CN109913647B - Wet processing method for recovering copper and zinc in bismuth middling - Google Patents
Wet processing method for recovering copper and zinc in bismuth middling Download PDFInfo
- Publication number
- CN109913647B CN109913647B CN201910337857.5A CN201910337857A CN109913647B CN 109913647 B CN109913647 B CN 109913647B CN 201910337857 A CN201910337857 A CN 201910337857A CN 109913647 B CN109913647 B CN 109913647B
- Authority
- CN
- China
- Prior art keywords
- zinc
- copper
- bismuth
- extraction
- solution
- Prior art date
- Legal status (The legal status is an assumption and is not a legal conclusion. Google has not performed a legal analysis and makes no representation as to the accuracy of the status listed.)
- Active
Links
Images
Classifications
-
- Y—GENERAL 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
- Y02—TECHNOLOGIES OR APPLICATIONS FOR MITIGATION OR ADAPTATION AGAINST CLIMATE CHANGE
- Y02P—CLIMATE CHANGE MITIGATION TECHNOLOGIES IN THE PRODUCTION OR PROCESSING OF GOODS
- Y02P10/00—Technologies related to metal processing
- Y02P10/20—Recycling
Landscapes
- Manufacture And Refinement Of Metals (AREA)
Abstract
The invention relates to a wet processing method for recovering copper and zinc in bismuth middlings, which comprises the steps of wet grinding the bismuth middlings to powder with minus 160 meshes accounting for 98%, slurrying and presoaking the powder and diluted zinc electrolysis waste liquid, and filtering to obtain filtrate and filter residue with the final acid of 20-30 g/L. Roasting the filter residue at high temperature to desulfurize the filter residue into oxide, and mixing and smelting bismuth oxide and bismuth concentrate to form crude bismuth; adjusting the pH of the filtrate to 1.5-2.0, performing 3-stage countercurrent extraction, 2-stage washing and 2-stage back extraction on 20% +80% N902 kerosene to obtain a copper sulfate solution, and feeding the copper sulfate solution into a copper electrodeposition system to separate out 99.9% cathode copper at a cathode; and neutralizing the copper extraction raffinate to remove iron, removing impurity elements such as copper, cadmium and the like by using zinc powder, extracting and purifying the solution by using 40% +60% P204 kerosene, and performing 4-stage countercurrent extraction, 3-stage washing and 3-stage back extraction to obtain a zinc sulfate solution. The zinc sulfate solution absorbs the oil in the solution by an oil remover and is sent to a zinc electrodeposition system for electrodeposition of zinc. The material consumption is less, the automation degree is high, no waste water, waste and waste residue are generated in the production process, and the process is a pollution-free cleaning process.
Description
Technical Field
The invention relates to the field of nonferrous metallurgy, in particular to a wet processing method for recovering copper and zinc in bismuth middling.
Background
Bismuth rarely produces ore deposits independently, and most of bismuth is symbiotic with metal elements such as copper, zinc, lead, tin, iron and the like. For bismuth-containing minerals, a beneficiation process is generally adopted to produce bismuth concentrate with higher bismuth content and low-grade bismuth middlings. The bismuth concentrate is reduced and smelted to generate crude bismuth, and the crude bismuth is refined through fire refining and electrolytic refining. Along with the improvement of the crude refining technology, the quality of crude bismuth is improved, impurities are reduced, the electrolytic refining is gradually replaced by the fire refining, and only a few bismuth refining enterprises use less electrolytic refining at present.
The bismuth middling contains low bismuth, the content of impurity metal elements is high, the quantity of slag smelted by a pyrogenic process is large, the production capacity of equipment is reduced, the production capacity of the equipment is reduced, the metal recovery rate is low, the cost is high, and the bismuth middling is generally treated by a hydrometallurgical method in China. In industry, ferric chloride hydrochloride solution is mostly adopted for leaching, and leaching solution is refined into refined bismuth after being replaced by iron. The method has been practiced for many years and has good economic and technical indexes.
