Method for recovering valuable metals from waste lithium batteries

Abstract

The invention relates to the technical field of battery recovery, particularly provides a method for recovering valuable metals from waste lithium batteries, and aims to solve the problem of complex recovery procedures caused by high residual acid concentration and high impurity content in the existing recovery process. The method for recycling the valuable metal from the waste lithium battery comprises the following steps: S1, performing sulfating roasting on positive electrode powder of the waste lithium battery to obtain a roasted material; s2, adding water into the roasted material for dissolving, and carrying out solid-liquid separation to obtain lithium extraction slag and a lithium-rich solution; s3, a leaching agent solution is added into the lithium extraction slag, valuable metal leaching is carried out, leaching liquid is obtained, and the leaching agent solution at least comprises ammonium salt and a reducing agent; and S4, an oxidizing agent is added into the lixivium, impurity ions in the lixivium form precipitates, filter residues are removed, and the valuable metal solution is obtained. According to the method, valuable metal can be efficiently leached, the problems of too high residual acid content and too high impurity content can be avoided, and the recovery process is simplified.

Description

Technical Field

[0001] The invention relates to the technical field of battery recycling, and specifically provides a method for recycling valuable metals from waste lithium batteries. Background Art

[0002] In recent years, lithium-ion batteries have been increasingly used in mobile phones, computers, energy storage, power tools, and electric vehicles, resulting in a rapid increase in the amount of scrapped lithium-ion batteries.

[0003] The wet recycling process of waste lithium-ion batteries has the advantages of high comprehensive metal recovery rate, good product purity, low energy consumption and less pollution, and is favored by researchers. Based on the reality of the high price of lithium metal, researchers have developed a process of roasting-water leaching of lithium-acid leaching of nickel, cobalt and manganese. However, this process has problems such as high residual acid concentration and a large amount of impurities entering the solution during the acid leaching process. A large amount of alkali is consumed in the process of neutralizing the residual acid. At the same time, due to the large amount of impurities entering the solution, the impurity concentration is too high, which increases the difficulty of impurity removal.

[0004] Therefore, this field needs a new technical solution to solve the above problems. Summary of the invention

[0005] The present invention aims to solve the above technical problems at least to a certain extent, that is, to solve at least to a certain extent the problem that the existing valuable metal recovery process has a complicated recovery process due to high residual acid concentration and high impurity content.

[0006] In a first aspect, the present invention provides a method for recovering valuable metals from waste lithium batteries, the method comprising the following steps: S1: subjecting the positive electrode powder of the waste lithium batteries to sulfuric acid calcination to obtain a calcined material; S2: dissolving the calcined material with water and performing solid-liquid separation to obtain lithium extraction slag and a lithium-rich solution; S3: adding a leaching agent solution to the lithium extraction slag to leach valuable metals to obtain a leachate, wherein the leaching agent solution comprises at least an ammonium salt solution and a reducing agent solution; S4: adding an oxidant to the leachate to cause the impurity ions in the leachate to form a precipitate, removing the filter residue, and obtaining a valuable metal solution.

[0007] In the preferred technical solution of the above method for recovering valuable metals from waste lithium batteries, in the above step S3, the leaching agent solution satisfies at least one of the following conditions: the ammonium salt includes at least one of ammonium sulfate, ammonium chloride, and ammonium nitrate; the reducing agent includes at least one of sodium sulfite, sodium thiosulfate, starch, and glucose; the liquid-solid ratio of the leaching agent solution to the lithium-extracted residue is 8 ml / g - 12 ml / g; the concentration of the ammonium salt in the leaching agent solution is 2 mol / L - 4 mol / L; the concentration of the reducing agent in the leaching agent solution is 0.6 mol / L - 1.0 mol / L; the concentration of the reducing agent in the leaching agent solution is 1.5 mol / L - 2.0 mol / L.

[0008] In the preferred technical solution of the above method for recovering valuable metals from waste lithium batteries, in the above step S3, the leaching agent solution satisfies at least one of the following conditions: the ammonium salt is ammonium sulfate; the reducing agent is sodium sulfite; the liquid-solid ratio of the leaching agent solution to the lithium-extracted residue is 10 ml / g - 12 ml / g; the concentration of ammonium sulfate in the leaching agent solution is 3 mol / L - 4 mol / L; the concentration of sodium sulfite in the leaching agent solution is 0.8 mol / L - 1.0 mol / L; the concentration of sodium sulfite in the leaching agent solution is 1.5 mol / L - 1.8 mol / L.

