CN115418486A - Method for recovering cobalt and manganese in zinc purification slag by combining acid leaching-precipitation flotation method - Google Patents
Method for recovering cobalt and manganese in zinc purification slag by combining acid leaching-precipitation flotation method Download PDFInfo
- Publication number
- CN115418486A CN115418486A CN202211158120.5A CN202211158120A CN115418486A CN 115418486 A CN115418486 A CN 115418486A CN 202211158120 A CN202211158120 A CN 202211158120A CN 115418486 A CN115418486 A CN 115418486A
- Authority
- CN
- China
- Prior art keywords
- leaching
- manganese
- cobalt
- solution
- zinc
- 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
- 239000011572 manganese Substances 0.000 title claims abstract description 57
- 229910017052 cobalt Inorganic materials 0.000 title claims abstract description 55
- 239000010941 cobalt Substances 0.000 title claims abstract description 55
- GUTLYIVDDKVIGB-UHFFFAOYSA-N cobalt atom Chemical compound [Co] GUTLYIVDDKVIGB-UHFFFAOYSA-N 0.000 title claims abstract description 55
- 229910052748 manganese Inorganic materials 0.000 title claims abstract description 55
- 238000000034 method Methods 0.000 title claims abstract description 55
- 238000005188 flotation Methods 0.000 title claims abstract description 51
- 239000002893 slag Substances 0.000 title claims abstract description 47
- 239000002253 acid Substances 0.000 title claims abstract description 45
- 238000001556 precipitation Methods 0.000 title claims abstract description 43
- 239000011701 zinc Substances 0.000 title claims abstract description 43
- 229910052725 zinc Inorganic materials 0.000 title claims abstract description 43
- HCHKCACWOHOZIP-UHFFFAOYSA-N Zinc Chemical compound [Zn] HCHKCACWOHOZIP-UHFFFAOYSA-N 0.000 title claims abstract description 38
- PWHULOQIROXLJO-UHFFFAOYSA-N Manganese Chemical compound [Mn] PWHULOQIROXLJO-UHFFFAOYSA-N 0.000 title claims abstract description 37
- 238000000746 purification Methods 0.000 title claims abstract description 30
- 238000002386 leaching Methods 0.000 claims abstract description 132
- QAOWNCQODCNURD-UHFFFAOYSA-N Sulfuric acid Chemical compound OS(O)(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-N 0.000 claims abstract description 79
- MZZUATUOLXMCEY-UHFFFAOYSA-N cobalt manganese Chemical compound [Mn].[Co] MZZUATUOLXMCEY-UHFFFAOYSA-N 0.000 claims abstract description 28
- 239000006260 foam Substances 0.000 claims abstract description 20
- 239000004094 surface-active agent Substances 0.000 claims abstract description 19
- 230000008569 process Effects 0.000 claims abstract description 17
- 230000001376 precipitating effect Effects 0.000 claims abstract description 4
- 230000003311 flocculating effect Effects 0.000 claims abstract description 3
- 239000003795 chemical substances by application Substances 0.000 claims description 31
- 238000006243 chemical reaction Methods 0.000 claims description 24
- 239000007787 solid Substances 0.000 claims description 22
- 239000008394 flocculating agent Substances 0.000 claims description 20
- 229920002401 polyacrylamide Polymers 0.000 claims description 20
- 230000009467 reduction Effects 0.000 claims description 18
- 239000003638 chemical reducing agent Substances 0.000 claims description 16
- YXAOOTNFFAQIPZ-UHFFFAOYSA-N 1-nitrosonaphthalen-2-ol Chemical compound C1=CC=CC2=C(N=O)C(O)=CC=C21 YXAOOTNFFAQIPZ-UHFFFAOYSA-N 0.000 claims description 10
- KRKNYBCHXYNGOX-UHFFFAOYSA-N citric acid Chemical compound OC(=O)CC(O)(C(O)=O)CC(O)=O KRKNYBCHXYNGOX-UHFFFAOYSA-N 0.000 claims description 9
- LZZYPRNAOMGNLH-UHFFFAOYSA-M Cetrimonium bromide Chemical group [Br-].CCCCCCCCCCCCCCCC[N+](C)(C)C LZZYPRNAOMGNLH-UHFFFAOYSA-M 0.000 claims description 8
- 239000012716 precipitator Substances 0.000 claims description 7
- MHAJPDPJQMAIIY-UHFFFAOYSA-N Hydrogen peroxide Chemical compound OO MHAJPDPJQMAIIY-UHFFFAOYSA-N 0.000 claims description 6
- 229910001429 cobalt ion Inorganic materials 0.000 claims description 6
- XLJKHNWPARRRJB-UHFFFAOYSA-N cobalt(2+) Chemical compound [Co+2] XLJKHNWPARRRJB-UHFFFAOYSA-N 0.000 claims description 6
- 230000002378 acidificating effect Effects 0.000 claims description 4
- 238000003723 Smelting Methods 0.000 claims description 3
- CZMRCDWAGMRECN-UGDNZRGBSA-N Sucrose Chemical compound O[C@H]1[C@H](O)[C@@H](CO)O[C@@]1(CO)O[C@@H]1[C@H](O)[C@@H](O)[C@H](O)[C@@H](CO)O1 CZMRCDWAGMRECN-UGDNZRGBSA-N 0.000 claims description 3
- 229930006000 Sucrose Natural products 0.000 claims description 3
- 239000005720 sucrose Substances 0.000 claims description 3
- 238000000926 separation method Methods 0.000 abstract description 26
- 238000011084 recovery Methods 0.000 abstract description 15
- 238000009854 hydrometallurgy Methods 0.000 abstract description 8
- 230000007613 environmental effect Effects 0.000 abstract description 3
- HEMHJVSKTPXQMS-UHFFFAOYSA-M Sodium hydroxide Chemical compound [OH-].[Na+] HEMHJVSKTPXQMS-UHFFFAOYSA-M 0.000 description 18
- 239000002245 particle Substances 0.000 description 18
- 239000003814 drug Substances 0.000 description 17
- 229910052751 metal Inorganic materials 0.000 description 16
- 239000002184 metal Substances 0.000 description 16
- 238000002156 mixing Methods 0.000 description 16
- 150000002739 metals Chemical class 0.000 description 15
- 238000010907 mechanical stirring Methods 0.000 description 14
- 239000000203 mixture Substances 0.000 description 14
- 239000007788 liquid Substances 0.000 description 12
- 238000003756 stirring Methods 0.000 description 11
- 229910000914 Mn alloy Inorganic materials 0.000 description 10
- 239000000047 product Substances 0.000 description 8
- 238000003760 magnetic stirring Methods 0.000 description 7
- 238000003828 vacuum filtration Methods 0.000 description 7
- 238000001035 drying Methods 0.000 description 6
- 238000002474 experimental method Methods 0.000 description 6
- 238000000227 grinding Methods 0.000 description 6
- 230000007935 neutral effect Effects 0.000 description 6
- 238000012216 screening Methods 0.000 description 6
- 238000005406 washing Methods 0.000 description 6
- 239000003153 chemical reaction reagent Substances 0.000 description 5
- 238000001914 filtration Methods 0.000 description 5
- 238000010438 heat treatment Methods 0.000 description 5
