Method for recycling waste lithium ion battery black powder

Abstract

The application belongs to the field of waste battery material recycling, and specifically discloses a recycling method of waste lithium ion battery black powder. The waste lithium ion battery black powder is mixed with aluminum cerium additives for pressure acid leaching, and then lithium leaching solution and delithiated residue enriched with transition metals and graphite in the black powder are obtained. The aluminum cerium additives contain an aluminum source and a cerium source. The delithiated residue is subjected to reduction acid leaching treatment, and graphite and acid leaching solution enriched with transition metals are obtained through solid-liquid separation. The acid leaching solution is subjected to first-stage extraction copper removal treatment, and first-stage impurity removal solution is obtained. The first-stage impurity removal solution and calcium impurity removal agent are subjected to second-stage impurity removal treatment, and second-stage impurity removal solution is obtained. Third-stage extraction impurity removal is performed on the second-stage impurity removal solution in cooperation with an extractant, and purified third-stage impurity removal solution enriched with transition metals is obtained. The application can realize selective separation of transition metal elements and impurities such as calcium, aluminum, fluorine and cerium.

Description

Technical Field

[0001] This invention belongs to the field of waste battery recycling technology, specifically relating to the recycling of waste lithium-ion battery black powder. Background Technology

[0002] With the peak of power battery retirement approaching, the large-scale disposal of used batteries has become an urgent industrial problem to be solved. Used lithium-ion batteries are rich in key metals such as nickel, cobalt, manganese, and lithium, and achieving efficient and clean recycling of these resources is crucial to alleviating supply pressure.

[0003] Currently, the industry generally adopts the "pretreatment-wet separation" route, which uses sulfuric acid and reducing agents to perform non-selective full leaching of battery black powder, followed by the separation of various metals through precipitation and extraction, with lithium remaining in the final raffinate for recovery. However, this process is lengthy, and lithium is prone to co-extraction loss or entrainment into the sodium sulfate crystals formed by evaporation and concentration during the multi-step separation process, resulting in a final lithium recovery rate that is usually less than 85%.

[0004] To improve lithium recovery rates and avoid lithium dispersion and loss in long downstream processes, the industry tends to adopt "priority lithium extraction" processes, but these often rely on energy-intensive roasting techniques. For example, patent document CN106129511A discloses a carbon reduction roasting-water leaching lithium extraction method. Although it achieves priority lithium extraction, the high-temperature treatment leads to high energy consumption and harmful flue gas problems. Furthermore, the reduced reaction efficiency during scale-up production and the residue of lithium-containing byproducts still limit the lithium recovery rate.

[0005] Furthermore, the deep removal of impurities and the efficient separation of valuable metals are also pain points in hydrometallurgy. Complex impurities such as aluminum, fluorine, calcium, and copper in black powder easily interfere with the main process in acidic systems. For example, although patent document CN112374511A achieves valuable metal recovery, it generates a large amount of copper slag, iron-aluminum slag, and fluoride slag during the impurity removal process, resulting in the loss of valuable metals and significant solid waste treatment pressure. In terms of nickel, cobalt, and manganese recovery, traditional processes such as Chinese patent document CN120664569A typically use extractants such as P204 and P507 for cumbersome stepwise separation, requiring frequent saponification and back-extraction operations. This not only consumes large amounts of acid and alkali reagents but also generates massive amounts of high-salt wastewater, making it difficult to meet the demand for low-cost, short-process green recycling. Summary of the Invention

[0006] In view of the shortcomings of the existing technology, the purpose of this invention is to provide a method for recycling waste lithium-ion battery black powder, which aims to efficiently recover graphite and valuable metal elements therein.

[0007] The black powder contains graphite anode, cathode material current collector, electrolyte, and other components, resulting in a complex composition. Metal elements are deeply embedded in the graphite structure, forming a dense SEI on the surface. Furthermore, the high fluorine content in the black powder not only causes some lithium to precipitate as insoluble lithium fluoride, but also easily triggers further fluorination and precipitation reactions during subsequent separation and recovery. This further exacerbates the difficulty of efficiently and selectively separating and recovering lithium and transition metals. Based on in-depth research, this invention provides the following improvement:

[0008] The recycling method for black powder from spent lithium-ion batteries includes the following steps:

[0009] Step a:

[0010] Waste lithium-ion battery black powder and aluminum-cerium additives are mixed and subjected to pressurized acid leaching, followed by separation to obtain lithium leachate and delithiation slag enriched with transition metals from the black powder and graphite; the aluminum-cerium additives contain aluminum source and cerium source.

[0011] Step b:

[0012] The lithium-depleted slag is subjected to reduction acid leaching treatment, and solid-liquid separation is performed to obtain graphite and acid leaching solution enriched with transition metals.

[0013] Step c:

[0014] The acid leaching solution is subjected to a first-stage extraction to remove copper (first-stage impurity removal) to obtain a first-stage impurity-removed solution.

[0015] Step d:

[0016] The first-stage impurity removal solution and calcium impurity removal agent are subjected to a second-stage impurity removal treatment (second impurity removal) to obtain a second-stage impurity removal solution;

[0017] Step e:

[0018] The second-stage impurity removal solution was subjected to a third-stage extraction and impurity removal using a synergistic extractant (third-stage impurity removal) to obtain a purified third-stage impurity removal solution enriched with transition metal elements.

