CN115818676A - Method for extracting lithium and collecting fluorine by roasting lithium-containing waste aluminum electrolyte at low temperature - Google Patents
Method for extracting lithium and collecting fluorine by roasting lithium-containing waste aluminum electrolyte at low temperature Download PDFInfo
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- 229910052744 lithium Inorganic materials 0.000 title claims abstract description 192
- WHXSMMKQMYFTQS-UHFFFAOYSA-N Lithium Chemical compound [Li] WHXSMMKQMYFTQS-UHFFFAOYSA-N 0.000 title claims abstract description 188
- 239000003792 electrolyte Substances 0.000 title claims abstract description 120
- 229910052782 aluminium Inorganic materials 0.000 title claims abstract description 112
- XAGFODPZIPBFFR-UHFFFAOYSA-N aluminium Chemical compound [Al] XAGFODPZIPBFFR-UHFFFAOYSA-N 0.000 title claims abstract description 112
- 239000002699 waste material Substances 0.000 title claims abstract description 89
- 229910052731 fluorine Inorganic materials 0.000 title claims abstract description 73
- 239000011737 fluorine Substances 0.000 title claims abstract description 73
- 238000000034 method Methods 0.000 title claims abstract description 62
- PXGOKWXKJXAPGV-UHFFFAOYSA-N Fluorine Chemical compound FF PXGOKWXKJXAPGV-UHFFFAOYSA-N 0.000 title claims 2
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Chemical compound O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 claims abstract description 126
- 238000002386 leaching Methods 0.000 claims abstract description 103
- 239000002253 acid Substances 0.000 claims abstract description 79
- 239000013078 crystal Substances 0.000 claims abstract description 48
- CURLTUGMZLYLDI-UHFFFAOYSA-N Carbon dioxide Chemical compound O=C=O CURLTUGMZLYLDI-UHFFFAOYSA-N 0.000 claims abstract description 42
- 239000003513 alkali Substances 0.000 claims abstract description 42
- 239000012535 impurity Substances 0.000 claims abstract description 29
- XGZVUEUWXADBQD-UHFFFAOYSA-L lithium carbonate Chemical compound [Li+].[Li+].[O-]C([O-])=O XGZVUEUWXADBQD-UHFFFAOYSA-L 0.000 claims abstract description 27
- 229910052808 lithium carbonate Inorganic materials 0.000 claims abstract description 27
- 238000000926 separation method Methods 0.000 claims abstract description 27
- 239000001569 carbon dioxide Substances 0.000 claims abstract description 21
- 229910002092 carbon dioxide Inorganic materials 0.000 claims abstract description 21
- HEMHJVSKTPXQMS-UHFFFAOYSA-M Sodium hydroxide Chemical compound [OH-].[Na+] HEMHJVSKTPXQMS-UHFFFAOYSA-M 0.000 claims description 153
- 238000005406 washing Methods 0.000 claims description 102
- CDBYLPFSWZWCQE-UHFFFAOYSA-L Sodium Carbonate Chemical compound [Na+].[Na+].[O-]C([O-])=O CDBYLPFSWZWCQE-UHFFFAOYSA-L 0.000 claims description 100
- 239000007787 solid Substances 0.000 claims description 67
- 238000002156 mixing Methods 0.000 claims description 51
- 229910000029 sodium carbonate Inorganic materials 0.000 claims description 50
- QAOWNCQODCNURD-UHFFFAOYSA-N Sulfuric acid Chemical compound OS(O)(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-N 0.000 claims description 48
- 235000017550 sodium carbonate Nutrition 0.000 claims description 46
- PUZPDOWCWNUUKD-UHFFFAOYSA-M sodium fluoride Chemical compound [F-].[Na+] PUZPDOWCWNUUKD-UHFFFAOYSA-M 0.000 claims description 46
- 239000000243 solution Substances 0.000 claims description 45
- 238000001704 evaporation Methods 0.000 claims description 44
- 238000001914 filtration Methods 0.000 claims description 42
- 239000007788 liquid Substances 0.000 claims description 37
- KWYUFKZDYYNOTN-UHFFFAOYSA-M Potassium hydroxide Chemical compound [OH-].[K+] KWYUFKZDYYNOTN-UHFFFAOYSA-M 0.000 claims description 36
- 239000002893 slag Substances 0.000 claims description 34
- 239000000463 material Substances 0.000 claims description 33
- 230000008020 evaporation Effects 0.000 claims description 24
- BWHMMNNQKKPAPP-UHFFFAOYSA-L potassium carbonate Chemical compound [K+].[K+].[O-]C([O-])=O BWHMMNNQKKPAPP-UHFFFAOYSA-L 0.000 claims description 24
- 230000001376 precipitating effect Effects 0.000 claims description 24
- 239000008367 deionised water Substances 0.000 claims description 22
- 229910021641 deionized water Inorganic materials 0.000 claims description 22
- 239000012452 mother liquor Substances 0.000 claims description 22
- 239000000203 mixture Substances 0.000 claims description 13
- 229910000027 potassium carbonate Inorganic materials 0.000 claims description 12
- 235000011181 potassium carbonates Nutrition 0.000 claims description 12
- 229910001610 cryolite Inorganic materials 0.000 claims description 9
- UIIMBOGNXHQVGW-UHFFFAOYSA-M Sodium bicarbonate Chemical compound [Na+].OC([O-])=O UIIMBOGNXHQVGW-UHFFFAOYSA-M 0.000 claims description 8
- 239000000706 filtrate Substances 0.000 claims description 8
- QTBSBXVTEAMEQO-UHFFFAOYSA-N Acetic acid Chemical compound CC(O)=O QTBSBXVTEAMEQO-UHFFFAOYSA-N 0.000 claims description 6
- BVKZGUZCCUSVTD-UHFFFAOYSA-L Carbonate Chemical compound [O-]C([O-])=O BVKZGUZCCUSVTD-UHFFFAOYSA-L 0.000 claims description 6
- MUBZPKHOEPUJKR-UHFFFAOYSA-N Oxalic acid Chemical compound OC(=O)C(O)=O MUBZPKHOEPUJKR-UHFFFAOYSA-N 0.000 claims description 6
- VEXZGXHMUGYJMC-UHFFFAOYSA-N Hydrochloric acid Chemical compound Cl VEXZGXHMUGYJMC-UHFFFAOYSA-N 0.000 claims description 4
- NBIIXXVUZAFLBC-UHFFFAOYSA-N Phosphoric acid Chemical compound OP(O)(O)=O NBIIXXVUZAFLBC-UHFFFAOYSA-N 0.000 claims description 4
- 150000004673 fluoride salts Chemical class 0.000 claims description 4
- 238000007654 immersion Methods 0.000 claims description 4
- 239000011259 mixed solution Substances 0.000 claims description 4
- 229910000030 sodium bicarbonate Inorganic materials 0.000 claims description 4
- 235000017557 sodium bicarbonate Nutrition 0.000 claims description 4
- GRYLNZFGIOXLOG-UHFFFAOYSA-N Nitric acid Chemical compound O[N+]([O-])=O GRYLNZFGIOXLOG-UHFFFAOYSA-N 0.000 claims description 3
- 238000001354 calcination Methods 0.000 claims description 3
- AXCZMVOFGPJBDE-UHFFFAOYSA-L calcium dihydroxide Chemical compound [OH-].[OH-].[Ca+2] AXCZMVOFGPJBDE-UHFFFAOYSA-L 0.000 claims description 3
- 239000000920 calcium hydroxide Substances 0.000 claims description 3
- 229910001861 calcium hydroxide Inorganic materials 0.000 claims description 3
- 229910017604 nitric acid Inorganic materials 0.000 claims description 3
- 239000011736 potassium bicarbonate Substances 0.000 claims description 3
- 235000015497 potassium bicarbonate Nutrition 0.000 claims description 3
- 229910000028 potassium bicarbonate Inorganic materials 0.000 claims description 3
- TYJJADVDDVDEDZ-UHFFFAOYSA-M potassium hydrogencarbonate Chemical compound [K+].OC([O-])=O TYJJADVDDVDEDZ-UHFFFAOYSA-M 0.000 claims description 3
- 239000000843 powder Substances 0.000 claims description 3
- 229910000147 aluminium phosphate Inorganic materials 0.000 claims description 2
- 235000006408 oxalic acid Nutrition 0.000 claims description 2
- 238000007670 refining Methods 0.000 claims description 2
- 230000001105 regulatory effect Effects 0.000 claims description 2
- 238000000967 suction filtration Methods 0.000 claims description 2
- 238000001556 precipitation Methods 0.000 claims 1
- YCKRFDGAMUMZLT-UHFFFAOYSA-N Fluorine atom Chemical compound [F] YCKRFDGAMUMZLT-UHFFFAOYSA-N 0.000 abstract description 70
- 238000011084 recovery Methods 0.000 abstract description 57
- 230000008901 benefit Effects 0.000 abstract description 4
- 238000004519 manufacturing process Methods 0.000 abstract description 4
- 238000000227 grinding Methods 0.000 description 41
- 230000008569 process Effects 0.000 description 33
- HBBGRARXTFLTSG-UHFFFAOYSA-N Lithium ion Chemical compound [Li+] HBBGRARXTFLTSG-UHFFFAOYSA-N 0.000 description 25
- 229910001416 lithium ion Inorganic materials 0.000 description 21
- 238000006243 chemical reaction Methods 0.000 description 16
- PQXKHYXIUOZZFA-UHFFFAOYSA-M lithium fluoride Chemical compound [Li+].[F-] PQXKHYXIUOZZFA-UHFFFAOYSA-M 0.000 description 14
- 229910003002 lithium salt Inorganic materials 0.000 description 12
- 159000000002 lithium salts Chemical class 0.000 description 12
- KRHYYFGTRYWZRS-UHFFFAOYSA-N Fluorane Chemical compound F KRHYYFGTRYWZRS-UHFFFAOYSA-N 0.000 description 11
- 239000007789 gas Substances 0.000 description 11
- 229910000040 hydrogen fluoride Inorganic materials 0.000 description 11
- 239000011775 sodium fluoride Substances 0.000 description 11
- 235000013024 sodium fluoride Nutrition 0.000 description 11
- 230000000052 comparative effect Effects 0.000 description 10
- 230000009466 transformation Effects 0.000 description 9
- 230000002378 acidificating effect Effects 0.000 description 8
- 239000011734 sodium Substances 0.000 description 7
- 239000000654 additive Substances 0.000 description 6
- 230000000996 additive effect Effects 0.000 description 6
- ANBBXQWFNXMHLD-UHFFFAOYSA-N aluminum;sodium;oxygen(2-) Chemical compound [O-2].[O-2].[Na+].[Al+3] ANBBXQWFNXMHLD-UHFFFAOYSA-N 0.000 description 6
- VVNXEADCOVSAER-UHFFFAOYSA-N lithium sodium Chemical compound [Li].[Na] VVNXEADCOVSAER-UHFFFAOYSA-N 0.000 description 6
- 229910001388 sodium aluminate Inorganic materials 0.000 description 6
- 239000002585 base Substances 0.000 description 5
- 238000004064 recycling Methods 0.000 description 5
- 238000005265 energy consumption Methods 0.000 description 4
- NROKBHXJSPEDAR-UHFFFAOYSA-M potassium fluoride Chemical compound [F-].[K+] NROKBHXJSPEDAR-UHFFFAOYSA-M 0.000 description 4
- 239000002994 raw material Substances 0.000 description 4
- 230000000694 effects Effects 0.000 description 3
- 238000010438 heat treatment Methods 0.000 description 3