- 230000002411 adverse Effects 0.000 description 3
- 230000000052 comparative effect Effects 0.000 description 3
- 230000000694 effects Effects 0.000 description 3
- 239000002244 precipitate Substances 0.000 description 3
- 230000004075 alteration Effects 0.000 description 2
- 230000009286 beneficial effect Effects 0.000 description 2
- 230000001276 controlling effect Effects 0.000 description 2
- 238000004519 manufacturing process Methods 0.000 description 2
- 238000012986 modification Methods 0.000 description 2
- 230000004048 modification Effects 0.000 description 2
- 239000002994 raw material Substances 0.000 description 2
- 238000000638 solvent extraction Methods 0.000 description 2
- 239000000126 substance Substances 0.000 description 2
- 239000002699 waste material Substances 0.000 description 2
- CWYNVVGOOAEACU-UHFFFAOYSA-N Fe2+ Chemical compound [Fe+2] CWYNVVGOOAEACU-UHFFFAOYSA-N 0.000 description 1
- UCKMPCXJQFINFW-UHFFFAOYSA-N Sulphide Chemical compound [S-2] UCKMPCXJQFINFW-UHFFFAOYSA-N 0.000 description 1
- 239000003513 alkali Substances 0.000 description 1
- 229910045601 alloy Inorganic materials 0.000 description 1
- 239000000956 alloy Substances 0.000 description 1
- IDKDXKUZTLIGSY-UHFFFAOYSA-N cobalt;1-nitrosonaphthalen-2-ol Chemical compound [Co].C1=CC=CC2=C(N=O)C(O)=CC=C21 IDKDXKUZTLIGSY-UHFFFAOYSA-N 0.000 description 1
- 150000001875 compounds Chemical class 0.000 description 1
- 230000007547 defect Effects 0.000 description 1
- 238000011161 development Methods 0.000 description 1
- 238000010586 diagram Methods 0.000 description 1
- 238000004070 electrodeposition Methods 0.000 description 1
- 238000005516 engineering process Methods 0.000 description 1
- ZOOODBUHSVUZEM-UHFFFAOYSA-N ethoxymethanedithioic acid Chemical compound CCOC(S)=S ZOOODBUHSVUZEM-UHFFFAOYSA-N 0.000 description 1
- 238000000605 extraction Methods 0.000 description 1
- 238000005189 flocculation Methods 0.000 description 1
- 238000009291 froth flotation Methods 0.000 description 1
- 230000007062 hydrolysis Effects 0.000 description 1
- 238000006460 hydrolysis reaction Methods 0.000 description 1
- XLYOFNOQVPJJNP-UHFFFAOYSA-M hydroxide Chemical compound [OH-] XLYOFNOQVPJJNP-UHFFFAOYSA-M 0.000 description 1
- 230000008676 import Effects 0.000 description 1
- 239000012535 impurity Substances 0.000 description 1
- 229910052500 inorganic mineral Inorganic materials 0.000 description 1
- 229910001437 manganese ion Inorganic materials 0.000 description 1
- WPBNNNQJVZRUHP-UHFFFAOYSA-L manganese(2+);methyl n-[[2-(methoxycarbonylcarbamothioylamino)phenyl]carbamothioyl]carbamate;n-[2-(sulfidocarbothioylamino)ethyl]carbamodithioate Chemical compound [Mn+2].[S-]C(=S)NCCNC([S-])=S.COC(=O)NC(=S)NC1=CC=CC=C1NC(=S)NC(=O)OC WPBNNNQJVZRUHP-UHFFFAOYSA-L 0.000 description 1
- 239000000463 material Substances 0.000 description 1
- 229910021645 metal ion Inorganic materials 0.000 description 1
- 238000005272 metallurgy Methods 0.000 description 1
- 239000011707 mineral Substances 0.000 description 1
- 230000003647 oxidation Effects 0.000 description 1
- 238000007254 oxidation reaction Methods 0.000 description 1
- 230000035484 reaction time Effects 0.000 description 1
- 238000004064 recycling Methods 0.000 description 1
- 230000001105 regulatory effect Effects 0.000 description 1
- 238000006467 substitution reaction Methods 0.000 description 1
- 239000001117 sulphuric acid Substances 0.000 description 1
- 235000011149 sulphuric acid Nutrition 0.000 description 1
- 239000000725 suspension Substances 0.000 description 1
- 239000012991 xanthate Substances 0.000 description 1
Images
Classifications
-
- 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
-
- 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
- C22B23/00—Obtaining nickel or cobalt
- C22B23/04—Obtaining nickel or cobalt by wet processes
- C22B23/0407—Leaching processes
- C22B23/0415—Leaching processes with acids or salt solutions except ammonium salts solutions
- C22B23/043—Sulfurated acids or salts thereof
-
- 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
- C22B23/00—Obtaining nickel or cobalt
- C22B23/04—Obtaining nickel or cobalt by wet processes
- C22B23/0453—Treatment or purification of solutions, e.g. obtained by 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
- C22B23/00—Obtaining nickel or cobalt
- C22B23/04—Obtaining nickel or cobalt by wet processes
- C22B23/0453—Treatment or purification of solutions, e.g. obtained by leaching
- C22B23/0461—Treatment or purification of solutions, e.g. obtained by leaching by chemical methods
-
- 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
- C22B47/00—Obtaining manganese
-
- 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
- Chemical & Material Sciences (AREA)
- Engineering & Computer Science (AREA)
- Metallurgy (AREA)
- Manufacturing & Machinery (AREA)
- Materials Engineering (AREA)
- Mechanical Engineering (AREA)
- Organic Chemistry (AREA)
- Geology (AREA)
- General Life Sciences & Earth Sciences (AREA)
- Environmental & Geological Engineering (AREA)
- Life Sciences & Earth Sciences (AREA)
- Chemical Kinetics & Catalysis (AREA)
- General Chemical & Material Sciences (AREA)
- Manufacture And Refinement Of Metals (AREA)
Abstract
The invention discloses a method for recovering cobalt and manganese in zinc purification slag by combining an acid leaching-precipitation flotation method, which comprises the following steps: leaching the zinc purification slag by adopting sulfuric acid to obtain zinc-rich leachate and leaching slag; reducing and leaching the leaching residue by acid to obtain cobalt-manganese leaching solution; and sequentially precipitating and flocculating the cobalt-manganese leaching solution, adding a surfactant for foam flotation, and collecting cobalt-rich foam and a manganese-rich solution. The method has the advantages of short process flow, low operation difficulty, easily controlled technical conditions and environmental friendliness, can realize the high-efficiency separation and recovery of cobalt and manganese in the purified cobalt slag from the zinc hydrometallurgy, obviously improves the comprehensive utilization additional value of the purified slag, has strong adaptability to a main system of the zinc hydrometallurgy, and has good industrial application prospect.