[0019] The extractable active ingredient in the extracted organic phase comprises extractant A and extractant B in a volume ratio of 1:2~5; wherein extractant A is a compound with structure of formula 1; and extractant B is a compound with structure of formula 2.

[0020] Formula 1;

[0021] Formula 2;

[0022] In Formula 1, R1 is a C1~C8 alkyl group, C4~C6 ... 10 The alkoxy or phenoxy group; R2 is H, C1~C8 alkyl, C4~C 10The alkoxy or phenoxy group; M is H, Na, K or NH4;

[0023] In Formula 2, R3 and R4 are individually C1~C8 alkyl groups and C4~C6 alkyl groups. 10 alkoxy or phenoxy groups;

[0024] The extraction process is carried out at a temperature below 45℃ and at a pH of 2.0 to 4.0.

[0025] This invention innovatively employs aluminum-cerium additives in the pressurized acid leaching of black powder, which is tailored to the characteristics of black powder and facilitates the selective and preferential extraction of lithium. Furthermore, the combined three-stage treatment of the delithiation slag—reductive acid leaching, primary copper removal, secondary calcium precipitation, and synergistic impurity removal using formulas 1 and 2—achieves selective separation of transition metal elements and impurities such as calcium, aluminum, fluorine, and cerium, facilitating the efficient recovery of graphite and transition metal elements.

[0026] In this invention, the transition metal element in the waste lithium-ion battery black powder includes at least one of nickel, cobalt, and manganese.

[0027] Preferably, in the waste lithium-ion battery black powder, the content of transition metals is 5-35%; the lithium content is 3-5%; the content of anionic impurities (at least one of F and P) is 3-6%; and the content of at least one metallic impurity from Al and Cu is 0.5-2.5%. Further, the total content of transition metals is 25-30%; the lithium content is 3-5%; the total content of anionic impurities (at least one of F and P) is 4-6%; and the total content of at least one metallic impurity from Al and Cu is 1.5-2.5%.

[0028] In this invention, the aluminum-cerium additive contains at least one aluminum source selected from aluminum sulfate, aluminum hydroxide, aluminum oxide, aluminum chloride, and aluminum nitrate; and the cerium source selected from at least one combination of cerium sulfate and cerium hydroxide.

[0029] Preferably, in the aluminum-cerium additive, the molar ratio of aluminum to cerium is 1~15:1~10; further, it can be 1~15:1; even further, it can be 5~10:1.

[0030] Preferably, the aluminum-cerium additive is 3-15 wt% of the weight of waste lithium-ion battery black powder; more preferably, it is 4-10 wt%; and even more preferably, it is 5-6 wt%.

[0031] In this invention, the acid used in the pressure pickling process is an inorganic acid, preferably sulfuric acid;

[0032] Preferably, the concentration of acid in the pressure pickling starting slurry is 0.4~0.6M;

[0033] Preferably, in pressurized acid leaching, the H+ of the acid... +The molar ratio of Li in the waste lithium-ion battery black powder is 0.8~1.3:1; it can be further 0.9~1.05:1.

[0034] Preferably, the temperature of the pressurized acid leaching is above 180°C, and more preferably it is 180~240°C;

[0035] Preferably, the pressure pickling time is more than 1 hour, and considering efficiency, it can be further extended to 1 to 2 hours.

[0036] This invention also includes the steps of removing impurities from lithium leaching solution and preparing lithium carbonate by carbonation precipitation;

[0037] The impurity removal process includes oil removal and weight removal.

[0038] The acid used in the reducing acid leaching process is an inorganic acid, preferably sulfuric acid;

[0039] Preferably, the concentration of acid in the initial reducing acid leaching solution is 1~5M;

[0040] Preferably, the reducing agent added during the reducing acid leaching process includes one of hydrogen peroxide, sodium sulfite, sodium bisulfite, and sulfurous acid;

[0041] Preferably, the ratio of the reducing agent to the total molar amount of transition metals in the delithiation slag is 0.5~1.5:1, and considering the cost, it can be further 0.7~1:1;

[0042] Preferably, the reduction acid leaching reaction temperature is 70~90℃, the solid-liquid ratio is 200~600 g / L, and the reaction time is 0.5~3h, which can further be 1~2.5h.

[0043] In this invention, after reducing acid leaching, solid-liquid separation is continued to separate graphite and an acid leaching solution enriched with transition metals. The acid leaching solution typically contains impurities such as aluminum, copper, cerium, and fluorine, which can interfere with the efficient recovery of transition metals. To address the selective problem of separating transition metals and the aforementioned impurities, this invention proposes a three-stage deep impurity removal method. Based on the combination of the aforementioned impurity removal method and parameters, synergy can be achieved, enabling high recovery rates and high selectivity of transition metal elements.

[0044] The extractant used for copper removal in the first stage is extractant of formula 3.

[0045] Formula 3;

[0046] In Formula 3, R5 and R6 are individually H, hydroxyl, alkyl, or alkoxy; R7 is H or a C1-C4 alkyl group; and M is H, Na, K, or NH4.

[0047] The present invention demonstrates that using Formula 3 as an extractant can selectively extract and remove copper without significantly affecting the recovery of transition metal elements.

