CN104152705A - Process for producing cathode zinc by utilizing casting zinc dross tailings - Google Patents
Process for producing cathode zinc by utilizing casting zinc dross tailings Download PDFInfo
- Publication number
- CN104152705A CN104152705A CN201410388483.7A CN201410388483A CN104152705A CN 104152705 A CN104152705 A CN 104152705A CN 201410388483 A CN201410388483 A CN 201410388483A CN 104152705 A CN104152705 A CN 104152705A
- Authority
- CN
- China
- Prior art keywords
- zinc
- dross
- extraction
- mine tailing
- section
- 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.)
- Granted
Links
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 provides a process for producing cathode zinc by utilizing fusion-cast zinc dross tailings, which mainly comprises the following process steps: (1) first-stage leaching: pulping zinc dross tailings with water, adding sulfuric acid and a second-stage leaching solution, heating, stirring, and filtering to obtain a first-stage leaching solution and a first-stage leaching residue; (2) secondary leaching: pulping the first-stage leaching residue with water, adding sulfuric acid, heating, stirring, and filtering while hot to obtain a second-stage leaching solution and a second-stage leaching residue; (3) pumping the first-stage leachate into a mixing and clarifying extraction box, sequentially mixing the first-stage leachate with a dechlorination extractant which flows reversely, carrying out extraction reaction, standing the dechlorinated raffinate, adsorbing the dechlorinated raffinate through activated carbon to remove oil, and then merging the dechlorinated raffinate into a zinc smelting system to produce cathode zinc. The invention recovers zinc in the zinc dross tailings to the maximum extent, and the harmful element chlorine is recovered in a byproduct form, thereby realizing the aim of changing harm into benefit.
Description
Technical field
The invention belongs to Metal smelting Comprehensive Recovery Technology field, be specifically related to a kind of production technique of the zinc dross mine tailing producing in zinc hydrometallurgy cathode zinc fusion-casting process being carried out to comprehensive recovery.
Background technology
Flowsheet of wet zinc smelting is sulfuric acid, the calcining Leaching Zinc of generation, and zinc lixiviating solution purification and impurity removal, the heavy zinc of zinc purification fluid electrodeposition obtains cathode zinc, and cathode zinc founding obtains product zinc ingot metal.Cathode zinc founding is produced in zinc ingot metal process, and separated for the oxidational losses that reduces metallic zinc and oxidation reinforced scum silica frost and zinc melt need to add the ammonium chloride of 1.5～4kg/t cathode zinc; The zinc dross of while output 3～5%.The chief component of zinc dross is: zinc metal 40～50%, and zinc oxide approximately 50%, zinc chloride 2～3%, total zinc content is about 80%, total chlorine approximately 0.8～2%.In zinc dross, zinc content is high, for it has put on the label of high value nonferrous metal raw material to be recycled; But on the other hand, owing to containing, have a strong impact on the dangerous element chlorine that zinc is smelted production system, and the advanced smelting technology that there is no at present effective separating chlorine, cause the zinc dross that potential value is high not utilized completely, the zinc dross that store up in a large number slag field also can pollute surrounding enviroment under severe weather conditions, affects operation result and the Environmental indicators of enterprise.
Zinc hydrometallurgy enterprise mainly contains following several processing mode to the zinc dross of output at present: 1. the simple production mode that adopts ball milling-scalping, coarse grain metallic zinc in zinc dross is reclaimed for the production of electric furnace zinc powder, screen underflow is that zinc dross mine tailing is piled and had slag field, and this technique only can reclaim 30% left and right of total zinc amount in zinc dross; 2. using zinc dross as byproduct, undersell to some private entrepreneur, adopt production technique simple, that fall behind, sulfuric acid or Leaching in Hydrochloric Acid, recrystallization are produced zinc sulfate or zinc chloride; 3. with salt acid treatment zinc dross, sodium carbonate precipitates zinc subcarbonate, then produces zinc oxide with pyrogenic process calcining, this explained hereafter long flow path, and cost is high; 4. multiple hearth furnace pyrogenic attack zinc dross, but the chlorine of enrichment causes fatal corrosion to equipment, and the decreasing ratio long term maintenance of chlorine is 40～50%, and efficiency is low, and cost is high; 5. adopt wet method alkali wash, 717 resin dechlorinating process, Silver Nitrate dechlorinating process, cuprous chloride method etc., these methods all exist dechlorination efficiency to be not more than 60%, and zinc loses the weak points such as large, is difficult to apply on a large scale.