Description
Technical Field
The invention relates to a treatment method of zinc hydrometallurgy purification slag, in particular to a method for recovering cobalt and manganese in zinc purification slag by combining an acid leaching-precipitation flotation method, and belongs to the technical field of comprehensive recovery and utilization of non-ferrous metal metallurgy.
Background
Cobalt and compounds thereof have wide application in the key fields of new energy materials, special performance alloys and the like. With the rapid development of intelligent equipment and new energy automobile industry, the demand of cobalt will be greater and greater. The cobalt resource in China is mainly associated minerals, the reserves only account for 1.1 percent of the reserves of the cobalt resource in the world, and the import dependency on the cobalt resource is higher. The recovery of the secondary resource containing cobalt can relieve the current situation of shortage of cobalt resources in China, improve the utilization efficiency of the resources and reduce the environmental impact caused by the traditional stacking or burning. In the process of zinc hydrometallurgy, in order to eliminate the adverse effect of impurity metal ions such as cobalt and the like in a solution on a subsequent electrodeposition step, a leachate needs to be purified, and common cobalt removal processes comprise zinc powder replacement, alpha-nitroso-beta-naphthol cobalt removal, xanthate cobalt removal and the like, and meanwhile, a large amount of purification residues containing valuable metals such as zinc, cobalt, manganese and the like are generated. The zinc in the purification slag can be recycled to a zinc hydrometallurgy system after treatment, and other valuable metals can also be recycled after separation and enrichment, so that the method has high recycling value.
At present, the recovery of the zinc hydrometallurgy purification slag is mainly to dissolve all or part of valuable metals into a solution through acid leaching or alkali leaching, and to separate and enrich the valuable metals in the solution. The common separation methods comprise a hydroxide precipitation method, a sulfide precipitation method, an oxidation precipitation method, a solvent extraction method and the like, but the precipitates are easy to carry with each other in the implementation process of various precipitation methods, so that the deep separation is difficult to carry out, and the solvent extraction method has good separation effect but high use cost of an extractant. For purified cobalt slag containing more valuable metals, a single separation method cannot usually enrich and recover all the valuable metals, and a combined process is adopted to gradually separate and enrich the valuable metals.
Disclosure of Invention
Based on the defects of poor selectivity to valuable metals, low separation efficiency and incomplete separation in the conventional wet-process zinc smelting purification slag recovery technology, the invention aims to provide a method for recovering cobalt and manganese in zinc purification slag by combining an acid leaching-precipitation flotation method. The method adopts a selective acid leaching-reduction acid leaching-precipitation flotation combined process to obtain a cobalt-rich foam product and a manganese-rich solution, has short process flow and low treatment cost, and can realize green high-value utilization of purified cobalt slag.
In order to achieve the aim, the invention provides a method for recovering cobalt and manganese in zinc purification slag by combining an acid leaching-precipitation flotation method, which comprises the following steps:
1) Leaching zinc smelting purification slag by using sulfuric acid to obtain a leaching solution and leaching slag;
2) Carrying out acid reduction leaching on the leaching residue to obtain cobalt-manganese leaching solution;
3) And sequentially precipitating and flocculating the cobalt-manganese leaching solution, adding a surfactant for foam flotation, and collecting cobalt-rich foam and a manganese-rich solution.
The invention can gradually enrich and recover three valuable metals, namely cobalt, manganese and zinc in the purified slag by a combined process of selective acid leaching, reduction acid leaching and precipitation flotation, and has simple operation and strong controllability, wherein the reaction principle in each step is as follows:
sulfuric acid leaching: zn + H 2 SO 4 ＝ZnSO 4 +H 2
ZnO+H 2 SO 4 ＝ZnSO 4 +H 2 O
CoO+H 2 SO 4 ＝CoSO 4 +H 2 O
Reduction and acid leaching: mnO 2 +H 2 SO 4 +H 2 O 2 ＝MnSO 4 +2H 2 O+O 2 ↑
Co 2 O 3 +2H 2 SO 4 +H 2 O 2 ＝2CoSO 4 +3H 2 O+O 2 ↑
Precipitation flotation: coSO 4 +Na 2 S＝CoS+Na 2 SO 4
Co 3+ +3C 10 H 6 ONOH＝CO(C 10 H 6 ONO) 3 ↓+3H +
As a preferable scheme, in the step 1), the leaching conditions are as follows: the concentration of the sulfuric acid is 2-4 mol/L, the liquid-solid ratio is 10-20 mL:1g, the temperature is 20-40 ℃, and the time is 10-30 min. The leaching process adopts mechanical stirring with the rotating speed of 200-400 rpm to assist in leaching. The cobalt-containing particle suspension is transferred to a micro-bubble flotation column for air flotation.
