CN111926189A - Method for recovering zinc from high-acid zinc dipping slag - Google Patents
Method for recovering zinc from high-acid zinc dipping slag Download PDFInfo
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- CN111926189A CN111926189A CN202010902361.0A CN202010902361A CN111926189A CN 111926189 A CN111926189 A CN 111926189A CN 202010902361 A CN202010902361 A CN 202010902361A CN 111926189 A CN111926189 A CN 111926189A
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- HCHKCACWOHOZIP-UHFFFAOYSA-N Zinc Chemical compound [Zn] HCHKCACWOHOZIP-UHFFFAOYSA-N 0.000 title claims abstract description 103
- 229910052725 zinc Inorganic materials 0.000 title claims abstract description 100
- 239000011701 zinc Substances 0.000 title claims abstract description 100
- 238000000034 method Methods 0.000 title claims abstract description 52
- 239000002253 acid Substances 0.000 title claims abstract description 35
- 239000002893 slag Substances 0.000 title claims abstract description 29
- 238000007598 dipping method Methods 0.000 title claims abstract description 15
- 238000000227 grinding Methods 0.000 claims abstract description 25
- 239000012141 concentrate Substances 0.000 claims abstract description 17
- 238000000926 separation method Methods 0.000 claims abstract description 17
- 239000000706 filtrate Substances 0.000 claims abstract description 15
- 239000007788 liquid Substances 0.000 claims abstract description 15
- 238000005188 flotation Methods 0.000 claims abstract description 13
- 238000004070 electrodeposition Methods 0.000 claims abstract description 8
- 239000003795 chemical substances by application Substances 0.000 claims abstract description 5
- 238000002386 leaching Methods 0.000 claims description 40
- 238000000605 extraction Methods 0.000 claims description 30
- 238000009854 hydrometallurgy Methods 0.000 claims description 14
- ODINCKMPIJJUCX-UHFFFAOYSA-N Calcium oxide Chemical compound [Ca]=O ODINCKMPIJJUCX-UHFFFAOYSA-N 0.000 claims description 10
- XEEYBQQBJWHFJM-UHFFFAOYSA-N Iron Chemical compound [Fe] XEEYBQQBJWHFJM-UHFFFAOYSA-N 0.000 claims description 10
- XLOMVQKBTHCTTD-UHFFFAOYSA-N Zinc monoxide Chemical compound [Zn]=O XLOMVQKBTHCTTD-UHFFFAOYSA-N 0.000 claims description 10
- 239000002699 waste material Substances 0.000 claims description 8
- 239000012074 organic phase Substances 0.000 claims description 7
- 238000000746 purification Methods 0.000 claims description 7
- 239000003792 electrolyte Substances 0.000 claims description 6
- 229910052984 zinc sulfide Inorganic materials 0.000 claims description 6
- 239000000292 calcium oxide Substances 0.000 claims description 5
- 235000012255 calcium oxide Nutrition 0.000 claims description 5
- 229910052742 iron Inorganic materials 0.000 claims description 5
- 239000012528 membrane Substances 0.000 claims description 5
- 239000011787 zinc oxide Substances 0.000 claims description 5
- 238000004064 recycling Methods 0.000 claims description 3
- WGPCGCOKHWGKJJ-UHFFFAOYSA-N sulfanylidenezinc Chemical compound [Zn]=S WGPCGCOKHWGKJJ-UHFFFAOYSA-N 0.000 claims description 3
- 239000004744 fabric Substances 0.000 claims description 2
- 238000010408 sweeping Methods 0.000 claims description 2
- 229910052500 inorganic mineral Inorganic materials 0.000 abstract description 2
- 239000011707 mineral Substances 0.000 abstract description 2
- 238000002156 mixing Methods 0.000 description 15
- CDBYLPFSWZWCQE-UHFFFAOYSA-L Sodium Carbonate Chemical compound [Na+].[Na+].[O-]C([O-])=O CDBYLPFSWZWCQE-UHFFFAOYSA-L 0.000 description 8
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Substances O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 description 8
- 238000003723 Smelting Methods 0.000 description 7
- 238000011084 recovery Methods 0.000 description 6
- 229910000365 copper sulfate Inorganic materials 0.000 description 4
- ARUVKPQLZAKDPS-UHFFFAOYSA-L copper(II) sulfate Chemical compound [Cu+2].[O-][S+2]([O-])([O-])[O-] ARUVKPQLZAKDPS-UHFFFAOYSA-L 0.000 description 4
- 238000004519 manufacturing process Methods 0.000 description 4
- 239000012071 phase Substances 0.000 description 4
