A method for recovering nickel, cobalt and manganese from waste ternary battery powder

Abstract

This invention belongs to the field of waste recycling technology, specifically involving a method for recovering nickel, cobalt, and manganese from waste ternary battery powder, including the following steps: (1) Pre-lithiation: Waste ternary battery powder is subjected to reduction roasting under oxygen-isolated conditions, water is added, hydrogenation is performed to remove impurities, and solid-liquid separation is carried out to obtain hydrogenated lithium liquid and nickel-cobalt-manganese slag. The hydrogenated lithium liquid is thermally decomposed to recover lithium carbonate; (2) Recovery of nickel, cobalt, and manganese: S1, Leaching: Crude nickel-cobalt hydroxide is mixed into the nickel-cobalt-manganese slag, water is added, concentrated sulfuric acid is added, acid leaching is performed, solid-liquid separation is carried out to obtain acid solution and acid leaching residue. The acid leaching residue is returned to step S1 for use; S2, Impurity removal: Neutralizing agent is added to the acid solution, solid-liquid separation is carried out to obtain impurity-removed residue and impurity-removed liquid; S3, Extraction with P204 and P507; S4, Manganese and cobalt recovery: Dilute sulfuric acid is added to the loaded phase to recover manganese and cobalt. The method of this invention can recover nickel, cobalt, and manganese from waste battery powder, and no bubbles are generated during the acid leaching process.

Description

Technical Field

[0001] This invention belongs to the field of waste recycling technology, specifically relating to a method for recovering nickel, cobalt, and manganese from waste ternary battery powder. Background Technology

[0002] In recent years, numerous researchers have conducted extensive work on the recovery of nickel, cobalt, and manganese from retired ternary lithium-ion battery materials. The most common method for recovering nickel, cobalt, and manganese involves wet leaching to directly extract all valuable elements such as lithium, nickel, cobalt, and manganese from the battery powder. This process requires the addition of a reducing agent, followed by the separation of impurities from the leachate to prepare a nickel-cobalt-manganese salt solution. However, this traditional wet process suffers from complex purification and separation procedures, difficulty in lithium recovery, and low recovery rates. Alternatively, a combination of pyrometallurgical and wet methods can be used to recover nickel, cobalt, and manganese. A more widely studied approach involves first roasting or reducing roasting the battery powder, followed by wet extraction of nickel, cobalt, and manganese. When using the reducing roasting method, the battery powder is typically mixed with a reducing agent. The high-valence nickel, cobalt, and manganese in the battery powder are reduced to a mixture of low-valence elemental metals and low-valence oxides, which are then recovered through wet leaching.

[0003] Chinese invention patent CN112374511B discloses a method for preparing lithium carbonate and ternary precursors from recycled waste ternary lithium batteries. The method utilizes graphite and binders inherent in the black powder obtained after pretreatment of waste ternary batteries as reducing agents, destroying the structure of the ternary material through self-reduction phase transformation. A carbonation-water leaching method is used to convert lithium carbonate into easily leached lithium bicarbonate, achieving preferential lithium leaching. The lithium-extracted material does not require the addition of a reducing agent and is directly leached with sulfuric acid to obtain a nickel-cobalt-manganese solution. This solution is then purified by precipitation and extraction to obtain a nickel-cobalt-manganese sulfate solution. After adjusting the metal ratio, a high-quality ternary cathode material precursor can be prepared. After reduction roasting, nickel, cobalt, and manganese are converted into metallic or low-valence oxide forms, reducing the difficulty of acid leaching. Under milder leaching conditions without the need for a reducing agent, a leaching rate of over 98% for nickel, cobalt, and manganese can be achieved. Reduction roasting can decompose or volatilize some of the organic matter in battery powder, preventing it from generating a large amount of foam during acid leaching and thus affecting production efficiency. However, extensive experiments and practices by the inventors have shown that while prioritizing lithium extraction through reduction roasting of battery powder can largely eliminate the difficulty in acid leaching caused by organic matter in the battery powder, it cannot eliminate the problem of poor acid leaching caused by the reaction of nickel and cobalt with sulfuric acid, which generates a large amount of bubbles. It is precisely because nickel and cobalt are converted into metallic elements during reduction roasting that a thick layer of viscous foam is generated at the top of the acid leaching tank, constantly overflowing. This is because when sulfuric acid is added directly for acid leaching, the reaction of nickel and cobalt with sulfuric acid generates a large amount of sulfur dioxide and hydrogen gas that is difficult to dissipate, thus hindering the normal progress of acid leaching and reaction. The rate of sulfuric acid addition during acid leaching is difficult to control, resulting in poor process stability and low production efficiency.

[0004] Chinese invention patent CN106129511B discloses a method for comprehensively recovering valuable metals from waste lithium-ion battery materials. The method involves mixing waste lithium-ion battery cathode materials with a reducing agent for reduction roasting, followed by lithium extraction through water leaching with CO2. The water leaching residue is then leached with oxidative acid leaching or ammonia leaching to extract valuable elements such as cobalt, nickel, and manganese. After extraction and purification, corresponding compound products are obtained. Although a reduction roasting method is used, the acid leaching process still requires the separate use of an oxidant or an ammonia-containing reagent. The leaching agents for oxidative acid leaching can be H2SO4, HCl, or HNO3, while the oxidants can be air, oxygen, or hydrogen peroxide. However, no solution is proposed to address the difficulty of adding acid during the acid leaching process.

[0005] Therefore, it is essential to research and develop a method that can solve the problems of adding an oxidant separately after reduction roasting and lithium extraction and adding acid to the nickel-cobalt-manganese acid dissolution process during the recovery of nickel-cobalt-manganese from waste ternary battery powder, as well as the problem of adding acid and foaming during the process. Summary of the Invention

[0006] The present invention aims to solve the above-mentioned technical problems and provides a method for recovering nickel, cobalt and manganese from waste ternary battery powder. This method can recover nickel, cobalt and manganese from waste ternary battery powder. After reduction roasting and lithium extraction, no additional oxidant needs to be added. No bubbles or polluting gases are generated during the acid leaching process.

[0007] The technical solution of this invention is as follows:

[0008] A method for recovering nickel, cobalt, and manganese from waste ternary lithium battery powder includes the following steps:

[0009] (1) Pre-extraction of lithium: waste ternary battery powder is reduced and roasted under oxygen-isolated conditions to obtain roasted battery powder. Water is added to the roasted battery powder to make slurry, and hydrogenation is carried out to remove impurities. Solid-liquid separation is performed to obtain hydrogenated lithium liquid and nickel-cobalt-manganese slag. The hydrogenated lithium liquid is thermally decomposed to recover lithium carbonate.

[0010] (2) Recovery of nickel, cobalt, and manganese:

[0011] S1, Leaching: Crude nickel-cobalt hydroxide is mixed into the nickel-cobalt-manganese slag obtained in step (1), water is added to make a slurry at a liquid-solid mass ratio of 3.0-6.5:1, concentrated sulfuric acid is added, acid leaching is performed, solid and liquid are separated, and acid solution and acid leaching residue are obtained. The acid leaching residue is returned to step S1 for use.

[0012] S2. Impurity removal: Add a neutralizing agent to the acid solution, and separate the solid and liquid to obtain the impurity-removed residue and the impurity-removed liquid;

[0013] S3. Extraction: After filtering the impurity-removed liquid, add P204 to extract impurities and separate the phases to obtain P204 raffinate and loaded organic phase I; add P507 to P204 raffinate to extract cobalt and magnesium and separate the phases to obtain purified nickel sulfate solution and loaded organic phase IV.

[0014] S4. Recovery of manganese and cobalt: Add dilute sulfuric acid to the supported organic phase I, back-extract and separate the phases to obtain a manganese sulfate solution. Add dilute sulfuric acid to the supported organic phase IV, back-extract and separate the phases to obtain a refined cobalt sulfate solution.

[0015] In S1 of this invention, there are no strict requirements for the concentration of concentrated sulfuric acid; a mass concentration of 50-98% is acceptable.