In the leaching reaction process, the concentration of sulfuric acid and the liquid-solid ratio need to be controlled within a proper range, wherein the leaching rate of zinc is low due to the over-low concentration of sulfuric acid or liquid-solid ratio, and selective leaching cannot be realized; too high a concentration of sulfuric acid will cause waste of reagents, while too high a liquid-solid ratio will reduce the enrichment degree of zinc. The leaching efficiency is also influenced by the reaction temperature and time, and the leaching rate of cobalt is increased due to overhigh temperature or overlong reaction time, so that the cobalt in the leaching slag is reduced.
As a preferred embodiment, the acid reduction leaching employs a leaching agent comprising sulphuric acid and a reducing agent. The reducing agent is at least one of hydrogen peroxide, sucrose and citric acid. Hydrogen peroxide, sucrose, citric acid, etc. are used as reducing agents, and Co can be added 3+ Reduction to Co 2+ ，Mn 4+ Reduction to Mn 2+ Thus dissolving cobalt and manganese in the solution and facilitating the subsequent separation of cobalt and manganese.
As a preferable scheme, in the step 2), the conditions of the acidic reduction leaching are as follows: the concentration of sulfuric acid in the leaching agent is 0.6-1.4 mol/L, the concentration of a reducing agent is 0.8-1.2 mol/L, and the liquid-solid ratio is 8-12 mL:1g of the total weight of the composition.
As a preferable scheme, in the step 2), the conditions of the acidic reduction leaching are as follows: the temperature is 20-40 ℃, and the time is 10-30 min. And when the leaching slag is leached by adopting a reducing acid solution, mechanical stirring with the rotating speed of 200-400 rpm is adopted for assisting leaching.
In the reduction leaching reaction process, the selective leaching efficiency can be improved by controlling the concentration of sulfuric acid, the concentration (addition amount) of a reducing agent and the liquid-solid ratio in appropriate ranges. The leaching rate of cobalt and manganese is low due to the over-low concentration of sulfuric acid, the concentration (addition amount) of a reducing agent or the liquid-solid ratio; too high concentration of sulfuric acid or reducing agent (addition amount) results in waste of reagents and increase of production cost, while too high liquid-solid ratio reduces the enrichment degree of cobalt and manganese.
As a preferred scheme, the precipitating agent is an agent A, and the agent A contains alpha-nitroso-beta-naphthol and Na 2 S, alpha-nitroso-beta-naphthol and Na 2 The mol ratio of S is 1-3: 1. the dosage of the precipitator is 0.8 to 1.2 times of the molar weight of the cobalt ions in the solution.
Cobalt ions in the solution can be selectively precipitated by using the precipitant component. In the selective precipitation reaction process, the dosage of the precipitator needs to be controlled within a proper range, incomplete cobalt ion precipitation reaction can be caused by too little dosage of the precipitator, and precipitation can be generated by reaction of excessive precipitator and manganese and selective precipitation can not be realized by too much dosage of the precipitator.
As a preferred embodiment, the flocculant is polyacrylamide. The concentration of the flocculant in the solution is 10-100 mg/L. In the flocculation process, magnetic stirring with the rotating speed of 200-400 rpm is adopted to assist leaching.
The flocculant is adopted to lead the cobalt precipitated particles to be aggregated, and simultaneously, the subsequent flotation separation efficiency can be improved by controlling the flocculant in a proper range. If the addition amount of the flocculating agent is too small (the concentration is too low), precipitate particles cannot be aggregated, and the subsequent flotation recovery rate is adversely affected; if the addition amount of the flocculating agent is too large (the concentration is too high), the reagent is wasted, and the production cost is increased.
As a preferred embodiment, the surfactant is CTAB. The concentration of the surfactant in the solution is 10-50 mg/L.
The foam stability can be enhanced and the floatability of the particles can be improved by adding the surfactant. The concentration of the surfactant in the solution is regulated and controlled within a proper range, so that the efficient separation of cobalt and manganese is facilitated, and the poor floatability of particles caused by the low addition amount (low concentration) of the surfactant affects the flotation recovery rate; excessive addition of surfactant (too high concentration) can result in surfactant residue in the solution, which can affect separation efficiency.
As a preferable scheme, in the step 3), the precipitation reaction conditions are: the pH of the solution system is 2-4. The temperature is 20-40 ℃, and the time is 10-30 min.
In the selective precipitation reaction process, the pH value of the solution system is controlled in a proper range, which is beneficial to the efficient separation and recovery of cobalt and manganese, the hydrolysis and precipitation of cobalt ions and manganese ions can be caused by the overhigh pH value of the solution system, and the adverse effect on the subsequent precipitation flotation reaction can be caused by the overlow pH value of the solution system.
Compared with the prior art, the invention has the following beneficial effects:
(1) In the method, firstly, sulfuric acid is adopted to selectively leach zinc, so that more than 95% of zinc is leached into a leaching solution, and the leaching solution can return to a zinc hydrometallurgy main system to recover zinc; secondly, the leaching solution rich in cobalt and manganese is obtained by acid reduction leaching, cobalt ions are selectively precipitated by adding a specific precipitator, and cobalt-containing precipitate particles and a manganese-rich solution are completely separated after froth flotation by adding a flocculating agent and a surfactant, so that the high-efficiency separation and high-value recovery of zinc, cobalt and manganese in the purified slag are gradually realized.
(2) The method has the advantages of short process flow, low operation difficulty, easily controlled technical conditions, environmental friendliness, strong adaptability to a zinc hydrometallurgy main system and good industrial application prospect.
Drawings
FIG. 1 is a schematic process flow diagram of the present invention.
Detailed Description
The present invention is further illustrated by the following examples and the accompanying drawings, but the present invention is not limited thereto in any way, and any modifications or alterations based on the teaching of the present invention are within the scope of the present invention.
Unless otherwise specifically stated, various raw materials, reagents, instruments, equipment and the like used in the present invention are commercially available or can be prepared by existing methods.
