CN116903002B - A full-element recovery method for nickel-cobalt-lithium manganese oxide waste - Google Patents
A full-element recovery method for nickel-cobalt-lithium manganese oxide waste Download PDFInfo
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Abstract
本发明公开了一种镍钴锰酸锂废料的全元素回收方法，属于废旧锂电池回收领域，解决现有回收方法存在的药剂消耗量大、工艺复杂、会产生大量的十水硫酸钠影响生产的问题。本发明方法：物料与碳粉混合煅烧；调浆碳化；热解；循环酸浸，利用溶解度特性，通过温度控制，分步结晶析出不同的硫酸盐。本发明大大地减少了硫酸钠的产生，提高了生产的流畅性和连续性。由于没有了大量的十水硫酸钠的产生，避免十水硫酸钠中的结晶水带走金属锂，锂损失明显减少。本发明方法流程短，简单易操作，不使用有机溶剂，节约生产成本；不使用双氧水，不存在存储问题、安全问题、环境问题、设备问题。The present invention discloses a method for recovering all elements of nickel-cobalt-manganese-oxide lithium waste, which belongs to the field of waste lithium battery recovery and solves the problems of large reagent consumption, complex process and large amount of sodium sulfate decahydrate that affects production in existing recovery methods. The method of the present invention comprises the following steps: mixing and calcining materials with carbon powder; slurry carbonization; pyrolysis; cyclic acid leaching, utilizing solubility characteristics, and controlling temperature to crystallize and precipitate different sulfates in steps. The present invention greatly reduces the generation of sodium sulfate and improves the fluency and continuity of production. Since there is no large amount of sodium sulfate decahydrate generated, the crystal water in the sodium sulfate decahydrate is prevented from taking away metallic lithium, and lithium loss is significantly reduced. The method of the present invention has a short process, is simple and easy to operate, does not use organic solvents, and saves production costs; does not use hydrogen peroxide, and does not have storage problems, safety problems, environmental problems, and equipment problems.
Description
Technical Field
The invention belongs to the field of waste lithium batteries, and particularly relates to a full-element recovery method of nickel cobalt lithium manganate waste.
Background
At present, waste materials for recovering nickel, cobalt and manganese generally adopt a fire method and a wet method. The wet recovery of nickel cobalt manganese waste mainly comprises two recovery techniques:
Firstly, dissolving valuable metal nickel cobalt manganese lithium in nickel cobalt lithium manganate waste into a solution by using acid and hydrogen peroxide, then adding sodium hydroxide to adjust the pH value to 12 to obtain a ternary precursor and a lithium-containing solution, and adding sodium carbonate into the lithium-containing solution to obtain lithium carbonate. The nickel, cobalt and manganese are recovered as ternary precursors and the lithium is recovered as lithium carbonate. The method needs to consume a large amount of acid, alkali and hydrogen peroxide, and has large medicament consumption and high storage risk coefficient. In addition, the biggest problem lies in the production of a large amount of sodium sulfate decahydrate, in the cold north of winter temperature, the production of a large amount of sodium sulfate decahydrate can lead to the jam of whole production system pipeline, and the solution difficult problem of all slow sodium sulfate decahydrate for production stagnates, and the crystal in the system is very hard and is difficult to clear, can only cut the pipeline, and the influence that causes the production is huge.
And secondly, dissolving valuable metals nickel cobalt manganese lithium in the nickel cobalt lithium manganate waste into a solution by using acid and hydrogen peroxide, saponifying an extracting agent by using a large amount of sodium hydroxide, extracting the valuable metals in the nickel cobalt lithium manganate waste in multiple stages, separating an organic phase from an inorganic phase, stripping the organic phase to prepare corresponding sulfate, such as nickel sulfate, cobalt sulfate and manganese sulfate, and adding the inorganic phase into sodium carbonate to prepare lithium carbonate. The method needs a large amount of organic solvents, has complex acid-base process and long process flow. In addition, this method also produces a large amount of sodium sulfate decahydrate in the stripping stage and the lithium carbonate precipitation stage as in the first method, and the above-described problems are also present.
Disclosure of Invention
The invention aims to provide a full-element recovery method of nickel cobalt lithium manganate waste, which aims to solve the problems of large medicament consumption, complex process and production influence caused by a large amount of sodium sulfate decahydrate in the existing recovery method.