- 230000007246 mechanism Effects 0.000 description 3
- 239000000047 product Substances 0.000 description 3
- 239000000376 reactant Substances 0.000 description 3
- 239000004575 stone Substances 0.000 description 3
- 239000000126 substance Substances 0.000 description 3
- UIIMBOGNXHQVGW-DEQYMQKBSA-M Sodium bicarbonate-14C Chemical compound [Na+].O[14C]([O-])=O UIIMBOGNXHQVGW-DEQYMQKBSA-M 0.000 description 2
- 238000002441 X-ray diffraction Methods 0.000 description 2
- JFBZPFYRPYOZCQ-UHFFFAOYSA-N [Li].[Al] Chemical compound [Li].[Al] JFBZPFYRPYOZCQ-UHFFFAOYSA-N 0.000 description 2
- 229910001515 alkali metal fluoride Inorganic materials 0.000 description 2
- 229910000272 alkali metal oxide Inorganic materials 0.000 description 2
- 238000005868 electrolysis reaction Methods 0.000 description 2
- 238000003912 environmental pollution Methods 0.000 description 2
- 150000002500 ions Chemical class 0.000 description 2
- VUZPPFZMUPKLLV-UHFFFAOYSA-N methane;hydrate Chemical compound C.O VUZPPFZMUPKLLV-UHFFFAOYSA-N 0.000 description 2
- 239000010413 mother solution Substances 0.000 description 2
- 230000001590 oxidative effect Effects 0.000 description 2
- 239000011698 potassium fluoride Substances 0.000 description 2
- 235000003270 potassium fluoride Nutrition 0.000 description 2
- 239000002910 solid waste Substances 0.000 description 2
- IRPGOXJVTQTAAN-UHFFFAOYSA-N 2,2,3,3,3-pentafluoropropanal Chemical compound FC(F)(F)C(F)(F)C=O IRPGOXJVTQTAAN-UHFFFAOYSA-N 0.000 description 1
- KLZUFWVZNOTSEM-UHFFFAOYSA-K Aluminum fluoride Inorganic materials F[Al](F)F KLZUFWVZNOTSEM-UHFFFAOYSA-K 0.000 description 1
- 229910010199 LiAl Inorganic materials 0.000 description 1
- VMHLLURERBWHNL-UHFFFAOYSA-M Sodium acetate Chemical compound [Na+].CC([O-])=O VMHLLURERBWHNL-UHFFFAOYSA-M 0.000 description 1
- XSQUKJJJFZCRTK-UHFFFAOYSA-N Urea Chemical compound NC(N)=O XSQUKJJJFZCRTK-UHFFFAOYSA-N 0.000 description 1
- 238000010306 acid treatment Methods 0.000 description 1
- 230000009471 action Effects 0.000 description 1
- 229910052783 alkali metal Inorganic materials 0.000 description 1
- 150000001340 alkali metals Chemical class 0.000 description 1
- 239000004411 aluminium Substances 0.000 description 1
- PNEYBMLMFCGWSK-UHFFFAOYSA-N aluminium oxide Inorganic materials [O-2].[O-2].[O-2].[Al+3].[Al+3] PNEYBMLMFCGWSK-UHFFFAOYSA-N 0.000 description 1
- 229910001570 bauxite Inorganic materials 0.000 description 1
- 230000009286 beneficial effect Effects 0.000 description 1
- 238000007664 blowing Methods 0.000 description 1
- BRPQOXSCLDDYGP-UHFFFAOYSA-N calcium oxide Chemical compound [O-2].[Ca+2] BRPQOXSCLDDYGP-UHFFFAOYSA-N 0.000 description 1
- 239000000292 calcium oxide Substances 0.000 description 1
- ODINCKMPIJJUCX-UHFFFAOYSA-N calcium oxide Inorganic materials [Ca]=O ODINCKMPIJJUCX-UHFFFAOYSA-N 0.000 description 1
- 239000004202 carbamide Substances 0.000 description 1
- 230000008859 change Effects 0.000 description 1
- 238000002485 combustion reaction Methods 0.000 description 1
- 125000004122 cyclic group Chemical group 0.000 description 1
- 238000010586 diagram Methods 0.000 description 1
- 238000004090 dissolution Methods 0.000 description 1
- 230000009977 dual effect Effects 0.000 description 1
- 238000000605 extraction Methods 0.000 description 1
- 238000010304 firing Methods 0.000 description 1
- -1 fluorine salt Chemical class 0.000 description 1
- 230000006872 improvement Effects 0.000 description 1
- 239000002244 precipitate Substances 0.000 description 1
- 238000011085 pressure filtration Methods 0.000 description 1
- 238000007789 sealing Methods 0.000 description 1
- 239000001632 sodium acetate Substances 0.000 description 1
- 235000017281 sodium acetate Nutrition 0.000 description 1
- ZNCPFRVNHGOPAG-UHFFFAOYSA-L sodium oxalate Chemical compound [Na+].[Na+].[O-]C(=O)C([O-])=O ZNCPFRVNHGOPAG-UHFFFAOYSA-L 0.000 description 1
- 229940039790 sodium oxalate Drugs 0.000 description 1
- 159000000000 sodium salts Chemical class 0.000 description 1
- 238000003756 stirring Methods 0.000 description 1
- 238000011426 transformation method Methods 0.000 description 1
- 230000001131 transforming effect Effects 0.000 description 1
- 230000007306 turnover Effects 0.000 description 1
- 238000009736 wetting Methods 0.000 description 1
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- 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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- Manufacture And Refinement Of Metals (AREA)
Abstract
The invention discloses a method for extracting lithium and collecting fluorine by roasting lithium-containing waste aluminum electrolyte at low temperature, which comprises the step of roasting the lithium-containing waste aluminum electrolyte and mixed alkali at low temperature in a double-atmosphere environment of carbon dioxide and water vapor to obtain a lithium-containing stable crystal form (Li) 2 Al 4 (CO 3 )(OH) 12 ·3H 2 O), the crystal form is insoluble in alkali and water, can be completely leached under the conditions of low temperature and weak acid, is mild in leaching conditions, convenient and efficient in fluorine and lithium separation, high in lithium leaching rate and less in impurities, and further ensures the purity of lithium carbonate obtained by subsequent recovery and reduces the recovery difficulty while efficiently recovering fluorine. Compared with the prior art, the method has the advantages of high separation and recovery efficiency of lithium and fluorine elements, high added value of products, low production cost and the like.
Description
Technical Field
The invention relates to treatment of lithium-containing waste aluminum electrolyte, in particular to a method for extracting lithium and collecting fluorine by roasting the lithium-containing waste aluminum electrolyte at low temperature, and belongs to the technical field of recycling treatment of industrial solid waste.
Background
The lithium salt plays an important role in the electrolytic aluminum industry as an additive of the aluminum electrolyte, and can obviously reduce the primary crystal temperature of the electrolyte, improve the conductivity, save energy and reduce consumption. Generally, the LiF content in the aluminum electrolyte is about 1-3%, but because the low-grade bauxite contains lithium, the lithium content in the electrolyte is continuously enriched along with the continuous enrichment of the alumina raw material in the electrolytic cell, the lithium content in some aluminum plants even exceeds 8%, and the high-lithium electrolyte can generate negative effects on the electrolytic process and seriously affect the normal operation of the electrolytic cell. Therefore, enterprises usually move the high-lithium electrolyte out of the electrolyte circulation, but the high-lithium electrolyte can cause a large amount of complex aluminum electrolyte to be stockpiled and wasted, so that the environment is polluted, the fluorine resource is seriously wasted, and the environment is greatly polluted. At present, lithium resources are in short supply, the demand amount is short of the demand, the price is high, if the lithium element in the waste aluminum electrolyte is extracted through a reasonable process, the recycling of the lithium resources is expanded, the influence of the lithium element on the aluminum electrolysis production is removed, and the method has important significance for the development of lithium salt industry and the improvement of the aluminum electrolysis economic benefit.
The aluminum electrolyte in China is acidic electrolyte, and lithium in the electrolyte is easily combined with excessive aluminum fluoride to form Na 2 LiAlF 6 So that lithium in the waste aluminum electrolyte exists mainly in the form of NaLi 2 AlF 6 。Na 2 LiAlF 6 Has the same effect as Na 3 AlF 6 Similar properties, being insoluble lithium salts, increase the difficulty of extracting lithium from spent electrolytes. At present, lithium salts and other substances in the aluminum electrolyte are dissolved by adopting a method of strong acid leaching or water leaching after strong acid roasting, but a large amount of HF gas is generated in the acid leaching or roasting process, so that the method is not environment-friendly, impurity ions can be leached together, and the difficulty of subsequent separation operation is high. Some existing measures are to transform insoluble lithium salt by high-temperature roasting, for example, CN107587167B proposes a method for changing crystal form of lithium-containing aluminum electrolyte, which utilizes additive to convert lithium sodium cryolite into LiF, li under high-temperature roasting 2 O，Li 3 AlF 6 And the additive comprises one or more of alkali metal oxide except lithium, alkali metal oxysalt except lithium capable of being converted into alkali metal oxide under high-temperature roasting condition, and mixed materialsThe aluminum electrolyte contains alkali metal fluoride and alkali metal fluoride which can be converted by an additive under the high-temperature roasting condition, and the converted crystal form still needs to use a large amount of concentrated and strong acid to leach lithium element, but other components can also be leached simultaneously. CN113684369B discloses a method for treating waste industrial lithium-containing aluminum electrolyte, which comprises roasting the lithium-containing aluminum electrolyte and additive at high temperature, wherein the additive is sodium hydroxide or potassium hydroxide powder, and lithium sodium cryolite is converted into LiAl 2 (OH) 7 And LiF, but HF gas is still generated during the LiF leaching process.
