Method for recovering valuable elements from nickel-cobalt-manganese slag containing lithium

Abstract

The invention belongs to the technical field of smelting recovery, and particularly relates to a method for recovering valuable elements from lithium-containing nickel-cobalt-manganese slag, which comprises the following steps: adding water into nickel-cobalt-manganese slag containing hydroxide of nickel, cobalt and manganese and lithium to prepare first slurry, leaching at the temperature of more than or equal to 140 DEG C and the oxygen partial pressure of more than or equal to 0.1 MPa, and carrying out solid-liquid separation to obtain a lithium-containing solution and first oxidizing slag; mixing the nickel-cobalt-manganese slag and the lithium-containing solution to obtain second slurry, leaching at the temperature of more than or equal to 140 DEG C and the oxygen partial pressure of more than or equal to 0.1 MPa, and performing solid-liquid separation to obtain a lithium-rich solution and second oxidizing slag; and after two times of oxygen pressure leaching, the leaching rate of lithium is greater than or equal to 99%. According to the technical scheme, the adsorbability of the nickel-cobalt-manganese slag to lithium is reduced through oxygen pressure leaching, release and recovery of lithium are achieved, and no high-temperature decomposed waste gas pollutes the environment; no acid chemical reagent is added in the leaching process, and no extra salt-containing wastewater is generated; the concentration of lithium in the leachate is improved through cyclic leaching, the concentration cost of lithium is reduced, and the yield of lithium is improved; and lithium is recycled through oxygen pressure leaching and resin impurity removal, and the lithium recycling process is short.

Description

Technical Field

[0001] This application belongs to the field of smelting and recycling technology, specifically a method for recovering valuable elements from lithium-containing nickel-cobalt-manganese slag. Background Technology

[0002] With the rapid development of the new energy vehicle industry, related industries have also entered a period of high-speed growth. Lithium, cobalt, nickel, manganese, and other valuable resources are key raw materials for the new energy vehicle industry, and domestic demand continues to rise. Domestic resource supply is insufficient, leading to a high dependence on imports. Therefore, recovering valuable resources from lithium-containing nickel-cobalt-manganese slag is extremely important, as it helps ensure resource supply and create economic benefits.

[0003] Currently, methods for recovering lithium from nickel-cobalt-manganese-lithium waste generally employ pyrometallurgical reduction decomposition and wet inorganic acid decomposition to achieve lithium leaching and separation. Patent CN107994288B discloses a method for recovering valuable metals from the positive electrode material of waste nickel-cobalt-manganese-lithium ternary batteries. This method involves uniformly mixing the disassembled and ground nickel-cobalt-manganese-lithium ternary battery positive electrode material with carbon powder, followed by calcination and reduction at a controlled temperature of 700-950℃. The calcined material is placed in a stirring device, pure water is added, and dilute acid is added dropwise to adjust the pH to 4.5-8. After soaking, the mixture is filtered. The filtrate is collected, and the pH is adjusted to 7.0-10.0 with sodium hydroxide. After filtration to remove impurities, soluble carbonate is added to precipitate lithium carbonate. The lithium carbonate precipitate is then filtered and washed, thus achieving the recovery of lithium metal. The pyrometallurgical reduction decomposition method suffers from high temperature, high energy consumption, and waste gas problems; the wet inorganic acid decomposition method suffers from acid consumption and the generation of saline wastewater; and the lithium recovery process is lengthy. Furthermore, after removing impurities from the lithium-containing filtrate with sodium hydroxide, the method yields a lithium-containing liquid and impurity-removed nickel-cobalt-manganese slag, but the treatment of the nickel-cobalt-manganese slag is not described.