The main chemical components of the purified slag used in the examples are shown in table 1:
TABLE 1 main chemical composition of the purification slag
Example 1
The embodiment provides a method for recovering cobalt and manganese in zinc purification slag by combining an acid leaching-precipitation flotation method, which comprises the following steps:
(1) Selective acid leaching: crushing, grinding and screening the purification slag until the mass of particles with the particle size of less than 0.074mm accounts for more than 90%, and using the obtained sample for an acid leaching experiment; preparing a sulfuric acid leaching agent with the concentration of 4mol/L, and mixing the sulfuric acid leaching agent with the treated purified slag according to the liquid-solid ratio of 20mL: mixing 1g, heating to 30 deg.C, leaching under mechanical stirring at 300rpm for 10min, and vacuum filtering for solid-liquid separation to obtain leachate and leaching residue. The leaching rates of three metals of Co, mn and Zn are respectively 9.3%, 0.1% and 96.6%.
(2) Reduction and acid leaching: washing the leaching residue obtained in the step (1) to be neutral, drying, preparing a sulfuric acid leaching agent with the concentration of 1mol/L, and mixing the sulfuric acid leaching agent with the leaching residue according to the liquid-solid ratio of 10mL:1g of the mixture is mixed, heated to 30 ℃ and the reducing agent H is added 2 O 2 The concentration of the cobalt-manganese alloy in the solution is 1mol/L, the cobalt-manganese alloy is leached for 20min under the mechanical stirring of 300rpm, the solid-liquid separation is carried out by vacuum filtration, and the leaching rates of Co and Mn are respectively 99.3 percent and 98.7 percent, thus obtaining the cobalt-manganese solution.
(3) Precipitation reaction: adjusting the pH value of the cobalt-manganese solution obtained in the step (2) to 3 by using NaOH solution, adding a medicament A with the reaction amount 1.2 times of the theoretical reaction amount of cobalt, wherein the molar ratio of the medicament A to the reagent A is 1:1 alpha-nitroso-beta-naphthol and Na 2 S; the mixture was heated to 30 ℃ and reacted for 10min under magnetic stirring at 300rpm, and the Co and Mn precipitations were 96.3% and 5.4%, respectively. Then adding flocculating agent polyacrylamide to ensure that the concentration of the flocculating agent polyacrylamide in the solution is 50mg/L, adding surfactant CTAB after uniformly stirring to ensure that the concentration of the flocculating agent polyacrylamide in the solution is 50mg/L,and stirred uniformly.
(4) Foam flotation: and (4) transferring the solution system obtained in the step (3) to a micro-bubble flotation column for air flotation, and collecting cobalt-rich foam products and manganese-rich solutions, wherein the flotation recovery rates of Co and Mn are 92.5% and 2.7% respectively.
Example 2
(1) Selective acid leaching: crushing, grinding and screening the purification slag until the mass of particles with the particle size of less than 0.074mm accounts for more than 90%, and using the obtained sample for an acid leaching experiment; preparing a sulfuric acid leaching agent with the concentration of 2mol/L, and mixing the sulfuric acid leaching agent with the treated purified slag according to the liquid-solid ratio of 10mL: mixing 1g, heating to 40 deg.C, leaching under mechanical stirring at 300rpm for 20min, and vacuum filtering for solid-liquid separation to obtain leachate and leaching residue. The leaching rates of three metals of Co, mn and Zn are respectively 8.9%, 0.1% and 93.4%.
(2) Reduction and acid leaching: washing the leaching residue obtained in the step (1) to be neutral, drying, preparing a sulfuric acid leaching agent with the concentration of 0.8mol/L, and mixing the sulfuric acid leaching agent with the leaching residue according to the liquid-solid ratio of 8mL:1g of the mixture is mixed, heated to 40 ℃ and the reducing agent H is added 2 O 2 The concentration of the cobalt-manganese alloy in the solution is 0.8mol/L, the cobalt-manganese alloy is leached for 30min under the mechanical stirring of 300rpm, and the solid-liquid separation is carried out by vacuum filtration, wherein the leaching rates of Co and Mn are respectively 98.6% and 98.1%, and the cobalt-manganese solution is obtained.
(3) Precipitation reaction: adjusting the pH value of the cobalt-manganese solution obtained in the step (2) to 2.5 by using a NaOH solution, adding a medicament A with the theoretical reaction amount being 1 time, wherein the molar ratio of the medicament A to the medicament A is 3:1 alpha-nitroso-beta-naphthol and Na 2 S; the mixture was heated to 40 ℃ and reacted for 20min under magnetic stirring at 300rpm, and the Co and Mn precipitation rates were 95.2% and 3.5%, respectively. Then adding a flocculating agent polyacrylamide to ensure that the concentration of the polyacrylamide in the solution is 30mg/L, stirring uniformly, adding a surfactant CTAB to ensure that the concentration of the polyacrylamide in the solution is 30mg/L, and stirring uniformly.
(4) Foam flotation: and (4) transferring the solution system obtained in the step (3) to a micro-bubble flotation column for air flotation, and collecting cobalt-rich foam products and manganese-rich solutions, wherein the flotation recovery rates of Co and Mn are 92.3% and 2.4% respectively.
Example 3
(1) Selective acid leaching: crushing, grinding and screening the purification slag until the mass of particles with the particle size of less than 0.074mm accounts for more than 90%, and using the obtained sample for an acid leaching experiment; preparing a sulfuric acid leaching agent with the concentration of 2mol/L, and mixing the sulfuric acid leaching agent with the treated purified slag according to the liquid-solid ratio of 15mL: mixing 1g, heating to 35 deg.C, leaching under mechanical stirring at 300rpm for 15min, and vacuum filtering for solid-liquid separation to obtain leachate and leaching residue. The leaching rates of the three metals of Co, mn and Zn are respectively 9.1%, 0.1% and 95.8%.
(2) Reduction and acid leaching: washing the leaching residue obtained in the step (1) to be neutral, drying, preparing a sulfuric acid leaching agent with the concentration of 1.2mol/L, and mixing the sulfuric acid leaching agent with the leaching residue according to the liquid-solid ratio of 12mL:1g of the mixture is mixed, heated to 35 ℃ and the reducing agent H is added 2 O 2 The concentration of the cobalt-manganese alloy in the solution is 1.2mol/L, the cobalt-manganese alloy is leached for 25min under the mechanical stirring of 300rpm, and solid-liquid separation is carried out by vacuum filtration, wherein the leaching rates of Co and Mn are respectively 99.5% and 99.1%, and the cobalt-manganese solution is obtained.