[0009] In the preferred technical solution of the above method for recovering valuable metals from waste lithium batteries, in the above step S3, during the leaching process of valuable metals, the leaching temperature is 140°C - 180°C; and / or, the leaching time is 1.5 h - 2.5 h.

[0010] In the preferred technical solution of the above method for recovering valuable metals from waste lithium batteries, in the above step S3, during the leaching process of valuable metals, the leaching temperature is 150°C - 180°C; and / or, the leaching time is 2 h - 2.5 h.

[0011] In the preferred technical solution of the above method for recovering valuable metals from waste lithium batteries, in the above step S1, the step of sulfating and roasting the positive electrode powder of the lithium battery specifically includes: adding concentrated sulfuric acid to the positive electrode powder of the lithium battery for sulfating and roasting; wherein, the addition amount of concentrated sulfuric acid is 1.02 times - 1.06 times the theoretical amount.

[0012] In the preferred technical solution of the above method for recovering valuable metals from waste lithium batteries, the roasting temperature for sulfating and roasting the positive electrode powder of the lithium battery is 550°C - 700°C, and / or, the roasting time is 2 h - 3 h.

[0013] In the preferred technical solution of the method for recovering valuable metals from waste lithium batteries, the roasting temperature for sulfation roasting of the positive electrode powder of the lithium battery is 600°C - 700°C, and / or the roasting time is 2.5h - 3h.

[0014] In the preferred technical solution of the method for recovering valuable metals from waste lithium batteries, in the above step S4, the oxidant is hydrogen peroxide.

[0015] In the preferred technical solution of the method for recovering valuable metals from waste lithium batteries, when heating the leaching solution to 40°C - 60°C, hydrogen peroxide is added to the leaching solution.

[0016] In the case of adopting the above technical solution, the present invention first performs sulfation roasting on the positive electrode powder, which can convert the lithium element in the positive electrode powder into soluble lithium sulfate and enter the solution system, leaving valuable metals and other impurities in the slag phase, thereby realizing the separation and recovery of lithium. By adding an ammonium salt and a reducing agent to the lithium-extracted slag, the oxides of valuable metals (nickel, cobalt, manganese) can form complexes with ammonium ions and enter the solution, achieving efficient leaching of nickel, cobalt, and manganese. Moreover, the present invention uses an ammonium salt and a reducing agent as the leaching agent solution. Compared with the acid leaching process in the prior art, on the one hand, since the pH of the leaching agent solution is always maintained within the range of 5 - 6, the problem of excessive residual acid during the leaching process can be avoided. On the other hand, since aluminum, phosphorus, ferric ions, etc. are insoluble in the ammonium salt solution, during the leaching process, aluminum, phosphorus, and ferric ions basically do not enter the solution, and a large amount of impurities will not be introduced, completely solving the problem of excessive residual acid in the acid leaching process of the prior art, simplifying the subsequent impurity removal process, and shortening the recovery process of nickel, cobalt, and manganese. BRIEF DESCRIPTION OF THE DRAWINGS

[0017] The preferred embodiments of the present invention will be described below with reference to the accompanying drawings. In the drawings:

[0018] Figure 1 is a flowchart of the method for recovering valuable metals from waste lithium batteries of the present invention. DETAILED DESCRIPTION OF THE INVENTION

[0001] The preferred embodiments of the present invention will be described below with reference to the accompanying drawings. Those skilled in the art should understand that these embodiments are only used to explain the technical principle of the present invention and are not intended to limit the protection scope of the present invention.

[0002] Unless otherwise specified, the terms used in this application have the well-known meanings commonly understood by those skilled in the art. Unless otherwise specified, the numerical values of the various parameters mentioned in this application can be measured by various commonly used measurement methods in the art (for example, they can be tested according to the methods given in the embodiments of this application).

[0003] A list of items joined by the term "at least one of" or other similar terms can mean any combination of the listed items. For example, if items A and B are listed, then the phrase "at least one of A and B" means only A; only B; or A and B. In another example, if items A, B, and C are listed, then the phrase "at least one of A, B, and C" means only A; or only B; only C; A and B (excluding C); A and C (excluding B); B and C (excluding A); or all of A, B, and C. Item A can include a single component or multiple components. Item B can include a single component or multiple components. Item C can include a single component or multiple components.