[0004] The article "Process Study on Preparation of Li-ion Battery Ternary Cathode Material Precursor by Co-precipitation Method" indicates that the nickel-cobalt-manganese precipitate formed in air is thin-film, loosely packed on the surface, and has large pores. Phase analysis shows that the charge imbalance and increased interlayer spacing in this crystal structure may lead to the intercalation of impurities. When removing nickel-cobalt-manganese from lithium-containing solutions using sodium hydroxide neutralization precipitation, some lithium is lost and enters the nickel-cobalt-manganese slag. The main reasons for this lithium loss are: when nickel, cobalt, and manganese ions react with sodium hydroxide to form hydroxide precipitates, the rapid formation and growth of the precipitates causes lithium ions in the solution to be encapsulated by the rapidly growing precipitate colloids, resulting in physical entrainment loss of lithium ions in the solution; the newly formed nickel-cobalt-manganese hydroxide colloids have a large specific surface area, and lithium ions in the solution easily form strong electrostatic adsorption on the surface of the hydroxide colloids, causing lithium ions with smaller ionic radii to be firmly adsorbed in the layered structure of the nickel-cobalt-manganese hydroxide, resulting in chemical loss of lithium ions in the solution. Conventional washing methods are insufficient to disrupt the layered structure of nickel-cobalt-manganese hydroxide, thus failing to effectively release the adsorbed and trapped lithium ions. Therefore, there is an urgent need to develop a highly efficient method for recovering lithium from nickel-cobalt-manganese slag. Summary of the Invention

[0005] In response to the problems of high energy consumption, waste gas and wastewater generation, long lithium recovery process, and lithium loss due to the adsorption of some lithium by nickel cobalt manganese hydroxide formed by adding sodium hydroxide, this application provides a method for recovering valuable elements from lithium-containing nickel cobalt manganese slag. This nickel cobalt manganese slag is a neutralized slag produced in the lithium solution precipitation and impurity removal process, and its main chemical components are nickel, cobalt, manganese and lithium.

[0006] This application discloses a method for recovering valuable elements from lithium-containing nickel-cobalt-manganese slag, comprising the following steps: S1. Obtain nickel-cobalt-manganese slag, wherein the nickel-cobalt-manganese slag includes hydroxides of nickel, cobalt, and manganese and lithium. S2. The nickel-cobalt-manganese slag is mixed with water to form a first slurry, and a first leaching is carried out at a temperature ≥140℃ and an oxygen partial pressure ≥0.1MPa. Then, solid-liquid separation is performed to obtain a lithium-containing solution and a first oxide slag. S3. The nickel-cobalt-manganese slag and the lithium-containing solution are mixed to form a second slurry. A second leaching is carried out at a temperature ≥140℃ and an oxygen partial pressure ≥0.1MPa. Then, solid-liquid separation is performed to obtain a lithium-rich solution and a second oxide slag.

[0007] In the above technical solution, nickel-cobalt-manganese slag, comprising nickel, cobalt, and manganese hydroxides and lithium, undergoes two separate oxygen pressure leaching processes to enrich lithium in the solution. During leaching, under oxygen-rich conditions, nickel, cobalt, and manganese hydroxides (such as Ni(OH)₂, Co(OH)₂, and Mn(OH)₂) are transformed into oxides / hydroxyoxides of nickel, cobalt, and manganese (such as NiOOH, MnOOH, and Co₂Mn₃O₈) under the action of oxygen, thereby disrupting the structure of the nickel, cobalt, and manganese hydroxide colloids, reducing the adsorption of lithium, and releasing the lithium trapped within. Therefore, lithium can be effectively leached into the solution. When the temperature is below 140℃ or the oxygen partial pressure is less than 0.1 MPa, the lithium leaching rate is low. Therefore, this application controls the temperature to be ≥140℃ and the oxygen partial pressure to be ≥0.1 MPa during leaching to maintain efficient lithium leaching. In addition, in step S3, the lithium-containing solution is used to leach the nickel-cobalt-manganese slag in a cycle, which can enrich the lithium element in the solution, reduce the amount of water used, and make the subsequent purification and recycling processes more efficient.

[0008] Furthermore, the concentration of lithium in the lithium-rich solution is ≥15g / L.

[0009] Furthermore, the nickel content in the nickel-cobalt-manganese slag is 18.0wt%~22.0wt%, the cobalt content is 13.0wt%~17.0wt%, the manganese content is 20.0wt%~25.0wt%, and the lithium content is 4.5wt%~5.5wt%.

[0010] Furthermore, the pH of the first slurry is 8.0~10.0.

[0011] Furthermore, the liquid-to-solid ratio of the first slurry is (4~6):1mL / g, and the liquid-to-solid ratio of the second slurry is (4~6):1mL / g.