[0048] In this invention, Formula 3 can further be a structure of Formula 3A:

[0049] Formula 3A;

[0050] In the first stage of copper removal extraction, the volume percentage concentration of Formula 1 in the organic phase is 5% to 30%; more specifically, it can be 15% to 25%.

[0051] Preferably, the O / A volume ratio in the first stage of copper removal extraction is 1:2 to 0.5.

[0052] In this invention, a calcium impurity remover is added to a section of the impurity removal solution to further remove impurities such as iron, cerium, aluminum, fluorine, and phosphorus.

[0053] Furthermore, the calcium impurity remover includes at least one of calcium hydroxide, calcium oxide, and calcium carbonate;

[0054] Based on the first stage of impurity removal solution, the amount of calcium impurity removal agent added is 5~15 g / L;

[0055] Preferably, the temperature of the second stage of impurity removal is 70~90℃, and the reaction time is 0.5~1 h.

[0056] In this invention, the second-stage raffinate contains fluorine (F) impurities, which are difficult to remove and hinder the selective separation of fluorine and transition metals. To address this problem, this invention regulates the Al / F molar ratio in the system by adding an aluminum-cerium additive during pressurized acid leaching. This, combined with the specific synergistic extractant and the joint control of parameters such as extraction ratio and pH, achieves synergistic effects, enabling deep removal of F and facilitating the highly selective acquisition of transition metal raffinate.

[0057] In this invention, thanks to the effect of the aluminum-cerium additive in step a, the molar ratio of Al / F in the aqueous phase to be treated in step e can be controlled simultaneously to be 1~3:1; further to 1.5~2.5:1.

[0058] In this invention, in step e, the volume percentage concentration of the extractable active ingredient (total amount of extractant A and extractant B) in the organic phase is 30-60%; more specifically, it can be 35-45%.

[0059] In this invention, the volume ratio of extractant A to extractant B is 1:2~2.5.

[0060] In this invention, the extraction temperature in step e is 10~20℃.

[0061] In this invention, in step e, the extraction pH is 2.5~3.5.

[0062] In this invention, in step e, the volume ratio of extracted O / A can be 1 to 5:1.

[0063] Beneficial effects:

[0064] 1) This invention introduces aluminum-cerium mixed salt for pre-slurrying, which can effectively improve the selective leaching of lithium on the one hand, and oxidize and precipitate a small amount of co-leached nickel, cobalt and manganese on the other hand, thereby significantly improving the preferential leaching rate and leaching selectivity of lithium; in addition, it is also beneficial to the deep removal of F in subsequent co-extraction.

[0065] 2) The present invention, through the synergistic extractant, combined with the control of the extractant ratio, Al / F ratio and pH during the extraction process, can synergistically improve the deep removal of impurities such as transition metal elements and F.

[0066] 3) The process of this invention does not require cumbersome nickel-cobalt-manganese single-metal cascade separation to obtain a high-purity mixed solution, which greatly simplifies the recovery process, avoids the large amount of extraction and back-extraction operations required for nickel-cobalt-manganese separation, significantly reduces the consumption of acid and alkali reagents and the discharge of high-salt wastewater, and realizes low-cost and green metal resource recovery. Attached Figure Description

[0067] Figure 1 This is a process flow diagram of the present invention. Detailed Implementation

[0068] This invention effectively reduces the complexity of subsequent separation and purification processes and significantly improves the overall recovery rate of valuable metals such as lithium, nickel, cobalt, and manganese by constructing a combined system of "pre-slurrying, pressurized acid leaching, lithium liquid purification, reducing acid leaching, copper extraction, calcium salt impurity removal, and selective extraction." Simultaneously, this method avoids excessive consumption of acid and alkali reagents in traditional leaching and extraction processes, significantly reduces the discharge of high-salt wastewater, and lowers production costs. It demonstrates significant industrial application prospects and feasibility in achieving efficient and selective recovery of key metals.

[0069] This invention provides a method for recycling valuable metals from waste lithium-ion battery black powder, comprising the following steps:

[0070] (1) Add a certain amount of dilute sulfuric acid and aluminum-cerium mixed salt (aluminum-cerium additive) to the waste lithium-ion battery black powder for pre-slurrying;

[0071] (2) The slurry is transferred to a high-pressure autoclave for pressurized acid leaching to separate lithium leaching solution and delithiation residue;

[0072] (3) The lithium leaching solution is deoiled and deweighted to obtain a lithium sulfate solution;

[0073] (4) The lithium-free slag is subjected to two-stage reduction acid leaching to separate nickel-cobalt-manganese leaching solution and graphite slag;

[0074] (5) The nickel-cobalt-manganese leaching solution is mixed with the copper extraction organic phase for extraction to obtain copper raffinate and copper solution;

[0075] (6) Add calcium salt to the copper extraction residue to adjust the alkali and preliminarily remove cerium, aluminum and fluorine;

[0076] (7) A preliminary impurity removal solution containing an organic solvent containing a synergistic extractant is used for deep purification, and aluminum, calcium and fluorine are removed simultaneously to obtain a high-purity nickel-cobalt-manganese solution.