[0019] It should be noted that the present invention is for the recovery of valuable metals from waste lithium batteries, without any limitation on the types of waste lithium batteries. For example, the waste lithium battery can be a primary lithium battery, or the waste lithium battery can also be a secondary lithium battery, or the waste lithium battery can also be a ternary lithium battery, etc. Such adjustments and changes to the types of waste lithium batteries do not deviate from the principle and scope of the present invention and should all be included within the protection scope of the present invention.

[0020] It also should be noted that the method for recovering valuable metals from waste lithium batteries according to the present invention takes the positive electrode powder of the lithium battery as the treatment object, and the present invention does not impose any limitation on the specific acquisition method of the positive electrode powder. The waste lithium battery can be disassembled by conventional disassembly methods in the art to obtain the positive electrode powder.

[0021] The following further elaborates the present application in conjunction with specific embodiments. It should be understood that these specific embodiments are only used to illustrate the present application and not to limit the scope of the present application.

[0022] The present application provides a method for recovering valuable metals from waste batteries, and the method includes the following steps:

[0023] S1: Sulfating roasting the positive electrode powder of the waste lithium battery to obtain a roasted material;

[0024] S2: Adding water to dissolve the roasted material and performing solid-liquid separation to obtain a lithium-extracted residue and a lithium-rich solution;

[0025] S3: Adding a leaching agent solution to the lithium-extracted residue to leach valuable metals to obtain a leachate, wherein the leaching agent solution includes at least an ammonium salt and a reducing agent;

[0026] S4: Adding an oxidizing agent to the leachate to form precipitates of impurity ions in the leachate, removing the filter residue to obtain a valuable metal solution.

[0027] By subjecting the positive electrode powder to sulfation roasting, the lithium element in the positive electrode powder can be converted into soluble lithium sulfate and enter the solution system, while the valuable metals and other impurities remain in the slag phase, thus realizing the separation and recovery of lithium. By adding a leaching agent solution (ammonium salt and reducing agent) to the lithium extraction slag for leaching the valuable metals, the oxides of the valuable metals (nickel, cobalt, manganese) can form complexes with ammonium ions and enter the solution, thereby enabling the efficient leaching of the valuable metals.

[0028] Meanwhile, since the ability of trivalent cobalt to form a complex with ammonium salt is lower than that of divalent cobalt to form a complex with ammonium sulfate, the reducing agent can first reduce trivalent cobalt to divalent cobalt, and then enable divalent cobalt to form a complex with ammonium salt and enter the solution system, thereby increasing the leaching rate of cobalt.

[0029] Specifically, the chemical equation for the reaction between the cobalt oxide in the lithium extraction slag and the reducing agent is as follows:

[0030] Co2O3+SO3 2- +4H + =2Co 2+ +SO4 2- +2H2O.

[0031] In the present invention, the ammonium salt and the reducing agent are used as the leaching agent solution. Compared with the acid leaching process in the prior art, on the one hand, since the pH of the leaching agent solution is always maintained within the range of 5-6, the problem of excessive residual acid during the leaching process can be avoided. On the other hand, since aluminum, phosphorus, trivalent iron, etc. are insoluble in the ammonium salt solution, aluminum, phosphorus, and trivalent iron basically do not enter the solution during the leaching process, and a large amount of impurities are not introduced, completely solving the problem of excessive residual acid in the acid leaching process of the prior art, simplifying the subsequent impurity removal process, and shortening the recovery process of nickel, cobalt, and manganese.

[0032] In the above step S1, the present invention does not impose any restrictions on the specific roasting method of sulfation roasting of the positive electrode powder. For example, when performing sulfation roasting on the positive electrode powder, only a sulfate or concentrated sulfuric acid is added to the positive electrode powder for sulfation roasting to obtain the roasted material. Alternatively, when performing sulfation roasting on the positive electrode powder, a mixture of concentrated sulfuric acid and sulfate can also be added to the positive electrode powder for sulfation roasting to obtain the roasted material. Such adjustments and changes in the types of reagents added during the sulfation roasting of the positive electrode powder do not deviate from the principle and scope of the present invention and should all be included within the protection scope of the present invention.

[0033] In a specific embodiment, when performing sulfation roasting on the positive electrode powder, concentrated sulfuric acid is added to the positive electrode powder for sulfation roasting to obtain the roasted material.