- 229910000029 sodium carbonate Inorganic materials 0.000 description 4
- CONMNFZLRNYHIQ-UHFFFAOYSA-N 3-methylbutoxymethanedithioic acid Chemical compound CC(C)CCOC(S)=S CONMNFZLRNYHIQ-UHFFFAOYSA-N 0.000 description 3
- 241000196324 Embryophyta Species 0.000 description 3
- LFQSCWFLJHTTHZ-UHFFFAOYSA-N Ethanol Chemical compound CCO LFQSCWFLJHTTHZ-UHFFFAOYSA-N 0.000 description 3
- 235000008331 Pinus X rigitaeda Nutrition 0.000 description 3
- 235000011613 Pinus brutia Nutrition 0.000 description 3
- 241000018646 Pinus brutia Species 0.000 description 3
- HEMHJVSKTPXQMS-UHFFFAOYSA-M Sodium hydroxide Chemical compound [OH-].[Na+] HEMHJVSKTPXQMS-UHFFFAOYSA-M 0.000 description 3
- 239000005083 Zinc sulfide Substances 0.000 description 3
- TUZCOAQWCRRVIP-UHFFFAOYSA-N butoxymethanedithioic acid Chemical compound CCCCOC(S)=S TUZCOAQWCRRVIP-UHFFFAOYSA-N 0.000 description 3
- 238000006243 chemical reaction Methods 0.000 description 3
- 239000000047 product Substances 0.000 description 3
- 230000002000 scavenging effect Effects 0.000 description 3
- DRDVZXDWVBGGMH-UHFFFAOYSA-N zinc;sulfide Chemical compound [S-2].[Zn+2] DRDVZXDWVBGGMH-UHFFFAOYSA-N 0.000 description 3
- 230000008901 benefit Effects 0.000 description 2
- 238000005034 decoration Methods 0.000 description 2
- 238000012986 modification Methods 0.000 description 2
- 230000004048 modification Effects 0.000 description 2
- 238000001556 precipitation Methods 0.000 description 2
- 239000002351 wastewater Substances 0.000 description 2
- 230000009286 beneficial effect Effects 0.000 description 1
- 238000009388 chemical precipitation Methods 0.000 description 1
- 230000007547 defect Effects 0.000 description 1
- 238000010586 diagram Methods 0.000 description 1
- 238000007865 diluting Methods 0.000 description 1
- 239000006185 dispersion Substances 0.000 description 1
- 230000000694 effects Effects 0.000 description 1
- 239000000284 extract Substances 0.000 description 1
- 239000004088 foaming agent Substances 0.000 description 1
- 231100000053 low toxicity Toxicity 0.000 description 1
- 239000002184 metal Substances 0.000 description 1
- 229910052751 metal Inorganic materials 0.000 description 1
- 235000010755 mineral Nutrition 0.000 description 1
- 238000004537 pulping Methods 0.000 description 1
- 235000011121 sodium hydroxide Nutrition 0.000 description 1
- 239000002689 soil Substances 0.000 description 1
- 238000005406 washing Methods 0.000 description 1
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- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B7/00—Working 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
- C22B7/006—Wet processes
- C22B7/007—Wet processes by acid leaching
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B19/00—Obtaining zinc or zinc oxide
- C22B19/20—Obtaining zinc otherwise than by distilling
- C22B19/22—Obtaining zinc otherwise than by distilling with leaching with acids
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B19/00—Obtaining zinc or zinc oxide
- C22B19/30—Obtaining zinc or zinc oxide from metallic residues or scraps
-
- 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
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- Engineering & Computer Science (AREA)
- Chemical & Material Sciences (AREA)
- Manufacturing & Machinery (AREA)
- Materials Engineering (AREA)
- Mechanical Engineering (AREA)
- Metallurgy (AREA)
- Organic Chemistry (AREA)
- Life Sciences & Earth Sciences (AREA)
- Environmental & Geological Engineering (AREA)
- General Life Sciences & Earth Sciences (AREA)
- Geology (AREA)
- Manufacture And Refinement Of Metals (AREA)
Abstract
The invention discloses a method for recovering zinc from high-acid zinc-dipping slag, which comprises the steps of grinding the high-acid zinc-dipping slag, carrying out solid-liquid separation on ore pulp after grinding, enabling filter residues to enter a flotation tank, adding a mineral separation agent, carrying out one-step four-step fine scanning operation to obtain zinc concentrate, and carrying out purification-extraction-back extraction-electrodeposition on filtrate after solid-liquid separation to obtain zinc ingots.