(3) Precipitation reaction: adjusting the pH value of the cobalt-manganese solution obtained in the step (2) to 3.5 by using a NaOH solution, adding a medicament A with the theoretical reaction amount of 0.8 time, wherein the molar ratio of the medicament A to the medicament A is 2:1 alpha-nitroso-beta-naphthol and Na 2 S; the mixture was heated to 35 ℃ and reacted for 15min with magnetic stirring at 300rpm, and the Co and Mn precipitation rates were 94.6% and 3.2%, respectively. Then adding a flocculating agent polyacrylamide to ensure that the concentration of the flocculating agent polyacrylamide in the solution is 60mg/L, stirring uniformly, adding a surfactant CTAB to ensure that the concentration of the flocculating agent polyacrylamide in the solution is 40mg/L, and stirring uniformly.
(4) Foam flotation: and (4) transferring the solution system obtained in the step (3) to a micro-bubble flotation column for air flotation, and collecting cobalt-rich foam products and manganese-rich solutions, wherein the flotation recovery rates of Co and Mn are 91.7% and 2.1% respectively.
Comparative example 1
(1) Selective acid leaching: crushing, grinding and screening the purification slag until the mass of particles with the particle size of less than 0.074mm accounts for more than 90%, and using the obtained sample for an acid leaching experiment; preparing a sulfuric acid leaching agent with the concentration of 4mol/L, and mixing the sulfuric acid leaching agent with the treated purified slag according to the liquid-solid ratio of 20mL: mixing 1g, heating to 30 deg.C, leaching under mechanical stirring at 300rpm for 10min, and vacuum filtering for solid-liquid separation to obtain leachate and leaching residue. The leaching rates of three metals of Co, mn and Zn are respectively 9.3%, 0.1% and 96.6%.
(2) Reduction and acid leaching: washing the leaching residue obtained in the step (1) to be neutral, drying, preparing a sulfuric acid leaching agent with the concentration of 1mol/L, and mixing the sulfuric acid leaching agent with the leaching residue according to the liquid-solid ratio of 10mL:1g of the mixture is mixed, heated to 30 ℃ and the reducing agent H is added 2 O 2 The concentration of the cobalt-manganese alloy in the solution was adjusted to 0.1mol/L, the cobalt-manganese alloy was extracted for 20min under mechanical stirring at 300rpm, and solid-liquid separation was performed by vacuum filtration, whereby the extraction rates of Co and Mn were 67.3% and 51.7%, respectively, to obtain a cobalt-manganese solution.
(3) Precipitation reaction: adjusting the pH value of the cobalt-manganese solution obtained in the step (2) to 3 by using a NaOH solution, adding a medicament A with the theoretical reaction amount of 1.2 times, wherein the molar ratio of the medicament A to the medicament A is 1:1 alpha-nitroso-beta-naphthol and Na 2 S; the mixture was heated to 30 ℃ and reacted for 10min under magnetic stirring at 300rpm, and the Co and Mn precipitation rates were 97.2% and 7.3%, respectively. Then adding a flocculating agent polyacrylamide to ensure that the concentration of the flocculating agent polyacrylamide in the solution is 50mg/L, stirring uniformly, adding a surfactant CTAB to ensure that the concentration of the flocculating agent polyacrylamide in the solution is 50mg/L, and stirring uniformly.
(4) Foam flotation: and (4) transferring the solution system obtained in the step (3) to a micro-bubble flotation column for air flotation, and collecting cobalt-rich foam products and manganese-rich solutions, wherein the flotation recovery rates of Co and Mn are 90.2% and 4.3% respectively.
Comparative example 2
(1) Selective acid leaching: crushing, grinding and screening the purification slag until the mass of particles with the particle size of less than 0.074mm accounts for more than 90%, and using the obtained sample for an acid leaching experiment; preparing a sulfuric acid leaching agent with the concentration of 1mol/L, and mixing the sulfuric acid leaching agent with the treated purified slag according to the liquid-solid ratio of 10mL: mixing 1g, heating to 40 deg.C, leaching under mechanical stirring at 300rpm for 20min, and vacuum filtering for solid-liquid separation to obtain leachate and leaching residue. The leaching rates of three metals of Co, mn and Zn are respectively 8.9%, 0.1% and 93.4%.
(2) Reduction and acid leaching: washing the leaching residue obtained in the step (1) to be neutral, drying, and preparing the concentration0.8mol/L sulfuric acid leaching agent, and leaching residues are mixed according to the liquid-solid ratio of 8mL:1g of the mixture is heated to 40 ℃, and a reducing agent H is added 2 O 2 The concentration of the cobalt-manganese alloy in the solution is 0.8mol/L, the cobalt-manganese alloy is leached for 30min under the mechanical stirring of 300rpm, and the solid-liquid separation is carried out by vacuum filtration, wherein the leaching rates of Co and Mn are respectively 98.6% and 98.1%, and the cobalt-manganese solution is obtained.
(3) Precipitation reaction: adjusting the pH value of the cobalt-manganese solution obtained in the step (2) to 2.5 by using a NaOH solution, adding a medicament A with the theoretical reaction amount of 0.5 time, wherein the molar ratio of the medicament A to the medicament A is 3:1 alpha-nitroso-beta-naphthol and Na 2 S; the mixture was heated to 40 ℃ and reacted for 20min under magnetic stirring at 300rpm, and the Co and Mn precipitation rates were 47.3% and 0.6%, respectively. Then adding a flocculating agent polyacrylamide to ensure that the concentration of the flocculating agent polyacrylamide in the solution is 30mg/L, stirring uniformly, adding a surfactant CTAB to ensure that the concentration of the flocculating agent polyacrylamide in the solution is 30mg/L, and stirring uniformly.
(4) Foam flotation: and (4) transferring the solution system obtained in the step (3) to a micro-bubble flotation column for air flotation, and collecting cobalt-rich foam products and manganese-rich solutions, wherein the flotation recovery rates of Co and Mn are 93.4% and 2.4% respectively.
Comparative example 3
(1) Selective acid leaching: crushing, grinding and screening the purification slag until the mass of particles with the particle size of less than 0.074mm accounts for more than 90%, and using the obtained sample for an acid leaching experiment; preparing a sulfuric acid leaching agent with the concentration of 2mol/L, and mixing the sulfuric acid leaching agent with the treated purified slag according to the liquid-solid ratio of 15mL:1g of the raw materials are mixed, heated to 35 ℃, leached for 15min under the mechanical stirring of 300rpm, and subjected to solid-liquid separation by vacuum filtration to obtain leachate and leaching residues. The leaching rates of the three metals of Co, mn and Zn are respectively 9.1%, 0.1% and 95.8%.