[0016] The main valuable metal elements in the waste NCM ternary battery powder of this invention are nickel, cobalt, manganese, and lithium. During the battery dismantling process, due to the presence of fluorine-containing organic matter and positive and negative electrodes, the final battery powder often contains small amounts of impurities such as copper, aluminum, and fluorine. Among them, copper and aluminum mainly exist in elemental form. During reduction roasting, copper, aluminum, and carbon powder participate in the redox reaction as reducing agents. The main reactions during reduction roasting are as follows:

[0017] 2LiCoO2+2A1=A12O3+2Co+Li2O (1)

[0018] 2LiCoO2+3Cu=3CuO+2Co+Li2O (2)

[0019] 2LiNiO2+2A1=A12O3+2Ni+Li2O (3)

[0020] 2LiNiO2+3Cu=3CuO+2Ni+Li2O (4)

[0021] 2LiMn2O4+2A1=A12O3+4MnO+Li2O (5)

[0022] 2LiMn2O4+3Cu=3CuO+4MnO+Li2O (6)

[0023] 4LiNiO2+3C=2Li2CO3+4Ni+CO2 (7)

[0024] 4LiCoO2+3C=2Li2CO3+4Co+CO2 (8)

[0025] 4LiMn2O4+7C=2Li2CO3+8Mn+5CO2 (9)

[0026] The amount of reducing agent should not be too much or too little. Too little will lead to incomplete reduction, which will further reduce the lithium leaching rate. When there is too much reducing agent, the utilization rate of the reducing agent is low. Excessive reducing agent will remain in the calcined battery powder, which is not conducive to the subsequent extraction of nickel, cobalt and manganese. Therefore, the present invention preferably adds 5-20% of the weight of waste ternary battery powder as carbon powder for reduction calcination.

[0027] Preferably, the reduction roasting conditions of this invention are: temperature 700-1200℃ and time 1-3h. During the reduction roasting process, if the reduction roasting temperature is below 700℃, the reduction is incomplete, further leading to a decrease in lithium leaching rate. Nickel, cobalt, and manganese remain in the form of high-valence oxides, and lithium is not completely converted into lithium oxide or lithium carbonate, which is detrimental to improving the recovery rate of nickel, cobalt, manganese, and lithium. If the reduction temperature is too high, energy consumption is too high, which is not conducive to cost savings. If the reduction roasting time is too short, the reduction is incomplete, which is not conducive to improving the subsequent recovery rate of nickel, cobalt, manganese, and lithium. If the reduction time is too long, it is not conducive to energy savings. Under the same conditions, the reduction effect of 1-3h is similar. To save energy and comprehensively consider the reduction effect, the reduction roasting time should be controlled within 1-3h.

[0028] After reduction roasting, in addition to the formation of lithium carbonate and lithium oxide, nickel, cobalt, and manganese in the battery powder are reduced to elemental nickel, cobalt, and manganese, respectively, as well as some manganese oxide. Simultaneously, copper and aluminum, which participate in the redox reaction, are converted into their corresponding oxides. During reduction roasting, because the battery powder itself contains reducing substances, carbon powder is added as a reducing agent. After roasting, there may be an excess of reducing agent, leaving a small amount of unreacted carbon powder in the battery powder. To improve reduction efficiency, oxygen should be isolated during roasting, such as using a vacuum, carbon dioxide, or nitrogen atmosphere. Reduction roasting not only increases the reduction rate of nickel, cobalt, and manganese in the battery powder but also converts lithium into lithium carbonate and lithium oxide, which is more conducive to the subsequent extraction of nickel, cobalt, manganese, and lithium.

[0029] If the liquid-to-solid ratio is too low during hydrogenation and impurity removal, lithium will become supersaturated, resulting in incomplete dissolution of lithium carbonate and reduced lithium recovery rate. If the liquid-to-solid ratio is too high, the lithium concentration in the hydrogenated lithium solution will be too low, increasing water evaporation during thermal decomposition, which is not conducive to reducing energy consumption and improving production efficiency. Therefore, preferably, in step (1) of this invention, water is added to the slurry at a lithium concentration of 5-10 g / L during hydrogenation and impurity removal. The hydrogenation time can be determined according to the pH of the hydrogenation process. Before the introduction of carbon dioxide, the slurry is alkaline with a pH of 11-13. After the introduction of carbon dioxide, the hydrogenation reaction begins, and the alkalinity of the slurry decreases until the pH no longer changes, at which point the hydrogenation reaction ends. Generally, the hydrogenation process can be completed in more than 1 hour. To improve production efficiency, the hydrogenation time should be controlled within 3 hours. If the amount of carbon dioxide introduced during hydrogenation is too small, hydrogenation will be incomplete; if too much, reagents will be wasted. The preferred excess carbon dioxide coefficient is 1.5-3.0.

[0030] During hydrogenation for impurity removal, lithium carbonate and lithium oxide are soluble in water or react chemically with water, while other oxides and metallic elements are insoluble in water. This allows for the initial separation of lithium from other impurities. Lithium oxide reacts with water to form lithium hydroxide. The introduced carbon dioxide converts lithium carbonate and lithium hydroxide (the latter is first converted to lithium carbonate, and then to lithium bicarbonate under the action of excess carbon dioxide) into lithium bicarbonate, which has higher solubility. After hydrogenation, lithium in the lithium solution exists as lithium bicarbonate.

[0031] Li₂O + H₂O = 2LiOH (10)

[0032] 2LiOH + CO2 = Li2CO3 + H2O (11)

[0033] Li2CO3+H2O+CO2=2LiHCO3 (12)

[0034] Because the hydrogenation process takes place in a neutral to slightly alkaline environment (pH 7-9), even if the battery powder contains small amounts of impurities such as silicon, aluminum, copper, nickel, cobalt, manganese, and zinc, these impurities will hydrolyze and precipitate during hydrogenation. The resulting hydrogenated lithium solution after solid-liquid separation will not contain these impurities. During thermal decomposition, lithium bicarbonate decomposes into lithium carbonate. The mother liquor obtained during solid-liquid separation mainly consists of lithium bicarbonate. The thermal decomposition conditions are: temperature 85-95℃ and time 1-3 hours. The following reaction occurs during thermal decomposition:

[0035] 2LiHCO3=Li2CO3+H2O+CO2↑(13)

[0036] Since the battery powder contains elemental nickel, cobalt, and manganese after reduction roasting, and may also contain unreacted carbon powder, the resulting nickel-cobalt-manganese slag is obtained after hydrogenation to remove impurities. This process, apart from dissolving lithium, does not cause any changes in the chemical properties of the nickel, cobalt, and manganese. Therefore, when only an inorganic acid, such as concentrated sulfuric acid, is added to the nickel-cobalt-manganese slag without any other reagents, the following main reactions occur:

[0037] M + 2H₂SO₄(conc.) = MSO₄ + SO₂↑ + 2H₂O (14)

[0038] M+H2SO4(dilute)=MSO4+H2↑ (15)

[0039] (M = Ni, Co, Mn)

[0040] C + H₂SO₄(conc.) = CO₂↑ + SO₂↑ + 2H₂O (16)

[0041] MnO + H₂SO₄ = MnSO₄ + H₂O (17)

[0042] A12O3+3H2SO4=A12(SO4)3+3H2O (18)

[0043] CuO + H₂SO₄ = CuSO₄ + H₂O (19)