[0034] In another specific embodiment, when subjecting the positive electrode powder material to sulfation roasting, sulfates are added to the positive electrode powder material for sulfation roasting to obtain the roasted material, wherein the sulfates include at least one of potassium sulfate, sodium sulfate, and ammonium sulfate.

[0035] In another specific embodiment, when subjecting the positive electrode powder material to sulfation roasting, a mixture of concentrated sulfuric acid and sulfates is added to the positive electrode powder material for sulfation roasting to obtain the roasted material, wherein the sulfates include at least one of potassium sulfate, sodium sulfate, and ammonium sulfate.

[0036] Preferably, in the above step S1, the step of subjecting the positive electrode powder material of the waste lithium battery to sulfation roasting to obtain the roasted material specifically includes:

[0037] S11: Adding concentrated sulfuric acid to the positive electrode powder material for sulfation roasting to obtain the roasted material.

[0038] In some embodiments, in the above step S1, when subjecting the positive electrode powder material to sulfation roasting, the addition amount of concentrated sulfuric acid is 1.02 - 1.06 times the theoretical amount. For example, the addition amount of concentrated sulfuric acid is 1.02 times, 1.03 times, 1.04 times, 1.05 times, 1.06 times the theoretical amount, or the range composed of any two of these values. In some embodiments, the addition amount of concentrated sulfuric acid is 1.04 times the theoretical amount.

[0039] By controlling the addition amount of concentrated sulfuric acid, on the one hand, it can avoid the situation that the addition amount of concentrated sulfuric acid is too small to fully convert the lithium element in the positive electrode powder material into lithium sulfate, thus avoiding the reduction of lithium recovery rate. On the other hand, it can also avoid the situation that the addition amount of concentrated sulfuric acid is too large, resulting in the conversion of valuable metals (such as nickel, cobalt, and manganese) in the positive electrode powder material into soluble sulfates and entering the solution system, thus avoiding the increase in the difficulty of lithium recovery due to excessive impurities in the lithium element.

[0040] It should be noted that the theoretical addition amount of concentrated sulfuric acid is determined according to the content of lithium element in the positive electrode powder material. The following uses the following reaction equation to illustrate the theoretical addition amount of concentrated sulfuric acid.

[0041] Exemplarily, when the positive electrode powder material is subjected to sulfation roasting, the reaction of LiMeO2 in the positive electrode powder material with concentrated sulfuric acid can be represented by the following reaction equation:

[0042] 2LiMeO2 + H2SO4 = Li2SO4 + Me2O3 + H2O,

[0043] where Me represents at least one of nickel, cobalt, and manganese.

[0044] It should be noted that since the reaction between the positive electrode powder and concentrated sulfuric acid is relatively complex, only the above reaction equation is used as an example for illustration, which does not mean that only the above reaction occurs during the sulfuric acid roasting of the positive electrode powder.

[0045] From the above reaction equation, it can be seen that the molar ratio of lithium element to concentrated sulfuric acid in the positive electrode powder is 2:1. Assuming the molar amount of lithium element in the positive electrode powder to be sulfuric acid roasted is A1 mol, then the theoretical addition amount B1 of concentrated sulfuric acid = 0.5 × A1.

[0046] The molar amount A1 of lithium element can be obtained by testing and calculating the positive electrode powder. Specifically, first take x1 g of the positive electrode powder, dissolve and dilute it in aqua regia, and use ICP to detect the molar amount A2 mol of lithium element in the positive electrode powder. Weigh the mass of the positive electrode powder to be sulfuric acid roasted as x2 g.

[0047] The molar amount A1 of lithium element in the positive electrode powder to be sulfuric acid roasted is calculated according to the following formula:

[0048] A1 = (A2 × x2 ÷ x1) mol;

[0049] Then the theoretical addition amount B1 of concentrated sulfuric acid = (0.5 × A2 × x2 ÷ x1) mol.

[0050] Exemplarily, the actual addition amount B2 of concentrated sulfuric acid = 1.04 × B1.

[0051] That is: the actual addition amount B2 of concentrated sulfuric acid = (0.52 × A2 × x2 ÷ x1) mol.

[0052] In some embodiments, in the above step S1, the roasting temperature during the sulfuric acid roasting of the positive electrode powder is 550°C - 700°C. For example, the roasting temperature during the sulfuric acid roasting is 550°C, 580°C, 600°C, 630°C, 650°C, 680°C, 700°C or the range composed of any two of these values. In some embodiments, the roasting temperature during the sulfuric acid roasting is 600°C - 700°C.

