obtain a briquette;
s2, oxygen-enriched sulfur-fixing reduction smelting: mixing the briquettes with coke in a mass ratio of 100: 12 add 7.8m2Carrying out slagging reaction in the oxygen-enriched sulfur-fixing reduction smelting furnace, wherein the oxygen-enriched sulfur-fixing reduction smelting conditions are as follows: coke rate 12% and blast intensity 45m3/min·m218kPa blast pressure, slag type Fe/SiO2CaO is 26: 30: 20, obtaining slag with the oxygen-enriched concentration of 25 percent;
s3, converting in a fuming furnace: mixing the furnace slag with low-grade waste slag containing lead and zinc to obtain a mixture, adding the mixture into a fuming furnace for converting, wherein the converting conditions of the fuming furnace are as follows: the zinc content of the mixture is 12 to 15 percent, and the total blast volume is 23.6Km3H; total wind pressure: 58kPa, primary air pressure: 51 kPa; secondary air pressure: 58 kPa; negative pressure of a tertiary air port: -80 Pa; at 1300 ℃, generating metal oxide from heavy metal gas in the material and oxygen in the furnace gas, entering the furnace gas, cooling by waste heat, and collecting dust by a cloth bag to recover secondary zinc oxide smoke dust; granulating the secondary zinc oxide smoke dust and the high-indium secondary zinc oxide, and feeding the granulated secondary zinc oxide smoke dust and furnace gas into a rotary kiln in a countercurrent manner for roasting at the roasting temperature of 1300 ℃ to obtain secondary zinc oxide calcine;
s4, neutral leaching: adding roasted secondary zinc oxide into mixed waste acid liquor with initial acid concentration of 70-100 g/L for neutral leaching at 60 ℃, adding manganese dioxide with the mass of 1.2 times of ferrous iron solution at a liquid-solid ratio of 8:1, and when the pH value reaches 5-5.2, adding Fe (OH) into the leachate3Hydrolyzing, coagulating and settling with impurity ions, and performing solid-liquid separation to obtain neutral leaching residue and neutral leaching solution;
s5, neutral leaching and purifying: a first stage: adding zinc powder when the temperature of the neutral leachate is 55 ℃, wherein the adding amount of the zinc powder is 4g/L, reacting for 1h, and filtering by a filter press to obtain copper-cadmium-containing slag and filtrate 1; and a second stage: heating the filtrate 1 to 90 ℃ by using steam, adding antimonate and zinc powder to remove impurities, wherein the adding amount of the zinc powder is 6g/L, the adding amount of the antimonate is 1 time of the mass of cobalt in the solution, reacting for 3 hours, filtering by using a filter press to obtain cobalt-containing nickel slag and filtrate 2, and sending the cobalt-containing nickel slag to a cadmium recovery workshop; a third stage: cooling the filtrate 2 to below 70 ℃, adding zinc powder to remove Cd, wherein the adding amount of the zinc powder is 2g/L, reacting for 1h, filtering by a filter press to obtain dregs and a filtrate 3, and returning the dregs to the first section of purification tank; mixing the filtrate 3 and the waste electrolyte at the outlet of a cooling tower according to the volume ratio of 1:20, adding the mixture into an electrodeposition tank which takes a lead, silver, calcium and strontium quaternary alloy plate as an anode and an aluminum plate as a cathode through a chute for electrodeposition, wherein the electrolyte has the following main components of sulfuric acid: 150g/L, Zn50g/L, other elements strictly controlled, and current density of 160A/m2The period is 24h, the bath voltage is 3.4V, cathode zinc and waste liquid are obtained after electrodeposition is finished, and the cathode zinc is stripped to obtain separated zinc sheets; melting and separating out zinc sheets by adopting a power frequency induction furnace, adding a clarifying agent, wherein the clarifying agent is caustic soda, the adding amount of the caustic soda is 30-40 kg/t of zinc, and casting into zinc ingots while maintaining the temperature of zinc liquid in the furnace at 490 ℃;