The technical scheme of the invention is that the full-element recovery method of the nickel cobalt lithium manganate waste comprises the following steps:
A. Uniformly mixing nickel cobalt lithium manganate waste with carbon powder, and calcining for 2-3 hours at 600-650 ℃ under the condition of nitrogen flow;
B. Adding water into the mixture obtained after calcining in the step A to prepare slurry, then introducing carbon dioxide to carbonize for 2-3h, and carrying out solid-liquid separation after carbonization to obtain carbonized slag and carbonized liquid;
C. Heating the carbonized liquid obtained in the step B to be more than or equal to 90 ℃ for pyrolysis, and reacting for 1-2h to obtain lithium carbonate and pyrolyzed liquid;
D. Adding dilute sulfuric acid with the acid concentration of 100-120g/L into the carbonized slag obtained in the step B for acid leaching, wherein the reaction temperature is not lower than 95 ℃, the reaction time is 40-90min, solid-liquid separation is carried out to obtain a mixed solution I of lithium sulfate and manganese sulfate and acid leaching slag I, adding concentrated sulfuric acid into the mixed solution I to adjust the acid concentration to 100-120g/L, adding the carbonized slag, continuing the acid leaching operation according to the conditions, carrying out cyclic leaching in such a way that the manganese concentration reaches 110g/L, carrying out cyclic leaching at the temperature of 25-35 ℃ to enable the manganese content in the leaching solution to reach 180-190g/L, then carrying out solid-liquid separation at normal temperature to obtain a mixed solution II and acid leaching slag II, heating the mixed solution II to be higher than or equal to 95 ℃, separating the separated solution into manganese sulfate crystals and separated solution, keeping the temperature of the separated solution to be not lower than 90 ℃, and continuously carrying out acid leaching operation when the lithium content in the solution reaches 15-20g/L, adding sodium hydroxide into the solution to adjust the pH to 11-12, and precipitating lithium carbonate solution;
E. Washing the acid leaching slag I and the acid leaching slag II in the step D by pure water at normal temperature, then carrying out solid-liquid separation to obtain washing slag and washing liquid, and returning the washing liquid to the step D for continuous leaching;
F. Supplementing carbon powder into the washing slag obtained in the step E, calcining for 2-3 hours at 650-750 ℃ under the condition of nitrogen flow, and then continuously calcining for 2-3 hours at 650-700 ℃ under the condition of air flow to obtain a calcined material;
G. Adding dilute sulfuric acid with the acid concentration of 180-200g/L into the calcined material obtained in the step F for acid leaching, wherein the reaction temperature is not less than 95 ℃, the reaction time is 40-90min, after solid-liquid separation, adding concentrated sulfuric acid into the obtained leaching solution to adjust the acid concentration to 180-200g/L, adding the calcined material, continuing the acid leaching operation according to the conditions, circularly leaching in such a way that the total concentration of nickel and cobalt in the leaching solution reaches 360-400g/L, then keeping the temperature not less than 90 ℃ and performing hot solid-liquid separation to obtain a mixed solution III and leached slag III, and cooling the mixed solution III to normal temperature to separate out cobalt sulfate and nickel sulfate crystals;
H. And (C) washing the leaching residue III obtained in the step (G) by using pure water, wherein the washing temperature is more than 90 ℃, separating the hot liquid, and returning the obtained washing liquid to the step (G) for leaching.
As a further improvement of the invention, in the step A, the mass ratio of the carbon powder is 10% -15% of that of the mixed powder.
As a further improvement of the invention, in the step B, when preparing the slurry, the ratio of the calcined material to water is prepared according to the lithium content of 7.5-8g/L in the carbonized liquid.
As a further improvement of the invention, in the step D, the solid-liquid mass ratio of the carbide slag is 1:5-6 when the acid leaching operation is carried out.
As a further improvement of the invention, in the step E, the washing solid-liquid mass ratio is 1:3-8, and the washing time is 30-50min.
As a further improvement of the invention, in the step F, the carbon powder is supplemented with 9-12% by mass.
As a further improvement of the invention, in the step G, the solid-liquid mass ratio of the calcined material is 1:5-6 when the acid leaching operation is carried out.
As a further improvement of the invention, in the step H, the washing solid-liquid mass ratio is 1:4-8, and the washing time is 30-50min.