Description
Technical Field
The invention belongs to the technical field of dressing and smelting combination, and particularly relates to a method for recovering zinc from high-acid zinc dipping slag.
Background
In the production of zinc hydrometallurgy, a roasting ore in-hot acid leaching process and a traditional roasting ore in-hot acid leaching process, medium-temperature acid leaching and high-temperature roasting process are generally adopted, and in the roasting process, zinc blende or zinc ferrosphalerite in zinc concentrate is difficult to be completely converted into zinc oxide, so that part of zinc sulfide after roasting remains in zinc leaching residues, and great waste of zinc resources is caused. In addition, in the production process of zinc hydrometallurgy, 0.4t of leaching slag is generated when 1t of metal zinc is produced, the leaching slag is limited to the existing process flow and solid-liquid separation equipment, 25% -30% of leaching solution still remains in the slag, and finally 4% -5% of water-soluble zinc remains in the slag.
The total mass fraction of zinc in the leached residues is relatively low (below 10 percent), the cost of zinc oxide production by returning to a rotary kiln for high-temperature roasting is high, the zinc oxide has to be used as waste residues for landfill treatment, so that a large amount of zinc resources are wasted, and meanwhile, due to the existence of water-soluble zinc, rivers and soil are easily polluted due to improper curing treatment, and great harm is brought to the ecological environment. At present, some domestic zinc smelting enterprises adopt the mode of adding clean water to carry out pulping, diluting, washing and concentrating for multiple times to recover water-soluble zinc in slag, but the process is complicated, the water consumption is large, and a zinc smelting system has low solution receiving capacity and limited recovery amount due to the problem of volume expansion. Chemical precipitation method is also used at home and abroad to recover water-soluble zinc in the leaching residue, but the zinc precipitation waste water in the method can not be reused, and the cost for treating the waste water is very high; in addition, the high-acid leaching residue has large caustic soda consumption when the pH value is adjusted, and the zinc precipitation cost is higher.
Disclosure of Invention
In order to solve the problems, the invention aims to provide a method for recovering zinc from high-acid zinc-leaching residues so as to overcome various defects of the prior art for recovering the zinc from the zinc-leaching residues. According to the invention, zinc in the high-acid zinc-leaching residue is recovered by a combined selection and smelting method, so that the recovery rate of zinc in the zinc hydrometallurgy process is improved, the waste of zinc resources is reduced, secondary resources are utilized to the greatest extent, and the economic benefit is improved.
The technical scheme of the invention is as follows: grinding the high-acid zinc-leaching residue, performing solid-liquid separation on the ground ore pulp, feeding the filter residue into a flotation tank, adding a mineral separation agent, performing rough four-fine scanning operation to obtain zinc concentrate, and performing purification-extraction-back extraction-electrodeposition flow on the filtrate after solid-liquid separation to obtain zinc ingots.
A method for recovering zinc from high-acid zinc-dipping slag is characterized by comprising the following specific steps:
(1) grinding the high-acid zinc-leaching residues to eliminate the caking phenomenon of the leaching residues, slurrying the leaching residues to be treated, and performing solid-liquid separation on ore pulp after grinding to obtain filtrate and filter residues;
(2) carrying out flotation on the filter residue obtained in the step (1), and recovering zinc blende which is not converted into zinc oxide in the filter residue;
(3) recovering water-soluble zinc in the filtrate obtained in the step (1) by adopting the processes of purification, extraction, back extraction and electrodeposition;
(4) raffinate is obtained after extraction in the step (3), and the raffinate is returned to the step (1) to be recycled as return water;
(5) and (4) obtaining an organic phase after the back extraction in the step (3), and returning the organic phase to the extraction process for recycling.