Summary of the invention
The object of the present invention is to provide a kind of novel process of the zinc dross mine tailing producing in zinc hydrometallurgy cathode zinc fusion-casting process being carried out to remanufacture cathode zinc, this technique can reclaim the zinc in zinc dross mine tailing to greatest extent, and harmful element chlorine is recycled with byproduct form, realize the comprehensive recovery of zinc dross mine tailing.
For achieving the above object, the present invention takes following technical scheme:
Utilize founding zinc dross mine tailing to produce a cathode zinc technique, comprise following processing step:
(1) one section of leaching: will add in one section of pickling tank after the pulp of zinc dross mine tailing water, then add sulfuric acid and two sections of leach liquors, the stirring of heating, control liquid-solid ratio is 6～8:1, it is 180～250g/L that control sulphuric acid makes sulfuric acid starting point concentration in solution, 50～60 ℃ of temperature of reaction, 40～60 minutes reaction times, filter, obtain one section of leach liquor and one section of leached mud;
(2) two sections of leachings: will join in two-stage acid leaching groove after the pulp of one section of leached mud water, then add sulfuric acid, the stirring of heating, control liquid-solid ratio is 4～5:1, and it is 250～350g/L that control sulphuric acid makes sulfuric acid starting point concentration in solution, 80～90 ℃ of temperature of reaction, 60～90 minutes reaction times, filtered while hot, obtains two sections of leach liquors and two sections of leached muds, and two sections of leach liquors return and enter in one section of pickling tank;
(3) chloride one section is leached solution of zinc sulfate extraction dechlorination: one section of leach liquor is pumped into and mixes clarification extraction box, with the raw extractive reaction of the dechlorination extraction agent organic phase of countercurrent flow successively hybrid concurrency, raffinate after dechlorination carries out standing and passes through activated carbon adsorption oil removing, then raffinate after dechlorination is incorporated to zinc smelting system for the production of cathode zinc.
Further, in step (2), it is 50-80ppm that reaction end adds flocculation agent to concentration, to strengthen solution filter performance.
Further, in step (2), using two sections of leached muds as food ingredient, add in zinc sulfide concentrates and carry out roasting, utilize the hot conditions of roasting that lead sulfate in leached mud is decomposed, discharge the metallic zinc of activation, improve the leaching rate of recovery of zinc.
Further, in step (3), after extractive reaction finishes, carrying chlorine organic phase and reverse-extraction agent ammonia-sal volatile carries out countercurrent reextraction and is regenerated, after back extraction, solution is again for back extraction after supplementing reverse-extraction agent solute concentration, and regularly after the back extraction of extracting part partial volume, solution carries out crystallization output byproduct ammonium chloride.
Further, in step (3), the consisting of of extracted organic phase (V%):
Sulfonated kerosene 60～75%
R
3n(tri-iso-octylamine) 20～35%
Secondary octanol 5～10%;
The physical method that described extracted organic phase adopts common mechanical to stir evenly mixes.
Further, in step (3), the consisting of of reverse-extraction agent:
The concentration of free ammonia 60～100g/L
Volatile salt concentration 150～210g/L;
Reverse-extraction agent adopts and adds the solvent of regulation proportional concentration and evenly mix in the aqueous solution.
Further, in step (3), the processing condition of extractive reaction are: compare O/A=1～1.5:1,40～50 ℃ of temperature, time 2-3min, 2～3 grades of progression.
Further, in step (3), the processing condition of back extraction are: compare O/A=1.5～1.8:1,40～50 ℃ of temperature, time 2-3min, 2～3 grades of progression.