The invention provides a recovery method with less alkali consumption, no use of organic solvent and little generation of sodium sulfate decahydrate, which has the following advantages compared with the prior art:
1. The invention greatly reduces the generation of sodium sulfate, solves the problem that a large amount of sodium sulfate is easy to generate in the conventional method, greatly reduces the production cost, improves the smoothness and continuity of production, and does not cause the collapse of a production system due to the problem of sodium sulfate crystallization.
2. The recovery rate of lithium is high, and because a large amount of sodium sulfate decahydrate is not generated, the crystallization water in the sodium sulfate decahydrate is prevented from taking out metal lithium, and the lithium loss is obviously reduced.
3. The method has short flow, is simple and easy to operate, and only utilizes the characteristic that the solubility of different sulfates is different along with the change of temperature, and separates out different sulfates through fractional crystallization by temperature control.
4. The method does not use an organic solvent, and saves the production cost. The method does not use hydrogen peroxide, and has no storage problem, safety problem, environmental problem and equipment problem.
Detailed Description
The present invention will be described in detail with reference to the following specific embodiments.
Example 1,
The full-element recovery method of the lithium nickel cobalt manganese oxide waste material comprises the following steps:
A. 85g of ternary powder and 15g of carbon powder are uniformly mixed, and then are calcined for 2 hours at 650 ℃ under the nitrogen flow condition.
B. And C, taking 80g of the calcined mixture in the step A, adding 640ml of pure water to prepare slurry, then introducing carbon dioxide to carbonize for 2h, and carrying out solid-liquid separation after carbonization to obtain carbonized slag and carbonized liquid, wherein the lithium content of the carbonized liquid is 7.5g/L.
C. and C, heating the carbonized liquid obtained in the step B to above 90 ℃ for pyrolysis, and reacting for 1h to obtain lithium carbonate and pyrolyzed liquid.
D. Adding dilute sulfuric acid with the acid concentration of 120g/L into the carbonized slag obtained in the step B for acid leaching, wherein the solid-liquid ratio of the carbonized slag is 1:5, the reaction temperature is above 95 ℃, the reaction time is 80min, solid-liquid separation is carried out to obtain a mixed solution I of lithium sulfate and manganese sulfate and acid leaching slag I, the pH value of the mixed solution I is about 5, concentrated sulfuric acid is added into the mixed solution I to adjust the acid concentration to 120g/L, the carbonized slag is added, acid leaching operation is continuously carried out according to the conditions, the circulating leaching is carried out in such a way that the manganese concentration reaches at least 110g/L, the circulating leaching is carried out at the temperature of 30 ℃ to enable the manganese content in the leaching solution to reach 180g/L, then the mixed solution II and the acid leaching slag II are obtained through solid-liquid separation at normal temperature, the temperature is raised to be above 95 ℃, the mixed solution II is separated into manganese sulfate crystals and separated liquid after the heat setting, the temperature of the separated liquid is kept to be not less than 90 ℃, the separated liquid is continuously used for the previous acid leaching operation, the manganese sulfate crystals are continuously separated out, when the lithium content in the solution reaches 15g/L, the lithium content is added into sodium hydroxide solution, the lithium carbonate solution is added into the solution, and the lithium carbonate solution is adjusted to have the lithium content of 9 times, and the total mass of lithium is obtained.
Leaching is carried out under the condition that the acidity of the low acid is 100-120g/L and hydrogen peroxide is not added, only manganese in the carbide slag is leached, and nickel and cobalt are not leached. The solubility of manganese sulfate increases and then decreases with increasing temperature.
E. and D, washing the acid leaching slag I and the acid leaching slag II in the step D by using pure water, wherein the mass ratio of washing solid to liquid is 1:5, the washing time is 40min, the washing temperature is normal temperature, then, the washing slag and the washing liquid are obtained by solid-liquid separation, and the washing liquid returns to the step D to be leached continuously.
F. And E, continuously supplementing carbon powder into the washing slag obtained in the step E, wherein the supplementing mass ratio of the carbon powder is 5%, calcining for 2 hours at 650 ℃ under the condition of nitrogen flow, and then calcining for 2 hours at 800 ℃ under the condition of air flow to obtain a calcined material.