Copper, zinc, lead, tin, bismuth and the like in the bismuth middling mainly exist in the form of sulfides, ferric trichloride leaching or chlorine leaching is mostly adopted in the wet treatment of the bismuth middling, particularly, the bismuth leaching rate of the bismuth middling leached by ferric trichloride hydrochloride is high, and the leaching temperature is 98-106 ℃; leaching time is 2 hours; the granularity of the concentrate is-200 meshes and accounts for more than 90 percent; fe3+The concentration is 70-90 g/L; liquid-solid ratio: 3-2: 1; the acidity of the leaching solution is 20-30 g/L of HCl concentration. The disadvantages of leaching with ferric trichloride are that the content of ferrous ions in the leachate is large, which brings difficulty to the separation and purification of the leachate, and the high concentration of chloride ions in the solution also requires the equipment to have the anti-corrosion capability.
In the prior art, for example, cn201110297150.x "process for separating tin, bismuth, copper and zinc from tin-bismuth waste", a solid waste produced by recovering tin-bismuth alloy electroplating waste liquid is called tin-bismuth waste, four elements of tin, bismuth, copper and zinc are efficiently separated through roasting, crushing, acid leaching, bismuth precipitation by neutralization, copper precipitation by displacement and zinc precipitation, the yield is high and reaches 99.2%, 97.8%, 95.4% and 93.2% respectively, and the product quality is good.
For example, CN201410362043.4, "a method for recovering bismuth from precious lead", performs fire refining on the obtained crude bismuth, adds sulfur to remove copper, oxidizes to remove arsenic and antimony, introduces chlorine to remove lead, adds zinc to remove silver, and finally adds caustic soda and niter to perform refining to produce a national standard bismuth ingot product. The method has concise process flow.
Then, as in CN201710334589.2 "A Process for recovering valuable elements from complex low-grade sulfide ores by all-wet method", crushing and grinding raw materials, sieving, mixing with sodium hydroxide solution, and pressure oxidizing leaching to obtain alkaline leaching solution and alkaline leaching residue; acidifying the alkaline leaching solution to obtain a lead sulfate product and a lead precipitation solution, and reducing the lead precipitation solution by controlling the potential step by step to obtain coarse selenium powder, coarse tellurium powder and a reduced solution; leaching copper from the alkaline leaching residue by using a sulfuric acid solution to obtain copper leaching residue and a copper leaching solution; soaking the copper-leaching residue in acidic solution containing chloride ions to obtain antimony-leaching solution and antimony-leaching residue, recovering noble metals from the antimony-leaching residue, and regulating pH of the antimony-leaching solution step by step to obtain antimony residue and bismuth residue; adjusting the pH of the copper leaching solution to obtain iron slag and iron removing solution, wherein the iron slag is a wastewater treatment additive, and the iron removing solution is replaced by zinc powder to obtain copper powder and replaced solution; and combining the reduced solution and the displaced solution, and concentrating and crystallizing to obtain zinc sulfate heptahydrate.
Disclosure of Invention
The invention aims to solve the defects in the prior art and provides a wet treatment method for recovering copper and zinc in bismuth middling.
In order to achieve the purpose, the technical scheme adopted by the invention is as follows:
a wet processing method for recovering copper and zinc in bismuth middlings comprises the following steps:
(1) coarse grinding
The bismuth middlings are subjected to ball milling by an overflow ball mill and a cyclone classifier to obtain ore pulp with-160 meshes accounting for 98%. Pumping the ore pulp into a settling tank for concentration, pumping the underflow into a ceramic filter by using an underflow pump for solid-liquid separation, and obtaining powder with the moisture content of about 18-20%.
(2) Oxygen pressure leaching
The powder and the zinc electrolyte enter a slurrying tank together for slurrying, 0.5-1% of calcium lignosulfonate is added at the same time, the mixture is fully mixed and then pumped into a pre-leaching tank for presoaking and heating to 80-85 ℃. Pumping the slurry after temperature rise into a horizontal reaction kettle through a diaphragm pump, supplying oxygen-enriched gas into the reaction kettle for reverse pressure leaching, flashing and depressurizing the leached slurry to discharge, performing solid-liquid separation through a plate-and-frame filter press, washing leached residues, and then sending to smelting. And the leaching solution enters a copper and zinc recovery process.