Principle of the present invention is: adopt full wet processing, with vitriol oil Leaching Zinc scum silica frost mine tailing, zinc wherein is all dissolved in solution, press filtration obtains the leach liquor that contains Zn, chlorine and a small amount of impurity iron.Leach liquor and dechlorination extraction agent extract dechlorination, obtain the raffinate of the high zinc of low chlorine, return to zinc and smelt host system for the production of cathode zinc after gac oil removal treatment; Harmful element chlorine in zinc dross mine tailing is reclaimed with byproduct ammonium chloride form.
Principal reaction equation of the present invention is as follows:
Zn+H
2SO
4=ZnSO
4+H
2↑
ZnCl
2(S)=ZnCl
2(l)
ZnO+H
2SO
4=ZnSO
4+H
2O
NH
4Cl=NH
4 ++Cl
-
Fe
2O
3+H
2SO
4=Fe
2(SO
4)
3+H
2O
2R
3N+H
2SO
4=（R
3NH）
2SO
4
（R
3NH）
2SO
4+ZnCl
2=2R
3NHCl+ZnSO
4
2R
3NHCl+(NH
4)
2CO
3=2R
3N+2NH
4Cl+H
2O+CO
2↑
The invention has the beneficial effects as follows: reclaimed to greatest extent the zinc in zinc dross mine tailing, harmful element chlorine is recycled with byproduct form, has realized the target of turning bane into boon; Two sections that in smelting process, produce are leached the quantity of slag seldom, be about 2.8～3.5% of raw material zinc dross mine tailing total amount, main component is still zinc, grade reaches 50% left and right, containing a small amount of chlorine (being about 0.15～0.2%), can be used as food ingredient and add and in zinc sulfide concentrates, carry out roasting, thereby this technique do not have solid waste to produce, can not produce the problem that waste is stored up; Leach liquor and dechlorination extraction agent extract dechlorination, strip liquor is cycled to used in back extraction by supplementing solute concentration, and the solution crystallization output byproduct of regular extracting part partial volume, can realize closed circuit operation and the recycle of liquid, there is no waste water, exhaust gas emission, realize cleaner production requirement, solved the technical barrier of long-term puzzlement industry.
Accompanying drawing explanation
Fig. 1 is process flow sheet of the present invention.
Embodiment
Below principle of the present invention and feature are described, illustrated embodiment, only for explaining the present invention, is not intended to limit scope of the present invention.
Embodiment 1: under the condition that is 6:1 in liquid-solid ratio by zinc dross mine tailing, adding the activator vitriol oil is 180 g/L to sulfuric acid starting point concentration in solution, and 50 ℃ of extraction temperatures stir and react and after 40 minutes, carry out one section of leaching; Filtered filtration residue is proceeded two sections of leachings, and control liquid-solid ratio is 4:1, adds the activator vitriol oil to starting point concentration 250g/L, 80 ℃ of extraction temperatures, 60 minutes reaction times.It is 50ppm that reaction end adds 0.2% 3# flocculation agent (polyacrylamide) to concentration, filtered while hot; By analysis, in one section of leach liquor, zinc concentration is 125g/L, and chlorine is 1.85 g/L; Two sections of leached mud productive rates are 3.25%, containing zinc 53.4%, add in zinc sulfide concentrates, carry out roasting using two sections of leached muds as food ingredient.Total leaching yield of zinc is 97.7%.
Embodiment 2: under the condition that is 8:1 in liquid-solid ratio by zinc dross mine tailing, add the activator vitriol oil to starting point concentration 200 g/L, and 55 ℃ of extraction temperatures, 60 minutes stirring reaction time was carried out one section of leaching; Filtered filtration residue is proceeded two sections of leachings, and control liquid-solid ratio is 5:1, adds the activator vitriol oil to starting point concentration 350g/L, 90 ℃ of extraction temperatures, 90 minutes reaction times.It is 60ppm that reaction end adds 0.2% 3# flocculation agent to concentration, filtered while hot; By analysis, in one section of leach liquor, zinc concentration is 102g/L, and chlorine is 1.42 g/L; Two sections of leached mud productive rates are 2.80%, containing zinc 50.3%.Total leaching yield of zinc is 98.2%.