G. Adding dilute sulfuric acid with the acid concentration of 180g/L into the calcined material obtained in the step F for acid leaching, wherein the reaction temperature is above 95 ℃, the reaction time is 80min, adding concentrated sulfuric acid into the obtained leaching solution to adjust the acid concentration to 200g/L after solid-liquid separation, adding the calcined material, continuing acid leaching operation according to the conditions, circularly leaching in such a way that the total concentration of nickel and cobalt in the leaching solution reaches 360g/L, then keeping the temperature above 90 ℃ for hot solid-liquid separation to obtain a mixed solution III and leached slag III, and cooling the mixed solution III to normal temperature to separate out cobalt sulfate and nickel sulfate crystals. The liquid is returned to the leaching operation of the step, and mixed crystals of nickel sulfate and cobalt sulfate are continuously precipitated.
Leaching the calcined material in an acid solution with the acidity of 180-200g/L to obtain a sulfate solution containing nickel and cobalt. The solubility of nickel sulfate and cobalt sulfate increases with increasing temperature.
H. And C, washing the leached slag III by pure water, wherein the content of nickel and cobalt in the leached slag III obtained in the step G is below 4%, the mass ratio of washing solid to liquid is 1:5, the washing time is 40min, the washing temperature is above 90 ℃, and the obtained washing liquid returns to the step G to be leached continuously after the hot solid is separated.
Examples 2 to 4,
Examples 2 to 4 differ from example 1 in that in step A the calcination temperature is different. The effect of calcination temperature on lithium recovery is shown in table 1.
As can be seen from table 1, when the calcination temperature was 650 ℃, the recovery rate of lithium was high, and the continuous increase in temperature had little effect on the recovery rate of lithium. Therefore, the calcination temperature is preferably 650 ℃.
Examples 5 to 10,
Examples 5 to 10 differ from example 1 in that in step B the carbonization times are different. The effect of carbonization time on lithium recovery is shown in table 2.
As is clear from table 2, when the carbonization time is 2 hours or longer, the recovery rate of lithium is high, and the continuous extension of the carbonization time has little influence on the recovery rate of lithium. Therefore, the carbonization time is preferably 2h.
Examples 11 to 13,
Examples 11-13 differ from example 1 in that in step C, the pyrolysis time is different. The effect of pyrolysis time on recovery of lithium carbonate is shown in table 3.
As is clear from Table 3, when the carbonization time is 1 hour or more, the recovery rate of lithium carbonate is high, and the continuous extension of the pyrolysis time has little effect on the recovery rate of lithium carbonate. Therefore, the pyrolysis time is preferably 1h.
Examples 14 to 22,
Examples 14-22 differ from example 1 in that in step D the acid concentration is different. The effect of acid concentration on manganese recovery is shown in Table 4.
As can be seen from Table 4, the recovery of manganese is higher when the concentration of the acid is 100-120g/L, and the leaching of manganese is not favored by too high or too low concentration of the acid. Therefore, the concentration of the preferential leaching acid is 100-120g/L.
Examples 23 to 35,
Examples 23-35 differ from example 1 in that in step D the concentration of manganese is different at high temperature (not less than 95 ℃) cycle leaching. The effect on the crystallization of manganese sulfate during high temperature cycle leaching is shown in Table 5.
As shown in Table 5, when the concentration of manganese ions is more than 110g/L, manganese sulfate is precipitated, which is disadvantageous for the subsequent low-temperature cycle leaching to continue to increase the concentration of manganese. Therefore, the content of the high-temperature circulating leached manganese is preferably controlled to be smaller than 110g/L, and the normal-temperature (25-35 ℃) circulating leaching is carried out when the manganese content reaches 110 g/L.
Examples 36 to 48,
Examples 36-48 differ from example 1 in that in step D, the concentration of manganese is different at normal temperature (25-35 ℃) cycle leaching. The effect on crystallization of manganese sulfate during normal temperature cyclic leaching is shown in Table 6.
As shown in Table 6, when the concentration of manganese ions is more than 190g/L in the normal-temperature cycle leaching of manganese, manganese sulfate is precipitated, which is unfavorable for the subsequent normal-temperature cycle leaching to continuously increase the concentration of manganese. Therefore, the content of the nickel leached by normal-temperature circulation is controlled to be less than 190g/L, and 180g/L is preferred.
Examples 49 to 62,
Examples 49 to 62 differ from example 1 in that the amount of carbon powder added in step F was different. The effect of carbon powder addition on the leaching rates of nickel and cobalt is shown in Table 7.