[0077] In this invention, the used lithium-ion battery black powder includes at least one of the following: waste ternary lithium-ion battery positive electrode material and waste ternary lithium-ion battery positive and negative electrode mixed powder.

[0078] In this invention, in the slurry treatment described in step (1), the molar ratio of hydrogen ions in dilute sulfuric acid to lithium ions in the waste lithium-ion battery black powder is 0.8~1.2:1, the amount of aluminum-cerium mixed salt is 1%~10% of the mass of the black powder, the reaction temperature is 70~90℃, and the solid-liquid ratio is 100~400 g / L.

[0079] Preferably, the reaction time of the pulping treatment is not less than 0.5 h, and more preferably 0.5 to 1 h.

[0080] Through the pulping treatment in step (1), the pH of the slurry can be controlled at 2.5~3.5, at which point the content of hydrofluoric acid and residual acid in the system is extremely low.

[0081] The aluminum-cerium mixed salt is a combination of at least one of aluminum sulfate, aluminum hydroxide, aluminum oxide, aluminum chloride, and aluminum nitrate, and at least one of cerium sulfate and cerium hydroxide.

[0082] The molar ratio of aluminum to cerium in the aluminum-cerium mixed salt is 1~15:1~10.

[0083] In this invention, the temperature of the pressurized acid leaching in step (2) is preferably 180~240℃, and more preferably 220~240℃.

[0084] In this invention, the reaction time of the pressurized acid leaching in step (2) is preferably 1 to 4 hours, and more preferably 3 to 4 hours.

[0085] The leaching endpoint pH value of the pressurized acid leaching process in step (2) is 6.5~7.5; under this condition, the leaching rate of elements other than lithium is extremely low, achieving highly selective extraction of lithium.

[0086] In this invention, the oil removal in step (3) uses granular activated carbon with a dosage of 10~40 g / L and a reaction time of 0.5~1 h.

[0087] The weight removal process uses 10 mol / L liquid alkali at a concentration of 10 mL / L, with a reaction time of 0.5–1 h.

[0088] The small amount of precipitate residue obtained after weight removal is combined and incorporated into the subsequent reduction acid leaching process to recover nickel, cobalt, and manganese.

[0089] In this invention, the acid used in the two-stage reduction acid leaching in step (4) is sulfuric acid, the concentration of which is 1.2~3 mol / L, the reaction temperature is 70~90℃, the solid-liquid ratio is 200~600 g / L, and the reaction time is 0.5~3 h.

[0090] The specific operation of the two-stage reduction acid leaching is as follows: First, the nickel-cobalt-manganese delithiation slag undergoes a first-stage leaching, and the resulting leaching residue is used as the raw material for the second-stage leaching; then, acid and reducing agent are added to the second-stage leaching solution, and a second-stage leaching is performed using it to finally obtain the nickel-cobalt-manganese leaching solution. This method can control the pH value of the final nickel-cobalt-manganese leaching solution at 2.5~3.5 and keep the content of impurities such as aluminum, iron, and fluorine at a low level.

[0091] In this invention, the reducing agent used in the two-stage reduction acid leaching in step (4) is one of hydrogen peroxide, sodium sulfite, sodium bisulfite, and sulfurous acid. The ratio of the reducing agent to the total molar amount of nickel, cobalt, and manganese in the lithium removal slag is preferably 0.5~1.5:1, and more preferably 0.5~1:1.

[0092] In this invention, the copper-extracted organic phase in step (5) includes an extractant and a diluent. The extractant is the extractant of formula 3, and the diluent is at least one of sulfonated kerosene, No. 260 solvent oil, and C8-13 higher alcohols.

[0093] Preferably, the volume percentage of the extractant in the copper-extracted organic phase is 5% to 20%.

[0094] In this invention, the calcium salt in step (6) is at least one of calcium hydroxide, calcium oxide, and calcium carbonate, with an addition amount of 5~15 g / L, a reaction temperature of 70~90℃, and a reaction time of 0.5~1 h.

[0095] In this invention, the synergistic extractant in step (7) includes compound 1 and compound 2, wherein the synergistic extractant includes compound 1 (formula 1) and compound 2 (formula 3).

[0096] The volume percentage of compound 1 to compound 2 in the synergistic extractant is 1:1~5;

[0097] The organic phase also contains a diluent, which is at least one of sulfonated kerosene, No. 260 solvent oil, and C8-13 higher alcohols.

[0098] The volume percentage of compound 1 in the organic phase is 10% to 20%.

[0099] The aqueous phase equilibrium pH during the extraction process is 2.5~3.5.

[0100] The pH balance of the aqueous phase during the extraction process is controlled by saponifying the organic phase. The saponification process uses 10 mol / L liquid alkali and a nickel sulfate solution containing 2 mol / L nickel for the conversion. The saponification rate of the organic phase is 10%–90%, preferably 20%–60%.