[0053] By controlling the roasting temperature of the sulfuric acid roasting, on the one hand, if the roasting temperature is too low, the reaction between concentrated sulfuric acid and the positive electrode powder will be insufficient, reducing the dissolution efficiency of lithium and the recovery rate of lithium. On the other hand, if the roasting temperature is too high, not only will it cause energy waste, but also manganese in the positive electrode powder can enter the solution system, making manganese mixed in the lithium-rich solution and increasing the difficulty of lithium recovery. By reasonably controlling the temperature of the sulfuric acid roasting, the recovery rate of lithium can be improved and the difficulty of lithium recovery can be reduced.

[0054] In some embodiments, in the above step S1, the roasting time for the positive electrode powder during sulfation roasting is 2 h - 3 h. For example, the roasting time during sulfation roasting is 2 h, 2.1 h, 2.2 h, 2.3 h, 2.4 h, 2.5 h, 2.6 h, 2.7 h, 2.8 h, 2.9 h, 3 h, or the range formed by any two of these values. In some embodiments, the roasting time during sulfation roasting is 2.5 h - 3 h.

[0055] By controlling the roasting time of sulfation roasting, on the one hand, it can avoid the insufficient reaction between concentrated sulfuric acid and the positive electrode powder due to too short roasting time, thereby reducing the lithium recovery rate. On the other hand, it can avoid energy waste caused by too long roasting time, and can also avoid the problem of too low recovery efficiency of valuable metals due to too long roasting time.

[0056] In some embodiments, the liquid-solid ratio of the leaching agent solution to the lithium extraction residue is 8 ml / g - 12 ml / g. For example, the liquid-solid ratio of the leaching agent solution to the lithium extraction residue is 8 ml / g, 9 ml / g, 10 ml / g, 11 ml / g, 12 ml / g, or the range formed by any two of these values. In some embodiments, the liquid-solid ratio of the leaching agent solution to the lithium extraction residue is 10 ml / g - 12 ml / g.

[0057] By controlling the liquid-solid ratio of the leaching agent solution to the lithium extraction residue, on the one hand, it can avoid the insufficient leaching of valuable components in the lithium extraction residue due to too low liquid-solid ratio of the leaching agent solution to the lithium extraction residue, resulting in too low leaching efficiency of valuable metals. On the other hand, it can also avoid the problems of too high recovery cost and too high impurity content in the valuable metal solution due to too high liquid-solid ratio of the leaching agent solution to the lithium extraction residue.

[0058] It should be noted that in the above step S3, the present invention does not impose any restrictions on the specific type of ammonium salt. For example, the ammonium salt includes at least one of ammonium sulfate, ammonium chloride, and ammonium nitrate, etc. Such adjustments and changes to the specific type of ammonium salt do not deviate from the principle and scope of the present invention and should be included in the protection scope of the present invention.

[0059] In some embodiments, in the above step S3, the ammonium salt is ammonium sulfate.

[0060] Selecting ammonium sulfate as the ammonium salt can, on the basis of ensuring the leaching of valuable metals, not introduce other impurity ions, and there is no need for impurity removal when preparing the precursor solution of valuable metals, further simplifying the recovery process of valuable metals.

[0061] In some embodiments, the concentration of ammonium sulfate in the leaching agent solution is 2 mol / L - 4 mol / L. For example, the concentration of ammonium sulfate is 2 mol / L, 2.5 mol / L, 3 mol / L, 3.5 mol / L, 4 mol / L, or a range composed of any two of these values. In some embodiments, the concentration of ammonium sulfate is 3 mol / L - 4 mol / L.

[0062] By controlling the concentration of ammonium sulfate in the leaching agent solution within 2 mol / L - 4 mol / L, the concentration of ammonium sulfate can be reasonably controlled, enabling the oxides of valuable metals to better form complexes with ammonium sulfate, and then leaching from the lithium-extracted slag, improving the leaching rate of valuable metals, and further increasing the recovery rate of valuable metals. At the same time, it can avoid the precipitation of the complex formed by manganese and ammonium sulfate from the leaching solution due to excessive addition of ammonium sulfate, thus preventing the leaching rate of manganese from decreasing due to too high a concentration of ammonium sulfate.