[0044] In other words, when only sulfuric acid is added to nickel-cobalt-manganese slag, although the nickel-cobalt-manganese metals and oxides can be converted into ionic forms and enter the acid solution, the slag contains a large amount of elemental nickel-cobalt-manganese under the action of concentrated sulfuric acid. Therefore, reactions will occur where the elemental substances react with concentrated sulfuric acid to generate a large amount of sulfur dioxide, and react with dilute sulfuric acid to generate a large amount of hydrogen gas. At the same time, if there is unreacted carbon powder in the battery powder, the carbon powder can also undergo redox reactions with concentrated sulfuric acid to generate carbon dioxide and sulfur dioxide gases. Therefore, in actual production practice, if inorganic acid is added alone to dissolve nickel-cobalt-manganese slag, a large number of bubbles are often generated during the acid dissolution process, which prevents the reaction from proceeding normally. Because overflowing is likely to occur and the large number of bubbles generated are difficult to eliminate, the acid addition time will be greatly extended, and the process will be unstable. To address the aforementioned problems, this invention employs a combined leaching method using crude nickel-cobalt hydroxide and nickel-cobalt-manganese slag. Preferably, in step S1, the amount of crude nickel-cobalt hydroxide added is 15-50 times the weight of the nickel-cobalt-manganese slag. If the proportion of crude nickel-cobalt hydroxide is too low, the low manganese dioxide content weakens its inhibitory effect on the displacement reaction that generates a large amount of gas from the reaction between nickel-cobalt and sulfuric acid during the redox reaction, thus failing to effectively improve the overflow phenomenon during acid leaching. If the proportion of crude nickel-cobalt hydroxide is too high, on the one hand, the excessive manganese dioxide results in a large amount of manganese dioxide slag remaining after acid leaching, reducing the overall manganese leaching rate. On the other hand, with an excessive proportion of crude nickel-cobalt hydroxide, the final acid-leached slag is mainly composed of manganese dioxide. The higher the manganese dioxide content, the worse the filtration performance of the acid-leached slurry. To balance the manganese leaching rate, solve the difficulties in the acid leaching process, and ensure good filtration performance of the acid-leached slurry, the amount of crude nickel-cobalt hydroxide added is 15-50 times the weight of the nickel-cobalt-manganese slag.

[0045] Preferably, the crude nickel-cobalt hydroxide of the present invention contains 2-10 wt.% manganese. If the manganese dioxide content is too low, the amount of crude nickel-cobalt hydroxide used will be too large, and it will not play a role in oxidizing and inhibiting the generation of hydrogen and sulfur dioxide gases. The manganese content in common crude nickel-cobalt hydroxide is 2-10%, of which the +4 valence manganese mainly exists in the form of manganese dioxide. When only sulfuric acid is added and no other leaching agent is added, it does not leach out with the acid dissolution process, but exists in the form of manganese dioxide slag. The proportion of +4 valence manganese in crude nickel-cobalt hydroxide with a manganese content of 2-10% is generally 50-80%, which can be used in the present invention for combined acid dissolution leaching with nickel-cobalt-manganese slag to achieve the effect of inhibiting overflow during the acid dissolution leaching process.

[0046] This invention utilizes the strong oxidizing properties of manganese dioxide in crude nickel-cobalt hydroxide to suppress the aforementioned reaction that generates a large amount of gas. It can simultaneously recover valuable metals from both the crude nickel-cobalt hydroxide and the nickel-cobalt-manganese slag. Both crude nickel-cobalt hydroxide and the nickel-cobalt-manganese slag contain valuable metals such as nickel, cobalt, and manganese; therefore, recovering these valuable metals together does not conflict with the process and does not increase production complexity. In the presence of sulfuric acid, the nickel-cobalt-manganese hydroxide in the crude nickel-cobalt hydroxide undergoes a neutralization reaction with the sulfuric acid. The high-valence manganese present in the hydroxide exists as manganese dioxide. Without the addition of a reducing agent, the manganese dioxide will not dissolve and convert to Mn along with the hydroxide. 2+ Experiments and production practice have proven that manganese dioxide has strong oxidizing properties. It preferentially reacts with the elemental metals in nickel-cobalt-manganese slag under acidic conditions to undergo redox reactions, thereby inhibiting the reaction between nickel-cobalt and sulfuric acid to generate a large amount of gas. The reaction in the acid leaching system is transformed into a reaction between nickel-cobalt and manganese dioxide and sulfuric acid, while the reaction between nickel-cobalt and concentrated or dilute sulfuric acid to generate a large amount of gas is weakened. The leaching system is dominated by the following reactions:

[0047] M(OH)2+2H2SO4=MSO4+2H2O (20)

[0048] N+2H2SO4+MnO2=NSO4+MnSO4+2H2O (21)

[0049] (M = Ni, Co, Mn, N = Ni, Co)

[0050] Other reactions that occur simultaneously are as follows:

[0051] C + H₂SO₄(conc.) = CO₂↑ + SO₂↑ + 2H₂O (22)

[0052] MnO + H₂SO₄ = MnSO₄ + H₂O (23)

[0053] A12O3+3H2SO4=A12(SO4)3+3H2O (24)

[0054] CuO + H₂SO₄ = CuSO₄ + H₂O (25)

[0055] Therefore, by mixing crude nickel-cobalt hydroxide with nickel-cobalt-manganese slag and then combining it with sulfuric acid leaching, the acid leaching process no longer generates a large number of bubbles. The only way bubbles are generated is through the reaction of a small amount of excess carbon powder with concentrated sulfuric acid to produce sulfur dioxide and carbon dioxide. This essentially eliminates or reduces the overflow phenomenon of nickel-cobalt-manganese slag during the acid leaching process, making the acid leaching process more stable, easier to operate, and improving the working environment. Preferably, the endpoint pH of the acid leaching process in this invention is 1.0-2.5. Too low a pH during acid leaching will lead to the consumption of a large amount of neutralizing agent in subsequent impurity removal, resulting in a large amount of impurity residue. Too high a pH will lead to incomplete leaching of valuable metals nickel, cobalt, and manganese, thus reducing the recovery rate of nickel, cobalt, and manganese. Under the condition of pH=1.0-2.5, not only can nickel, cobalt, and manganese be completely leached, but it also allows manganese dioxide to fully exert its oxidizing effect, thereby inhibiting the generation of hydrogen and sulfur dioxide.

[0056] The main component of the acid-dissolving slag in step S1 of this invention is the remaining manganese dioxide, which can be used as an oxidant to replace the crude nickel-cobalt hydroxide and be acid-dissolved together with the nickel-cobalt-manganese slag. It can also suppress the generation of hydrogen and sulfur dioxide, so as to eliminate or reduce the overflow phenomenon of nickel-cobalt-manganese slag during the acid dissolution process.

[0057] In step S2 of this invention, the neutralizing agent can hydrolyze and precipitate the main impurities such as iron, aluminum, silicon, copper, chromium, and silicon in the acid solution, and remove them through solid-liquid separation. The main reactions that occur during the hydrolysis process are as follows:

[0058] Fe 3+ +3H₂O=Fe(OH)₃(s)+3H + (26)

[0059] Fe 2+ +2H₂O=Fe(OH)₂(s)+2H + (27)

[0060] 4Fe(OH)2+O2+2H2O=4Fe(OH)3(28)

[0061] Al 3+ +3H2O - =Al(OH)3(s)+3H + (29)

[0062] Cu 2+ +2H₂O=Cu(OH)₂(s)+2H + (30)

[0063] Cr 3+ +3H₂O=Cr(OH)₃(s)+3H + (31)

[0064] SiO3 2- +H₂O=H₂SiO₃(s)+3H+ (32)