(2) Reduction acid leaching: washing the leaching residue obtained in the step (1) to be neutral, drying, preparing a sulfuric acid leaching agent with the concentration of 1.2mol/L, and mixing the sulfuric acid leaching agent with the leaching residue according to the liquid-solid ratio of 12mL:1g of the mixture is mixed, heated to 35 ℃ and the reducing agent H is added 2 O 2 Making the concentration of the product in the solution to be 1.2mol/L, leaching for 25min under mechanical stirring at 300rpm, performing solid-liquid separation by vacuum filtration, wherein the leaching rates of Co and Mn are respectively 99.5% and 99.1%,obtaining the cobalt-manganese solution.
(3) And (3) performing precipitation reaction, namely adjusting the pH value of the cobalt-manganese solution obtained in the step (2) to 3.5 by using a NaOH solution, adding a medicament A with the theoretical reaction amount of 0.8 time, wherein the molar ratio of the medicament A to the medicament A is 2:1 alpha-nitroso-beta-naphthol and Na 2 S; the mixture was heated to 35 ℃ and reacted for 15min with magnetic stirring at 300rpm, and the Co and Mn precipitation rates were 94.6% and 3.2%, respectively. And then adding a flocculating agent polyacrylamide to ensure that the concentration of the flocculating agent polyacrylamide in the solution is 1mg/L, uniformly stirring, adding a surfactant CTAB to ensure that the concentration of the flocculating agent polyacrylamide in the solution is 40mg/L, and uniformly stirring.
(4) Foam flotation: and (4) transferring the solution system obtained in the step (3) to a micro-bubble flotation column for air flotation, and collecting cobalt-rich foam products and manganese-rich solution, wherein the flotation recovery rates of Co and Mn are 84.2% and 1.1% respectively.
The embodiments of the present invention have been described in detail with reference to the examples, but the present invention is not limited to the described embodiments. It will be apparent to those skilled in the art that various changes, modifications, substitutions and alterations can be made in these embodiments without departing from the principles and spirit of the invention, and the scope of protection is still within the scope of the invention.
Claims (8)
1. A method for recovering cobalt and manganese in zinc purification slag by combining an acid leaching-precipitation flotation method is characterized by comprising the following steps: the method comprises the following steps:
1) Leaching zinc smelting purification slag by using sulfuric acid to obtain a leaching solution and leaching slag;
2) Reducing and leaching the leaching residue by acid to obtain cobalt-manganese leaching solution;
3) And sequentially precipitating and flocculating the cobalt-manganese leaching solution, adding a surfactant for foam flotation, and collecting cobalt-rich foam and a manganese-rich solution.
2. The method for recovering cobalt and manganese in zinc purification slag by acid leaching-precipitation flotation process in combination according to claim 1, is characterized in that: in the step 1), the leaching conditions are as follows: the concentration of the sulfuric acid is 2-4 mol/L, the liquid-solid ratio is 10-20 mL:1g, the temperature is 20-40 ℃, and the time is 10-30 min.
3. The method for recovering cobalt and manganese in zinc purification slag by combining acid leaching and precipitation flotation according to claim 1 or 2, characterized by comprising the following steps:
the acidic reduction leaching adopts a leaching agent containing sulfuric acid and a reducing agent;
the reducing agent is at least one of hydrogen peroxide, sucrose and citric acid.
4. The method for recovering cobalt and manganese in zinc purification slag by acid leaching-precipitation flotation process in combination according to claim 3, is characterized in that: in the step 2), the acidic reduction leaching conditions are as follows: the concentration of sulfuric acid in the leaching agent is 0.6-1.4 mol/L, the concentration of a reducing agent is 0.8-1.2 mol/L, and the liquid-solid ratio is 8-12 mL:1g; the temperature is 20-40 ℃, and the time is 10-30 min.
5. The method for recovering cobalt and manganese in zinc purification slag by acid leaching-precipitation flotation process in combination according to claim 1, characterized in that:
the precipitant is agent A containing alpha-nitroso-beta-naphthol and Na 2 S, alpha-nitroso-beta-naphthol and Na 2 The mol ratio of S is 1-3: 1;
the dosage of the precipitator is 0.8 to 1.2 times of the molar weight of the cobalt ions in the solution.
6. The method for recovering cobalt and manganese in zinc purification slag by acid leaching-precipitation flotation process according to claim 1 or 5, characterized in that:
the flocculating agent is polyacrylamide;
the concentration of the flocculant in the solution is 10-100 mg/L.
7. The method for recovering cobalt and manganese in zinc purification slag by acid leaching-precipitation flotation process according to claim 1 or 5, characterized in that:
the surfactant is CTAB;
the concentration of the surfactant in the solution is 10-50 mg/L.
8. The method for recovering cobalt and manganese in zinc purification slag by acid leaching-precipitation flotation process in combination according to claim 1, is characterized in that: in the step 3), the precipitation reaction conditions are as follows: the pH value of the solution system is 2-4;
the temperature is 20-40 ℃, and the time is 10-30 min.