In summary, in the existing lithium extraction process, if the insoluble lithium salt is not transformed or the transformation is not thorough, the problems of low lithium leaching rate or the need of adopting extremely harsh conditions such as concentration and strong acid leaching and difficult subsequent treatment and separation of the lithium-containing leachate exist, and if the lithium-containing leachate is roasted at high temperature or roasted by strong acid and strong base, the energy consumption is high, the requirement on equipment is high, and the problem that the strong acid reacts with the fluorine-containing electrolyte to generate HF still exists. Therefore, it is important to find a mild transformation method for completely transforming the insoluble lithium salt and improve the leaching conditions of the transformed product.
Disclosure of Invention
Aiming at the problems of high energy consumption, low lithium recovery rate, easy generation of HF (hydrogen fluoride) to pollute the environment, difficult impurity separation, ineffective recovery of fluorine element and the like existing in the process of recycling lithium-containing waste aluminum electrolyte in the prior art, the invention provides a method for extracting lithium and recovering fluorine by roasting the lithium-containing waste aluminum electrolyte at low temperature, wherein the lithium-containing waste aluminum electrolyte and mixed alkali are roasted at low temperature in a double-atmosphere environment of carbon dioxide and water vapor to obtain a lithium-containing stable crystal form (Li is insoluble in alkali and can leach lithium at low temperature under a weak acid condition) 2 Al 4 (CO 3 )(OH) 12 ·3H 2 O), still almost all change into soluble villiaumite with the fluorine that contains in the lithium-containing aluminium scrap electrolyte simultaneously, only need through simple washing, can realize the high-efficient recovery of fluorine, further ensured the purity of follow-up recovery acquisition lithium carbonate and reduced the recovery degree of difficulty when high-efficient recovery fluorine. Compared with the prior art, the invention has the advantages of high separation and recovery efficiency of lithium and fluorine elements, high added value of products and productionLow cost and the like.
In order to achieve the above purpose, the technical scheme adopted by the invention is as follows:
a method for extracting lithium and collecting fluorine by roasting lithium-containing waste aluminum electrolyte at low temperature specifically comprises the following steps:
1) Firstly, mixing lithium-containing waste aluminum electrolyte powder with alkali to obtain a mixture, and then roasting the mixture in a double-atmosphere environment of carbon dioxide and water vapor to obtain a roasted material.
2) Refining the roasted material, then carrying out water leaching treatment, and carrying out solid-liquid separation to obtain filter residue and filtrate. The filter residue is processed in the next stage, and the filtrate is separately disposed.
3) Washing the filter residue to obtain washing residue and washing water. And concentrating and crystallizing the washing water to obtain the high-purity villiaumite. Acid leaching treatment is carried out on the washing slag by adopting acid, and acid leaching slag and leachate are obtained after solid-liquid separation. The leachate is treated in the next stage, and the acid leaching residue is separately treated. In the invention, the leachate can be optionally subjected to cyclic leaching treatment or not, and the adjustment is carried out according to the actual working condition.
4) And (4) adjusting the leaching solution to be alkaline, filtering out impurities, concentrating, adding soluble carbonate, finally precipitating crystals to obtain lithium carbonate, and recycling the mother solution for precipitating lithium (namely, recycling the mother solution for the step 4 after mixing with the leaching solution)).
Preferably, the method further comprises:
5) Mixing the filtrate obtained in the step 2) with the residual liquid after the fluoride salt is separated out from the washing water in the step 3) to obtain a mixed solution, and using the mixed solution to prepare and obtain the cryolite.
Preferably, in step 1), the base comprises a mixture of one or more of sodium hydroxide, potassium hydroxide, calcium hydroxide and one or more of sodium carbonate, potassium carbonate, sodium bicarbonate, potassium bicarbonate. Preferably a mixed base consisting of sodium hydroxide and sodium carbonate, or a mixed base consisting of potassium hydroxide and potassium carbonate.
Preferably, in the step 1), the mixing mass ratio of sodium hydroxide to sodium carbonate or potassium hydroxide to potassium carbonate in the mixed alkali is 1 to 10, preferably 1.5 to 8, and more preferably 2 to 6. 1, 1.2, 1.5.
Preferably, in step 1), the amount of the mixed alkali added is 10 to 50%, preferably 15 to 40%, more preferably 20 to 35%, for example, 10%, 12%, 15%, 18%, 20%, 22%, 24%, 26%, 28%, 30%, 33%, 35%, 38%, 40%, 45%, 50% of the total mass of the lithium-containing waste aluminum electrolyte.
Preferably, in step 1), the dual atmosphere environment of carbon dioxide and water vapor is specifically to fill carbon dioxide and water vapor into the roasting equipment, and make the atmosphere pressure in the roasting equipment be positive pressure, preferably the atmosphere pressure in the roasting equipment is 1.1-2 times of the standard atmospheric pressure, and more preferably 1.3-1.8 times of the standard atmospheric pressure. For example, 1.1 times, 1.2 times, 1.3 times, 1.4 times, 1.5 times, 1.6 times, 1.7 times, 1.8 times, 1.9 times, or 2 times the standard atmospheric pressure.
Preferably, in step 1), the temperature of the baking treatment is 100 to 275 ℃, preferably 120 to 270 ℃, more preferably 150 to 260 ℃, for example 120 ℃, 125 ℃, 130 ℃, 135 ℃, 140 ℃, 145 ℃, 150 ℃, 155 ℃, 160 ℃, 165 ℃, 170 ℃, 175 ℃, 180 ℃, 185 ℃, 190 ℃, 200 ℃, 210 ℃, 220 ℃, 230 ℃, 240 ℃, 250 ℃, 260 ℃, 270 ℃.
Preferably, in step 1), the time of the calcination treatment is 1 to 10 hours, preferably 1.5 to 8 hours, more preferably 2 to 5 hours, for example, one of 1 hour, 1.2 hours, 1.5 hours, 1.8 hours, 2.5 hours, 3 hours, 3.5 hours, 4 hours, 5 hours, 6 hours, 7 hours, 8 hours, 9 hours and 10 hours.
Preferably, in step 2), the water for water immersion treatment is deionized water, and the liquid-solid ratio in water immersion treatment is 3 to 10.
Preferably, in step 2), the time for the water immersion treatment is 1 to 8 hours, preferably 1.5 to 7 hours, more preferably 2 to 6 hours, for example, one of 1 hour, 1.5 hours, 1.8 hours, 2.0 hours, 2.5 hours, 3 hours, 3.5 hours, 4 hours, 4.5 hours, 5 hours, 5.5 hours, 6 hours, 6.5 hours, 7 hours, 7.5 hours and 8 hours.
Preferably, in step 2), the leaching temperature of the water leaching treatment is 50 to 90 ℃, preferably 55 to 85 ℃, more preferably 60 to 80 ℃, for example, one of 55 ℃, 60 ℃, 65 ℃, 70 ℃, 75 ℃, 80 ℃, 85 ℃ and 90 ℃.
Preferably, in step 3), the filter residue is washed 1 to 10 times, preferably 2 to 8 times, more preferably 3 to 5 times, for example, 2 times, 3 times, 4 times, 5 times, 6 times, 7 times, or 8 times.
As a preference, in step 3), the liquid-solid ratio at a single washing of the residue is 1-5, preferably 1.5-4, and more preferably 2-3.5, for example 1. The washing method is generally a multi-stage (e.g., 3, 4, 5, 6, etc.) counter-current washing.
Preferably, in step 3), the acid for acid leaching treatment is a dilute acid solution with a pH of less than 4, preferably one or more of a sulfuric acid solution, a hydrochloric acid solution, a nitric acid solution, a phosphoric acid solution, an oxalic acid solution and an acetic acid solution with a pH of 1-3.5.
Preferably, in step 3), the liquid-solid ratio of the acid leaching treatment is 2 to 8, preferably 2.5 to 6, and more preferably 3 to 5, and is, for example, 2.
Preferably, in step 3), the time of the acid leaching treatment is 1.5 to 8 hours, preferably 2 to 6 hours, more preferably 3 to 5 hours, for example, one of 1.5 hours, 1.8 hours, 2.0 hours, 2.2 hours, 2.5 hours, 2.8 hours, 3 hours, 3.2 hours, 3.5 hours, 3.8 hours, 4 hours, 4.2 hours, 4.5 hours, 4.8 hours, 5 hours, 5.5 hours, 6 hours, 6.5 hours, 7 hours, 7.5 hours, and 8 hours.
Preferably, in step 3), the temperature of the acid leaching treatment is 30 to 90 ℃, preferably 40 to 80 ℃, more preferably 50 to 70 ℃, for example, 30 ℃, 35 ℃, 40 ℃, 45 ℃, 50 ℃, 55 ℃, 60 ℃, 65 ℃, 70 ℃, 75 ℃, 80 ℃, 85 ℃ or 90 ℃.
Preferably, in step 4), the leachate is adjusted to alkaline by adjusting the pH of the leachate to 7.5 to 10, preferably 8 to 9.5, more preferably 8.5 to 9, using sodium hydroxide and/or potassium hydroxide.
Preferably, in step 4), the soluble carbonate is sodium carbonate and/or potassium carbonate, preferably sodium carbonate.
Preferably, in the step 1-4), the solid-liquid separation is one or more of pressure filtration, negative pressure suction filtration and centrifugal separation.
Preferably, in step 1-4), the concentration is an evaporative concentration with a concentration factor of 3-10 times, preferably 4-8 times. The water vapor generated by evaporation and concentration is recycled for roasting treatment.