[0065] Under the action of the neutralizing agent, the sulfuric acid in the acid solution is neutralized, the pH of the system increases, and the impurity ions in the solution are hydrolyzed to form hydroxide colloids or precipitates. Hydrogen ions are also generated when precipitates or colloids are formed. However, with the addition of the neutralizing agent, the neutralizing agent can also consume the hydrogen ions generated by hydrolysis, thereby promoting the forward hydrolysis reaction. Therefore, the above impurities can be removed by solid-liquid separation. The generated ferric hydroxide colloid has an adsorption effect and can also adsorb and remove silicon in the solution. The neutralizing agent of this invention is preferably limestone or zirconium hydroxide. When limestone is used as the neutralizing agent, while impurity ions are hydrolyzed, calcium in calcium carbonate undergoes a metathesis reaction with sulfate ions in the acid solution, partially converting into calcium sulfate precipitate. If the pH is too low, the impurity ions will not be completely hydrolyzed, increasing the burden on subsequent extraction. During extraction, emulsification is easily generated, reducing the extraction effect and increasing the difficulty of phase separation. If the pH is too high, nickel, cobalt, and manganese in the acid solution will also be partially hydrolyzed to form hydroxide precipitates, and the amount of neutralizing agent used will increase dramatically. In order to balance the impurity removal effect, reduce the amount of neutralizing agent used, and improve the recovery rate of nickel, cobalt, and manganese, the impurity removal pH should be controlled between 5.0 and 6.0. When zirconium hydroxide is used as a neutralizing agent, similarly, if the pH is too low, the removal of aluminum, chromium, and silicon is incomplete. When the pH is above 6.0, a larger amount of zirconium hydroxide is required, and the impurity removal effect is not significantly improved. Furthermore, due to the introduction of excessive zirconium hydroxide, some valuable metal ions will hydrolyze and precipitate into the impurity removal residue, which is detrimental to improving the recovery rate of nickel, cobalt, and manganese. At pH = 5.0-6.0, the impurity removal effect is similar, achieving deep impurity removal. Therefore, the endpoint pH for impurity removal should be controlled between 5.0 and 6.0. There are no strict requirements for the impurity removal temperature; room temperature or heating can achieve similar impurity removal effects. Within the temperature range of 20-95℃, under the same conditions, the impurity removal effect is similar, and the impurity removal time should be more than 2 hours. To improve the impurity removal efficiency, the impurity removal time should be controlled within 4 hours.

[0066] Since manganese dioxide is an extremely fine particle, if it does not react completely, it can easily penetrate the filter cloth during the solid-liquid separation step. Therefore, the liquid after impurity removal needs to be precisely filtered to prevent the generation of suspended solids and emulsions during the extraction and impurity removal process. The manganese slag obtained by precise filtration can replace a portion of the crude nickel-cobalt hydroxide. The manganese slag is acid-dissolved together with the nickel-cobalt-manganese slag, which can also inhibit the generation of hydrogen and sulfur dioxide, thereby eliminating or reducing the overflow phenomenon of nickel-cobalt-manganese slag during the acid dissolution process.

[0067] After impurity removal, most impurities have been removed, but small amounts of calcium, zinc, and other impurities remain in the purified solution. In step S3, P204 extraction removes these impurities, allowing calcium and zinc to be extracted into the P204 organic phase, thus separating them from nickel and cobalt. While extracting calcium and zinc, manganese in the solution is also extracted by P204. Therefore, to recover the manganese, the organic-loaded P204 needs to be back-extracted to obtain a manganese sulfate solution, which can then be further recovered. The P204 raffinate after P204 extraction mainly consists of a nickel-cobalt sulfate solution, and also contains a small amount of magnesium impurities. To remove magnesium, the P204 raffinate can be extracted using P507. However, during magnesium extraction, cobalt in the solution is also extracted. Therefore, to recover the cobalt, the organic-loaded P507 needs to be back-extracted to obtain a refined cobalt sulfate solution, which can then be further recovered. The P507 raffinate is a refined nickel sulfate solution. Refined nickel sulfate and cobalt sulfate solutions can be used to produce battery-grade nickel sulfate and cobalt sulfate crystals. The main impurities in manganese sulfate solution are calcium and a small amount of zinc. After impurity removal, it can be used to produce battery-grade manganese sulfate crystals.

[0068] In step S3 of this invention, there are no strict requirements on the volume concentration of the extractant. When the extractant concentration is too high, the viscosity is too large, which is not conducive to phase separation after extraction. When the concentration is too low, it is not conducive to improving the extraction efficiency. It is advisable to control the volume concentration of P204 at 20-25% and the volume concentration of P507 at 10-15%. A single-stage extraction time of more than 10 minutes can achieve the effect of extracting the corresponding impurities. In order to improve the working efficiency, the extraction time is controlled at 10-20 minutes. There are no strict requirements on the extraction temperature. When the temperature is too low, the phase separation after extraction is slow. In order to save energy and achieve rapid phase separation, the extraction temperature is advisable to control it at 20-50℃.

[0069] To recover valuable metals such as manganese and cobalt, and to reuse the P204 and P507 extractants loaded with impurities, it is necessary to back-extract the loaded impurity ions to achieve extractant regeneration and recycling. The main impurities in the P204 organic phase are manganese, zinc, and calcium, which can be back-extracted using dilute sulfuric acid. To improve nickel recovery, before back-extraction to remove impurities, the loaded organic phase can be washed with pure water to remove nickel entrained in the loaded organic phase. The resulting nickel washing solution can be reused as the P204 extraction stock solution, and then dilute sulfuric acid is added to it for back-extraction of the loaded impurities. Specifically:

[0070] (1) First wash nickel with pure water in the loaded organic phase I, then back-extract manganese with 1.5-2.5N dilute sulfuric acid, controlling the pH of the aqueous phase to 2.0-2.5, to obtain manganese sulfate solution and loaded organic phase II. Then add 1.5-2.5N dilute sulfuric acid to loaded organic phase II to back-extract zinc, controlling the pH of the aqueous phase to 1.0-1.8, to obtain zinc sulfate solution and loaded organic phase III. Loaded organic phase III is the regenerated P2O4, which can be returned to the extraction for impurity removal. In the above-mentioned back-extraction process, if the pH of the first-stage back-extraction is higher than 2.5, the manganese back-extraction will be incomplete, causing the remaining manganese in the loaded organic phase II to enter the zinc sulfate solution, resulting in incomplete separation of manganese and zinc. If the pH of the first-stage back-extraction is lower than 2.0, the excessive acidity will cause zinc to be partially back-extracted first, resulting in a large amount of zinc entering the manganese sulfate solution, again resulting in incomplete separation of manganese and zinc. When back-extracting zinc, if the pH is higher than 1.8, the zinc back-extraction will be incomplete, reducing the extraction effect when the extractant is reused. If the pH is lower than 1.0, the resulting zinc sulfate solution will have a high residual acid content, which is not conducive to cost savings when processing the zinc sulfate solution. There are no strict requirements on the concentration of dilute sulfuric acid used in the above-mentioned back-extraction process, as long as it is conducive to operation (facilitating pH control in each stage of back-extraction).

[0071] (2) First, wash the nickel in the supported organic phase IV with pure water, then back-extract magnesium with 0.5-1.5N dilute sulfuric acid, controlling the pH of the aqueous phase to 4.0-4.5, to obtain a magnesium sulfate solution and a supported organic phase V. Then, add 1.5-2.5N dilute sulfuric acid to the supported organic phase V for a first back-extraction of cobalt, controlling the pH of the aqueous phase to 2.5-3.5, to obtain a refined cobalt sulfate solution and a supported organic phase VI. Then, add 1.5-2.5N dilute sulfuric acid to the supported organic phase VI for a second back-extraction of cobalt, controlling the pH of the aqueous phase to 1.0-1.5, to obtain the second cobalt back-extraction solution and an organic phase. organic phase This is regenerated P507, which can be reused for cobalt and magnesium extraction. During the aforementioned back-extraction process, if the pH is higher than 4.5 during magnesium back-extraction, the magnesium back-extraction will be incomplete, resulting in a large amount of residual magnesium from the loaded organic phase V entering the refined cobalt sulfate solution, leading to incomplete separation of magnesium and cobalt. If the pH is lower than 4.0 during magnesium back-extraction, excessive acidity will cause a small amount of cobalt to be back-extracted first, with a large amount of cobalt entering the magnesium sulfate solution, again resulting in incomplete separation of magnesium and cobalt. There are no strict requirements on the concentration of dilute sulfuric acid used for cobalt back-extraction, as long as it facilitates operation (making it easy to control the pH at each stage during back-extraction). If the pH is higher than 3.5 during one cobalt back-extraction process, the cobalt back-extraction rate will be low; if the pH is lower than 2.5, the residual acid in the refined cobalt sulfate solution will be high, which is detrimental to the subsequent processing of the refined cobalt sulfate solution. During the secondary cobalt back-extraction, if the pH is higher than 1.5, the cobalt and other impurities loaded in the organic material will not be completely back-extracted, reducing the extraction effect when the extractant is reused. If the pH is lower than 1.0, the residual acid in the liquid after the secondary cobalt back-extraction will be high, increasing the cost of subsequent processing of the secondary cobalt back-extraction liquid.