[0101] Example 1

[0102] This embodiment provides a method for recycling valuable metals from waste lithium-ion battery black powder. The raw material composition is shown in Table 1, and the process diagram is shown in [Figure 1]. Figure 1 The specific process is as follows:

[0103] Table 1. Composition of black powder from waste lithium-ion batteries (wt%)

[0104]

[0105] Step (1):

[0106] Waste lithium-ion battery black powder, aluminum-cerium mixed salt (aluminum-cerium additive), and dilute sulfuric acid were mixed and stirred for slurry treatment. The molar ratio of hydrogen ions to lithium ions in the dilute sulfuric acid was 0.9:1. The aluminum-cerium mixed salt used was a combination of aluminum sulfate and cerium sulfate, with an aluminum to cerium molar ratio of 10:1 and a mass of 5% of the black powder mass. The reaction temperature was 90℃, the solid-liquid ratio was 400 g / L, and the reaction time was 1 h. After slurry treatment, the pH of the slurry was 3.2.

[0107] Step (2):

[0108] The slurry obtained in step (1) was transferred to an autoclave for pressurized acid leaching. The reaction temperature was 220℃ and the time was 3h, yielding a lithium leachate and delithiated slag. The pH of the leachate was 6.76, the lithium leaching rate was 99.1%, and the leaching rates of nickel, cobalt, manganese, and fluorine were only 1.6%, 1.9%, 2.3%, and 1.1%, respectively. Copper, iron, and aluminum were not leached, achieving highly selective extraction of lithium.

[0109] Step (3):

[0110] Granular activated carbon was added to the lithium leaching solution for degreasing, with a dosage of 20 g / L and a treatment time of 1 h. After degreasing, 10 mol / L liquid alkali (sodium hydroxide aqueous solution) was added to precipitate and remove heavy metals. The amount of alkali added was not less than the theoretical amount of nickel, cobalt, and manganese to be precipitated (e.g., considering cost, the amount of liquid alkali used was 10 mL / L), and the reaction time was 1 h. The lithium loss rate during the degreasing and heavy metal removal processes was only 0.9%, while the removal rates of nickel, cobalt, and manganese were all greater than 99.9%. The concentrations of nickel, cobalt, and manganese impurities in the purified lithium solution were all below 0.001 g / L, meeting the requirements for preparing battery-grade lithium carbonate. The small amount of precipitated residue generated was incorporated into the subsequent reduction acid leaching process to recover nickel, cobalt, and manganese.

[0111] Step (4):

[0112] The delithiation slag underwent a first-stage reducing acid leaching. The sulfuric acid concentration was 1.5 mol / L, the reaction temperature was 80℃, the solid-liquid ratio was 400 g / L, the time was 2 h, and the reducing agent was hydrogen peroxide. The molar ratio of the reducing agent to the total nickel, cobalt, and manganese in the delithiation slag was 0.75:1. The leaching rates of nickel, cobalt, manganese, and copper in the first stage were 81.2%, 79.3%, 78.8%, and 79.2%, respectively. The resulting first-stage slag was used as the raw material for the second stage leaching.

[0113] A second-stage reducing acid leaching was performed on the first-stage slag. The sulfuric acid concentration was 1.5 mol / L, the reaction temperature was 80℃, the solid-liquid ratio was 800 g / L, the time was 2 h, and the reducing agent was hydrogen peroxide. The molar ratio of the reducing agent to the total nickel, cobalt, and manganese in the delithiation slag was 0.5:1. The content of nickel, cobalt, manganese, and copper in the second-stage slag was less than 0.1%, and the final total leaching rate was greater than 99.9%.

[0114] After adding acid to the second-stage leaching solution, it was returned as the leaching agent for the first-stage leaching to treat the fresh delithiated slag. The sulfuric acid concentration after acid addition was 1.5 mol / L, the reaction temperature was 80℃, the solid-liquid ratio was 400 g / L, the time was 2 h, and the reducing agent was hydrogen peroxide. The molar ratio of the reducing agent to the total molar amount of nickel, cobalt, and manganese in the delithiated slag was 0.75:1. The leaching rates of nickel, cobalt, manganese, and copper in the first-stage leaching were 79.3%, 78.5%, 78.0%, and 78.4%, respectively, and the pH of the leaching solution was 3.5. This effect is consistent with that of directly using fresh sulfuric acid for the first-stage leaching. The resulting leaching solution was used for subsequent copper extraction, and the resulting first-stage slag continued to undergo the second-stage reducing acid leaching.

[0115] Step (5):

[0116] Copper extraction was performed on the reduced leachate. The extractant in the organic phase of the copper extraction was a mixture of 5-nonylsalicylaldehyde oxime and 2-hydroxy-5-nonylacetophenone oxime (volume ratio 1:1), with a total volume percentage of 20%. Sulfonated kerosene was used as the diluent. The extraction conditions were: three-stage countercurrent extraction, an extraction phase O / A ratio of 2:1, a time of 10 min, a temperature of 40℃, and an equilibrium pH of 1.8. The results showed that the copper extraction rate was 99.6%, while the co-extraction rates of nickel, cobalt, and manganese were only 0.08%, 0.09%, and 0.04%, respectively, achieving selective extraction and separation of copper.

[0117] Step (6):

[0118] Calcium salt was added to the copper extraction residue for alkali adjustment and preliminary impurity removal, yielding a pre-clean solution. The calcium salt was calcium hydroxide, added at a concentration of 10 g / L. The reaction temperature was 70℃, and the reaction time was 1 h. The final solution pH was 5.13, at which point the removal rates of iron, cerium, aluminum, fluorine, and phosphorus were 100%, 100%, 89.6%, 92.5%, and 97.3%, respectively. The co-precipitation rates of nickel, cobalt, and manganese were only 3.63%, 3.67%, and 0.72%, respectively, and these metals could be recovered by acid washing of the precipitate residue.