[0063] The equation for the reaction of the complex formed by manganese and ammonium sulfate with excessive ammonium sulfate is as follows:

[0064]

[0065] It can be seen from this that excessive addition of ammonium sulfate will cause the complex formed by manganese and ammonium sulfate to precipitate from the leaching solution, reducing the leaching rate of manganese.

[0066] In the above step S3, the present invention does not impose any restrictions on the specific types of reducing agents. The reducing agents include at least one of sodium sulfite, sodium thiosulfate, starch, and glucose, etc. Such adjustments and changes to the specific types of reducing agents do not deviate from the principles and scope of the present invention and should all be included within the protection scope of the present invention.

[0067] Preferably, the reducing agent is sodium sulfite.

[0068] Compared with selecting other reducing agents, when the reducing agent is selected as sodium sulfite, after the reaction between sodium sulfite and cobalt oxide, sulfate ions are generated again, and no new impurities are introduced.

[0069] It should be noted that the complex formed by manganese and ammonium sulfate can react with excessive sulfite ions and ammonium ions. The chemical equation for the reaction is as follows:

[0070]

[0071] It can be seen from this that excessive sodium sulfite can cause the complex formed by manganese and ammonium sulfate to precipitate from the leaching solution. Therefore, the leaching rate and types of valuable metals in the leaching solution can be controlled by controlling the addition amount of sodium sulfite.

[0072] In some embodiments of the present invention, the concentration of sodium sulfite in the leaching agent solution is 0.6 mol / L - 1.0 mol / L. For example, the concentration of sodium sulfite in the leaching agent solution is 0.6 mol / L, 0.7 mol / L, 0.8 mol / L, 0.9 mol / L, 1.0 mol / L, or the range composed of any two of these values. In some embodiments, the concentration of sodium sulfite in the leaching agent solution is 0.8 mol / L - 1.0 mol / L.

[0073] By controlling the concentration of sodium sulfite in the leaching agent solution within 0.6 mol / L - 1.0 mol / L, it is possible to make the added sodium sulfite react completely with trivalent cobalt, so that the generated divalent cobalt can better form a complex with ammonium sulfate and be leached. At the same time, it can also avoid excessive sodium sulfite reacting with manganese and reducing the leaching rate of manganese, resulting in relatively high leaching rates of nickel, cobalt, and manganese, and obtaining a valuable metal solution containing nickel, cobalt, and manganese.

[0074] In other embodiments of the present invention, the concentration of sodium sulfite in the leaching agent solution is 1.5 mol / L - 2.0 mol / L. For example, the concentration of sodium sulfite in the leaching agent solution is 1.5 mol / L, 1.6 mol / L, 1.7 mol / L, 1.8 mol / L, 1.9 mol / L, 2.0 mol / L, or the range composed of any two of these values. In some embodiments, the concentration of sodium sulfite in the leaching agent solution is 1.5 mol / L - 1.8 mol / L.

[0075] By controlling the addition amount of sodium sulfite in the leaching agent solution within 1.5 mol / L - 2.0 mol / L, on the basis of reducing trivalent cobalt to divalent cobalt, it is also possible to make the complex formed by manganese and ammonium sulfate react with sufficient sodium sulfite and precipitate from the leaching solution, so that the leaching rates of nickel and cobalt are relatively high, and thus the lower-cost manganese can enter the slag phase, obtaining a valuable metal solution containing nickel and cobalt.

[0076] It can be seen from this that if it is necessary to separate and obtain a valuable metal solution containing nickel, cobalt, and manganese, controlling the concentration of sodium sulfite in the leaching agent solution within 0.6 mol / L - 1.0 mol / L can result in relatively high leaching rates of nickel, cobalt, and manganese; if it is necessary to separate and obtain a valuable metal solution containing nickel and cobalt, controlling the concentration of sodium sulfite in the leaching agent solution within 1.5 mol / L - 2.0 mol / L can result in relatively high leaching rates of nickel and cobalt and make manganese enter the slag phase.

[0077] In some embodiments, during the leaching process of valuable metals, the leaching temperature is 140°C - 180°C. For example, the leaching temperature is 140°C, 150°C, 160°C, 170°C, 180°C, or a range composed of any two of these values. In some embodiments, the leaching temperature is 150°C - 180°C.

[0078] By controlling the leaching temperature within the range of 140°C - 180°C, on the one hand, it can avoid the insufficient leaching of valuable metals due to too low temperature, reducing the leaching efficiency of valuable metals. On the other hand, it can also avoid the energy waste and equipment safety problems caused by too high leaching temperature.