[0072] Due to the adoption of the above technical solution, the beneficial effects of the present invention are as follows:

[0073] 1. The method of the present invention can recover nickel sulfate, cobalt sulfate and manganese sulfate solutions from waste ternary battery powder, and no bubbles or polluting gases are generated during the recovery process. No oxidant needs to be added separately during reduction roasting.

[0074] 2. In the method of this invention, the nickel-cobalt-manganese slag after lithium extraction is leached with crude nickel-cobalt hydroxide using a combined acid leaching method. The manganese dioxide in the crude nickel-cobalt hydroxide raw material is used to suppress the reaction between the nickel-cobalt elements in the nickel-cobalt-manganese slag and concentrated or dilute sulfuric acid to generate a large amount of sulfur dioxide and hydrogen gas. This achieves the effect of preventing the nickel-cobalt-manganese slag from overflowing during acid leaching. When the two are leached together, there are basically no bubbles generated. Since the generation of a large amount of gas is basically eliminated, the working environment is greatly improved, the sulfuric acid addition rate is easy to control, and the production efficiency is improved.

[0075] 3. In the method of the present invention, the extractants P204 and P507 can be recycled and reused, reducing extraction costs and further achieving the effect of energy saving and environmental protection. Attached Figure Description

[0076] Figure 1 This is a flowchart of the method for recovering nickel, cobalt, and manganese from waste ternary battery powder according to the present invention. Detailed Implementation

[0077] The technical solutions in the embodiments of the present invention will be clearly and completely described below. Obviously, the described embodiments are only some embodiments of the present invention, and not all embodiments. Based on the embodiments of the present invention, all other embodiments obtained by those skilled in the art without creative effort are within the scope of protection of the present invention.

[0078] Example 1

[0079] (1) Prioritize lithium extraction

[0080] Take 500g of No. 1 battery powder. The main components of the battery powder are as follows: nickel 36.44%, cobalt 13.15%, manganese 10.64%, lithium 3.43%, fluorine 0.78%, copper 0.72%, and aluminum 0.99%. The main components of the battery powder are shown in Table 1. Add 100g of carbon powder to No. 1 battery powder, mix well, and place in a vacuum tube furnace. Calcinate at 700℃ for 2 hours to obtain No. 1 calcined battery powder. Add water at a liquid-to-solid mass ratio of 6:1 to form a slurry. Hydrogenate by introducing carbon dioxide at room temperature for 1 hour. The excess carbon dioxide coefficient is 1.5. Separate the solid and liquid of the hydrogenated slurry. The lithium concentration of the No. 1 hydrogenated lithium liquid is 8.62g / L. At the same time, No. 1 nickel-cobalt-manganese slag is obtained. Thermally decompose the No. 1 hydrogenated lithium liquid at 85℃ for 1 hour. Separate the solid and liquid to obtain No. 1 mother liquor and No. 1 lithium carbonate crystals. The grade of No. 1 lithium carbonate crystals meets the industrial grade requirements.

[0081] (2) Leaching

[0082] The No. 1 nickel-cobalt-manganese slag was mixed with the No. 1 crude nickel-cobalt hydroxide. The main dry components of the crude nickel-cobalt hydroxide are shown in Table 2. The mass ratio of the dry crude nickel-cobalt hydroxide to the nickel-cobalt-manganese slag was 20:1. Water was added to make a slurry according to the liquid-solid mass ratio of 5.5:1. After the slurry was completely formed, concentrated sulfuric acid with a mass concentration of 50% was slowly added. No obvious bubbles were generated during the acid addition process. The pH of the leaching endpoint was controlled at 1.52. After the pH stabilized for 1 hour, the solid and liquid were separated to obtain the No. 1 acid solution and the No. 1 acid slag. The main components of the acid solution are shown in Table 3.

[0083] (3) Removing impurities

[0084] Zirconium hydroxide was added to the No. 1 acid solution at 20℃ (the amount was controlled according to the final pH=5.00). The impurity removal time was 1h. After the impurity removal was completed, the solid and liquid were separated to obtain the No. 1 impurity-removed solution. Fe, Al, SiO2, F, Cr and Cu in the impurity-removed solution were effectively removed. The concentrations of the remaining components did not change much before and after the impurity removal. The main impurities and concentrations of the impurity-removed solution are shown in Table 4.

[0085] (4) Extraction

[0086] After precise filtration of the No. 1 acid solution, 20% (v / v) P204 nickel soap was added for extraction and impurity removal (the P204 blank organic solution was first diluted with kerosene to a volume concentration of 20%, then 30% (w / w) liquid alkali was added according to a saponification rate of 50% for sodium soaping. After sodium soaping, a low-impurity nickel sulfate solution (nickel ion concentration of 10-100 g / L) was added for nickel soaping for 15 min, with two stages and a ratio of 1:1. The purpose was to wash off the sodium introduced by the sodium soap so as not to increase the sodium ion concentration in the refined nickel sulfate solution). The extraction temperature was 25℃, the single-stage extraction time was 20 min, the ratio was 1:3 (v / v), and 12 stages of countercurrent extraction were performed. After phase separation, the No. 1 P204 raffinate and the No. 1 loaded organic phase I were obtained. The P204 raffinate contained calcium, The concentrations of manganese and zinc were as low as 0.8-1.5 mg / L. Except for a slight increase in nickel concentration in the P204 raffinate due to the use of nickel soap, the concentrations of other elements remained basically unchanged. The specific composition is shown in Table 5. Further extraction of cobalt and magnesium was carried out by adding 15% P507 nickel soap (using the same saponification method as P204 nickel soap) to the 1# P204 raffinate. The extraction temperature was 25℃, the single-stage extraction time was 20 min, the ratio was 1:3 (v / v), and the extraction method was countercurrent extraction for 12 stages. After phase separation, 1# refined nickel sulfate solution and 1# supported organic phase IV were obtained. The concentration of cobalt and magnesium in the 1# refined nickel sulfate solution decreased to 1.2 mg / L. Except for a slight increase in nickel concentration compared to the P204 raffinate due to the use of nickel soap, the concentrations of other elements remained basically unchanged. The specific composition is shown in Table 6.

[0087] (5) Recovery of manganese and cobalt

[0088] The #1 loaded organic phase I was washed twice with water at room temperature. Then, 2.5N dilute sulfuric acid was added to the washed P204 organic solution to control the pH of the aqueous phase at 2.09 for manganese back-extraction, resulting in a #1 manganese sulfate solution and a #1 loaded organic phase II. Zinc was then back-extracted from the #1 loaded organic phase II with 2.5N dilute sulfuric acid, controlling the pH of the aqueous phase at 1.51. After zinc back-extraction, the #1 organic phase III was obtained. The #1 organic phase III can be returned for extraction and impurity removal.

[0089] Organic phase IV (1#) was washed twice with water at room temperature, and then magnesium was back-extracted using 1.0N dilute sulfuric acid to control the pH of the aqueous phase at 4.02, yielding organic phase V (1#). Organic phase V (1#) was then back-extracted for cobalt once using 1.5N dilute sulfuric acid to control the pH of the aqueous phase at 3.00, yielding organic phase VI (1#) and a purified cobalt sulfate solution (1#). Organic phase VI (1#) was then back-extracted for cobalt a second time using 1.5N dilute sulfuric acid to control the pH of the aqueous phase at 1.06, yielding organic phase VI (1#). And the No. 1 secondary cobalt back-extraction solution, No. 1 organic phase It can be reused for extracting cobalt and magnesium. The main components of manganese sulfate solution and refined cobalt sulfate solution are shown in Tables 7 and 8, respectively.