[0119] Step (7):

[0120] The initial purified solution (A / F ratio of 2:1 in the aqueous phase) underwent deep impurity removal. An organic phase containing a synergistic extractant was used, wherein compound 1 is (formula a1: Compound 1), with a volume percentage of 15%; Compound 2 is (Formula b1: The volume percentage of the organic phase was 30%; the diluent was sulfonated kerosene. Extraction conditions were: 4-stage countercurrent extraction, O / A ratio of 2:1, saponification rate of 40%, time 10 min, temperature 20℃, equilibrium pH of 3.2, at which point the Al / F ratio in the aqueous phase was approximately 2:1. Results showed that the simultaneous extraction rates of aluminum, calcium, and fluorine were greater than 99.9%, and the concentrations of aluminum, calcium, and fluorine were less than 0.002 g / L. After acid washing of the loaded organic phase, the co-extraction rates of nickel, cobalt, and manganese were only 0%, 0%, and 2.4%, respectively, achieving deep purification of the nickel-cobalt-manganese solution and obtaining a high-purity nickel-cobalt-manganese solution.

[0121] Example 2

[0122] Compared with Example 1, the only difference is that the amount of aluminum cerium additive added in step 1 is changed; other conditions are the same as in Example 1. The experimental group is as follows:

[0123] Group A: The amount of aluminum cerium additive added is set to 7.5% of the mass of black powder.

[0124] Group B: The amount of aluminum cerium additive added is set to 10% of the mass of black powder.

[0125] The results of step 2 are as follows:

[0126] Group A: The leaching rate of lithium was 99.1%, while the leaching rates of nickel, cobalt, manganese, and fluorine were only 2.0%, 2.1%, 2.4%, and 1.0%, respectively. Copper, iron, and aluminum were not leached. At this point, the preferential leaching rate of lithium remained basically unchanged, but the leaching rates of nickel, cobalt, and manganese all increased, and the leaching selectivity decreased.

[0127] Group B: The leaching rate of lithium was 99.3%, while the leaching rates of nickel, cobalt, manganese, and fluorine were 2.5%, 2.6%, 3.2%, and 0.7%, respectively. Copper, iron, and aluminum were not leached. At this point, the leaching rates of lithium, nickel, cobalt, and manganese all increased, but the leaching selectivity decreased.

[0128] Example 3

[0129] Compared with Example 1, the only difference is that the ratio of aluminum to cerium in the aluminum-cerium additive in step 1 is changed. The total amount of aluminum-cerium additive and other operating conditions are the same as in Example 1. The experimental group is:

[0130] Group A: The molar ratio of aluminum to cerium in the aluminum-cerium additive is 1:1.

[0131] Group B: The molar ratio of aluminum to cerium in the aluminum-cerium additive is 1:10.

[0132] The results of step 2 are as follows:

[0133] Group A: The leaching rate of lithium was 99.0%, while the leaching rates of nickel, cobalt, manganese, and fluorine were only 1.5%, 1.8%, 2.2%, and 1.3%, respectively. Copper, iron, and aluminum were not leached. At this point, the leaching rates of lithium, nickel, cobalt, and manganese did not change significantly.

[0134] Group B: The leaching rate of lithium was 98.1%, while the leaching rates of nickel, cobalt, manganese, and fluorine were 1.2%, 1.4%, 1.3%, and 1.4%, respectively. Copper, iron, and aluminum were not leached. At this point, the leaching rates of lithium, nickel, cobalt, and manganese all decreased to some extent.

[0135] Example 4

[0136] Compared to Example 1, the only difference is that the conditions in steps 1 and 2 are changed. The experimental group is as follows:

[0137] Group A: The molar ratio of hydrogen ions in dilute sulfuric acid to lithium ions in black powder is set to 0.8:1; in step 2, the reaction temperature is 240℃;

[0138] Group B: The molar ratio of hydrogen ions in dilute sulfuric acid to lithium ions in black powder is set to 1.2:1; in step 2, the reaction temperature is 200℃;

[0139] The results of step 2 are as follows:

[0140] Group A: The leaching rate of lithium was 92.1%, while the leaching rates of nickel, cobalt, manganese and fluorine were only 1.1%, 1.2%, 1.9% and 1.3% respectively. Copper, iron and aluminum were not leached. Although the preferential leaching rate of lithium decreased, the leaching selectivity was improved.

[0141] Group B: The leaching rate of lithium was 99.5%, while the leaching rates of nickel, cobalt, manganese, and fluorine were 3.5%, 3.4%, 4.9%, and 1.0%, respectively. Copper, iron, and aluminum were not leached. At this point, the leaching rates of lithium, nickel, cobalt, and manganese were all improved.

[0142] Example 5

[0143] Compared to Example 1, the conditions in step 7 were changed, and the experimental groups were as follows:

[0144] Group A: Change compound 1 to (formula a2: ), change compound 2 to (formula b2: );

[0145] Group B: Control the initial purified liquid; change the volume percentage of compound 1 to 10% and the volume percentage of compound 2 to 25% in the extracted organic phase; change the equilibrium pH to 3.0 and the temperature to 15℃.