As shown in Table 7, when the carbon powder was used in an amount of 8-12%, the leaching rates of nickel and cobalt reached 99%. Therefore, the preferred use ratio of the carbon powder is 8 percent.
Examples 63 to 69,
Examples 63 to 69 differ from example 1 in that in step F, the calcination temperature is different under nitrogen flow conditions. The effect of calcination temperature on nickel and cobalt leaching rates is shown in table 8.
As is clear from Table 8, the leaching rates of nickel and cobalt reached 99% at the calcination temperatures of 650-850 ℃. Therefore, the calcination temperature of the nitrogen stream is preferably 650 ℃.
Examples 70 to 76,
Examples 70-76 differ from example 1 in the solid-liquid mass ratio of the carbide slag when the acid leaching operation is performed in step D. The effect of the solid to liquid ratio on nickel and cobalt leaching rates is shown in table 9.
As shown in Table 9, when the solid-liquid ratio is 1:5-1:8, the leaching rate of manganese reaches 89%, the solid-liquid ratio is 1:5-1:6, the pH value of the leaching solution is 5-6, and the pH value of the leaching solution is 2-4, so that the leaching solid-liquid ratio is preferably 1:5-1:6 in view of the utilization rate of acid.
Examples 77 to 83,
Examples 77-83 differ from example 1 in that in step E the washing solid-liquid mass ratio is different. The effect of the wash solids to manganese content in the wash slag is shown in Table 10.
As is clear from Table 10, the content of Mn in the washing slag was about 0.2% when the ratio of the washing solid to the liquid was 1:3 to 1:8. Therefore, the preferred washing solid-liquid ratio is 1:3.
Examples 84 to 90,
Examples 84-90 example 1 differ in that in step G, the solid-liquid mass ratio of the carbide slag is different when the acid leaching operation is performed. The effect of the solid-liquid mass ratio on nickel and cobalt leaching rates is shown in table 11.
As shown in Table 11, when the solid-liquid ratio is 1:5-1:8, the leaching rate of nickel and cobalt reaches 99%, the solid-liquid ratio is 1:5-1:6, the pH value of the leaching solution is 5-6, and the pH value of the leaching solution is 2-4, considering the utilization rate of acid. Therefore, the preferred leaching solid-liquid ratio is 1:5-1:6.
Examples 77 to 83,
Examples 77-83 differ from example 1 in that in step E the washing solid-liquid mass ratio is different. The effect of the washing solid-liquid mass ratio on the manganese content in the washing slag is shown in Table 12.
As is clear from Table 12, the content of nickel in the washing slag was 0.5% and the content of cobalt was about 0.4% when the washing solid-liquid ratio was 1:4 to 1:8. Therefore, the mass ratio of washing solid to liquid is preferably 1:4.
Claims (8)
1. The full-element recovery method of the lithium nickel cobalt manganese oxide waste is characterized by comprising the following steps of:
A. Uniformly mixing nickel cobalt lithium manganate waste with carbon powder, and calcining for 2-3 hours at 600-650 ℃ under the condition of nitrogen flow;
B. Adding water into the mixture obtained after calcining in the step A to prepare slurry, then introducing carbon dioxide to carbonize for 2-3h, and carrying out solid-liquid separation after carbonization to obtain carbonized slag and carbonized liquid;
C. Heating the carbonized liquid obtained in the step B to be more than or equal to 90 ℃ for pyrolysis, and reacting for 1-2h to obtain lithium carbonate and pyrolyzed liquid;
D. Adding dilute sulfuric acid with the acid concentration of 100-120g/L into the carbonized slag obtained in the step B for acid leaching, wherein the reaction temperature is not lower than 95 ℃, the reaction time is 40-90min, solid-liquid separation is carried out to obtain a mixed solution I of lithium sulfate and manganese sulfate and acid leaching slag I, adding concentrated sulfuric acid into the mixed solution I to adjust the acid concentration to 100-120g/L, adding the carbonized slag, continuing the acid leaching operation according to the conditions, carrying out cyclic leaching in such a way that the manganese concentration reaches 110g/L, carrying out cyclic leaching at the temperature of 25-35 ℃ to enable the manganese content in the leaching solution to reach 180-190g/L, then carrying out solid-liquid separation at normal temperature to obtain a mixed solution II and acid leaching slag II, heating the mixed solution II to be higher than or equal to 95 ℃, separating the separated solution into manganese sulfate crystals and separated solution, keeping the temperature of the separated solution to be not lower than 90 ℃, and continuously carrying out acid leaching operation when the lithium content in the solution reaches 15-20g/L, adding sodium hydroxide into the solution to adjust the pH to 11-12, and precipitating lithium carbonate solution;