[0090] Example 2

[0091] (1) Prioritize lithium extraction

[0092] Take 5 kg of No. 2 battery powder. The main components of the battery powder are as follows: nickel 48.22%, cobalt 9.86%, manganese 15.25%, lithium 1.45%, fluorine 0.17%, copper 2.04%, and aluminum 1.37%. The main components of the battery powder are shown in Table 1. Add 500 g of carbon powder to No. 2 battery powder, mix well, and place in a calcining furnace. Introduce carbon dioxide and calcine at 800℃ for 3 hours to obtain No. 2 calcined battery powder. Add water to form a slurry at a liquid-to-solid mass ratio of 4.5:1. Under normal temperature conditions, introduce carbon dioxide for hydrogenation for 2 hours. The excess carbon dioxide coefficient is 2.0. Separate the solid and liquid of the hydrogenated slurry to obtain No. 2 hydrogenated lithium liquid with a lithium concentration of 5.25 g / L. Thermally decompose No. 2 hydrogenated lithium liquid at 90℃ for 2 hours and separate the solid and liquid to obtain No. 2 mother liquor and No. 2 lithium carbonate crystals. The grade of No. 2 lithium carbonate crystals meets the industrial grade requirements.

[0093] (2) Leaching

[0094] The No. 2 nickel-cobalt-manganese slag was mixed with the No. 2 crude nickel-cobalt hydroxide. The main dry components of the crude nickel-cobalt hydroxide are shown in Table 2. The mass ratio of the dry crude nickel-cobalt hydroxide to the nickel-cobalt-manganese slag was 15:1. Water was added to make a slurry according to the liquid-solid mass ratio of 6.5:1. After the slurry was completely formed, concentrated sulfuric acid with a mass concentration of 70% was slowly added. No obvious bubbles were generated during the acid addition process. The pH of the leaching endpoint was controlled at 2.50. After the pH stabilized for 2 hours, the solid and liquid were separated to obtain the No. 2 acid solution and the No. 2 acid slag. The main components of the acid solution are shown in Table 3.

[0095] (3) Removing impurities

[0096] Zirconium hydroxide was added to acid solution #2 at 60℃ (the amount was controlled according to the endpoint pH=5.51). The total impurity removal time was 2h. After the impurity removal was completed, solid and liquid were separated to obtain impurity-removed solution #2. Fe, Al, SiO2, F, Cr and Cu in the impurity-removed solution were effectively removed, and the concentration of other components did not change much. The main impurities and concentrations in the impurity-removed solution are shown in Table 4.

[0097] (4) Extraction

[0098] After precise filtration of the #2 acid solution, 25% P204 nickel soap was added for extraction and impurity removal (the P204 blank organic solution was first diluted with kerosene to a volume concentration of 25%, then 30% liquid alkali was added according to a saponification rate of 50% for sodium soaping. After sodium soaping, a low-impurity nickel sulfate solution was added for nickel soaping for 15 minutes, with two stages and a ratio of 1:1). The extraction temperature was 35℃, the single-stage extraction time was 15 minutes, the ratio was 1:3, and 12 stages of countercurrent extraction were performed. After phase separation, the #2 P204 raffinate and the #2 loaded organic phase I were obtained. The concentrations of calcium, manganese, and zinc in the raffinate were all as low as 0.1-1.3 mg / L, except that due to the use of nickel soap, The nickel concentration in the P204 raffinate increased slightly, while the concentrations of other elements remained essentially unchanged. The specific composition is shown in Table 5. Further extraction of cobalt and magnesium was carried out by adding 10% P507 nickel soap (using the same saponification method as P204 nickel soap) to the #2 P204 raffinate. The extraction temperature was 35℃, the single-stage extraction time was 15 min, the ratio was 1:3, and the extraction method was countercurrent extraction for 12 stages. After phase separation, two purified nickel sulfate solutions and a #2 supported organic phase IV were obtained. The cobalt and magnesium concentrations in the purified nickel sulfate solution decreased to 1.0-1.3 mg / L. Except for a slight increase in nickel concentration compared to the P204 raffinate due to the use of nickel soap, the concentrations of other elements remained essentially unchanged. The specific composition is shown in Table 6.

[0099] (5) Recovery of manganese and cobalt

[0100] The No. 2 loaded organic phase I was washed twice with water at room temperature. Then, 2.0N dilute sulfuric acid was added to the washed P204 organic solution to control the pH of the aqueous phase at 2.19 for manganese back-extraction, resulting in the No. 2 manganese sulfate solution and the No. 2 loaded organic phase II. Zinc was then back-extracted from the No. 2 loaded organic phase II with 2.0N dilute sulfuric acid, controlling the pH of the aqueous phase at 1.03. After zinc back-extraction, the No. 2 organic phase III was obtained. The No. 2 organic phase III can be returned for extraction and impurity removal.

[0101] Organic phase IV (2#) was washed twice with water at room temperature, and then magnesium was back-extracted using 0.5N dilute sulfuric acid to control the pH of the aqueous phase at 4.55, yielding organic phase V (2#). Organic phase V (2#) was then back-extracted for cobalt once using 2.5N dilute sulfuric acid to control the pH of the aqueous phase at 3.22, yielding organic phase VI (2#) and purified cobalt sulfate solution (1#). Organic phase VI (2#) was then back-extracted for cobalt a second time using 2.5N dilute sulfuric acid to control the pH of the aqueous phase at 1.50, yielding organic phase VII (2#) and a secondary cobalt aqueous phase (2#). Organic phase VII (2#) can be reused for extracting cobalt and magnesium. The main components of the manganese sulfate solution and the purified cobalt sulfate solution are shown in Tables 7 and 8, respectively.

[0102] Example 3

[0103] (1) Prioritize lithium extraction

[0104] 100 kg of No. 3 battery powder was taken. The main components of the battery powder are as follows: nickel 41.10%, cobalt 8.16%, manganese 6.65%, lithium 6.93%, fluorine 0.30%, copper 0.46%, and aluminum 1.80%, as shown in Table 1. 15 kg of carbon powder was added to the No. 3 battery powder, mixed well, and placed in a calcining furnace. Nitrogen gas was introduced, and the mixture was calcined at 1000℃ for 1.5 h to obtain No. 3 calcined battery powder. Water was added to the mixture at a liquid-to-solid mass ratio of 10:1 to form a slurry. Carbon dioxide was introduced at room temperature for hydrogenation for 3 h, with an excess carbon dioxide coefficient of 2.5. The hydrogenated slurry was separated into solid and liquid components to obtain No. 3 hydrogenated lithium liquid with a lithium concentration of 10.06 g / L. The No. 3 hydrogenated lithium liquid was thermally decomposed at 95℃ for 3 h, and the solid and liquid components were separated to obtain No. 3 mother liquor and No. 3 lithium carbonate crystals. The grade of the No. 3 lithium carbonate crystals met the industrial grade requirements.

[0105] (2) Leaching

[0106] The No. 3 nickel-cobalt-manganese slag was mixed with the No. 3 crude nickel-cobalt hydroxide. The main dry basis components of the crude nickel-cobalt hydroxide are shown in Table 2. The mass ratio of the crude nickel-cobalt hydroxide dry basis to the nickel-cobalt-manganese slag was 30:1. Water was added to make a slurry according to the liquid-solid mass ratio of 4:1. After the slurry was completely formed, concentrated sulfuric acid with a mass concentration of 98% was slowly added. No obvious bubbles were generated during the acid addition process. The pH of the leaching endpoint was controlled at 2.17. After the pH stabilized for 2 hours, the solid and liquid were separated to obtain the No. 3 acid solution and the No. 3 acid slag. The main components of the acid solution are shown in Table 3.

[0107] (3) Removing impurities

[0108] At 95℃, calcium carbonate was added to acid solution #3 (the amount was controlled according to the endpoint pH=5.68). The total impurity removal time was 4h. After the impurity removal was completed, solid and liquid were separated to obtain impurity-removed solution #3. Fe, Al, SiO2, F, Cr and Cu in the impurity-removed solution were effectively removed, and the concentration of other components did not change much. The main impurities and concentrations in the impurity-removed solution are shown in Table 4.