In the prior art, the lithium-containing aluminum scrap electrolyte has high content of insoluble lithium salts, and needs to adopt a harsher condition to leach lithium (for example, high and strong acid is used for leaching, and leaching is carried out after high-temperature roasting), so that impurity ions can be completely leached while lithium is leached, particularly, HF gas is generated to overflow into air or be partially dissolved into the leachate, and the problems of difficult subsequent treatment and difficult impurity separation of the lithium-containing leachate exist. Most of the alkaline roasting or alkaline leaching is to convert insoluble lithium salt into LiF, the LiF is difficult to dissolve in water, high and strong acid (sulfuric acid and nitric acid) is still needed to dissolve, a large amount of HF gas is generated during dissolution, and the operation environment is severe. In addition, in the prior art, the problem that the transformation difficulty of insoluble lithium salt in the lithium-containing waste aluminum electrolyte is high is generally solved by adopting a high-temperature roasting method or a strong acid and strong base roasting method, the energy consumption is high, the requirement on equipment is high, and particularly, a large amount of HF gas is generated during acid roasting, so that the environmental pollution is serious. Furthermore, in the prior art, in the process of extracting lithium from lithium-containing waste aluminum electrolyte, the recovery treatment of fluorine element cannot be effectively carried out (HF is generated or fluorine-containing precipitate is directly removed).
In the invention, mixed alkali composed of strong alkali (sodium hydroxide or potassium hydroxide, etc.) and basic carbonate (sodium carbonate or potassium carbonate, etc.) is used as roasting reactant to mix with lithium-containing waste aluminum electrolyte for roasting, and carbon dioxide gas and water are introduced during the roasting processThe double steam atmosphere is used as transformation promoter, and the insoluble lithium salt can be transformed into lithium-containing stable crystal form (Li) by roasting at low temperature (100-275 ℃) in a roasting device 2 Al 4 (CO 3 )(OH) 12 ·3H 2 O) and converting elemental fluorine to a soluble fluoride salt (sodium fluoride or potassium fluoride). The lithium-containing stable crystal form is insoluble in alkali and water, and the soluble villiaumite in the roasting slag can be almost completely washed out by directly adopting a water leaching and water washing mode after the roasting slag is ground, so that the high-efficiency separation of fluorine and lithium is realized, and simultaneously, the high-purity sodium fluoride or potassium fluoride can be recovered and obtained, and almost no HF is generated under an alkaline condition. The defluorinated roasting residue is mainly in a lithium-containing stable crystal form, so that the lithium-containing acidic leaching solution can be obtained only by carrying out acid leaching treatment under the weak acidic condition, the alkalinity of the lithium-containing acidic leaching solution is weak after water washing, the acid consumption in the subsequent acid leaching treatment is low, and finally the lithium carbonate is recovered after removing residual impurities by regulating the pH to be alkaline, evaporating, concentrating and precipitating lithium.
In the invention, the mixed alkali comprises a mixture of one or more of sodium hydroxide, potassium hydroxide, calcium hydroxide and calcium oxide and one or more of sodium carbonate, potassium carbonate, sodium bicarbonate and potassium bicarbonate, and the mixing mass ratio of the two is 1-10. The addition amount of the mixed alkali is 10-50% of the total mass of the lithium-containing waste aluminum electrolyte. The lithium sodium cryolite is directly subjected to transformation reaction through alkaline roasting, so that the lithium sodium cryolite which is difficult to transform at normal temperature is transformed into Li 2 Al 4 (CO 3 )(OH) 12 ·3H 2 O, the crystal form can be leached under mild weak acidic condition without being leached under strong acid like lithium fluoride, so that the acid consumption is greatly reduced, and the leaching rate of lithium element is improved. The carbon dioxide and the water vapor are introduced to mainly provide partial carbonate ions and moisture required by a new crystal form for the transformation of the lithium sodium cryolite to participate in the transformation reaction, and the reactant contact area can be increased by blowing in the gaseous form, so that the reaction of the lithium sodium cryolite is promoted to be more sufficient. Wherein the mild weakly acidic conditions are those in which the molar concentration of the acid is less than 0.5mol/L andthe diluted acid solution of water is preferably a diluted acid solution of acid and water with a molar concentration of less than 0.3mol/L, more preferably a diluted acid solution of acid and water with a molar concentration of less than 0.1mol/L, for example, so that the pH of the diluted acid solution of acid and water is between 1 and 3.5. Of course, the above-mentioned mild, slightly acidic conditions should not be considered as limiting the technical solution of the present invention, which is merely a preferred example of the present invention, and the corresponding technical objects of the present invention can be achieved even under high concentration of strong acid, only with relative danger and higher cost.
It should be noted that, in the present invention, the introduction of carbon dioxide and water vapor requires that the reaction environment is a certain positive pressure atmosphere (for example, the atmospheric pressure in the roasting equipment is 1.1 to 2 times of the standard atmospheric pressure), and under a certain positive pressure, carbon dioxide and water vapor molecules can easily enter the interior of the stockpile, so as to increase the contact area with the material and promote the rapid progress of the crystal transformation reaction. The atmosphere pressure is low or only a small amount of carbon dioxide or water is introduced, which easily causes incomplete crystal transformation reaction, and further influences the recovery rate of lithium and fluorine. In addition, the introduction of carbon dioxide and water vapor in the invention also comprises the step of mixing substances which are decomposed by heating and generate carbon dioxide and water into the roasted material (namely, most of carbon dioxide gas and water vapor can be directly introduced by pure gas, and a small part of carbon dioxide gas and water vapor can be obtained by chemical reaction, such as the operations of introducing water gas for combustion, decomposing sodium bicarbonate, oxidizing sodium acetate, oxidizing sodium oxalate, pyrolyzing part of organic matters, urea and the like, wetting the mixture and the like), so that the reaction efficiency in the material is improved.
In the present invention, as the equipment for the roasting reaction, it should be a roasting equipment with good sealing performance, and in addition, if necessary, it may also have a mechanism (including but not limited to a stirring mechanism or a turnover vibration mechanism) for dispersing the material or breaking up the material, so as to improve the contact with the atmosphere during the roasting process of the material, improve the reaction efficiency, practically ensure the proceeding of the crystal transformation reaction, ensure the recovery rate of the final lithium and fluorine, reduce the time required for the reaction, improve the production efficiency, and reduce the production cost. In the present invention, the main body of the roasting equipment can be a fixed bed, a moving bed, a fluidized bed, a reverberatory furnace, a rotary kiln, a fluidized bed furnace, a microwave furnace, etc., and can achieve the purpose of roasting, and is preferably a fluidized bed furnace.
In the invention, li is not suitable for being excessively high after the roasting temperature is not too high, generally exceeds 275 ℃ by passing the lithium-containing aluminum scrap electrolyte at low temperature (100-275 ℃ C.) in the presence of mixed alkali and in the environment of carbon dioxide and water vapor atmosphere 2 Al 4 (CO 3 )(OH) 12 ·3H 2 The O crystal form can be thermally decomposed, the temperature is lower than 100 ℃, the reaction rate and the yield of the new crystal form are lower), the crystal form conversion of the lithium-containing substance is realized by roasting, the subsequent water treatment and weak acid treatment are combined, the high-efficiency separation and recovery of lithium and fluorine in the lithium-containing waste aluminum electrolyte can be realized (almost all lithium elements in the lithium-containing waste aluminum electrolyte are Li) 2 Al 4 (CO 3 )(OH) 12 ·3H 2 The crystal form O exists, and the fluorine element exists almost completely in the form of soluble fluorine salt). The reaction formula involved in the calcination process is (taking sodium salt as an example):
(1)Na 3 AlF 6 +4NaOH＝NaAl(OH) 4 +6NaF
(2)2Na 2 LiAlF 6 +6NaOH+CO 2 +2H 2 O+2NaAl(OH) 4 ＝Li 2 Al 4 (CO 3 )(OH) 12 ·3H 2 O+12NaF
(3)2LiF+CO 2 +2H 2 O+4NaAl(OH) 4 ＝Li 2 Al 4 (CO 3 )(OH) 12 ·3H 2 O+2NaF+2NaOH
(4)4Na 3 AlF 6 +2LiF+CO 2 +14NaOH+2H 2 O＝Li 2 Al 4 (CO 3 )(OH) 12 ·3H 2 O+26NaF
(5)2Na 2 LiAlF 6 +4NaOH+Na 2 CO 3 +3H 2 O+2NaAl(OH) 4 ＝Li 2 Al 4 (CO 3 )(OH) 12 ·3H 2 O+12NaF
(6)2NaHCO 3 ＝Na 2 CO 3 +CO 2 +H 2 o (this reaction is only present when part of the sodium bicarbonate is incorporated in the baked mix)
During the roasting process, most of cryolite (Na) in the lithium-containing waste aluminum electrolyte 3 AlF 6 ) The sodium aluminate and the sodium fluoride can be generated by reacting and decomposing with alkali, so that the roasted slag is soaked in water and washed with water respectively before being subjected to acid leaching treatment, the liquid-solid mass ratio (3-10) during water leaching treatment is reasonably controlled, the sodium aluminate can be completely leached, part of the sodium fluoride is leached, water leaching filtrate containing the sodium aluminate and the sodium fluoride is obtained, and the sodium aluminate and the sodium fluoride are raw materials for preparing ice crystal stone, so that the water leaching filtrate is used as a raw material for preparing the ice crystal stone, and the prepared ice crystal stone is also used as a raw material for preparing the aluminum electrolyte, thereby realizing forward circulation in a system and realizing zero emission of solid waste. Meanwhile, because the solubility of the sodium fluoride is low, the liquid-solid ratio controlled during water leaching treatment is mainly based on the premise of completely leaching the sodium aluminate, and the yield of the sodium fluoride is far greater than that of the sodium aluminate in the whole roasting process, namely, a large amount of sodium fluoride still exists in filter residue after water leaching, so that the filter residue after water leaching is washed for many times, the sodium fluoride in the filter residue is further dissolved out, the separation of lithium and fluorine is realized, and a high-purity sodium fluoride product (the purity is more than 98%) can be recovered and obtained; and the alkalinity of the filter residue can be further weakened, so that the acid consumption is saved in the subsequent acid leaching treatment.