E. Washing the acid leaching slag I and the acid leaching slag II in the step D by pure water at normal temperature, then carrying out solid-liquid separation to obtain washing slag and washing liquid, and returning the washing liquid to the step D for continuous leaching;
F. Supplementing carbon powder into the washing slag obtained in the step E, calcining for 2-3 hours at 650-750 ℃ under the condition of nitrogen flow, and then continuously calcining for 2-3 hours at 650-700 ℃ under the condition of air flow to obtain a calcined material;
G. Adding dilute sulfuric acid with the acid concentration of 180-200g/L into the calcined material obtained in the step F for acid leaching, wherein the reaction temperature is not less than 95 ℃, the reaction time is 40-90min, after solid-liquid separation, adding concentrated sulfuric acid into the obtained leaching solution to adjust the acid concentration to 180-200g/L, adding the calcined material, continuing the acid leaching operation according to the conditions, circularly leaching in such a way that the total concentration of nickel and cobalt in the leaching solution reaches 360-400g/L, then keeping the temperature not less than 90 ℃ and performing hot solid-liquid separation to obtain a mixed solution III and leached slag III, and cooling the mixed solution III to normal temperature to separate out cobalt sulfate and nickel sulfate crystals;
H. And (C) washing the leaching residue III obtained in the step (G) by using pure water, wherein the washing temperature is more than 90 ℃, separating the hot liquid, and returning the obtained washing liquid to the step (G) for leaching.
2. The method for recycling all elements of lithium nickel cobalt manganese oxide waste according to claim 1, wherein in the step A, the mass ratio of carbon powder is 10% -15% of that of the mixed powder.
3. The method for recycling all elements of lithium nickel cobalt manganate waste according to claim 1 or 2, wherein in the step B, when preparing slurry, the ratio of calcined material to water is prepared according to the content of lithium in the carbonized liquid of 7.5-8 g/L.
4. The method for recycling all elements of lithium nickel cobalt manganese oxide waste material, according to claim 3, wherein in the step D, the solid-liquid mass ratio of the carbide slag is 1:5-6 when the acid leaching operation is carried out.
5. The method for recycling all elements from lithium nickel cobalt manganese oxide waste material according to claim 4, wherein in the step E, the mass ratio of washing solid to liquid is 1:3-8, and the washing time is 30-50min.
6. The method for recycling all elements of lithium nickel cobalt manganese oxide waste material according to claim 5, wherein in the step F, the carbon powder is supplemented with 9% -12% by mass.
7. The method for recycling all elements of lithium nickel cobalt manganese oxide waste material according to claim 6, wherein in the step G, the solid-liquid mass ratio of the calcined material is 1:5-6 when the acid leaching operation is performed.
8. The method for recycling all elements from lithium nickel cobalt manganese oxide waste according to claim 7, wherein in the step H, the mass ratio of washing solid to liquid is 1:4-8, and the washing time is 30-50min.
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| CN107017443A (en) * | 2017-03-28 | 2017-08-04 | 北京科技大学 | A kind of method of the comprehensively recovering valuable metal from waste and old lithium ion battery |
| CN109082522A (en) * | 2018-08-15 | 2018-12-25 | 合肥国轩高科动力能源有限公司 | Method for recovering anode powder of waste ternary lithium battery |
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| CN112111651B (en) * | 2020-09-21 | 2022-05-20 | 天齐锂业（江苏）有限公司 | Pyrogenic process for recovering waste lithium ion battery powder |
| KR102562997B1 (en) * | 2021-01-12 | 2023-08-02 | 한국해양대학교 산학협력단 | Method for recovering useful metals by water leaching of sulfate roasted products of waste lithium secondary batteries |
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| CN109082522A (en) * | 2018-08-15 | 2018-12-25 | 合肥国轩高科动力能源有限公司 | Method for recovering anode powder of waste ternary lithium battery |
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