[0109] (4) Extraction

[0110] After precise filtration of the #3 acid solution, 22% P204 nickel soap was added for extraction and impurity removal (the P204 blank organic solution was first diluted with kerosene to a volume concentration of 22%, then 30% liquid alkali was added according to a saponification rate of 50% for sodium soaping. After sodium soaping, low-impurity nickel sulfate solution was added for nickel soaping for 15 minutes, with two stages and a ratio of 1:1). The extraction temperature was 40℃, the single-stage extraction time was 10 minutes, the ratio was 1:3, and 12 stages of countercurrent extraction were performed. After phase separation, the #3 P204 raffinate and the #3 loaded organic phase I were obtained. The concentrations of calcium, manganese, and zinc in the P204 raffinate were all as low as 0.5-1.5 mg / L, except for the fact that... The use of nickel soap resulted in a slight increase in the nickel concentration of the P204 raffinate, while the concentrations of other elements remained essentially unchanged. The specific composition is shown in Table 5. Further extraction of cobalt and magnesium was carried out by adding 15% P507 nickel soap (using the same saponification method as P204 nickel soap) to the #3 P204 raffinate. The extraction temperature was 40℃, the single-stage extraction time was 10 min, the phase ratio was 1:3, and 12 stages of countercurrent extraction were performed. After phase separation, the #3 refined nickel sulfate solution and the #3 supported organic phase IV were obtained. The cobalt and magnesium concentrations in the refined nickel sulfate solution decreased to below 2 mg / L. Except for a slight increase in nickel concentration compared to the P204 raffinate due to the use of nickel soap, the concentrations of other elements remained essentially unchanged. The specific composition is shown in Table 6.

[0111] (5) Recovery of manganese and cobalt

[0112] The No. 3 loaded organic phase I was washed twice with water at room temperature. Then, 1.5N dilute sulfuric acid was added to the washed P204 organic solution, and the pH of the aqueous phase was controlled at 2.52 for manganese back-extraction to obtain the No. 3 manganese sulfate solution and the No. 3 loaded organic phase II. Zinc was back-extracted from the No. 3 loaded organic phase II with 1.5N dilute sulfuric acid, and the pH of the aqueous phase was controlled at 1.80. After zinc back-extraction, the No. 3 organic phase III was obtained. The No. 3 organic phase III can be returned for extraction and impurity removal.

[0113] Organic phase IV (#3) was washed twice with water at room temperature, and then magnesium was back-extracted with 0.8N dilute sulfuric acid, controlling the pH of the aqueous phase at 4.21, yielding organic phase V (#3). Organic phase V (#3) was then back-extracted for cobalt once with 2.0N dilute sulfuric acid, controlling the pH of the aqueous phase at 3.49, yielding organic phase VI (#3) and purified cobalt sulfate solution (#1). Organic phase VI (#3) was then back-extracted for cobalt a second time with 2.0N dilute sulfuric acid, controlling the pH of the aqueous phase at 1.24, yielding organic phase VII (#3) and secondary cobalt aqueous phase (#3). Organic phase VII (#3) can be reused for extracting cobalt and magnesium. The main components of the manganese sulfate solution and the purified cobalt sulfate solution are shown in Tables 7 and 8, respectively.

[0114] Example 4

[0115] (1) Prioritize lithium extraction

[0116] 500 kg of No. 4 battery powder was taken. The main components of the battery powder are as follows: nickel 27.08%, cobalt 5.27%, manganese 4.36%, lithium 4.22%, fluorine 1.46%, copper 0.79%, and aluminum 6.14%, as shown in Table 1. 25 kg of carbon powder was added to the No. 4 battery powder, mixed well, and placed in a calcining furnace. Nitrogen gas was introduced, and the mixture was calcined at 1200℃ for 1 hour to obtain No. 4 calcined battery powder. Water was added at a liquid-to-solid mass ratio of 8:1 to form a slurry. Under normal temperature conditions, carbon dioxide was introduced for hydrogenation for 2 hours. The excess carbon dioxide coefficient was 3.0. The hydrogenated slurry was separated into solid and liquid components to obtain No. 4 hydrogenated lithium liquid with a lithium concentration of 8.19 g / L. The No. 4 hydrogenated lithium liquid was thermally decomposed at 92℃ for 2 hours, and the solid and liquid components were separated to obtain No. 4 mother liquor and No. 4 lithium carbonate crystals. The grade of No. 4 lithium carbonate crystals met the industrial grade requirements.

[0117] (2) Leaching

[0118] The No. 4 nickel-cobalt-manganese slag was mixed with the No. 4 crude nickel-cobalt hydroxide. The main dry components of the crude nickel-cobalt hydroxide are shown in Table 2. The mass ratio of the dry crude nickel-cobalt hydroxide to the nickel-cobalt-manganese slag was 50:1. Water was added to make a slurry according to the liquid-solid mass ratio of 3:1. After the slurry was completely formed, concentrated sulfuric acid with a mass concentration of 85% was slowly added. No obvious bubbles were generated during the acid addition process. The pH of the leaching endpoint was controlled at 1.03. After the pH stabilized for 1 hour, the solid and liquid were separated to obtain the No. 4 acid solution and the No. 4 acid slag. The main components of the acid solution are shown in Table 3.

[0119] (3) Removing impurities

[0120] At 70℃, calcium carbonate was added to acid solution #4 (the amount was controlled according to the endpoint pH=6.00). The impurity removal time was 3h. After the impurity removal was completed, solid and liquid were separated to obtain impurity-removed solution #4. Fe, Al, SiO2, F, Cr and Cu in the impurity-removed solution were effectively removed, and the concentration of other components did not change much. The main impurities and concentrations in the impurity-removed solution are shown in Table 4.

[0121] (4) Extraction

[0122] After precise filtration of the #4 acid solution, 20% P204 nickel soap was added for extraction and impurity removal (the P204 blank organic solution was first diluted with kerosene to a volume concentration of 20%, then sodium soap was performed by adding 30% liquid alkali according to a saponification rate of 50%, and after the sodium soap was completed, low-impurity nickel sulfate solution was added for nickel soap, with a nickel soaping time of 15 min, 2 stages, and a ratio of 1:1). The extraction temperature was 40℃, the single-stage extraction time was 10 min, the ratio was 1:4, and 10 stages of countercurrent extraction were performed. After phase separation, the #4 P204 raffinate and the #4 loaded organic phase I were obtained. The concentrations of calcium, manganese, and zinc in the P204 raffinate were all as low as 0.6-1.1 mg / L, except for those from... The use of nickel soap resulted in a slight increase in the nickel concentration of the P204 raffinate, while the concentrations of other elements remained essentially unchanged. The specific composition is shown in Table 5. Further extraction of cobalt and magnesium was carried out by adding 10% P507 nickel soap (using the same saponification method as P204 nickel soap) to the P204 raffinate. The extraction temperature was 40℃, the single-stage extraction time was 10 min, the phase ratio was 1:4, and 10 stages of countercurrent extraction were performed. After phase separation, a refined nickel sulfate solution (P204) and a loaded organic phase (P204) were obtained. The cobalt and magnesium concentrations in the refined nickel sulfate solution decreased to 1.5 mg / L. Except for a slight increase in nickel concentration compared to the P204 raffinate due to the use of nickel soap, the concentrations of other elements remained essentially unchanged. The specific composition is shown in Table 6.

[0123] (5) Recovery of manganese and cobalt

[0124] The No. 4 loaded organic phase I was washed twice with water at room temperature. Then, 1.8N dilute sulfuric acid was added to the washed P204 organic phase, and the pH of the aqueous phase was controlled at 2.35 for manganese back-extraction to obtain the No. 4 manganese sulfate solution and the No. 4 loaded organic phase II. Zinc was back-extracted from the No. 4 loaded organic phase II with 1.8N dilute sulfuric acid, and the pH of the aqueous phase was controlled at 1.33. After zinc back-extraction, the No. 4 organic phase III was obtained. The No. 4 organic phase III can be returned for extraction and impurity removal.

[0125] Organic phase IV (4#) was washed twice with water at room temperature, and then magnesium was back-extracted using 0.5N dilute sulfuric acid to control the pH of the aqueous phase at 4.10, yielding organic phase V (4#). Organic phase V (4#) was then back-extracted for cobalt once using 1.5N dilute sulfuric acid to control the pH of the aqueous phase at 3.16, yielding organic phase VI (4#) and a purified cobalt sulfate solution (4#). Organic phase VI (4#) was then back-extracted for cobalt a second time using 1.5N dilute sulfuric acid to control the pH of the aqueous phase at 1.37, yielding organic phase VII (4#) and a secondary cobalt aqueous phase (4#). Organic phase VII (4#) can be reused for extracting cobalt and magnesium. The main components of the manganese sulfate solution and the purified cobalt sulfate solution are shown in Tables 7 and 8, respectively.