Compared with the prior art, the invention has the following beneficial technical effects:
1: the invention carries out roasting treatment on lithium-containing waste aluminum electrolyte under the combined action of a mixed alkali roasting reactant and an atmosphere crystal form promoter of carbon dioxide and water vapor at a lower temperature (100-275 ℃), so that insoluble lithium salt which is difficult to transform can be transformed into lithium-containing crystal form Li 2 Al 4 (CO 3 )(OH) 12 ·3H 2 O, the crystal form is insoluble in alkali, lithium can be leached at low temperature under the weak acidic condition, and high-purity lithium carbonate and villiaumite can be simultaneously recovered and obtained while the high-efficiency separation of lithium and fluorine is realized.
2: the invention is used for roasting and leachingNo HF gas is generated in the process, no environmental pollution is caused, and simultaneously, the energy consumption of low-temperature roasting treatment is low, and the obtained lithium-containing crystal form Li 2 Al 4 (CO 3 )(OH) 12 ·3H 2 The method has the advantages of stable O, insolubility in alkali, insolubility in water, complete leaching under the weak acid condition, mild leaching condition, quick and efficient fluorine and lithium separation, high lithium leaching rate, and convenience for operation and large-scale industrial application.
Drawings
FIG. 1 is a process flow diagram of the present invention.
Fig. 2 is an XRD pattern of the lithium-aluminum containing spent electrolyte.
Fig. 3 is an XRD pattern of lithium-aluminum-containing spent electrolyte after firing.
Detailed Description
The technical solution of the present invention is illustrated below, and the claimed scope of the present invention includes, but is not limited to, the following examples.
Example 1
(1) Firstly, grinding the lithium-containing waste aluminum electrolyte, mixing the ground lithium-containing waste aluminum electrolyte with mixed alkali (sodium hydroxide: sodium carbonate = 2.5).
(2) And (3) grinding the roasted material, performing water leaching treatment at the temperature of 70 ℃ for 1.5h according to the liquid-solid mass ratio of 6.
(3) Washing the filter residue for 4 times by using deionized water (the mass ratio of the washing liquid to the solid in a single time is 2); and evaporating, concentrating, crystallizing and filtering the washing water to obtain the high-purity villiaumite.
(4) Carrying out acid leaching treatment on the washing slag for 1h at the temperature of 70 ℃ by adopting dilute sulfuric acid with the pH of 2-3 according to the liquid-solid mass ratio of 6.
(5) And (3) adjusting the pH value of the leachate to 8-8.5 by using sodium hydroxide, filtering out impurities after reacting for a period of time, then carrying out evaporation concentration, then adding sodium carbonate (the addition amount is 1.5 times of the molar amount of lithium ions in the concentrated solution), precipitating crystals to obtain lithium carbonate, and carrying out the step (5) by circularly mixing the mother liquor and the leachate.
The whole process is calculated by the lithium content and the fluorine content in the lithium-containing waste aluminum electrolyte, the lithium recovery rate is 81.1 percent, and the fluorine recovery rate is 82.0 percent.
Example 2
(1) Firstly, grinding the lithium-containing waste aluminum electrolyte, mixing the ground lithium-containing waste aluminum electrolyte with mixed alkali (sodium hydroxide: sodium carbonate = 2.8).
(2) And (3) grinding the roasted material, performing water leaching treatment at the temperature of 70 ℃ for 1.5h according to the liquid-solid mass ratio of 6.
(3) Washing the filter residue for 5 times by using deionized water (the mass ratio of the washing liquid to the washing liquid in a single time is 2; and evaporating, concentrating, crystallizing and filtering the washing water to obtain the high-purity villiaumite.
(4) Carrying out acid leaching treatment on the washing slag for 1h at the temperature of 70 ℃ by adopting dilute sulfuric acid with the pH of 2-3 according to the liquid-solid mass ratio of 6.
(5) And (3) adjusting the pH value of the leachate to 8-8.5 by using sodium hydroxide, filtering out impurities after reacting for a period of time, then carrying out evaporation concentration, then adding sodium carbonate (the addition amount is 1.8 times of the molar amount of lithium ions in the concentrated solution), precipitating crystals to obtain lithium carbonate, and carrying out the step (5) by circularly mixing the mother liquor and the leachate.
The whole process is calculated by the lithium content and the fluorine content in the lithium-containing waste aluminum electrolyte, the lithium recovery rate is 83.5 percent, and the fluorine recovery rate is 83.2 percent.
Example 3
(1) Firstly, grinding the lithium-containing waste aluminum electrolyte, mixing the ground lithium-containing waste aluminum electrolyte with mixed alkali (sodium hydroxide: sodium carbonate = 3).
(2) And (3) grinding the roasted material, performing water leaching treatment for 1.5h at the temperature of 75 ℃ according to the liquid-solid mass ratio of 6.
(3) Washing the filter residue for 5 times by using deionized water (the solid-to-solid mass ratio of a single washing liquid is 2); and evaporating, concentrating, crystallizing and filtering the washing water to obtain the high-purity villiaumite.
(4) Carrying out acid leaching treatment on the washing slag for 1h at the temperature of 75 ℃ by adopting dilute sulfuric acid with the pH of 2-3 according to the liquid-solid mass ratio of 6 to 1, and carrying out solid-liquid separation to obtain acid leaching slag (which can be used for preparing aluminum electrolyte additionally) and leachate.
(5) And (4) adjusting the pH value of the leachate to 8-8.5 by adopting sodium hydroxide, filtering out impurities after reacting for a period of time, then carrying out evaporation concentration, then adding sodium carbonate (the addition amount is 1.5 times of the molar amount of lithium ions in the concentrated solution), precipitating crystals to obtain lithium carbonate, and carrying out step (5) by circularly mixing the mother liquor and the leachate.
The whole process is calculated by the lithium content and the fluorine content in the lithium-containing waste aluminum electrolyte, the lithium recovery rate is 82.2 percent, and the fluorine recovery rate is 83.0 percent.
Example 4
(1) Firstly, grinding the lithium-containing waste aluminum electrolyte, mixing the ground lithium-containing waste aluminum electrolyte with mixed alkali (sodium hydroxide: sodium carbonate = 2.5).
(2) And (3) grinding the roasted material, performing water leaching treatment for 1.5h at the temperature of 75 ℃ according to the liquid-solid mass ratio of 6.
(3) Washing the filter residue for 4 times by using deionized water (the mass ratio of the washing liquid to the solid in a single time is 2); and evaporating, concentrating, crystallizing and filtering the washing water to obtain the high-purity villiaumite.
(4) Carrying out acid leaching treatment on the washing slag for 1h at the temperature of 75 ℃ by adopting dilute sulfuric acid with the pH of 2-3 according to the liquid-solid mass ratio of 6 to 1, and carrying out solid-liquid separation to obtain acid leaching slag (which can be used for preparing aluminum electrolyte additionally) and leachate.
(5) And (3) adjusting the pH value of the leachate to 8-8.5 by using sodium hydroxide, filtering out impurities after reacting for a period of time, then carrying out evaporation concentration, then adding sodium carbonate (the addition amount is 1.5 times of the molar amount of lithium ions in the concentrated solution), precipitating crystals to obtain lithium carbonate, and carrying out the step (5) by circularly mixing the mother liquor and the leachate.
The whole process is calculated by the lithium content and the fluorine content in the lithium-containing waste aluminum electrolyte, the lithium recovery rate is 83.9 percent, and the fluorine recovery rate is 82.5 percent.
Example 5
(1) Firstly, grinding the lithium-containing waste aluminum electrolyte, mixing the ground lithium-containing waste aluminum electrolyte with mixed alkali (potassium hydroxide: potassium carbonate = 2.5).
(2) And (3) grinding the roasted material, performing water leaching treatment for 1.5h at the temperature of 75 ℃ according to the liquid-solid mass ratio of 6.
(3) Washing the filter residue for 5 times by using deionized water (the mass ratio of the washing liquid to the washing liquid in a single time is 2; and evaporating, concentrating, crystallizing and filtering the washing water to obtain the high-purity villiaumite.
(4) Carrying out acid leaching treatment on the washing slag for 1h at the temperature of 75 ℃ by adopting dilute sulfuric acid with the pH of 2-3 according to the liquid-solid mass ratio of 6 to 1, and carrying out solid-liquid separation to obtain acid leaching slag (which can be used for preparing aluminum electrolyte additionally) and leachate.
(5) And (4) adjusting the pH value of the leachate to 8-8.5 by adopting sodium hydroxide, filtering out impurities after reacting for a period of time, then carrying out evaporation concentration, then adding potassium carbonate (the addition amount is 1.5 times of the molar amount of lithium ions in the concentrated solution), separating out crystals to obtain lithium carbonate, and carrying out step (5) by circularly mixing the mother liquor and the leachate.
The whole process is calculated by the lithium content and the fluorine content in the lithium-containing waste aluminum electrolyte, the lithium recovery rate is 82.4 percent, and the fluorine recovery rate is 81.8 percent.
Example 6
(1) Firstly, grinding the lithium-containing waste aluminum electrolyte, mixing the ground lithium-containing waste aluminum electrolyte with mixed alkali (sodium hydroxide: sodium bicarbonate = 2.5).
(2) And (3) grinding the roasted material, performing water leaching treatment at the temperature of 70 ℃ for 1.5h according to the liquid-solid mass ratio of 6.
(3) Washing the filter residue for 4 times by using deionized water (the mass ratio of the washing liquid to the solid in a single time is 2); and evaporating, concentrating, crystallizing and filtering the washing water to obtain the high-purity villiaumite.
(4) Carrying out acid leaching treatment on the washing slag for 1h at the temperature of 75 ℃ by adopting dilute sulfuric acid with the pH of 2-3 according to the liquid-solid mass ratio of 6 to 1, and carrying out solid-liquid separation to obtain acid leaching slag (which can be used for preparing aluminum electrolyte additionally) and leachate.
(5) And (3) adjusting the pH value of the leachate to 8-8.5 by using sodium hydroxide, filtering out impurities after reacting for a period of time, then carrying out evaporation concentration, then adding sodium carbonate (the addition amount is 1.5 times of the molar amount of lithium ions in the concentrated solution), precipitating crystals to obtain lithium carbonate, and carrying out the step (5) by circularly mixing the mother liquor and the leachate.
The whole process is calculated by the lithium content and the fluorine content in the lithium-containing waste aluminum electrolyte, the lithium recovery rate is 84.4 percent, and the fluorine recovery rate is 81.3 percent.
Example 7
(1) Firstly, grinding the lithium-containing waste aluminum electrolyte, mixing the ground lithium-containing waste aluminum electrolyte with mixed alkali (sodium hydroxide: sodium carbonate = 2.5).