[0012] Furthermore, in step S2, the temperature for the first leaching is 140~170℃ and the oxygen partial pressure is 0.1~0.5MPa; In step S3, the temperature for the second leaching is 140~170℃ and the oxygen partial pressure is 0.1~0.5MPa; When the temperature is above 170℃ or the oxygen partial pressure is greater than 0.5MPa, the effect on improving the lithium leaching rate is not significant. Controlling the leaching temperature to 140~170℃ and the oxygen partial pressure to 0.1~0.5MPa can ensure a high lithium leaching rate while reducing unnecessary energy consumption.

[0013] Furthermore, in step S2, the first leaching is performed simultaneously with the first stirring, the first leaching time is 1~3 hours, and the first stirring speed is 300~450 rpm. In step S3, the second leaching is performed simultaneously with the second stirring. The second leaching time is 1-3 hours, and the stirring speed is 300-450 rpm.

[0014] Furthermore, the lithium element in the nickel-cobalt-manganese slag, the first oxide slag, and the second oxide slag was detected and analyzed, and the lithium leaching rate was calculated to be ≥99%.

[0015] Furthermore, after step S3, the following is also included: S4. The lithium-rich solution is purified using resin, and then a precipitant is added to the purified lithium-rich solution and solid-liquid separation is performed to obtain lithium carbonate. S5. The first oxide residue and the second oxide residue are acid-leached to obtain an acid leaching solution. Then, the acid leaching solution is purified to obtain a purified solution. The purified solution includes nickel, cobalt and manganese. The purified solution is extracted to separate and recover nickel, cobalt and manganese, or the content of nickel, cobalt and manganese in the purified solution is adjusted to prepare a nickel-cobalt-manganese precursor.

[0016] Furthermore, the resin includes a macroporous chelating resin, and during the purification process, the flow rate of the lithium-rich solution is 2~4 BV / h.

[0017] Furthermore, the precipitant includes sodium carbonate, and the molar ratio of sodium carbonate to lithium in the purified lithium-rich solution is 0.525~0.575, and the temperature at which the precipitant is added is 90~97℃.

[0018] Furthermore, in step S5, the impurity removal includes: Ni(OH)2, Co(OH)2 and / or Mn(OH)2 are added to the acid leaching solution to adjust the pH to 3.0~4.0, and iron is precipitated to remove iron elements, resulting in an iron removal solution; then, calcium is removed from the iron removal solution by five-stage countercurrent extraction under the conditions of O / A=2:1 and 5 min, to obtain the purified solution.

[0019] This application provides a method for recovering valuable elements from lithium-containing nickel-cobalt-manganese slag, which produces the following beneficial effects: oxygen pressure leaching reduces the adsorption of lithium by the nickel-cobalt-manganese slag, thereby releasing and recovering lithium without high-temperature decomposition waste gas pollution; no acid chemical reagents are added during the leaching process, and no additional saline wastewater is generated; the lithium concentration in the leachate is increased through cyclic leaching, reducing the lithium concentration cost and increasing the lithium yield; and lithium is recovered through oxygen pressure leaching and resin impurity removal, resulting in a simplified lithium recovery process. Attached Figure Description

[0020] To more clearly illustrate the technical solutions in the embodiments of this application or the prior art, the drawings used in the description of the embodiments or the prior art will be briefly introduced below. Obviously, the drawings described below are only some embodiments of this application. For those skilled in the art, other drawings can be obtained based on the structures shown in these drawings without creative effort.

[0021] Figure 1 This is the XRD pattern of the nickel-cobalt-manganese slag in Example 1 of this application.

[0022] Figure 2 This is the XRD pattern of the first oxidized slag in Example 1 of this application.

[0023] The realization of the purpose, functional features and advantages of this application will be further explained in conjunction with the embodiments and with reference to the accompanying drawings. Detailed Implementation

[0024] The technical solutions in the embodiments will be clearly and completely described below. Obviously, the described embodiments are only some embodiments of this application, and not all embodiments. Based on the embodiments in this application, all other embodiments obtained by those of ordinary skill in the art without creative effort are within the scope of protection of this application.