In the step (1), the concentration of grinding ore is selected to be 40-60%.
In the step (1), solid-liquid separation is carried out by using a membrane filter press of monofilament filter cloth.
In the step (2), a coarse four-fine sweeping flow is adopted for flotation, and a pH regulator of 1000-.
And (3) performing flotation in the step (2) to obtain zinc concentrate, and returning the zinc concentrate to the zinc hydrometallurgy roasting operation.
The purification process in the step (3) is as follows: and (3) using quicklime as a neutralizer, adjusting the pH value to 4.0-4.5, and reacting for 0.5-1.5h to remove iron.
In the step (3), an extractant is used for extraction, and the extractant is unsaponifiable P204.
And (4) using the waste electrolyte of the zinc hydrometallurgy as the stripping agent in the step (3).
The ratio of unsaponified P204 to the water phase in the extraction phase ratio is 1-3: 1, and the ratio of the spent electrolyte to the water phase in the stripping phase ratio is 4-6: 1.
Compared with the prior art, the invention has the following advantages:
(1) the invention adopts ore grinding to the high-acid zinc leaching residues to replace the traditional water-soluble process, and is beneficial to dissolving out the water-soluble zinc in the caked leaching residues.
(2) The addition of sodium carbonate and copper sulfate in the flotation process plays roles in adjusting the pH value of zinc leaching residues, adjusting the dispersion degree of the zinc leaching residues and modifying the surface property of zinc sulfide, zinc sulfide in the leaching residues is recovered by selective size-adjusting flotation, and finally zinc concentrate can be returned to the original zinc hydrometallurgy system for recovery.
(3) The process of purification-extraction-back extraction-electrodeposition effectively extracts water-soluble zinc in the leaching residue and prepares zinc ingots.
The invention is a method for economically treating zinc leaching residue with low zinc content, and the method has simple process, low required investment and low production cost. All backwater in the process flow is recycled, the process is environment-friendly, and finally the high-acid leaching residue is subjected to low-toxicity treatment by the process, so that the pollution to the environment is reduced.
Drawings
FIG. 1 is a schematic process flow diagram of the present invention.
Detailed Description
In order to make the technical means, the creation characteristics, the achievement purposes and the effects of the invention easy to understand, the invention is further explained by combining the drawings and the embodiments.
Example 1: in the embodiment, the high-acid leaching residue (zinc content is 9.4%) of a certain wet-process zinc smelting plant is treated by the following specific steps:
(1) grinding the high-acid zinc leaching slag which is stored for a long time, eliminating the caking phenomenon of the leaching slag, and slurrying the leaching slag to be treated; wherein the concentration of ore grinding is 48%, the ore grinding is carried out until the ore grinding accounts for 92% of minus 400 meshes, and the ore pulp after ore grinding is subjected to solid-liquid separation by using a membrane filter press to obtain filtrate and filter residue;
(2) adding 1080g/t of sodium carbonate (size mixing for 3 min) as a pH regulator, 450g/t of copper sulfate (size mixing for 5 min), 130g/t of butyl xanthate (size mixing for 4 min) +180g/t of isoamyl xanthate (size mixing for 4 min) as a collecting agent and 45g/t of pine alcohol oil (size mixing for 1 min) as a foaming agent into the filter residue obtained in the step (1) in sequence, performing one roughing operation, four times of fine selection operation and one scavenging operation, wherein the concentration of ore pulp is 33% in the operation process, returning tailings to the previous operation in each operation, and forming closed cycle to obtain zinc concentrate and flotation tailings;
(3) recovering water-soluble zinc in the high-acid zinc leaching residue from the filtrate obtained after solid-liquid separation in the step (1) by adopting a purification-extraction-back extraction-electrodeposition process, wherein quicklime is used as a neutralizer in the purification process, the pH is adjusted to 4.3, iron is removed after reaction for 1h, an extractant used for extraction is unsaponified P204, a back extractant is waste electrolyte obtained by zinc hydrometallurgy, the extraction ratio is 1: 1, and the back extraction ratio is 4: 1;
(4) returning the raffinate in the step (3) to the step (1) for returning water;
(5) and (4) returning the organic phase subjected to the back extraction in the step (3) to the extraction process for repeated use.