(2) And (3) grinding the roasted material, performing water leaching treatment at the temperature of 70 ℃ for 1.5h according to the liquid-solid mass ratio of 6.
(3) Washing the filter residue for 4 times by using deionized water (the mass ratio of the washing liquid to the solid in a single time is 2); and evaporating, concentrating, crystallizing and filtering the washing water to obtain the high-purity villiaumite.
(4) Carrying out acid leaching treatment on the washing slag for 1h at the temperature of 75 ℃ by adopting dilute sulfuric acid with the pH of 2-3 according to the liquid-solid mass ratio of 6 to 1, and carrying out solid-liquid separation to obtain acid leaching slag (which can be used for preparing aluminum electrolyte additionally) and leachate.
(5) And (3) adjusting the pH value of the leachate to 8-8.5 by using sodium hydroxide, filtering out impurities after reacting for a period of time, then carrying out evaporation concentration, then adding sodium carbonate (the addition amount is 1.5 times of the molar amount of lithium ions in the concentrated solution), precipitating crystals to obtain lithium carbonate, and carrying out the step (5) by circularly mixing the mother liquor and the leachate.
The whole process is calculated by the lithium content and the fluorine content in the lithium-containing waste aluminum electrolyte, the lithium recovery rate is 86.3 percent, and the fluorine recovery rate is 82.5 percent.
Example 8
(1) Firstly, grinding the lithium-containing waste aluminum electrolyte, mixing the ground lithium-containing waste aluminum electrolyte with mixed alkali (sodium hydroxide: sodium carbonate = 2.5).
(2) And (3) grinding the roasted material, performing water leaching treatment for 1.5h at the temperature of 75 ℃ according to the liquid-solid mass ratio of 6.
(3) Washing the filter residue for 4 times by using deionized water (the mass ratio of the washing liquid to the solid in a single time is 2); and evaporating, concentrating, crystallizing and filtering the washing water to obtain the high-purity villiaumite.
(4) Carrying out acid leaching treatment on the washing slag for 1h at the temperature of 75 ℃ by adopting dilute sulfuric acid with the pH of 2-3 according to the liquid-solid mass ratio of 6 to 1, and carrying out solid-liquid separation to obtain acid leaching slag (which can be used for preparing aluminum electrolyte additionally) and leachate.
(5) And (3) adjusting the pH value of the leachate to 8-8.5 by using sodium hydroxide, filtering out impurities after reacting for a period of time, then carrying out evaporation concentration, then adding sodium carbonate (the addition amount is 1.5 times of the molar amount of lithium ions in the concentrated solution), precipitating crystals to obtain lithium carbonate, and carrying out the step (5) by circularly mixing the mother liquor and the leachate.
The whole process is calculated by the lithium content and the fluorine content in the lithium-containing waste aluminum electrolyte, the lithium recovery rate is 78.7 percent, and the fluorine recovery rate is 75.3 percent.
Example 9
(1) Firstly, grinding the lithium-containing waste aluminum electrolyte, mixing the ground lithium-containing waste aluminum electrolyte with mixed alkali (sodium hydroxide: sodium carbonate = 2.5).
(2) And (3) grinding the roasted material, performing water leaching treatment for 1.5h at the temperature of 75 ℃ according to the liquid-solid mass ratio of 6.
(3) Washing the filter residue for 4 times by using deionized water (the mass ratio of the washing liquid to the solid in a single time is 2); and evaporating, concentrating, crystallizing and filtering the washing water to obtain the high-purity villiaumite.
(4) Carrying out acid leaching treatment on the washing slag for 1h at the temperature of 75 ℃ by adopting dilute sulfuric acid with the pH of 2-3 according to the liquid-solid mass ratio of 6 to 1, and carrying out solid-liquid separation to obtain acid leaching slag (which can be used for preparing aluminum electrolyte additionally) and leachate.
(5) And (3) adjusting the pH value of the leachate to 8-8.5 by using sodium hydroxide, filtering out impurities after reacting for a period of time, then carrying out evaporation concentration, then adding sodium carbonate (the addition amount is 1.5 times of the molar amount of lithium ions in the concentrated solution), precipitating crystals to obtain lithium carbonate, and carrying out the step (5) by circularly mixing the mother liquor and the leachate.
The whole process is calculated by the lithium content and the fluorine content in the lithium-containing waste aluminum electrolyte, the lithium recovery rate is 75.6 percent, and the fluorine recovery rate is 73.2 percent.
Example 10
(1) Firstly, grinding the lithium-containing waste aluminum electrolyte, mixing the ground lithium-containing waste aluminum electrolyte with mixed alkali (sodium hydroxide: sodium carbonate = 2.5).
(2) And (3) grinding the roasted material, performing water leaching treatment at the temperature of 70 ℃ for 1.5h according to the liquid-solid mass ratio of 6.
(3) Washing the filter residue for 4 times by using deionized water (the mass ratio of the washing liquid to the solid in a single time is 2); and evaporating, concentrating, crystallizing and filtering the washing water to obtain the high-purity villiaumite.
(4) Carrying out acid leaching treatment on the washing slag for 1.5h at the temperature of 70 ℃ by adopting dilute sulfuric acid with the pH of 2-3 according to the liquid-solid mass ratio of 6.
(5) And (3) adjusting the pH value of the leachate to 8-8.5 by using sodium hydroxide, filtering out impurities after reacting for a period of time, then carrying out evaporation concentration, then adding sodium carbonate (the addition amount is 1.5 times of the molar amount of lithium ions in the concentrated solution), precipitating crystals to obtain lithium carbonate, and carrying out the step (5) by circularly mixing the mother liquor and the leachate.
The whole process is calculated by the lithium content and the fluorine content in the lithium-containing waste aluminum electrolyte, the lithium recovery rate is 72.3 percent, and the fluorine recovery rate is 73.5 percent.
Comparative example 1
(1) Firstly, grinding the lithium-containing waste aluminum electrolyte, mixing the ground lithium-containing waste aluminum electrolyte with mixed alkali (sodium hydroxide: sodium carbonate = 2.8).
(2) And (3) grinding the roasted material, performing water leaching treatment at the temperature of 70 ℃ for 1.5h according to the liquid-solid mass ratio of 6.
(3) Washing the filter residue for 4 times by using deionized water (the mass ratio of the washing liquid to the solid in a single time is 2); and evaporating, concentrating, crystallizing and filtering the washing water to obtain the high-purity villiaumite.
(4) Carrying out acid leaching treatment on the washing slag for 1h at the temperature of 70 ℃ by adopting dilute sulfuric acid with the pH of 2-3 according to the liquid-solid mass ratio of 6.
(5) And (3) adjusting the pH value of the leachate to 8-8.5 by using sodium hydroxide, filtering out impurities after reacting for a period of time, then carrying out evaporation concentration, then adding sodium carbonate (the addition amount is 1.5 times of the molar amount of lithium ions in the concentrated solution), precipitating crystals to obtain lithium carbonate, and carrying out the step (5) by circularly mixing the mother liquor and the leachate.
The whole process is calculated by the lithium content and the fluorine content in the lithium-containing waste aluminum electrolyte, the recovery rate of lithium is 55.3 percent, and the recovery rate of fluorine is 70.5 percent.
Comparative example 2
(1) Firstly, grinding the lithium-containing waste aluminum electrolyte, mixing the ground lithium-containing waste aluminum electrolyte with mixed alkali (sodium hydroxide: sodium carbonate = 2.8).
(2) And (3) grinding the roasted material, performing water leaching treatment at the temperature of 70 ℃ for 1.5h according to the liquid-solid mass ratio of 6.
(3) Washing the filter residue for 4 times by using deionized water (the mass ratio of the washing liquid to the solid in a single time is 2); and evaporating, concentrating, crystallizing and filtering the washing water to obtain the high-purity villiaumite.
(4) Carrying out acid leaching treatment on the washing slag for 1h at the temperature of 70 ℃ by adopting dilute sulfuric acid with the pH value of 2-3 according to the liquid-solid mass ratio of 6.
(5) And (3) adjusting the pH value of the leachate to 8-8.5 by using sodium hydroxide, filtering out impurities after reacting for a period of time, then carrying out evaporation concentration, then adding sodium carbonate (the addition amount is 1.5 times of the molar amount of lithium ions in the concentrated solution), precipitating crystals to obtain lithium carbonate, and carrying out the step (5) by circularly mixing the mother liquor and the leachate.
The whole process is calculated by the lithium content and the fluorine content in the lithium-containing waste aluminum electrolyte, the lithium recovery rate is 62.6 percent, and the fluorine recovery rate is 75.8 percent.
Comparative example 3
(1) Firstly, grinding the lithium-containing waste aluminum electrolyte, mixing the ground lithium-containing waste aluminum electrolyte with mixed alkali (sodium hydroxide: sodium carbonate = 2.8).
(2) Carrying out acid leaching treatment on the roasted material for 1h at the temperature of 70 ℃ by adopting dilute sulfuric acid with the pH of 2-3 according to the liquid-solid mass ratio of 6.
(3) And (3) adjusting the pH value of the leachate to 8-8.5 by using sodium hydroxide, filtering out impurities after reacting for a period of time, then carrying out evaporation concentration, then adding sodium carbonate (the addition amount is 1.5 times of the molar amount of lithium ions in the concentrated solution), precipitating crystals to obtain lithium carbonate, and carrying out the step (5) by circularly mixing the mother liquor and the leachate.
The whole process is calculated by the lithium content and the fluorine content in the lithium-containing waste aluminum electrolyte, the lithium recovery rate is 65.8 percent, and the fluorine recovery rate is 30.3 percent.
Comparative example 4
(1) Firstly, grinding the lithium-containing waste aluminum electrolyte, mixing the ground lithium-containing waste aluminum electrolyte with mixed alkali (sodium hydroxide: sodium carbonate = 2.8).
(2) And (3) grinding the roasted material, performing water leaching treatment at the temperature of 70 ℃ for 1.5h according to the liquid-solid mass ratio of 6.
(3) Washing the filter residue for 4 times by using deionized water (the mass ratio of the washing liquid to the solid in a single time is 2); and evaporating, concentrating, crystallizing and filtering the washing water to obtain the high-purity villiaumite.
(4) Carrying out acid leaching treatment on the washing slag for 1h at the temperature of 70 ℃ by adopting dilute sulfuric acid with the pH of 2-3 according to the liquid-solid mass ratio of 6.