Embodiment 3: under the condition that is 7:1 in liquid-solid ratio by zinc dross mine tailing, add the activator vitriol oil to starting point concentration 250 g/L, and 60 ℃ of extraction temperatures, 50 minutes stirring reaction time was carried out one section of leaching; Filtered filtration residue is proceeded two sections of leachings, and control liquid-solid ratio is 4.5:1, adds the activator vitriol oil to starting point concentration 300g/L, 90 ℃ of extraction temperatures, 80 minutes reaction times.It is 80ppm that reaction end adds 0.2% 3# flocculation agent to concentration, filtered while hot; By analysis, in one section of leach liquor, zinc concentration is 118g/L, and chlorine is 1.64 g/L; Two sections of leached mud productive rates are 3.20%, containing zinc 52.8%.Total leaching yield of zinc is 97.9%.
Embodiment 4: under the condition that is 6:1 in liquid-solid ratio by zinc dross mine tailing, add the activator vitriol oil to starting point concentration 220 g/L, and 60 ℃ of extraction temperatures, 60 minutes stirring reaction time was carried out one section of leaching; Filtered filtration residue is proceeded two sections of leachings, and control liquid-solid ratio is 5:1, adds the activator vitriol oil to starting point concentration 300g/L, 85 ℃ of extraction temperatures, 75 minutes reaction times.It is 50ppm that reaction end adds 0.2% 3# flocculation agent to concentration, filtered while hot; By analysis, in one section of leach liquor, zinc concentration is 131g/L, and chlorine is 1.85 g/L; Two sections of leached mud productive rates are 3.15%, containing zinc 51.2%.Total leaching yield of zinc is 98.0%.
Embodiment 5: the one section of leach liquor of take is subjects, includes zinc 118.6 g/L, chlorine 1.69 g/L, sulfuric acid 62 g/L; With consisting of sulfonated kerosene 60%, R
3n35%, the extracted organic phase of secondary octanol 8% is carried out counter-current extraction dechlorination, and extraction phase is than O/A=1/1,40 ℃ of temperature, time 2min, 3 grades of progression; Load organic phases the concentration of free ammonia 100g/L, the reverse-extraction agent of volatile salt concentration 200g/L is by comparing O/A=1.5/1,40 ℃ of temperature, time 2min, 3 grades of progression carry out back extraction dechlorination.By analysis, in dechlorination raffinate, zinc concentration is 117.4g/L, and chlorine is 0.042g/L; In strip liquor, zinc concentration is 0.065g/L, and chlorine is 2.42g/L.The extraction loss that calculates zinc is 1.01%, and the percentage extraction of chlorine and back extraction ratio are respectively 97.51% and 97.90%.
Embodiment 6: the one section of leach liquor of take is subjects, includes zinc 118.6 g/L, chlorine 1.69 g/L, sulfuric acid 62 g/L; With consisting of sulfonated kerosene 70%, R
3n20%, the extracted organic phase of secondary octanol 10% is carried out counter-current extraction dechlorination, and extraction phase is than O/A=1.2/1, temperature 45 C, time 3min, 3 grades of progression; Load organic phases the concentration of free ammonia 60g/L, the reverse-extraction agent of volatile salt concentration 150g/L is by comparing O/A=1.6/1,40 ℃ of temperature, time 3min, 3 grades of progression carry out back extraction dechlorination.By analysis, in dechlorination raffinate, zinc concentration is 117.8g/L, and chlorine is 0.13g/L; In strip liquor, zinc concentration is 0.36g/L, and chlorine is 1.86g/L.The extraction loss that calculates zinc is 0.67%, and the percentage extraction of chlorine and back extraction ratio are respectively 92.31% and 79.49%.