[0025] This application discloses a method for recovering valuable elements from lithium-containing nickel-cobalt-manganese slag, comprising the following steps: S1. Obtain nickel-cobalt-manganese slag, which includes hydroxides of nickel, cobalt, and manganese and lithium. Preferably, the nickel content in the nickel-cobalt-manganese slag is 18.0 wt%~22.0 wt%, the cobalt content is 13.0 wt%~17.0 wt%, the manganese content is 20.0 wt%~25.0 wt%, and the lithium content is 4.5 wt%~5.5 wt%. Specifically, the nickel content can be any one or a range between any two of the following: 18.0 wt%, 19.0 wt%, 20.0 wt%, 21.0 wt%, and 22.0 wt%; the cobalt content can be any one or a range between any two of the following: 13.0 wt%, 14.0 wt%, 15.0 wt%, 16.0 wt%, and 17.0 wt%; the manganese content can be any one or a range between any two of the following: 20.0 wt%, 21.0 wt%, 22.0 wt%, 23.0 wt%, 24.0 wt%, and 25.0 wt%; and the lithium content can be any one or a range between any two of the following: 4.5 wt%, 4.8 wt%, 5.0 wt%, 5.2 wt%, and 5.5 wt%.

[0026] S2. Add water to nickel-cobalt-manganese slag to prepare the first slurry, and carry out the first leaching at a temperature ≥140℃ and an oxygen partial pressure ≥0.1MPa. Then, separate the solid and liquid to obtain a lithium-containing solution and the first oxide slag. Preferably, the temperature for the first leaching is 140~170℃ and the oxygen partial pressure is 0.1~0.5MPa; the first leaching is carried out simultaneously with the first stirring, the first leaching time is 1~3h, the first stirring speed is 300~450rpm; the pH of the first slurry is 8.0~10.0; and the liquid-to-solid ratio of the first slurry is (4~6):1mL / g. Specifically, the temperature for the first leaching can be any one or a range between two of 140℃, 145℃, 150℃, 155℃, 160℃, 165℃, and 170℃; the oxygen partial pressure can be any one or a range between two of 0.1MPa, 0.2MPa, 0.3MPa, 0.4MPa, and 0.5MPa; and the leaching time can be any one or a range between two of 1h, 1.5h, 2h, 2.5h, and 3h. The stirring speed can be any one of 300 rpm, 350 rpm, 400 rpm, 450 rpm, or any combination thereof; the pH of the first slurry can be any one of 8.0, 8.5, 9.0, 9.5, 10.0, or any combination thereof; the liquid-to-solid ratio of the first slurry can be any one of 4:1 mL / g, 4.5:1 mL / g, 5:1 mL / g, 5.5:1 mL / g, 6:1 mL / g, or any combination thereof.

[0027] S3. Mix nickel-cobalt-manganese slag and lithium-containing solution to form a second slurry. Perform a second leaching at a temperature ≥140℃ and an oxygen partial pressure ≥0.1MPa. Then, separate the solid and liquid to obtain a lithium-rich solution and a second oxide slag.

[0028] Preferably, the temperature for the second leaching is 140~170℃ and the oxygen partial pressure is 0.1~0.5MPa; the second leaching is performed simultaneously with the second stirring, the second leaching time is 1~3h, and the stirring speed is 300~450rpm; the liquid-to-solid ratio of the second slurry is (4~6):1mL / g; the lithium concentration in the lithium-rich solution is ≥15g / L; the lithium element in the nickel-cobalt-manganese slag, the first oxide slag, and the second oxide slag is detected and analyzed, and the lithium leaching rate is calculated to be ≥99%, calculated as follows: Where m0 is the total mass of the nickel-cobalt-manganese slag, g; w0 is the mass fraction of lithium in the nickel-cobalt-manganese slag, %; m1 is the mass of the first oxide slag, g; w1 is the mass fraction of lithium in the first oxide slag, %; m2 is the mass of the second oxide slag, g; w2 is the mass fraction of lithium in the second oxide slag, % Specifically, the temperature for the second leaching can be any one or a range between 140℃, 145℃, 150℃, 155℃, 160℃, 165℃, and 170℃; the oxygen partial pressure can be any one or a range between 0.1MPa, 0.2MPa, 0.3MPa, 0.4MPa, and 0.5MPa; the second leaching time can be any one or a range between 1h, 1.5h, 2h, 2.5h, and 3h; the second stirring speed can be any one or a range between 300rpm, 350rpm, 400rpm, and 450rpm; and the liquid-to-solid ratio of the second slurry can be any one or a range between 4:1mL / g, 4.5:1mL / g, 5:1mL / g, 5.5:1mL / g, and 6:1mL / g.