The zinc concentrate obtained by the treatment by the method of the embodiment has the grade of 36.4% and the recovery rate of 44.38%, and the zinc concentrate is finally returned to the original zinc hydrometallurgy process; the filtrate is purified, extracted, back extracted and electrodeposited to obtain zinc ingot product with zinc recovering rate of 41.22% and zinc recovering rate of 85.6%.
Example 2: in the embodiment, the high-acid leaching residue (the zinc content is 8.5%) of a certain wet-process zinc smelting plant is treated by the following specific steps:
(1) grinding the high-acid zinc leaching slag which is stored for a long time, eliminating the caking phenomenon of the leaching slag, and slurrying the leaching slag to be treated; wherein the concentration of ore grinding is 40%, the ore grinding is carried out until the ore grinding is 88% of-400 meshes, and the ore pulp after ore grinding is subjected to solid-liquid separation by using a membrane filter press to obtain filtrate and filter residue;
(2) sequentially adding 1200g/t of sodium carbonate (size mixing for 3 min), 400g/t of copper sulfate (size mixing for 5 min), 100g/t of butyl xanthate (size mixing for 4 min) +150g/t of isoamyl xanthate (size mixing for 4 min) and 40g/t of pine alcohol oil (size mixing for 1 min) into the filter residue obtained in the step (1), performing one roughing operation, four times of fine selection operation and one time of scavenging operation, wherein the concentration of the ore pulp is 40% in the operation process, returning tailings in the previous operation every time, and forming closed cycle to obtain zinc concentrate and flotation tailings;
(3) recovering water-soluble zinc in the high-acid zinc leaching residue from the filtrate obtained after solid-liquid separation in the step (1) by adopting a purification-extraction-back extraction-electrodeposition process, wherein quicklime is used as a neutralizer in the purification process, the pH is adjusted to 4.0, iron is removed after reaction for 0.5h, an extractant used for extraction is unsaponifiable P204, a back extractant is waste electrolyte obtained by zinc hydrometallurgy, the extraction ratio is 2: 1, and the back extraction ratio is 5: 1;
(4) returning the raffinate in the step (3) to the step (1) for returning water;
(5) and (4) returning the organic phase subjected to the back extraction in the step (3) to the extraction process for repeated use.
The zinc concentrate obtained by the treatment by the method of the embodiment has the grade of 33.2% and the recovery rate of 41.20%, and the zinc concentrate is finally returned to the original zinc hydrometallurgy process; the filtrate is purified, extracted, back extracted and electrodeposited to obtain zinc ingot product with zinc recovering rate of 45.31% and zinc recovering rate of 86.51%.
Example 3: in the embodiment, the high-acid leaching residue (the zinc content is 8.8%) of a certain wet-process zinc smelting plant is treated by the following specific steps:
(1) grinding the high-acid zinc leaching slag which is stored for a long time, eliminating the caking phenomenon of the leaching slag, and slurrying the leaching slag to be treated; wherein the concentration of ore grinding is 60%, the ore grinding is carried out until the ore grinding accounts for 95% of minus 400 meshes, and the ore pulp after ore grinding is subjected to solid-liquid separation by using a membrane filter press to obtain filtrate and filter residue;
(2) adding 1100g/t of sodium carbonate (size mixing for 3 min), 500g/t of copper sulfate (size mixing for 5 min), 150g/t of butyl xanthate (size mixing for 4 min) +200g/t of isoamyl xanthate (size mixing for 4 min) and 60g/t of pine alcohol oil (size mixing for 1 min) into the filter residue obtained in the step (1) in sequence, performing one roughing operation, four times of fine selection operation and one time of scavenging operation, wherein the concentration of the ore pulp is 42% in the operation process, returning tailings in the previous operation every time, and forming closed cycle to obtain zinc concentrate and flotation tailings;
(3) recovering water-soluble zinc in the high-acid zinc leaching residue from the filtrate obtained after solid-liquid separation in the step (1) by adopting a purification-extraction-back extraction-electrodeposition process, wherein quicklime is used as a neutralizer in the purification process, the pH is adjusted to 4.5, iron is removed after reaction for 1.5h, an extractant used for extraction is unsaponifiable P204, a back extractant is waste electrolyte obtained by zinc hydrometallurgy, the extraction ratio is 3: 1, and the back extraction ratio is 6: 1;
(4) returning the raffinate in the step (3) to the step (1) for returning water;
(5) and (4) returning the organic phase subjected to the back extraction in the step (3) to the extraction process for repeated use.