Embodiment 7: the one section of leach liquor of take is subjects, includes zinc 118.6 g/L, chlorine 1.69 g/L, sulfuric acid 62 g/L; With consisting of sulfonated kerosene 75%, R
3n25%, the extracted organic phase of secondary octanol 5% is carried out counter-current extraction dechlorination, and extraction phase is than O/A=1/1,40 ℃ of temperature, time 2min, 2 grades of progression; Load organic phases the concentration of free ammonia 80g/L, the reverse-extraction agent of volatile salt concentration 180g/L is by comparing O/A=1.8/1, temperature 45 C, time 2min, 2 grades of progression carry out back extraction dechlorination.By analysis, in dechlorination raffinate, zinc concentration is 117.6g/L, and chlorine is 0.066g/L; In strip liquor, zinc concentration is 0.45g/L, and chlorine is 2.78g/L.The extraction loss that calculates zinc is 0.84%, and the percentage extraction of chlorine and back extraction ratio are respectively 96.09% and 95.10%.
Embodiment 8: the one section of leach liquor of take is subjects, includes zinc 118.6 g/L, chlorine 1.69 g/L, sulfuric acid 62 g/L; With consisting of sulfonated kerosene 65%, R
3n30%, the extracted organic phase of secondary octanol 5% is carried out counter-current extraction dechlorination, and extraction phase is than O/A=1.5/1, temperature 50 C, time 2min, 2 grades of progression; Load organic phases the concentration of free ammonia 100g/L, the reverse-extraction agent of volatile salt concentration 150g/L is by comparing O/A=1.8/1,40 ℃ of temperature, time 2min, 2 grades of progression carry out back extraction dechlorination.By analysis, in dechlorination raffinate, zinc concentration is 117.8g/L, and chlorine is 0.049g/L; In strip liquor, zinc concentration is 0.48g/L, and chlorine is 2.82g/L.The extraction loss that calculates zinc is 0.67%, and the percentage extraction of chlorine and back extraction ratio are respectively 97.10% and 95.47%.
The foregoing is only preferred embodiment of the present invention, not in order to limit the present invention, all any modifications of doing within the spirit and principles in the present invention, be equal to replacement, improvement etc., within all should being included in protection scope of the present invention.
Claims (8)
1. utilize founding zinc dross mine tailing to produce a cathode zinc technique, it is characterized in that, comprise following processing step:
(1) one section of leaching: will add in one section of pickling tank after the pulp of zinc dross mine tailing water, then add sulfuric acid and two sections of leach liquors, the stirring of heating, control liquid-solid ratio is 6～8:1, it is 180～250g/L that control sulphuric acid makes sulfuric acid starting point concentration in solution, 50～60 ℃ of temperature of reaction, 40～60 minutes reaction times, filter, obtain one section of leach liquor and one section of leached mud;
(2) two sections of leachings: will join in two-stage acid leaching groove after the pulp of one section of leached mud water, then add sulfuric acid, the stirring of heating, control liquid-solid ratio is 4～5:1, and it is 250～350g/L that control sulphuric acid makes sulfuric acid starting point concentration in solution, 80～90 ℃ of temperature of reaction, 60～90 minutes reaction times, filtered while hot, obtains two sections of leach liquors and two sections of leached muds, and two sections of leach liquors return and enter in one section of pickling tank;
(3) chloride one section is leached solution of zinc sulfate extraction dechlorination: one section of leach liquor is pumped into and mixes clarification extraction box, with the raw extractive reaction of the dechlorination extraction agent organic phase of countercurrent flow successively hybrid concurrency, raffinate after dechlorination carries out standing and passes through activated carbon adsorption oil removing, then raffinate after dechlorination is incorporated to zinc smelting system for the production of cathode zinc.
2. a kind of founding zinc dross mine tailing that utilizes according to claim 1 is produced cathode zinc technique, it is characterized in that, in step (2), it is 50-80ppm that reaction end adds flocculation agent to concentration.
3. a kind of founding zinc dross mine tailing that utilizes according to claim 1 is produced cathode zinc technique, it is characterized in that, in step (2), adds in zinc sulfide concentrates, carry out roasting recovery using two sections of leached muds as food ingredient.