[0126] Table 1 Main components of waste NCM ternary battery powder

[0127] unit:%

[0128]

[0129] Table 2 Main components of crude nickel-cobalt hydroxide (dry basis)

[0130] unit:%

[0131]

[0132] Table 3 Main components of acid solutions (unit: g / L)

[0133]

[0134] Table 4. Concentration of main impurities in the purified solution (unit: g / L)

[0135]

[0136] Table 5 Main components of P204 raffinate (unit: g / L)

[0137]

[0138] As shown in Table 3-5, the main impurities in the acid solution are Ca, Zn, Cu, Al, and Mg. After purification, except for Ca and Mg, the other impurity ions are thoroughly removed. Under the action of the neutralizing agent, the concentrations of Fe, SiO2, F, and Cr in the purified solution remain at a low level and will not affect the quality of subsequent products. After the purified solution is extracted with P204, the concentration of nickel ions in the extract increases slightly due to the use of nickel soap during extraction, while the concentrations of cobalt and magnesium ions do not change significantly. Manganese ions are loaded into the P204 extractant, thus achieving the separation of nickel ions from cobalt and magnesium ions.

[0139] 6. Composition of refined nickel sulfate solution (unit: g / L)

[0140]

[0141] As shown in Table 6, the refined nickel sulfate solution contains virtually no impurities and can be used to produce battery-grade nickel sulfate through degreasing, evaporation, and crystallization, or it can also be used to produce electrolytic nickel.

[0142] Table 7 Main components of manganese sulfate solution (unit: g / L)

[0143]

[0144] As shown in Table 7, the main impurities in manganese sulfate solution are small amounts of calcium and heavy metals such as nickel and zinc below 100 mg / L. Manganese sulfate crystals can be produced by removing calcium and heavy metals (e.g., by using fluoride precipitation to remove calcium and sulfide precipitation to remove heavy metals).

[0145] Table 8 Main components of refined cobalt sulfate solution (unit: g / L)

[0146]

[0147] As shown in Table 8, the refined cobalt sulfate solution has a low impurity content. After oil removal, it can be used to produce battery-grade cobalt sulfate crystals or electrolytic cobalt.

[0148] In summary, the main component of the manganese slag produced by acid dissolution and precision filtration is manganese dioxide. It can be used to partially replace crude nickel-cobalt hydroxide, making full use of the oxidant to save costs while solving the problem of difficult acid dissolution of nickel-cobalt-manganese slag and improving the recovery rate of manganese.

[0149] The above description is a detailed description of the preferred embodiments of the present invention. However, the embodiments are not intended to limit the scope of the patent application of the present invention. All equivalent changes or modifications made under the technical spirit of the present invention should fall within the patent scope covered by the present invention.

Claims

1. A method for recovering nickel, cobalt, and manganese from waste ternary lithium battery powder, characterized in that, Includes the following steps: (1) Pre-extraction of lithium: waste ternary battery powder is reduced and roasted under oxygen-isolated conditions to obtain roasted battery powder. Water is added to the roasted battery powder to make slurry, and hydrogenation is carried out to remove impurities. Solid-liquid separation is performed to obtain hydrogenated lithium liquid and nickel-cobalt-manganese slag. The hydrogenated lithium liquid is thermally decomposed to recover lithium carbonate. (2) Recovery of nickel, cobalt, and manganese: S1, Leaching: Crude nickel-cobalt hydroxide is mixed into the nickel-cobalt-manganese slag obtained in step (1), water is added to make a slurry at a liquid-solid mass ratio of 3.0-6.5:1, concentrated sulfuric acid is added, acid leaching is performed, solid and liquid are separated, and acid solution and acid leaching residue are obtained. The acid leaching residue is returned to step S1 for use. S2. Impurity removal: Add a neutralizing agent to the acid solution, and separate the solid and liquid to obtain the impurity-removed residue and the impurity-removed liquid; S3. Extraction: After filtering the impurity-removed liquid, add P204 to extract impurities and separate the phases to obtain P204 raffinate and loaded organic phase I; add P507 to P204 raffinate to extract cobalt and magnesium and separate the phases to obtain purified nickel sulfate solution and loaded organic phase IV. S4. Manganese and cobalt recovery: Add dilute sulfuric acid to the supported organic phase I for back-extraction and phase separation to obtain a manganese sulfate solution. Add dilute sulfuric acid to the supported organic phase IV for back-extraction and phase separation to obtain a refined cobalt sulfate solution. In step S1, the amount of crude nickel-cobalt hydroxide added is 15-50 times the weight of the nickel-cobalt-manganese slag. The crude nickel-cobalt hydroxide contains 2-10 wt.% manganese; In step S1, the final pH of the acid leaching is 1.0-2.

5.

2. The method for recovering nickel, cobalt, and manganese from waste ternary battery powder as described in claim 1, characterized in that: In step (1), 5-20% of the weight of waste ternary battery powder is added to carbon powder for reduction roasting. The reduction roasting conditions are: temperature 700-1200℃ and time 1-3h.

3. The method for recovering nickel, cobalt, and manganese from waste ternary battery powder as described in claim 1, characterized in that: In step (1), water is added to make a slurry with a lithium concentration of 5-10 g / L, and carbon dioxide is introduced for hydrogenation to remove impurities. The hydrogenation conditions are: hydrogenation time of 1-3 h and carbon dioxide excess coefficient of 1.5-3.0; thermal decomposition conditions are: temperature of 85-95 °C and time of 1-3 h.

4. The method for recovering nickel, cobalt, and manganese from waste ternary battery powder as described in claim 1, characterized in that: In step S2, the neutralizing agent is limestone or zirconium hydroxide, the pH is controlled at 5.0-6.0, and the time is 1-4 hours.

5. The method for recovering nickel, cobalt, and manganese from waste ternary battery powder as described in claim 1, characterized in that: In step S3, the volume concentration of P204 is controlled at 20-25% when extracting impurities; the volume concentration of P507 is controlled at 10-15% when extracting cobalt and magnesium; the single-stage extraction conditions are: time of 10-20 min and temperature of 20-50℃.

6. The method for recovering nickel, cobalt, and manganese from waste ternary battery powder as described in claim 1, characterized in that: Step (2) further includes step S5, which specifically involves: washing the nickel in the supported organic phase I with pure water, then adding 1.5-2.5N dilute sulfuric acid to back-extract manganese, controlling the pH of the aqueous phase to be 2.0-2.5, to obtain a manganese sulfate solution and the supported organic phase. Continue adding 1.5-2.5N of dilute sulfuric acid to the supported organic phase II for zinc back-extraction, controlling the pH of the aqueous phase to 1.0-1.8, to obtain a zinc sulfate solution and an organic phase. organic phase Return to step S3 for use.

7. The method for recovering nickel, cobalt, and manganese from waste ternary battery powder as described in claim 6, characterized in that: Step (2) further includes step S6, which specifically involves: loading the organic phase... First, nickel is washed with pure water, then magnesium is back-extracted with 0.5-1.5N dilute sulfuric acid, controlling the pH of the aqueous phase to 4.0-4.5, to obtain a magnesium sulfate solution and a supported organic phase. Continue to support the organic phase Cobalt was back-extracted once by adding 1.5-2.5N dilute sulfuric acid, controlling the pH of the aqueous phase to 2.5-3.5, to obtain a purified cobalt sulfate solution and a supported organic phase. The loaded organic phase VI is subjected to a second back-extraction of cobalt using 1.5-2.5N dilute sulfuric acid, while controlling the pH of the aqueous phase to 1.0-1.50, to obtain organic phase VII and the liquid after the second cobalt back-extraction. Organic phase VII is returned to step S3 for use.

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