[0146] The results of step 7 are as follows:

[0147] Group A: The simultaneous extraction rates of aluminum, calcium, and fluorine were 98.1%, 96.6%, and 97.9%, respectively. After acid washing of the supported organic phase, the co-extraction rates of nickel, cobalt, and manganese were 0%, 0%, and 4.2%, respectively.

[0148] Group B: The simultaneous extraction rates of aluminum, calcium, and fluorine were 97.1%, 99%, and 97.1%, respectively. After acid washing of the supported organic phase, the co-extraction rates of nickel, cobalt, and manganese were 0%, 0%, and 2.1%, respectively.

[0149] Comparative Example 1

[0150] Compared to Example 1, the only difference is that the conditions in step 1 are changed, and the experimental groups are as follows:

[0151] Group A: No aluminum cerium additive was added; all other operations and parameters were the same as in Example 1.

[0152] Group B: The aluminum cerium additive was replaced with aluminum sulfate alone, with the same amount added as the aluminum cerium additive, and other operations and parameters were the same as in Example 1;

[0153] Group C: The aluminum cerium additive was replaced with a single cerium sulfate, with the same amount added as the aluminum cerium additive, and other operations and parameters were the same as in Example 1;

[0154] The results of step 2 are as follows:

[0155] Group A: The leaching rate of lithium was 85.12%, while the leaching rates of nickel, cobalt, manganese, and fluorine were 1.5%, 1.7%, 2.1%, and 5.4%, respectively. Copper, iron, and aluminum were not leached. At this point, the preferential leaching rate of lithium decreased significantly.

[0156] Group B: The leaching rate of lithium was 96.5%, while the leaching rates of nickel, cobalt, manganese and fluorine were only 1.8%, 2.2%, 2.5% and 1.1% respectively. Copper, iron and aluminum were not leached. At this point, not only did the preferential leaching rate of lithium decrease, but the leaching rates of nickel, cobalt and manganese also increased, and the leaching selectivity decreased significantly.

[0157] Group C: The leaching rate of lithium was 96.8%, while the leaching rates of nickel, cobalt, manganese, and fluorine were only 1.1%, 1.2%, 1.2%, and 1.3%, respectively. Copper, iron, and aluminum were not leached. At this point, the preferential leaching rate of lithium decreased.

[0158] Comparative Example 2

[0159] Compared with Example 1, the only difference is that in the extractant of step 7, compound 1 (formula 1a) is missing, and the missing amount is made up by equal volume of formula b1. All other operations and parameters are the same as in Example 1.

[0160] The result of step 7 is:

[0161] The simultaneous extraction rates of aluminum, calcium, fluorine, nickel, cobalt, and manganese were all below 0.1%. This indicates that compound 2 alone has virtually no extraction effect on the elements in the solution, and it needs to be used in conjunction with compound 1 to achieve the desired impurity removal effect.

[0162] Comparative Example 3

[0163] Compared with Example 1, the only difference is that in the extractant of step 7, compound 2 (formula b1) is missing, and the missing amount is made up by equal volume of formula 1a. All other operations and parameters are the same as in Example 1.

[0164] The result of step 7 is:

[0165] The simultaneous extraction rates of aluminum, calcium, and fluorine were 96.3%, 95.2%, and 96.2%, respectively. After acid washing of the supported organic phase, the co-extraction rates of nickel, cobalt, and manganese were 2.6%, 2.1%, and 15.3%, respectively. This indicates that the reduction of compound 2 is detrimental to the extraction and separation of impurities and valuable components.

[0166] Comparative Example 4

[0167] Compared with Example 1, the only difference is that aluminum cerium additive was not added in step 1, while other operations and parameters are the same as in Example 1.

[0168] The result of step 7 is:

[0169] The Al / F ratio in the solution before extraction was approximately 0.5. The simultaneous extraction rates of aluminum, calcium, and fluorine were 82.2%, 93.3%, and 75.3%, respectively. This indicates that the lack of aluminum cerium additives resulted in a relatively low Al / F ratio, making it impossible to achieve deep extraction of aluminum and fluorine.

[0170] Comparative Example 5

[0171] Compared to Example 1, the conditions in step 7 were changed, and the experimental groups were as follows:

[0172] Group A: The total content of compound 1 and compound 2 in the extracted organic phase is the same as in Example 1, except that the volume ratio of compound 1 to compound 2 is 1:1;

[0173] Group B: Change the equilibrium pH to 5.0;

[0174] The results of step 7 are as follows:

[0175] Group A: The simultaneous extraction rates of aluminum, calcium, and fluorine were 35.1%, 35.7%, and 36.5%, respectively, indicating poor extraction and impurity removal effects.

[0176] Group B: The simultaneous extraction rates of aluminum, calcium, and fluorine were 97.1%, 99%, and 97.1%, respectively. After acid washing of the supported organic phase, the co-extraction rates of nickel, cobalt, and manganese were 1.2%, 1.5%, and 16.3%, respectively. At this point, there was a significant loss of nickel, cobalt, and manganese during co-extraction.