4. a kind of founding zinc dross mine tailing that utilizes according to claim 1 is produced cathode zinc technique, it is characterized in that, in step (3), after extractive reaction finishes, carrying chlorine organic phase and reverse-extraction agent ammonia-sal volatile carries out countercurrent reextraction and is regenerated, after back extraction, solution is again for back extraction after supplementing reverse-extraction agent solute concentration, and regularly after the back extraction of extracting part partial volume, solution carries out crystallization output byproduct ammonium chloride.
5. a kind of founding zinc dross mine tailing that utilizes according to claim 1 is produced cathode zinc technique, it is characterized in that, and in step (3), the consisting of of extraction agent organic phase (V%):
Sulfonated kerosene 60～75%
R
3N 20～35%
Secondary octanol 5～10%.
6. a kind of founding zinc dross mine tailing that utilizes according to claim 4 is produced cathode zinc technique, it is characterized in that, and in step (3), the consisting of of reverse-extraction agent:
The concentration of free ammonia 60～100g/L
Volatile salt concentration 150～210g/L.
7. a kind of founding zinc dross mine tailing that utilizes according to claim 1 is produced cathode zinc technique, it is characterized in that, in step (3), the processing condition of extractive reaction are: compare O/A=1～1.5:1,40～50 ℃ of temperature, time 2-3min, 2～3 grades of progression.
8. a kind of founding zinc dross mine tailing that utilizes according to claim 4 is produced cathode zinc technique, it is characterized in that, in step (3), the processing condition of back extraction are: compare O/A=1.5～1.8:1,40～50 ℃ of temperature, time 2-3min, 2～3 grades of progression.
Priority Applications (1)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| CN201410388483.7A CN104152705B (en) | 2014-08-08 | 2014-08-08 | Process for producing cathode zinc by utilizing casting zinc dross tailings |
Applications Claiming Priority (1)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| CN201410388483.7A CN104152705B (en) | 2014-08-08 | 2014-08-08 | Process for producing cathode zinc by utilizing casting zinc dross tailings |
Publications (2)
| Publication Number | Publication Date |
|---|---|
| CN104152705A true CN104152705A (en) | 2014-11-19 |
| CN104152705B CN104152705B (en) | 2016-08-17 |
Family
ID=51878328
Family Applications (1)
| Application Number | Title | Priority Date | Filing Date |
|---|---|---|---|
| CN201410388483.7A Active CN104152705B (en) | 2014-08-08 | 2014-08-08 | Process for producing cathode zinc by utilizing casting zinc dross tailings |
Country Status (1)
| Country | Link |
|---|---|
| CN (1) | CN104152705B (en) |
Citations (6)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN101560670A (en) * | 2009-05-20 | 2009-10-21 | 北京矿冶研究总院 | High nickel matte refining method |
| CN101643857A (en) * | 2009-08-25 | 2010-02-10 | 西部矿业股份有限公司 | Comprehensive recovery method of complex polymetal sulphide ore containing copper, lead and zinc |
| CN101818250A (en) * | 2009-02-27 | 2010-09-01 | 浙江华友钴业股份有限公司 | Method for processing cobalt-copper-iron alloy |
| CN101994008A (en) * | 2010-11-11 | 2011-03-30 | 西北矿冶研究院 | New process for separating zinc and cobalt from nickel and cobalt slag produced by zinc smelting and purification |
| CN102433439A (en) * | 2011-12-16 | 2012-05-02 | 湖南有色金属研究院 | Method for recovering rhenium from arsenic filter cake |
| CN102649998A (en) * | 2012-05-28 | 2012-08-29 | 陕西安圣美贸易发展有限公司 | Method for separating and enriching copper, cadmium and zinc in copper and cadmium sediment |
-
2014
- 2014-08-08 CN CN201410388483.7A patent/CN104152705B/en active Active
Patent Citations (6)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN101818250A (en) * | 2009-02-27 | 2010-09-01 | 浙江华友钴业股份有限公司 | Method for processing cobalt-copper-iron alloy |