[0029] Preferably, after step S3, the method further includes: S4. Use resin to purify the lithium-rich solution, then add a precipitant to the purified lithium-rich solution and perform solid-liquid separation to obtain lithium carbonate. S5. The first and second oxide residues are acid-leached to obtain an acid leaching solution. Then, impurities are removed from the acid leaching solution to obtain a purified solution containing nickel, cobalt, and manganese. The purified solution is extracted to separate and recover nickel, cobalt, and manganese, or the content of nickel, cobalt, and manganese in the purified solution is adjusted to prepare a nickel-cobalt-manganese precursor.

[0030] Preferably, the resin includes a macroporous chelating resin, and the flow rate of the lithium-rich solution during purification is 2~4 BV / h.

[0031] Preferably, the precipitant includes sodium carbonate, the molar ratio of sodium carbonate to lithium in the purified lithium-rich solution is 0.525~0.575, and the temperature at which the precipitant is added is 90~97℃.

[0032] Preferably, in step S5, the impurity removal includes: Ni(OH)2, Co(OH)2 and / or Mn(OH)2 are added to the acid leaching solution to adjust the pH to 3.0~4.0, and iron is precipitated to remove iron elements, resulting in an iron removal solution; then, calcium is removed from the iron removal solution by five-stage countercurrent extraction under the conditions of O / A=2:1 and 5 min, to obtain a purified solution.

[0033] The technical solution of this application will be further described below with reference to specific embodiments. The lithium-containing nickel-cobalt-manganese slag used in the following embodiments and comparative examples is a neutralized slag produced by the lithium-containing element solution precipitation and impurity removal process. Its main chemical components include Ni, Co, Mn and Li, with contents of Ni: 20.50wt%, Co: 15.10wt%, Mn: 24.70wt%, and Li: 4.78wt%, respectively.

[0034] Example 1 In a pressure reactor, nickel-cobalt-manganese slag was mixed with water to form a first slurry with a liquid-to-solid ratio of 6:1 mL / g. Oxygen was introduced and the mixture was stirred, with the oxygen partial pressure controlled at 0.3 MPa, the temperature at 170℃, and the rotation speed at 450 rpm. After leaching for 1 hour, solid-liquid separation was performed to obtain a lithium-containing solution and a first oxide slag. In the same pressure reactor, the nickel-cobalt-manganese slag and the lithium-containing solution were mixed to form a second slurry with a liquid-to-solid ratio of 6:1 mL / g. Oxygen was introduced and the mixture was stirred, with the oxygen partial pressure controlled at 0.3 MPa, the temperature at 170℃, and the rotation speed at 450 rpm. After leaching for 1 hour, solid-liquid separation was performed to obtain a lithium-rich solution with a lithium element concentration of 15.9 g / L and a second oxide slag. The calculated lithium leaching rate was 99.8%.

[0035] The lithium-rich solution was purified by passing it through a macroporous chelating resin at a flow rate of 2 BV / h. Sodium carbonate was then added to the purified lithium-rich solution at 95°C, with a molar ratio of sodium carbonate to lithium in the purified solution of 0.55. Solid-liquid separation yielded lithium carbonate. The first and second oxide residues were leached with sulfuric acid to obtain an acid leaching solution. Ni(OH)₂, Co(OH)₂, and / or Mn(OH)₂ were added to the acid leaching solution to adjust the pH to 3.5, precipitating and removing iron to obtain an iron-removing solution. Calcium was then removed from the iron-removing solution using a calcium extraction agent at a ratio of O / A = 2:1 and a time of 5 min through a 5-stage countercurrent extraction process, yielding a purified solution. The purified solution was then extracted to separate and recover nickel, cobalt, and manganese, or the nickel, cobalt, and manganese content in the purified solution was adjusted to prepare nickel-cobalt-manganese precursors.