Priority Applications (1)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| CN202211158120.5A CN115418486B (en) | 2022-09-22 | 2022-09-22 | Method for jointly recovering cobalt and manganese in zinc purification slag by acid leaching-precipitation flotation method |
Applications Claiming Priority (1)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| CN202211158120.5A CN115418486B (en) | 2022-09-22 | 2022-09-22 | Method for jointly recovering cobalt and manganese in zinc purification slag by acid leaching-precipitation flotation method |
Publications (2)
| Publication Number | Publication Date |
|---|---|
| CN115418486A true CN115418486A (en) | 2022-12-02 |
| CN115418486B CN115418486B (en) | 2024-02-23 |
Family
ID=84203997
Family Applications (1)
| Application Number | Title | Priority Date | Filing Date |
|---|---|---|---|
| CN202211158120.5A Active CN115418486B (en) | 2022-09-22 | 2022-09-22 | Method for jointly recovering cobalt and manganese in zinc purification slag by acid leaching-precipitation flotation method |
Country Status (1)
| Country | Link |
|---|---|
| CN (1) | CN115418486B (en) |
Citations (8)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| WO1998014623A1 (en) * | 1996-10-02 | 1998-04-09 | International Curator Resources Limited | Hydrometallurgical extraction of copper, zinc and cobalt from ores containing manganese dioxide |
| AU2008255245A1 (en) * | 2007-12-17 | 2009-07-09 | Bhp Billiton Ssm Development Pty Ltd | Selective recovery of cobalt |
| CN105039739A (en) * | 2015-08-12 | 2015-11-11 | 葫芦岛锌业股份有限公司 | Method for comprehensively recovering cobalt and zinc from purified cobalt residues of zinc hydrometallurgy |
| CN108203764A (en) * | 2018-01-09 | 2018-06-26 | 云南驰宏资源综合利用有限公司 | A kind of method of microwave calcination zinc hydrometallurgy purified cobalt nickel slag production cobalt concentrate |
| CN110747342A (en) * | 2019-12-04 | 2020-02-04 | 河南豫光锌业有限公司 | Method for producing pure cobalt manganese sulfate solution from zinc smelting cobalt slag |
| CN111466051A (en) * | 2017-12-19 | 2020-07-28 | 巴斯夫欧洲公司 | Battery pack recycling by treating leach liquor with metallic nickel |
| CN111850305A (en) * | 2020-07-28 | 2020-10-30 | 昆明理工大学 | Method for leaching cobalt and manganese from manganese-rich cobalt slag |
| CN113957248A (en) * | 2021-01-18 | 2022-01-21 | 郑州大学 | Zinc-cobalt separation method for selective precipitation flotation of cobalt ions in acid solution |
-
2022
- 2022-09-22 CN CN202211158120.5A patent/CN115418486B/en active Active
Patent Citations (8)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| WO1998014623A1 (en) * | 1996-10-02 | 1998-04-09 | International Curator Resources Limited | Hydrometallurgical extraction of copper, zinc and cobalt from ores containing manganese dioxide |
| AU2008255245A1 (en) * | 2007-12-17 | 2009-07-09 | Bhp Billiton Ssm Development Pty Ltd | Selective recovery of cobalt |
| CN105039739A (en) * | 2015-08-12 | 2015-11-11 | 葫芦岛锌业股份有限公司 | Method for comprehensively recovering cobalt and zinc from purified cobalt residues of zinc hydrometallurgy |
| CN111466051A (en) * | 2017-12-19 | 2020-07-28 | 巴斯夫欧洲公司 | Battery pack recycling by treating leach liquor with metallic nickel |
| CN108203764A (en) * | 2018-01-09 | 2018-06-26 | 云南驰宏资源综合利用有限公司 | A kind of method of microwave calcination zinc hydrometallurgy purified cobalt nickel slag production cobalt concentrate |
| CN110747342A (en) * | 2019-12-04 | 2020-02-04 | 河南豫光锌业有限公司 | Method for producing pure cobalt manganese sulfate solution from zinc smelting cobalt slag |
| CN111850305A (en) * | 2020-07-28 | 2020-10-30 | 昆明理工大学 | Method for leaching cobalt and manganese from manganese-rich cobalt slag |
| CN113957248A (en) * | 2021-01-18 | 2022-01-21 | 郑州大学 | Zinc-cobalt separation method for selective precipitation flotation of cobalt ions in acid solution |
Also Published As
| Publication number | Publication date |
|---|---|
| CN115418486B (en) | 2024-02-23 |
Similar Documents
| Publication | Publication Date | Title |
|---|---|---|
| CN107267759B (en) | A kind of comprehensive recovering process of anode material for lithium-ion batteries | |
| CN111270073A (en) | Method for recovering valuable metals from leachate of waste lithium ion battery electrode material | |
| CN112662877B (en) | Method for preparing high-purity nickel sulfate from electrolytic manganese sulfide slag | |
| CN112575208B (en) | Method for preparing high-purity manganese sulfate from electrolytic manganese sulfide slag | |
| CN109022793B (en) | Method for selectively leaching lithium from waste powder of cathode material containing at least one of cobalt, nickel and manganese | |
| CN110760680B (en) | Method for leaching, recovering and separating cobalt from manganese-sulfur purification waste residue | |
| CN113060712B (en) | Method for preparing iron phosphate and nickel cobalt manganese hydroxide battery precursor material from metal nickel cobalt iron powder | |
| CN112410555B (en) | Comprehensive recovery method for flotation silver concentrate from zinc hydrometallurgy acidic leaching residue | |
| CN111455189A (en) | Method for leaching copper from tin-copper slag | |
| CN113403477B (en) | Comprehensive utilization method of nickel sulfide concentrate | |
| CN112662878B (en) | Method for preparing high-purity cobalt sulfate from electrolytic manganese sulfide slag | |
| CN112342383B (en) | Method for separating and recovering nickel, cobalt, manganese and lithium in ternary waste | |
| CN116646633B (en) | Method for recycling active substances in lithium ion positive electrode material | |
| CN111500860B (en) | Process method for recovering copper from low-grade copper oxide ore | |
| CN113430369A (en) | Comprehensive utilization method of nickel sulfide concentrate | |
| CN112410568A (en) | Method for preparing cobalt ferrite from cobalt-containing slag | |
| CN115418486B (en) | Method for jointly recovering cobalt and manganese in zinc purification slag by acid leaching-precipitation flotation method | |
| CN114988382B (en) | Recovery method of waste lithium iron phosphate battery powder | |
| CN113416855B (en) | Method for preparing nickel sulfate from nickel sulfide concentrate leaching solution | |
| CN111455188B (en) | Process method for leaching copper from matte slag by alkaline wet method | |
| CN113621835A (en) | Method for efficiently removing molybdenum based on extraction-precipitation combination | |
| CN110699553B (en) | Method for leaching, recovering and separating nickel from manganese-sulfur purification waste residue | |
| CN115491496B (en) | Method for selectively separating cobalt, manganese, zinc and cadmium from purified cobalt slag of zinc hydrometallurgy | |
| CN114875241B (en) | Method for comprehensively recovering valuable metals from neodymium iron boron waste acid leaching residues under sulfuric acid system | |
| US20230416871A1 (en) | Method for Extracting and Recovering Gold from Aqueous Solution |
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 |