| CN101560670A (en) * | 2009-05-20 | 2009-10-21 | 北京矿冶研究总院 | High nickel matte refining method |
| CN101643857A (en) * | 2009-08-25 | 2010-02-10 | 西部矿业股份有限公司 | Comprehensive recovery method of complex polymetal sulphide ore containing copper, lead and zinc |
| CN101994008A (en) * | 2010-11-11 | 2011-03-30 | 西北矿冶研究院 | New process for separating zinc and cobalt from nickel and cobalt slag produced by zinc smelting and purification |
| CN102433439A (en) * | 2011-12-16 | 2012-05-02 | 湖南有色金属研究院 | Method for recovering rhenium from arsenic filter cake |
| CN102649998A (en) * | 2012-05-28 | 2012-08-29 | 陕西安圣美贸易发展有限公司 | Method for separating and enriching copper, cadmium and zinc in copper and cadmium sediment |
Also Published As
| Publication number | Publication date |
|---|---|
| CN104152705B (en) | 2016-08-17 |
Similar Documents
| Publication | Publication Date | Title |
|---|---|---|
| CN102312083A (en) | Method for extracting zinc indium and recovering iron from high-iron high indium zinc concentrate | |
| US8906220B2 (en) | Method for production of metallic cobalt from the nickel solvent extraction raffinate | |
| JP6176491B2 (en) | Method for removing copper from aqueous nickel chloride solution | |
| CN104060106B (en) | From bismuth-containing solution, extract bismuth with solvent extration and prepare the method for bismuth oxide | |
| EP2147128B1 (en) | Process for producing pure metallic indium from zinc oxide and/or solution containing the metal | |
| CN105339511A (en) | Method of recovering copper and precious metals | |
| CN104480325A (en) | Method for extracting cobalt from cobalt-containing raw material | |
| Agrawal et al. | Extractive separation of copper and nickel from copper bleed stream by solvent extraction route | |
| CN101338375A (en) | Technological process for extracting nickel and cobalt and comprehensively developing iron and magnesium from laterite-nickel ore | |
| CN106337134A (en) | Technique for recovering indium from indium-containing soot | |
| CN103243349A (en) | Comprehensive zinc hydrometallurgy recovery system technique | |
| CN112458280A (en) | Method for extracting valuable metals by leaching low grade nickel matte with acidic etching solution | |
| CN106222454A (en) | A kind of from the method containing indium flue dust recovery indium | |
| US8911533B2 (en) | Method of recovering gold from dilute gold solution | |
| CN104212976A (en) | Method for recovering silver from wet zinc smelting waste slag | |
| CN104164567A (en) | Method for enriching and recycling niobium and tantalum from waste high-temperature alloy | |
| CN102226236B (en) | Hydrometallurgical method for comprehensively recycling components in lateritic nickel ore as products | |
| CN102965520B (en) | Method for separating and enriching copper in leachate of acid leaching sulfide mine tailing | |
| CN103937975B (en) | The method of extracting directly silver from zinc hydrometallurgy flotation of silver concentrate | |
| EP1258535A1 (en) | Process for electrolytic production of highly pure zinc or zinc compounds from primary and secondary zinc raw materials | |
| CN101798639A (en) | Method for recovering indium from indium-containing neutralized dregs directly leached from zinc concentrate | |
| JP2008208441A (en) | Solvent extraction method for chloride aqueous solution | |
| CN105154681A (en) | Process for recovering fluorine and zinc from steelmaking soot | |
| CN105886797A (en) | Method for preparing sponge indium from polymetallic sulfide material | |
| CN114657387A (en) | Method for recovering copper by synergistic treatment of acidic etching waste liquid and zinc-containing copper dust |
Legal Events
| Date | Code | Title | Description |
|---|---|---|---|
| C06 | Publication | ||
| PB01 | Publication | ||
| C10 | Entry into substantive examination | ||
| SE01 | Entry into force of request for substantive examination | ||
| C14 | Grant of patent or utility model | ||
| GR01 | Patent grant |