The zinc concentrate obtained by the treatment by the method of the embodiment has the grade of 35.11% and the recovery rate of 46.67%, and the zinc concentrate is finally returned to the original zinc hydrometallurgy process; the filtrate is purified, extracted, back extracted and electrodeposited to obtain zinc ingot product with zinc recovering rate of 39.12% and total zinc recovering rate of 85.79%.
The foregoing is only a preferred embodiment of the present invention, and it should be noted that, for those skilled in the art, various modifications and decorations can be made without departing from the principle of the present invention, and these modifications and decorations should also be regarded as the protection scope of the present invention.
Claims (9)
1. A method for recovering zinc from high-acid zinc-dipping slag is characterized by comprising the following specific steps:
(1) grinding the high-acid zinc-leaching residue, and performing solid-liquid separation on ore pulp after grinding to obtain filtrate and filter residue;
(2) carrying out flotation on the filter residue obtained in the step (1), and recovering zinc blende which is not converted into zinc oxide in the filter residue;
(3) recovering water-soluble zinc in the filtrate obtained in the step (1) by adopting the processes of purification, extraction, back extraction and electrodeposition;
(4) raffinate is obtained after extraction in the step (3), and the raffinate is returned to the step (1) for recycling;
(5) and (4) obtaining an organic phase after the back extraction in the step (3), and returning the organic phase to the extraction process for recycling.
2. The method for recovering zinc from the high-acid zinc-dipping slag according to claim 1, which is characterized in that: and (2) grinding the ore in the step (1), wherein the concentration of the selected ore is 40-60%.
3. The method for recovering zinc from the high-acid zinc-dipping slag according to claim 1, which is characterized in that: in the step (1), solid-liquid separation is carried out by using a membrane filter press of monofilament filter cloth.
4. The method for recovering zinc from the high-acid zinc-dipping slag according to claim 1, which is characterized in that: in the step (2), a coarse four-fine sweeping flow is adopted for flotation, and a pH regulator of 1000-.
5. The method for recovering zinc from the high-acid zinc-dipping slag according to claim 1, which is characterized in that: and (3) performing flotation in the step (2) to obtain zinc concentrate, and returning the zinc concentrate to the zinc hydrometallurgy roasting operation.
6. The method for recovering zinc from the high-acid zinc-dipping slag according to claim 1, which is characterized in that: the purification process in the step (3) is as follows: and (3) using quicklime as a neutralizer, adjusting the pH value to 4.0-4.5, and reacting for 0.5-1.5h to remove iron.
7. The method for recovering zinc from the high-acid zinc-dipping slag according to claim 1, which is characterized in that: in the step (3), an extractant is used for extraction, and the extractant is unsaponifiable P204.
8. The method for recovering zinc from the high-acid zinc-dipping slag according to claim 1, which is characterized in that: and (4) using the waste electrolyte of the zinc hydrometallurgy as the stripping agent in the step (3).
9. The method for recovering zinc from the high-acid zinc-dipping slag according to claim 1, which is characterized in that: in the step (3), the extraction ratio is 1-3: 1, and the back extraction ratio is 4-6: 1.
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Cited By (1)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN113231191A (en) * | 2021-06-02 | 2021-08-10 | 昆明理工大学 | Method for comprehensively recovering zinc, silver and tin in zinc leaching residues |
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| CN101972703A (en) * | 2010-10-27 | 2011-02-16 | 株洲市湘麒科技开发有限公司 | Beneficiation method for recovering zinc, lead and silver from zinc leaching residue |
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2020
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| BG108826A (en) * | 2004-07-26 | 2006-03-31 | "Стефан Кънев" Ет | Method for the treatment of zinc slags |
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| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN113231191A (en) * | 2021-06-02 | 2021-08-10 | 昆明理工大学 | Method for comprehensively recovering zinc, silver and tin in zinc leaching residues |
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