[0036] The XRD pattern of the nickel-cobalt-manganese slag used in this embodiment is as follows: Figure 1 As shown, the XRD pattern of the obtained first oxidation slag is as follows. Figure 2 As shown, by Figure 1 It can be seen that nickel, cobalt, and manganese in the nickel-cobalt-manganese slag exist in the forms of Ni(OH)₂·2H₂O, Co(OH)₂, and Mn(OH)₂, respectively. After oxygen pressure leaching, under the combined action of oxygen and temperature, the structure of nickel, cobalt, and manganese hydroxides is destroyed and transformed into... Figure 2 Nickel, cobalt, and manganese are present in oxide / hydroxy oxide (NiOOH, MnOOH, and Co2Mn3O8) forms.

[0037] Example 2 In a pressure reactor, nickel-cobalt-manganese slag was mixed with water to form a first slurry with a liquid-to-solid ratio of 5:1 mL / g. Oxygen was introduced and the mixture was stirred, with the oxygen partial pressure controlled at 0.1 MPa, the temperature at 160℃, and the rotation speed at 300 rpm. After leaching for 2 hours, solid-liquid separation was performed to obtain a lithium-containing solution and a first oxide slag. In the same pressure reactor, the nickel-cobalt-manganese slag and the lithium-containing solution were mixed to form a second slurry with a liquid-to-solid ratio of 5:1 mL / g. Oxygen was introduced and the mixture was stirred, with the oxygen partial pressure controlled at 0.1 MPa, the temperature at 160℃, and the rotation speed at 300 rpm. After leaching for 2 hours, solid-liquid separation was performed to obtain a lithium-rich solution with a lithium element concentration of 19.0 g / L and a second oxide slag. The calculated lithium leaching rate was 99.4%.

[0038] The lithium-rich solution was purified by passing it through a macroporous chelating resin at a flow rate of 3 BV / h. Sodium carbonate was then added to the purified lithium-rich solution at 95°C, with a molar ratio of sodium carbonate to lithium in the purified solution of 0.55. Solid-liquid separation yielded lithium carbonate. The first and second oxide residues were leached with sulfuric acid to obtain an acid leaching solution. Ni(OH)₂, Co(OH)₂, and / or Mn(OH)₂ were added to the acid leaching solution to adjust the pH to 3.0, precipitating and removing iron to obtain an iron-removing solution. Calcium was then removed from the iron-removing solution using a calcium extraction agent at a ratio of O / A = 2:1 and a time of 5 min through a 5-stage countercurrent extraction process, yielding a purified solution. The purified solution was then extracted to separate and recover nickel, cobalt, and manganese, or the nickel, cobalt, and manganese content in the purified solution was adjusted to prepare nickel-cobalt-manganese precursors.

[0039] Example 3 In a pressure reactor, nickel-cobalt-manganese slag was mixed with water to form a first slurry with a liquid-to-solid ratio of 4:1 mL / g. Oxygen was introduced and the mixture was stirred. The oxygen partial pressure was controlled at 0.5 MPa, the temperature at 140℃, and the rotation speed at 400 rpm. After leaching for 3 hours, solid-liquid separation was performed to obtain a lithium-containing solution and a first oxide slag. In the same pressure reactor, nickel-cobalt-manganese slag and the lithium-containing solution were mixed to form a second slurry with a liquid-to-solid ratio of 4:1 mL / g. Oxygen was introduced and the mixture was stirred. The oxygen partial pressure was controlled at 0.5 MPa, the temperature at 140℃, and the rotation speed at 400 rpm. After leaching for 3 hours, solid-liquid separation was performed to obtain a lithium-rich solution with a lithium element concentration of 23.8 g / L and a second oxide slag. The calculated lithium leaching rate was 99.6%.