Claims

1. A method for recycling black powder from waste lithium-ion batteries, characterized by the following steps: include: Step a: Waste lithium-ion battery black powder and aluminum-cerium additives are mixed and subjected to pressurized acid leaching, followed by separation to obtain lithium leachate and delithiation slag enriched with transition metals from the black powder and graphite; the aluminum-cerium additives contain aluminum source and cerium source. Step b: The lithium-depleted slag is subjected to reduction acid leaching treatment, and solid-liquid separation is performed to obtain graphite and acid leaching solution enriched with transition metals. Step c: The acid leaching solution is subjected to a first-stage extraction to remove copper, resulting in a first-stage purified solution. Step d: The first-stage impurity removal solution and calcium impurity removal agent are subjected to a second-stage impurity removal treatment to obtain a second-stage impurity removal solution; Step e: The second-stage impurity removal solution was subjected to a third-stage extraction using a synergistic extractant to obtain a purified third-stage impurity removal solution enriched with transition metal elements. The extractable active ingredient in the extracted organic phase comprises extractant A and extractant B in a volume ratio of 1:2~5; wherein extractant A is a compound with structure of formula 1; and extractant B is a compound with structure of formula 2. Formula 1; Formula 2; In Formula 1, R1 is a C1~C8 alkyl group, C4~C6 ... 10 The alkoxy or phenoxy group; R2 is H, C1~C8 alkyl, C4~C 10 The alkoxy or phenoxy group; M is H, Na, K or NH4; In Formula 2, R3 and R4 are individually C1~C8 alkyl groups and C4~C6 alkyl groups. 10 The alkoxy or phenoxy groups are used; the temperature during the extraction process is below 45℃; the pH during the extraction process is 2.0~4.

0.

2. The method for recycling waste lithium-ion battery black powder as described in claim 1, characterized in that, The transition metal element in the waste lithium-ion battery black powder includes at least one of nickel, cobalt, and manganese; Preferably, in the waste lithium-ion battery black powder, the content of transition metals is 5-35%; the content of lithium is 3-5%; the content of at least one anionic impurity among F and P is 3-6%; and the content of at least one metallic impurity among Al and Cu is 0.5-2.5%.

3. The method for recycling waste lithium-ion battery black powder as described in claim 1, characterized in that, In aluminum-cerium additives, the aluminum source includes at least one of aluminum sulfate, aluminum hydroxide, aluminum oxide, aluminum chloride, and aluminum nitrate; the cerium source includes a combination of at least one of cerium sulfate and cerium hydroxide. Preferably, in the aluminum-cerium additive, the molar ratio of aluminum to cerium is 1~15:1~10; Preferably, the aluminum-cerium additive is 3-15 wt% of the waste lithium-ion battery black powder.

4. The method for recycling waste lithium-ion battery black powder as described in claim 1, characterized in that, The acid used in the pressure pickling process is an inorganic acid, preferably sulfuric acid; Preferably, the concentration of acid in the pressure pickling starting slurry is 0.4~0.6M; Preferably, in pressurized acid leaching, the H+ of the acid... + The molar ratio of Li in the waste lithium-ion battery black powder is 0.8~1.3:1; Preferably, the temperature of the pressurized acid leaching is above 180°C; Preferably, the pressure pickling time is more than 1 hour.

5. The method for recycling waste lithium-ion battery black powder as described in claim 1, characterized in that, It also includes the steps of removing impurities from lithium leaching solution and preparing lithium carbonate by carbonation precipitation; The impurity removal process includes oil removal and weight removal.

6. The method for recycling waste lithium-ion battery black powder as described in claim 1, characterized in that, The acid used in the reducing acid leaching process is an inorganic acid, preferably sulfuric acid; Preferably, the concentration of acid in the initial reducing acid leaching solution is 1~5M; Preferably, the reducing agent added during the reducing acid leaching process includes one of hydrogen peroxide, sodium sulfite, sodium bisulfite, and sulfurous acid; Preferably, the ratio of the reducing agent to the total molar amount of transition metals in the delithiation slag is 0.5~1.5:1; Preferably, the reduction acid leaching reaction temperature is 70~90℃, the solid-liquid ratio is 200~600 g / L, and the reaction time is 0.5~3 h.

7. The method for recycling waste lithium-ion battery black powder as described in claim 1, characterized in that, The extractant used for copper removal in the first stage is extractant of formula 3. Formula 3; In Formula 3, R5 and R6 are individually H, hydroxyl, alkyl, or alkoxy; R7 is H or a C1-C4 alkyl group; and M is H, Na, K, or NH4. During the first stage of copper removal extraction, the volume percentage concentration of Formula 1 in the organic phase is 5%~30%; Preferably, the O / A volume ratio in the first stage of copper removal extraction is 1:2 to 0.

5.

8. The method for recycling waste lithium-ion battery black powder as described in claim 1, characterized in that, Calcium impurity removers include at least one of calcium hydroxide, calcium oxide, and calcium carbonate; Based on the first stage of impurity removal solution, the amount of calcium impurity removal agent added is 5~15 g / L; Preferably, the temperature of the second stage of impurity removal is 70~90℃, and the reaction time is 0.5~1 h.

9. The method for recycling waste lithium-ion battery black powder as described in claim 1, characterized in that, In step e, the volume percentage concentration of the active ingredient in the synergistic extractant is 30-60%.

Images