(3) Extraction-electrodeposition recovery of copper
Adjusting the pH of the leachate obtained in the step (2) to 1.5-2.0 by using limestone powder, filtering to obtain a supernatant, performing 3-stage countercurrent extraction, 2-stage washing and 2-stage back extraction on the supernatant by using 20% +80% N902 kerosene to obtain a copper sulfate solution, concentrating and evaporating the copper sulfate solution to prepare copper sulfate pentahydrate, and feeding the copper sulfate solution into a copper electrodeposition system to separate out 99.9% cathode copper at a cathode.
(4) P204 extraction-back extraction electro-deposition zinc
And (4) adjusting the pH of the raffinate obtained in the step (3) to 3.5-4.0 by using limestone, and adding 0.5% of hydrogen peroxide by volume ratio to deeply remove iron. Adding trace impurity elements such as copper, cadmium and the like into the iron-removed solution, performing 4-stage countercurrent extraction, 3-stage washing and 3-stage back extraction on the purified solution by using 40% +60% P204 kerosene to obtain a zinc sulfate solution, and preparing zinc ingots by conveying the zinc to the zinc sulfate solution.
The wet treatment method for recovering copper and zinc in the bismuth middling comprises the following steps: the reaction temperature is 170-180 ℃, the total pressure is 1.4-1.6 Mpa, the reaction time is 4 hours, and the liquid-solid ratio is 6-8: 1. the initial acidity of the leaching agent is 60-80 g/L, and the addition amount of calcium lignosulfonate is 0.5-1% of the raw material.
According to the wet treatment method for recovering copper and zinc in the bismuth middling, the end-point acidity of the leachate is controlled to be 20-30 g/L, and the concentration of iron ions is lower than 1 g/L.
In the wet treatment method for recovering copper and zinc in bismuth middling, the organic phase of the copper extraction section of N902: the feed liquid is 1-1.5: 1; a washing section: the water phase is 10-15: 1; stripping section organic phase: the ratio of the counter copper acid is 4-5: 1.
In the wet treatment method for recovering copper and zinc in bismuth middling, the organic phase at the zinc extraction stage of P204: the feed liquid is 2-3: 1; a washing section: the water phase is 15-20: 1; stripping section organic phase: the ratio of the counter copper acid is 6-8: 1.
The oxygen pressure leaching slag contains about 10 percent of lead and about 30 percent of bismuth; the contents of copper and zinc are lower than 1%; the iron content is 10-15%. Lead and bismuth in leaching slag exist in the form of sulfate, high-temperature roasting is needed to desulfurize the lead and the bismuth into oxide, and the bismuth oxide and bismuth concentrate are mixed and smelted to form crude bismuth.
The invention has the beneficial effects that: 1. the metals such as copper, zinc and the like in the bismuth middling are leached by using sulfuric acid oxygen pressure, so that the separation of soluble zinc sulfate and copper sulfate from insoluble lead sulfate and bismuth sulfate is realized, the interference of elements such as copper, zinc and the like on the recovery of bismuth is avoided, and the aim of comprehensively recovering complex multi-metal minerals is fulfilled; 2. the process flow for treating the bismuth middling by oxygen pressure leaching is short, the raw material adaptability is strong, the recovery rate of valuable metals is high, the impurity removal load of a solution is small, the material consumption is low, the automation degree is high, no waste water, waste and waste residue are generated in the production process, and the process is a pollution-free clean process; 3. the invention aims to recover copper and zinc in bismuth middlings, and can obtain final products of cathode copper and zinc ingots, lead and bismuth in oxygen pressure leaching residues are enriched respectively, and the leaching residues can return to a bismuth smelting system by a pyrogenic process.
Drawings
FIG. 1 is a flow chart of oxygen pressure leaching of bismuth middlings according to the present invention;
FIG. 2 is a flow chart of the extraction of the copper-zinc leaching solution of the invention.
Detailed Description
The present invention will be further described by way of the following examples, but the scope of the present invention is not limited to the following examples.
Example 1: the operation was carried out according to the process shown in FIGS. 1-2.
The method is realized by the following 4 steps and technical processes: (1) ball milling; (2) carrying out acid leaching on copper and zinc by oxygen pressure; (3) n902 extraction-stripping electrodeposited copper; (4) p204 extraction-stripping electro-deposition zinc.