s6, indium and germanium enrichment and recovery: mixing the waste liquid, concentrated sulfuric acid and washing water to prepare mixed acid liquid, wherein the acid concentration is 150 g/L; adding the mixed acid solution and the neutral leaching residues into a leaching tank, wherein the liquid-solid ratio is 5:1, the leaching temperature is 80 ℃, the leaching time is 8 hours, the end-point residual acid is 18g/L, then carrying out solid-liquid separation to obtain acidic leaching residues and acidic leaching solution, sending the acidic leaching residues to an oxygen-enriched sulfur-fixing reduction smelting furnace for smelting, and sending the acidic leaching solution into an indium-germanium enrichment tank; when the temperature of the acid leachate is 75 ℃, adding 60Kg/g of indium into Zn powder, carrying out replacement for 4 hours, wherein the pH value of the replacement end point is 4.8-5.0, then filtering, and taking the filtrate as neutral leaching base solution of zinc hypoxide calcine to obtain indium and germanium enriched slag; acid leaching and separating the indium-germanium enriched slag to obtain acid leaching slag and acid leaching liquid; returning the acid leaching residue to a fuming furnace for roasting; extracting the pickle liquor to obtain an indium-rich organic phase and a germanium-containing raffinate, performing back extraction on the indium-rich organic phase by hydrochloric acid, and adding a zinc plate to the indium-containing liquid for replacement to obtain sponge indium; pressing into clusters, casting crude indium ingots at 260 ℃, washing germanium-containing raffinate with water, concentrating, adding oxalic acid with the mass concentration of 7% to precipitate germanium, and obtaining germanium-containing enriched substances;
the solvent for extracting the pickle liquor is 30 percent of P204And 70% of 260#Solvent oil, extracting with 120g/L H after 3 stages2SO4Washing at the 2 level; the concentration of the hydrochloric acid for back extraction is as follows: 6mol/L, the back extraction temperature is less than 40 ℃, and H with the concentration of 120g/L is used after 3 grades of back extraction2SO4Washing for 2 grades to obtain indium-containing liquid; the indium-containing solution is sent into a replacement box and a zinc-coated plate for indium replacement for a week at room temperature, the initial acidity pH is 1.5, and the In content of the replaced solution is less than or equal to 50 mg/L;
s7, the recovery operation of the copper-containing cadmium slag comprises the following steps: the cadmium is recovered by adopting sulfuric acid leaching, zinc powder replacement and sponge cadmium rectification processes, and the method specifically comprises the following steps: a sulfuric acid leaching stage: the leaching base solution is the washing slag water and the waste liquid obtained in the step S5, the solid-to-solid ratio of the copper-containing cadmium slag to the leaching base solution is 6:1, the initial leaching acid is 125g/L, the leaching temperature is 80 ℃, the addition amount of the oxidant manganese dioxide ore powder is 1.2 times of the divalent iron amount, the leaching time is 5 hours, the end point is when the residual acid is 3-5 g/L, the leached copper slag and the leaching liquid are obtained, and the leached copper slag is returned to the fuming furnace for volatilization and zinc extraction;
putting the leachate containing acid more than or equal to 5g/L into an iron removing tank to measure the content of ferrous iron, and when the ferrous iron does not reach the standard, adding hydrogen peroxide to the leachate until the pH value is 5-5.2, and adding C with the mass of 1.4 times that of cobalt2H5OCS2Removing Co from Na, adding the agent for 1h, filtering, returning filter residue as iron slag, and volatilizing in a rotary kiln for harmless treatment; the Cd content of the filtrate reaches 18kg/m3Then flows into the replacement groove;
replacing Cd with zinc powder at the replacement temperature of less than 60 ℃ for 2h, and performing filter pressing and hot water washing when the Cd content of the solution is less than 50mg/L to obtain sponge cadmium; the filtrate returns to the step of neutral leaching of the secondary zinc oxide calcine to be used as a supplementary solution; the sponge cadmium is rectified by continuous distillation and purification by a tray type electric heating vacuum rectifying furnace, so that the content of the rectified cadmium reaches 99.995 percent, the total amount of impurities is less than 0.005 percent, the rectified slag contains 27 to 30 percent of zinc, and the rectified slag is returned to a fuming furnace to volatilize and extract the zinc.
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