[0040] The lithium-rich solution was purified by passing it through a macroporous chelating resin at a flow rate of 3 BV / h. Sodium carbonate was then added to the purified lithium-rich solution at 95°C, with a molar ratio of sodium carbonate to lithium in the purified solution of 0.55. Solid-liquid separation yielded lithium carbonate. The first and second oxide residues were leached with sulfuric acid to obtain an acid leaching solution. Ni(OH)₂, Co(OH)₂, and / or Mn(OH)₂ were added to the acid leaching solution to adjust the pH to 4.0, precipitating and removing iron to obtain an iron-removing solution. Calcium was then removed from the iron-removing solution using a calcium extraction agent at a ratio of O / A = 2:1 and a time of 5 min through a 5-stage countercurrent extraction process, yielding a purified solution. The purified solution was then extracted to separate and recover nickel, cobalt, and manganese, or the nickel, cobalt, and manganese content in the purified solution was adjusted to prepare nickel-cobalt-manganese precursors.

[0041] Example 4 In a pressure reactor, nickel-cobalt-manganese slag was mixed with water to form a first slurry with a liquid-to-solid ratio of 6:1 mL / g. Oxygen was introduced and the mixture was stirred, with the oxygen partial pressure controlled at 0.6 MPa, the temperature at 180℃, and the rotation speed at 450 rpm. After leaching for 1 hour, solid-liquid separation was performed to obtain a lithium-containing solution and a first oxide slag. In the same pressure reactor, the nickel-cobalt-manganese slag and the lithium-containing solution were mixed to form a second slurry with a liquid-to-solid ratio of 6:1 mL / g. Oxygen was introduced and the mixture was stirred, with the oxygen partial pressure controlled at 0.6 MPa, the temperature at 180℃, and the rotation speed at 450 rpm. After leaching for 1 hour, solid-liquid separation was performed to obtain a lithium-rich solution with a lithium element concentration of 15.9 g / L and a second oxide slag. The calculated lithium leaching rate was 99.8%.

[0042] The lithium-rich solution was purified by passing it through a macroporous chelating resin at a flow rate of 2 BV / h. Sodium carbonate was then added to the purified lithium-rich solution at 95°C, with a molar ratio of sodium carbonate to lithium in the purified solution of 0.55. Solid-liquid separation yielded lithium carbonate. The first and second oxide residues were leached with sulfuric acid to obtain an acid leaching solution. Ni(OH)₂, Co(OH)₂, and / or Mn(OH)₂ were added to the acid leaching solution to adjust the pH to 3.5, precipitating and removing iron to obtain an iron-removing solution. Calcium was then removed from the iron-removing solution using a calcium extraction agent at a ratio of O / A = 2:1 and a time of 5 min through a 5-stage countercurrent extraction process, yielding a purified solution. The purified solution was then extracted to separate and recover nickel, cobalt, and manganese, or the nickel, cobalt, and manganese content in the purified solution was adjusted to prepare nickel-cobalt-manganese precursors.

[0043] Comparative Example 1 The only difference between this comparative example and Example 1 is that: During the leaching of the first and second slurries, the oxygen partial pressure was controlled at 0.05 MPa; the lithium concentration in the resulting lithium-rich solution was 14.1 g / L; and the calculated lithium leaching rate was 88.5%.

[0044] Comparative Example 2 The only difference between this comparative example and Example 1 is that: The temperature was controlled at 120℃ during the leaching of the first and second slurries; the lithium concentration in the resulting lithium-rich solution was 14.2 g / L; and the calculated lithium leaching rate was 89.1%.

[0045] Comparative Example 3 The only difference between this comparative example and Example 1 is that: The first and second slurries were leached under normal pressure and the temperature was controlled at 90℃; the lithium concentration in the resulting lithium-rich solution was 12.7 g / L; the lithium leaching rate was calculated to be 79.7%.

[0046] As can be seen from the above examples and comparative examples, when leaching the first slurry and the second slurry, if the oxygen partial pressure is too low or the leaching temperature is too low, the lithium leaching rate will decrease.

[0047] This application provides a method for recovering valuable elements from lithium-containing nickel-cobalt-manganese slag, which produces the following beneficial effects: oxygen pressure leaching reduces the adsorption of lithium by the nickel-cobalt-manganese slag, thereby releasing and recovering lithium without high-temperature decomposition waste gas pollution; no acid chemical reagents are added during the leaching process, and no additional saline wastewater is generated; the lithium concentration in the leachate is increased through cyclic leaching, reducing the lithium concentration cost and increasing the lithium yield; and lithium is recovered through oxygen pressure leaching and resin impurity removal, resulting in a simplified lithium recovery process.