100t of bismuth middling containing 18.63% of bismuth, 7.59% of copper, 7.72% of lead and 17.64% of zinc is subjected to ball milling by a wet ball mill, grading by a grader and filtering by a ceramic filter to obtain 98% of bismuth middling powder of-160 meshes.
Oxygen pressure leaching of copper and zinc: using diluted zinc post-electrolysis solution (H)2SO4180～200g/L、Zn2+50-60 g/L) and bismuth middling powder subjected to ball milling are slurried and presoaked, and the presoaked slurry is pumped into 100m by a diaphragm pump3Reacting for 4 hours in a horizontal reaction kettle under the conditions that the reaction temperature is 170-180 ℃ and the total pressure is 1.4-1.6 Mpa, wherein the flow of slurry is 20m3H; the concentration of copper ions in the leaching solution is 8-10 g/L, the concentration of zinc ions is 35-40 g/L, the concentration of iron ions is 0.78g/L, and the concentration of bismuth ions is 0.2 g/L.
20-30 g/l of residual acid in the leaching solution, adjusting the acidity to 1.5-2.0 by limestone powder, and pulping the limestone powder firstly in the acid adjusting process and then pumping the limestone powder into a neutralization tank. Carrying out 3-stage countercurrent extraction on copper by using 20% N902+80% kerosene for a neutralization solution with the pH of 1.5-2.0, wherein the ratio of an extracted N902 organic phase to a feed liquid flow is 1-1.5: 1, and the copper content of a copper extraction residual liquid obtained by extraction is lower than 0.2 g/L; carrying out 2-stage countercurrent washing on the extracted N902-loaded organic phase by using washing water, wherein the flow ratio of the N902-loaded organic phase to the water phase in the washing process is 10-15: 1, precipitating copper and zinc ions in the washed water by using lime, and returning the precipitated liquid to be used for repeatedly washing the N902-loaded organic phase; and (3) carrying out 2-stage back extraction on copper in the organic phase by using an electro-deposition copper-poor liquid after washing, wherein the liquid flow ratio of the N902 carrying organic phase to the copper-poor liquid is 8-10: 1, obtaining copper sulfate solution after back extraction, and enabling the copper sulfate solution to enter a copper electrodeposition system to be separated out into 99.9 percent cathode copper at a cathode.
The copper extraction raffinate contains 12-15 g/L of acid, pH is adjusted to 4.0-4.5 by limestone, and 0.5% hydrogen peroxide is added for oxidation and iron removal until iron ions are lower than 0.001 g/L. And removing trace impurity elements such as copper, cadmium and the like in the solution by using zinc powder after iron removal, wherein the using amount of the zinc powder is 1.0-1.2 times of the content of the copper and the cadmium.
The pH of the purified solution is 4.0-4.5, and the pH is adjusted to 2.0-2.5 by sulfuric acid. Extracting purified liquid by using kerosene with the substance content of 40% P204+60%, carrying out 4-stage countercurrent extraction on zinc, wherein the liquid-flow ratio of the extracted P204 organic phase to the purified liquid-flow is 2-3: 1, the extracted loaded P204 organic phase is washed by using washing water 3-stage countercurrent washing, and the ratio of the loaded P204 organic phase to the washing water-flow in the washing process is 15-20: 1, after washing, the P204 loaded organic phase is subjected to 3-level countercurrent reverse extraction of zinc in the organic phase by using the electro-deposition zinc-containing waste liquid, and the flow ratio of the P204 loaded organic phase to the zinc-containing waste liquid is 8-10: 1, obtaining zinc sulfate back extraction liquid after back extraction. The zinc sulfate solution absorbs the oil in the solution by an oil remover, the solution is sent to a zinc electrodeposition system for electrodeposition of zinc, and a zinc sheet is cast into a zinc ingot.
Further, the zinc extraction raffinate is added with sulfuric acid and returned to the oxygen pressure leaching process.