(5) And (3) adjusting the pH value of the leachate to 8-8.5 by using sodium hydroxide, filtering out impurities after reacting for a period of time, then carrying out evaporation concentration, then adding sodium carbonate (the addition amount is 1.5 times of the molar amount of lithium ions in the concentrated solution), precipitating crystals to obtain lithium carbonate, and carrying out the step (5) by circularly mixing the mother liquor and the leachate.
The whole process is calculated by the lithium content and the fluorine content in the lithium-containing waste aluminum electrolyte, the lithium recovery rate is 60.2 percent, and the fluorine recovery rate is 68.5 percent.
Comparative example 5
(1) Firstly, grinding the lithium-containing waste aluminum electrolyte, mixing the ground lithium-containing waste aluminum electrolyte with mixed alkali (sodium hydroxide: sodium carbonate = 2.8).
(2) And (3) grinding the roasted material, performing water leaching treatment at the temperature of 70 ℃ for 1.5h according to the liquid-solid mass ratio of 6.
(3) Washing filter residues for 4 times by using deionized water (the mass ratio of the washing liquid to the solid in a single washing is 2; and evaporating, concentrating, crystallizing and filtering the washing water to obtain the high-purity villiaumite.
(4) Carrying out acid leaching treatment on the washing slag for 1h at the temperature of 70 ℃ by adopting dilute sulfuric acid with the pH of 2-3 according to the liquid-solid mass ratio of 6.
(5) And (3) adjusting the pH value of the leachate to 8-8.5 by using sodium hydroxide, filtering out impurities after reacting for a period of time, then carrying out evaporation concentration, then adding sodium carbonate (the addition amount is 1.5 times of the molar amount of lithium ions in the concentrated solution), precipitating crystals to obtain lithium carbonate, and carrying out the step (5) by circularly mixing the mother liquor and the leachate.
The whole process is calculated by the lithium content and the fluorine content in the lithium-containing waste aluminum electrolyte, the lithium recovery rate is 57.3 percent, and the fluorine recovery rate is 60.2 percent.
Comparative example 6
(1) Firstly, grinding the lithium-containing waste aluminum electrolyte, mixing the ground lithium-containing waste aluminum electrolyte with mixed alkali (sodium hydroxide: sodium carbonate = 2.8).
(2) And (3) grinding the roasted material, performing water leaching treatment at the temperature of 70 ℃ for 1.5h according to the liquid-solid mass ratio of 6.
(3) Washing the filter residue for 4 times by using deionized water (the mass ratio of the washing liquid to the solid in a single time is 2); and evaporating, concentrating, crystallizing and filtering the washing water to obtain the high-purity villiaumite.
(4) Carrying out acid leaching treatment on the washing slag for 1h at the temperature of 70 ℃ by adopting dilute sulfuric acid with the pH of 2-3 according to the liquid-solid mass ratio of 6.
(5) And (3) adjusting the pH value of the leachate to 8-8.5 by using sodium hydroxide, filtering out impurities after reacting for a period of time, then carrying out evaporation concentration, then adding sodium carbonate (the addition amount is 1.5 times of the molar amount of lithium ions in the concentrated solution), precipitating crystals to obtain lithium carbonate, and carrying out the step (5) by circularly mixing the mother liquor and the leachate.
The whole process is calculated by the lithium content and the fluorine content in the lithium-containing waste aluminum electrolyte, the lithium recovery rate is 30.5 percent, and the fluorine recovery rate is 68.3 percent.
Comparative example 7
(1) Firstly, grinding the lithium-containing waste aluminum electrolyte, mixing the ground lithium-containing waste aluminum electrolyte with mixed alkali (sodium hydroxide: sodium carbonate = 2.5).
(2) And (3) grinding the roasted material, performing water leaching treatment at the temperature of 70 ℃ for 1.5h according to the liquid-solid mass ratio of 6.
(3) Washing the filter residue for 4 times by using deionized water (the mass ratio of the washing liquid to the solid in a single time is 2); and evaporating, concentrating, crystallizing and filtering the washing water to obtain the high-purity villiaumite.
(4) Carrying out acid leaching treatment on the washing slag for 1h at the temperature of 75 ℃ by adopting dilute sulfuric acid with the pH of 2-3 according to the liquid-solid mass ratio of 6.
(5) And (3) adjusting the pH value of the leachate to 8-8.5 by using sodium hydroxide, filtering out impurities after reacting for a period of time, then carrying out evaporation concentration, then adding sodium carbonate (the addition amount is 1.5 times of the molar amount of lithium ions in the concentrated solution), precipitating crystals to obtain lithium carbonate, and carrying out the step (5) by circularly mixing the mother liquor and the leachate.
The whole process is calculated by the lithium content and the fluorine content in the lithium-containing waste aluminum electrolyte, the lithium recovery rate is 46.8 percent, and the fluorine recovery rate is 65.3 percent.
Comparative example 8
(1) Firstly, grinding the lithium-containing waste aluminum electrolyte, mixing the ground lithium-containing waste aluminum electrolyte with mixed alkali (sodium hydroxide: sodium carbonate = 2.8).
(2) And (3) grinding the roasted material, performing water leaching treatment at the temperature of 70 ℃ for 1.5h according to the liquid-solid mass ratio of 6.
(3) Washing the filter residue for 4 times by using deionized water (the mass ratio of the washing liquid to the solid in a single time is 2); and evaporating, concentrating, crystallizing and filtering the washing water to obtain the high-purity villiaumite.
(4) Carrying out acid leaching treatment on the washing slag for 1h at the temperature of 70 ℃ by adopting dilute sulfuric acid with the pH of 2-3 according to the liquid-solid mass ratio of 6.
(5) And (3) adjusting the pH value of the leachate to 8-8.5 by using sodium hydroxide, filtering out impurities after reacting for a period of time, then carrying out evaporation concentration, then adding sodium carbonate (the addition amount is 1.5 times of the molar amount of lithium ions in the concentrated solution), precipitating crystals to obtain lithium carbonate, and carrying out the step (5) by circularly mixing the mother liquor and the leachate.
The whole process is calculated by the lithium content and the fluorine content in the lithium-containing waste aluminum electrolyte, the recovery rate of lithium is 36.3 percent, and the recovery rate of fluorine is 63.2 percent.
Comparative example 9
(1) Firstly, grinding the lithium-containing waste aluminum electrolyte, mixing the ground lithium-containing waste aluminum electrolyte with mixed alkali (sodium hydroxide: sodium carbonate = 2.8).
(2) And (3) grinding the roasted material, performing water leaching treatment at the temperature of 70 ℃ for 1.5h according to the liquid-solid mass ratio of 6.
(3) Washing the filter residue for 4 times by using deionized water (the mass ratio of the washing liquid to the solid in a single time is 2); and evaporating, concentrating, crystallizing and filtering the washing water to obtain the high-purity villiaumite.
(4) Carrying out acid leaching treatment on the washing slag for 1h at the temperature of 70 ℃ by adopting dilute sulfuric acid with the pH value of 2-3 according to the liquid-solid mass ratio of 6.
(5) And (3) adjusting the pH value of the leachate to 8-8.5 by using sodium hydroxide, filtering out impurities after reacting for a period of time, then carrying out evaporation concentration, then adding sodium carbonate (the addition amount is 1.5 times of the molar amount of lithium ions in the concentrated solution), precipitating crystals to obtain lithium carbonate, and carrying out the step (5) by circularly mixing the mother liquor and the leachate.
The whole process is calculated by the lithium content and the fluorine content in the lithium-containing waste aluminum electrolyte, the lithium recovery rate is 31.4 percent, and the fluorine recovery rate is 62.5 percent.
Comparative example 10
(1) Firstly, grinding the lithium-containing waste aluminum electrolyte, mixing the ground lithium-containing waste aluminum electrolyte with sodium hydroxide according to the mass ratio of 100 to obtain a mixture, then placing the mixture into a roasting furnace, introducing carbon dioxide and steam to ensure that the air pressure in the furnace is 1.1-1.3 times of the standard atmospheric pressure, and then heating to 200 ℃ for roasting treatment for 4 hours to obtain a roasted material.
(2) And (3) grinding the roasted material, performing water leaching treatment at the temperature of 70 ℃ for 1.5h according to the liquid-solid mass ratio of 6.
(3) Washing the filter residue for 5 times by using deionized water (the solid-to-solid mass ratio of a single washing liquid is 2); and evaporating, concentrating, crystallizing and filtering the washing water to obtain the high-purity villiaumite.
(4) Carrying out acid leaching treatment on the washing slag for 1h at the temperature of 70 ℃ by adopting dilute sulfuric acid with the pH of 2-3 according to the liquid-solid mass ratio of 6.
(5) And (3) adjusting the pH value of the leachate to 8-8.5 by using sodium hydroxide, filtering out impurities after reacting for a period of time, then carrying out evaporation concentration, then adding sodium carbonate (the addition amount is 1.8 times of the molar amount of lithium ions in the concentrated solution), precipitating crystals to obtain lithium carbonate, and carrying out the step (5) by circularly mixing the mother liquor and the leachate.
The whole process is calculated by the lithium content and the fluorine content in the lithium-containing waste aluminum electrolyte, the lithium recovery rate is 43.2 percent, and the fluorine recovery rate is 62.3 percent.
Example 11
(1) Firstly, grinding the lithium-containing waste aluminum electrolyte, mixing the ground lithium-containing waste aluminum electrolyte with sodium carbonate according to the mass ratio of 100 to obtain a mixture, then placing the mixture into a roasting furnace, introducing carbon dioxide and steam to ensure that the air pressure in the furnace is 1.1-1.3 times of the standard atmospheric pressure, and then heating to 200 ℃ for roasting treatment for 4 hours to obtain a roasted material.
(2) And (3) grinding the roasted material, performing water leaching treatment at the temperature of 70 ℃ for 1.5h according to the liquid-solid mass ratio of 6.
(3) Washing the filter residue for 5 times by using deionized water (the solid-to-solid mass ratio of a single washing liquid is 2); the fluoride salt can not be obtained after the washing water is evaporated, concentrated, crystallized and filtered.
(4) Carrying out acid leaching treatment on the washing slag for 1h at the temperature of 70 ℃ by adopting dilute sulfuric acid with the pH of 2-3 according to the liquid-solid mass ratio of 6.