[0048] The above description is only a preferred embodiment of this application and does not limit the patent scope of this application. All equivalent structural transformations made using the content of this application's specification under the inventive concept of this application, or direct / indirect applications in other related technical fields, are included within the patent protection scope of this application.

Claims

1. A method for recovering valuable elements from lithium-containing nickel-cobalt-manganese slag, characterized in that, Includes the following steps: S1. Obtain nickel-cobalt-manganese slag, wherein the nickel-cobalt-manganese slag includes hydroxides of nickel, cobalt, and manganese and lithium. S2. The nickel-cobalt-manganese slag is mixed with water to form a first slurry, and a first leaching is carried out at a temperature ≥140℃ and an oxygen partial pressure ≥0.1MPa. Then, solid-liquid separation is performed to obtain a lithium-containing solution and a first oxide slag. S3. The nickel-cobalt-manganese slag and the lithium-containing solution are mixed to form a second slurry. A second leaching is carried out at a temperature ≥140℃ and an oxygen partial pressure ≥0.1MPa. Then, solid-liquid separation is performed to obtain a lithium-rich solution and a second oxide slag.

2. The method for recovering valuable elements from lithium-containing nickel-cobalt-manganese slag according to claim 1, characterized in that, The lithium concentration in the lithium-rich solution is ≥15g / L.

3. The method for recovering valuable elements from lithium-containing nickel-cobalt-manganese slag according to claim 1, characterized in that, The pH of the first slurry is 8.0~10.

0.

4. The method for recovering valuable elements from lithium-containing nickel-cobalt-manganese slag according to claim 1, characterized in that, The liquid-to-solid ratio of the first slurry is (4~6):1mL / g, and the liquid-to-solid ratio of the second slurry is (4~6):1mL / g.

5. The method for recovering valuable elements from lithium-containing nickel-cobalt-manganese slag according to claim 1, characterized in that, In step S2, the temperature for the first leaching is 140~170℃ and the oxygen partial pressure is 0.1~0.5MPa; In step S3, the temperature for the second leaching is 140~170℃ and the oxygen partial pressure is 0.1~0.5MPa.

6. The method for recovering valuable elements from lithium-containing nickel-cobalt-manganese slag according to claim 1, characterized in that, In step S2, the first leaching is performed simultaneously with the first stirring. The first leaching time is 1-3 hours, and the first stirring speed is 300-450 rpm. In step S3, the second leaching is performed simultaneously with the second stirring. The second leaching time is 1-3 hours, and the stirring speed is 300-450 rpm.

7. The method for recovering valuable elements from lithium-containing nickel-cobalt-manganese slag according to claim 1, characterized in that, The lithium element in the nickel-cobalt-manganese slag, the first oxide slag, and the second oxide slag was detected and analyzed, and the lithium leaching rate was calculated to be ≥99%.

8. The method for recovering valuable elements from lithium-containing nickel-cobalt-manganese slag according to claim 1, characterized in that, The process after step S3 also includes: S4. The lithium-rich solution is purified using resin, and then a precipitant is added to the purified lithium-rich solution and solid-liquid separation is performed to obtain lithium carbonate. S5. The first oxide residue and the second oxide residue are acid-leached to obtain an acid leaching solution. Then, the acid leaching solution is purified to obtain a purified solution. The purified solution includes nickel, cobalt and manganese. The purified solution is extracted to separate and recover nickel, cobalt and manganese, or the content of nickel, cobalt and manganese in the purified solution is adjusted to prepare a nickel-cobalt-manganese precursor.

9. A method for recovering valuable elements from lithium-containing nickel-cobalt-manganese slag according to claim 8, characterized in that, The resin includes a macroporous chelating resin, and during the purification process, the flow rate of the lithium-rich solution is 2~4 BV / h.

10. A method for recovering valuable elements from lithium-containing nickel-cobalt-manganese slag according to claim 8, characterized in that, The precipitant includes sodium carbonate, and the molar ratio of sodium carbonate to lithium in the purified lithium-rich solution is 0.525~0.

575. The temperature at which the precipitant is added is 90~97℃.

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