The method is used for leaching metals such as copper and zinc in the bismuth middling ore by oxygen pressure, so that the separation of copper and zinc from lead and bismuth is realized, the interference of elements such as copper and zinc on the recovery of bismuth is avoided, and the aim of comprehensively recovering complex multi-metal minerals is fulfilled; 2. the process flow for treating the bismuth middling by oxygen pressure leaching is short, the raw material adaptability is strong, the recovery rate of valuable metals is high, the impurity removal load of the solution is small, the material consumption is low, the automation degree is high, no waste water, waste and waste residue are generated in the production process, and the process is a pollution-free clean process.
The foregoing shows and describes the general principles, essential features, and advantages of the invention. It will be understood by those skilled in the art that the present invention is not limited to the embodiments described above, which are merely illustrative of the principles of the invention, but that various changes and modifications may be made without departing from the spirit and scope of the invention, which fall within the scope of the invention as claimed. The scope of the invention is defined by the appended claims and equivalents thereof.
Claims (4)
1. A wet processing method for recovering copper and zinc in bismuth middling is characterized by comprising the following steps:
ball milling: bismuth middling is subjected to ball milling and a spiral classifier to obtain ore pulp with bismuth middling particles of-160 meshes, the ore pulp is pumped into a settling tank for concentration, and underflow is pumped into a ceramic filter by using an underflow pump for solid-liquid separation to obtain powder with the water content of 18-20% by mass;
oxygen pressure leaching: slurrying the ball-milled powder with zinc electrolyte, adding 0.5-1% by mass of calcium lignosulfonate at the same time, fully mixing, pumping into a pre-leaching tank for presoaking and heating to 80-85 ℃, pumping the slurry after heating into a horizontal reaction kettle through a diaphragm pump, supplying oxygen-enriched gas into the reaction kettle for pressure leaching, flashing and depressurizing the slurry after leaching, discharging, performing solid-liquid separation through a plate and frame filter press, washing leaching residues, and sending to a fire process;
n902 extraction-electrodeposition recovery of copper: adjusting the pH of the leachate to 1.5-2.0 by using limestone powder, filtering, performing 3-stage countercurrent extraction, 2-stage washing and 2-stage back extraction on 20% +80% N902 kerosene to obtain a copper sulfate solution, and conveying the copper sulfate solution into a copper electrodeposition system to electrodeposit 99.9% of cathode copper;
p204 extraction-electrodeposition of zinc: adjusting the pH of the copper extraction residual liquid to 3.5-4.0, adding 0.5% hydrogen peroxide for neutralization and deironing, adding zinc powder into the liquid after deironing to purify the solution, performing 4-stage countercurrent extraction, 3-stage washing and 3-stage back extraction on the purified liquid by using 40% +60% P204 kerosene to obtain a zinc sulfate solution, absorbing oil in the solution by using an oil remover, and sending the solution into a zinc electrodeposition system to prepare zinc ingots; the zinc extraction raffinate is added with sulfuric acid and returned to the oxygen pressure leaching process.
2. The wet processing method for recovering copper and zinc in bismuth middlings as claimed in claim 1, wherein the oxygen pressure leaching conditions are as follows: the reaction temperature is 170-180 ℃, the total pressure is 1.4-1.6 MPa, the reaction time is 4 hours, and the liquid-solid ratio is 6-8: 1. the initial acidity of the leaching agent is 60-80 g/L, and the addition amount of calcium lignosulfonate is 0.5-1% of the raw material.
3. The wet processing method for recovering copper and zinc from bismuth middlings according to claim 2, wherein the end-point acidity of the leachate is controlled to be 20-30 g/L, and the concentration of iron ions is lower than 1 g/L.