(5) And (3) adjusting the pH value of the leachate to 8-8.5 by using sodium hydroxide, filtering out impurities after reacting for a period of time, then carrying out evaporation concentration, then adding sodium carbonate (the addition amount is 1.8 times of the molar amount of lithium ions in the concentrated solution), precipitating crystals to obtain lithium carbonate, and carrying out the step (5) by circularly mixing the mother liquor and the leachate. The whole process is calculated by the lithium content and the fluorine content in the lithium-containing waste aluminum electrolyte, the lithium recovery rate is 15.3 percent, and the fluorine recovery rate is 0 percent (fluorine element has no recovery value, so the fluorine element is not recovered).
Claims (10)
1. A method for extracting lithium and collecting fluorine by roasting lithium-containing waste aluminum electrolyte at low temperature is characterized by comprising the following steps: the method specifically comprises the following steps:
1) Firstly, mixing lithium-containing waste aluminum electrolyte powder with alkali to obtain a mixture, and then roasting the mixture in a double-atmosphere environment of carbon dioxide and water vapor to obtain a roasted material;
2) Refining the roasted material, then carrying out water leaching treatment, and carrying out solid-liquid separation to obtain filter residue and filtrate; performing next-stage treatment on the filter residue, and separately treating the filtrate;
3) Washing the filter residue to obtain washing residue and washing water; concentrating and crystallizing the washing water to obtain high-purity villiaumite; acid leaching treatment is carried out on the washing slag by adopting acid, and acid leaching slag and leachate are obtained after solid-liquid separation; performing next-stage treatment on the leachate, and separately treating acid leaching residues;
4) And regulating the leachate to be alkaline, filtering out impurities, concentrating, adding soluble carbonate, finally precipitating crystals to obtain lithium carbonate, and circularly using the mother liquor for lithium precipitation.
2. The method of claim 1, wherein: the method further comprises the following steps:
5) Mixing the filtrate obtained in the step 2) with the residual liquid after the fluoride salt is separated out from the washing water in the step 3) to obtain a mixed solution, and preparing the mixed solution to obtain the cryolite.
3. The method according to claim 1 or 2, characterized in that: in step 1), the alkali comprises a mixture of one or more of sodium hydroxide, potassium hydroxide and calcium hydroxide and one or more of sodium carbonate, potassium carbonate, sodium bicarbonate and potassium bicarbonate; preferably a mixed alkali composed of sodium hydroxide and sodium carbonate or a mixed alkali composed of potassium hydroxide and potassium carbonate;
preferably, in the mixed alkali, the mixing mass ratio of sodium hydroxide to sodium carbonate and potassium hydroxide to potassium carbonate is 1-10, preferably 1.5-8; 1, 1.2, 1.5;
preferably, the amount of the mixed alkali added is 10 to 50%, preferably 15 to 40%, more preferably 20 to 35%, for example, 10%, 12%, 15%, 18%, 20%, 22%, 24%, 26%, 28%, 30%, 33%, 35%, 38%, 40%, 45%, 50% of the total mass of the lithium-containing waste aluminum electrolyte.
4. The method according to any one of claims 1-3, wherein: in the step 1), the dual-atmosphere environment of carbon dioxide and water vapor is specifically to fill carbon dioxide and water vapor into the roasting equipment, and make the atmosphere pressure in the roasting equipment be positive pressure, preferably the atmosphere pressure in the roasting equipment is 1.1-2 times of the standard atmospheric pressure, and more preferably 1.3-1.8 times of the standard atmospheric pressure; for example, 1.1 times, 1.2 times, 1.3 times, 1.4 times, 1.5 times, 1.6 times, 1.7 times, 1.8 times, 1.9 times, or 2 times the standard atmospheric pressure.
5. The method according to any one of claims 1-4, wherein: in step 1), the temperature of the calcination treatment is 100 to 275 ℃, preferably 120 to 270 ℃, more preferably 150 to 260 ℃, for example, 120 ℃, 125 ℃, 130 ℃, 135 ℃, 140 ℃, 145 ℃, 150 ℃, 155 ℃, 160 ℃, 165 ℃, 170 ℃, 175 ℃, 180 ℃, 185 ℃, 190 ℃, 200 ℃, 210 ℃, 220 ℃, 230 ℃, 240 ℃, 250 ℃, 260 ℃, 270 ℃; and/or
The time of the roasting treatment is 1 to 10 hours, preferably 1.5 to 8 hours, more preferably 2 to 5 hours, for example, 1 hour, 1.2 hours, 1.5 hours, 1.8 hours, 2.5 hours, 3 hours, 3.5 hours, 4 hours, 5 hours, 6 hours, 7 hours, 8 hours, 9 hours, 10 hours.
6. The method according to any one of claims 1-5, wherein: in step 2), the water for water leaching treatment is deionized water, and the liquid-solid ratio in water leaching treatment is 3 to 10, preferably 4 to 9, and more preferably 5 to 8, and is, for example, 3; and/or
The time of the water immersion treatment is 1-8h, preferably 1.5-7h, more preferably 2-6h, for example, one of 1h, 1.5h, 1.8h, 2.0h, 2.5h, 3h, 3.5h, 4h, 4.5h, 5h, 5.5h, 6h, 6.5h, 7h, 7.5h and 8 h; and/or
The leaching temperature of the water leaching treatment is 50-90 deg.C, preferably 55-85 deg.C, more preferably 60-80 deg.C, such as one of 55 deg.C, 60 deg.C, 65 deg.C, 70 deg.C, 75 deg.C, 80 deg.C, 85 deg.C, and 90 deg.C.
7. The method according to any one of claims 1-6, wherein: in step 3), the filter residue is washed for 1 to 10 times, preferably 2 to 8 times, more preferably 3 to 5 times, for example, one of 2 times, 3 times, 4 times, 5 times, 6 times, 7 times and 8 times;
as a preference, the liquid-solid ratio at a single washing of the filter residue is 1-5, preferably 1.5-4, and more preferably 2-3.5, for example, 1.
8. The method according to any one of claims 1-7, wherein: in the step 3), the acid for acid leaching treatment is a dilute acid solution with the pH value lower than 4, preferably one or more of a sulfuric acid solution, a hydrochloric acid solution, a nitric acid solution, a phosphoric acid solution, an oxalic acid solution and an acetic acid solution with the pH value of 1-3.5; and/or
The liquid-solid ratio of the acid leaching treatment is 2 to 8, preferably 2.5 to 6, and more preferably 3 to 5, and is, for example, one of 2; and/or
The acid leaching time is 1.5-8h, preferably 2-6h, more preferably 3-5h, for example, 1.5h, 1.8h, 2.0h, 2.2h, 2.5h, 2.8h, 3h, 3.2h, 3.5h, 3.8h, 4h, 4.2h, 4.5h, 4.8h, 5h, 5.5h, 6h, 6.5h, 7h, 7.5h, and 8 h; and/or
The temperature of the acid leaching treatment is 30-90 deg.C, preferably 40-80 deg.C, more preferably 50-70 deg.C, such as 30 deg.C, 35 deg.C, 40 deg.C, 45 deg.C, 50 deg.C, 55 deg.C, 60 deg.C, 65 deg.C, 70 deg.C, 75 deg.C, 80 deg.C, 85 deg.C, 90 deg.C.
9. The method according to any one of claims 1-8, wherein: in the step 4), the leachate is adjusted to be alkaline, namely, the pH of the leachate is adjusted to be 7.5-10, preferably 8-9.5, and more preferably 8.5-9 by using sodium hydroxide and/or potassium hydroxide; and/or
The soluble carbonate is sodium carbonate and/or potassium carbonate, and sodium carbonate is preferred.
10. The method of claim 9, wherein: the solid-liquid separation is one or more of filter pressing, negative pressure suction filtration and centrifugal separation; and/or
The concentration is evaporation concentration, and the concentration multiple is 3-10 times, preferably 4-8 times; the water vapor generated by evaporation and concentration is recycled for roasting treatment.
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Citations (6)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| GB934447A (en) * | 1960-02-09 | 1963-08-21 | Mini Richesses Nature | Lithium carbonate production |
| US3380802A (en) * | 1963-09-30 | 1968-04-30 | Mini Richesses Nature | Carbonatizing roast of lithiumbearing ores |
| CN107587167A (en) * | 2017-10-27 | 2018-01-16 | 东北大学 | A kind of crystal formation of aluminium electrolyte containing lithium changes method |
| CN109777960A (en) * | 2019-02-25 | 2019-05-21 | 河北工程大学 | A method of separating and recovering lithium, aluminium from flyash |
| CN111268705A (en) * | 2020-03-06 | 2020-06-12 | 中国地质大学（北京） | Method for preparing lithium carbonate by using lepidolite powder |
| CN113684369A (en) * | 2021-09-16 | 2021-11-23 | 中南大学 | Method for treating waste industrial lithium-containing aluminum electrolyte |
-
2022
- 2022-12-13 CN CN202211601031.3A patent/CN115818676B/en active Active
Patent Citations (6)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| GB934447A (en) * | 1960-02-09 | 1963-08-21 | Mini Richesses Nature | Lithium carbonate production |
| US3380802A (en) * | 1963-09-30 | 1968-04-30 | Mini Richesses Nature | Carbonatizing roast of lithiumbearing ores |
| CN107587167A (en) * | 2017-10-27 | 2018-01-16 | 东北大学 | A kind of crystal formation of aluminium electrolyte containing lithium changes method |
| CN109777960A (en) * | 2019-02-25 | 2019-05-21 | 河北工程大学 | A method of separating and recovering lithium, aluminium from flyash |
| CN111268705A (en) * | 2020-03-06 | 2020-06-12 | 中国地质大学（北京） | Method for preparing lithium carbonate by using lepidolite powder |
| CN113684369A (en) * | 2021-09-16 | 2021-11-23 | 中南大学 | Method for treating waste industrial lithium-containing aluminum electrolyte |
Cited By (2)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN117285054A (en) * | 2023-11-02 | 2023-12-26 | 巩义市红旗炉料有限公司 | Method for preparing lithium carbonate from lithium extracted from aluminum electrolyte |
| CN117285054B (en) * | 2023-11-02 | 2024-03-29 | 巩义市红旗炉料有限公司 | Method for preparing lithium carbonate from lithium extracted from aluminum electrolyte |
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