4. The wet processing method for recovering copper and zinc in bismuth middling according to claim 1, characterized in that 100t of bismuth middling containing 18.63% of bismuth, 7.59% of copper, 7.72% of lead and 17.64% of zinc is ball-milled by a wet ball mill, classified by a classifier and filtered by a ceramic filter to obtain 98% of bismuth middling powder of-160 meshes;
oxygen pressure leaching of copper and zinc: using diluted zinc electrolysis solution, wherein H2SO4180～200g/L、Zn2+50-60 g/L; slurrying and presoaking with the milled bismuth middling powder, and pumping the presoaked slurry into a pump of 100m3Reacting for 4 hours in a horizontal reaction kettle under the conditions that the reaction temperature is 170-180 ℃ and the total pressure is 1.4-1.6 Mpa, wherein the flow of slurry is 20m3H; the concentration of copper ions in the leachate is 8-10 g/L, the concentration of zinc ions is 35-40 g/L, the concentration of iron ions is 0.78g/L, and the concentration of bismuth ions is 0.2 g/L;
20-30 g/l of residual acid in the leaching solution, adjusting the acidity to 1.5-2.0 by limestone powder, slurrying the limestone powder in the acid adjusting process, and pumping into a neutralization tank; carrying out 3-stage countercurrent extraction on copper by using 20% N902+80% kerosene for a neutralization solution with the pH of 1.5-2.0, wherein the ratio of an extracted N902 organic phase to a feed liquid flow is 1-1.5: 1, and the copper content of a copper extraction residual liquid obtained by extraction is lower than 0.2 g/L; carrying out 2-stage countercurrent washing on the extracted N902-loaded organic phase by using washing water, wherein the flow ratio of the N902-loaded organic phase to the water phase in the washing process is 10-15: 1, precipitating copper and zinc ions in the washed water by using lime, and returning the precipitated liquid to be used for repeatedly washing the N902-loaded organic phase; and (3) carrying out 2-stage back extraction on copper in the organic phase by using an electro-deposition copper-poor liquid after washing, wherein the liquid flow ratio of the N902 carrying organic phase to the copper-poor liquid is 8-10: 1, obtaining a copper sulfate solution after back extraction, and separating the copper sulfate solution into 99.9 percent cathode copper in a copper electrodeposition system at a cathode;
the copper extraction raffinate contains 12-15 g/L of acid, pH is adjusted to 4.0-4.5 by limestone, and 0.5% hydrogen peroxide is added for oxidation and iron removal until iron ions are lower than 0.001 g/L; removing trace copper, cadmium and impurity elements in the solution by using zinc powder, wherein the using amount of the zinc powder is 1.0-1.2 times of the content of the copper and the cadmium;
the pH value of the purified liquid is 4.0-4.5, and the pH value is adjusted to 2.0-2.5 by sulfuric acid; extracting purified liquid by using kerosene with the substance content of 40% P204+60%, carrying out 4-stage countercurrent extraction on zinc, wherein the liquid-flow ratio of the extracted P204 organic phase to the purified liquid-flow is 2-3: 1, the extracted loaded P204 organic phase is washed by using washing water 3-stage countercurrent washing, and the ratio of the loaded P204 organic phase to the washing water-flow in the washing process is 15-20: 1, after washing, the P204 loaded organic phase is subjected to 3-level countercurrent reverse extraction of zinc in the organic phase by using the electro-deposition zinc-containing waste liquid, and the flow ratio of the P204 loaded organic phase to the zinc-containing waste liquid is 8-10: 1, obtaining zinc sulfate back extraction liquid after back extraction; the zinc sulfate solution absorbs the oil in the solution by an oil remover, the solution is sent to a zinc electrodeposition system for electrodeposition of zinc, and a zinc sheet is cast into a zinc ingot.
Priority Applications (1)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| CN201910337857.5A CN109913647B (en) | 2019-04-25 | 2019-04-25 | Wet processing method for recovering copper and zinc in bismuth middling |
Applications Claiming Priority (1)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| CN201910337857.5A CN109913647B (en) | 2019-04-25 | 2019-04-25 | Wet processing method for recovering copper and zinc in bismuth middling |
Publications (2)
| Publication Number | Publication Date |
|---|---|
| CN109913647A CN109913647A (en) | 2019-06-21 |
| CN109913647B true CN109913647B (en) | 2020-10-27 |
Family
ID=66978358
Family Applications (1)
| Application Number | Title | Priority Date | Filing Date |
|---|---|---|---|
| CN201910337857.5A Active CN109913647B (en) | 2019-04-25 | 2019-04-25 | Wet processing method for recovering copper and zinc in bismuth middling |
Country Status (1)
| Country | Link |
|---|---|
| CN (1) | CN109913647B (en) |
Families Citing this family (2)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN110396608B (en) * | 2019-08-02 | 2021-06-18 | 中南大学 | Oxygen pressure leaching method for bismuth sulfide concentrate in methanesulfonic acid system |
| CN111647748A (en) * | 2020-05-08 | 2020-09-11 | 成都中建材光电材料有限公司 | Method for separating and purifying crude cuprous telluride slag |
Family Cites Families (3)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| JP2011231366A (en) * | 2010-04-27 | 2011-11-17 | Ibaraki Univ | Valuable metal recovery method |
| CN102851693B (en) * | 2012-08-29 | 2013-08-21 | 江西自立资源再生有限公司 | Technology for recovering production of electrolytic copper and zinc from smelting ash |
| CN109554539A (en) * | 2018-12-26 | 2019-04-02 | 湖南柿竹园有色金属有限责任公司 | A kind of method of bismuth and copper and iron in pressurized acid leaching separation of Bismuth concentrate |
-
2019
- 2019-04-25 CN CN201910337857.5A patent/CN109913647B/en active Active
Also Published As
| Publication number | Publication date |
|---|---|
| CN109913647A (en) | 2019-06-21 |
Similar Documents
| Publication | Publication Date | Title |
|---|---|---|
| CN109234522B (en) | Comprehensive recovery processing method for cobalt-sulfur concentrate | |
| CN102586600B (en) | Process for recycling valuable metal from lead copper matte | |
| CN107190143B (en) | The technique of valuable element in a kind of low-grade sulphide ore of Whote-wet method recycling complexity | |
| CN106086440B (en) | Method for separating and recovering valuable metals in Kaldo furnace smelting slag by wet method | |
| CN107630146B (en) | Nickel recovery process | |
| CN102994747B (en) | Technology for recovering metallic copper from high-lead copper matte | |
| CA2454821C (en) | Process for direct electrowinning of copper | |
| CN102690946B (en) | Method for comprehensively extracting valuable metals from tellurium-containing polymetallic materials | |
| CN111647754A (en) | Comprehensive utilization method of zinc-containing dust and sludge in steel plant | |
| CN102392136B (en) | Process for comprehensively recovering high arsenic polymetallic material | |
| CN101994008A (en) | New process for separating zinc and cobalt from nickel and cobalt slag produced by zinc smelting and purification | |
| CN109913647B (en) | Wet processing method for recovering copper and zinc in bismuth middling | |
| CN110846496B (en) | Sulfating roasting smelting method of sulfur-containing copper-cobalt concentrate | |
| CN113088710A (en) | Method for separating copper and germanium from copper and germanium replacement slag | |
| CN113846214B (en) | Method for treating zinc-containing material in zinc hydrometallurgy production | |
| EP3739069B1 (en) | Method of extracting metals from polymetallic sulphide ores or concentrates | |
| CN114502752B (en) | Method for processing copper sulfide and nickel sulfide materials | |
| CA1097507A (en) | Process of treating residues from the leaching of roasted zinc blende | |
| CN113481371A (en) | Method for efficiently recovering antimony, bismuth, copper and silver from silver separating slag of lead anode slime | |
| CN1237641A (en) | Technological process for extracting Ni, Cu, Co and Mg from nickel sulfide preparation concentrate and making nickelferrite | |
| WO2020149122A1 (en) | Method for manufacturing nickel/cobalt mixed sulfide from nickel oxide ore by wet smelting method | |
| CN103993170A (en) | Method for recycling metals from copper-lead-zinc-arsenic-antimony bulk concentrate | |
| JPH04311541A (en) | Wet-type treating method for zinc concentration and zinc leaching residue at the same time | |
| CN114214520B (en) | Copper-containing refractory material waste-free environment-friendly recovery method | |
| CN111411224B (en) | Beneficiation method for comprehensively recovering and combining silver and copper from low-grade manganese-containing ore |
Legal Events
| Date | Code | Title | Description |
|---|---|---|---|
| PB01 | Publication | ||
| PB01 | Publication | ||
| SE01 | Entry into force of request for substantive examination | ||
| SE01 | Entry into force of request for substantive examination | ||
| GR01 | Patent grant | ||
| GR01 | Patent grant |