[0079] It should be noted that since the leaching system needs to be pressurized when the leaching temperature is raised to the range of 140°C - 180°C, the present invention does not impose any restrictions on the pressure conditions during the leaching of valuable metals. As long as the leaching temperature can be raised to 140°C - 180°C, etc., such adjustments and changes to the pressure conditions during the leaching of valuable metals do not deviate from the principle and scope of the present invention and should be included within the protection scope of the present invention.

[0080] In some embodiments, during the leaching process of valuable metals, the leaching time is 1.5h - 2.5h. For example, the leaching time is 1.5h, 1.7h, 1.9h, 2h, 2.2h, 2.5h, or a range composed of any two of these values. In some embodiments, during the leaching process of valuable metals, the leaching time is 2h - 2.5h.

[0081] By controlling the leaching time within the range of 1.5 - 2.5h, it can avoid the insufficient leaching of valuable metals due to too short leaching time, thereby reducing the leaching rate of valuable metals, and can also avoid the too low recovery efficiency of valuable metals due to too long leaching time.

[0082] Further preferably, the leaching time is 2h.

[0083] In the above step S4, an oxidant is added to the leaching solution to form a precipitate of impurity ions in the leaching solution, and the filter residue is removed to obtain a valuable metal solution.

[0084] Since the leaching agent solution in step S3 includes a reducing agent, this will cause a small amount of trivalent iron in the lithium extraction residue to be reduced to divalent iron and enter the leaching solution. By adding an oxidant to the leaching solution, the divalent iron in the leaching solution can be oxidized to trivalent iron, thereby forming ferric hydroxide precipitate. Then, the filter residue is removed to obtain a valuable metal solution containing reduced content of impurity iron in the valuable metal solution.

[0085] In the above step S4, the present invention does not impose any restrictions on the type of oxidant. For example, the oxidant can be hydrogen peroxide, sodium hypochlorite, or any other possible oxidant, etc. Such adjustments and changes to the type of oxidant do not deviate from the principle and scope of the present invention and should be included within the protection scope of the present invention.

[0086] Preferably, the oxidant is hydrogen peroxide.

[0087] By adding hydrogen peroxide to the filtrate, not only can the divalent iron in the solution be oxidized to trivalent iron to remove iron impurities, but also no new impurities will be introduced, further reducing the impurity content in the valuable metal solution and simplifying the recovery process of valuable metals.

[0088] It should be noted that in the above step S4, hydrogen peroxide can be directly added to the leachate, or the leachate can be first heated to 40°C - 60°C and then hydrogen peroxide can be added to the leachate, etc. Such adjustments and changes to the specific addition method of hydrogen peroxide do not deviate from the principle and scope of the present invention and should be included within the protection scope of the present invention.

[0089] Preferably, the leachate is first heated to 40°C - 60°C and then hydrogen peroxide is added to the leachate. For example, the leachate can be first heated to 40°C, 45°C, 46°C, 48°C, 50°C, 52°C, 54°C, 55°C, 60°C or the range composed of any two of these values. In some embodiments, the leachate is first heated to 50°C and then hydrogen peroxide is added to the leachate.

[0090] By first heating the leachate to an appropriate temperature and then adding hydrogen peroxide to the leachate, on the one hand, it can avoid the reaction rate between hydrogen peroxide and divalent iron being too slow due to the too low temperature of the leachate, which affects the recovery efficiency of valuable metals. On the other hand, it can also avoid the decomposition of hydrogen peroxide accelerating due to the too high temperature of the leachate, resulting in a decrease in the utilization rate of hydrogen peroxide.

[0091] In some embodiments, the actual addition amount of hydrogen peroxide is 1.3 times - 2.0 times the theoretical amount. For example, the actual addition amount of hydrogen peroxide is 1.3 times, 1.4 times, 1.5 times, 1.6 times, 1.7 times, 1.8 times, 1.9 times, 2.0 times the theoretical amount or the range composed of any two of these values. In some embodiments, the actual addition amount of hydrogen peroxide is 1.5 times the theoretical amount.

[0092] By controlling the actual addition amount of hydrogen peroxide within 1.3 times - 2.0 times the theoretical amount, hydrogen peroxide can be made excessive, so that hydrogen peroxide can react more thoroughly with the divalent iron in the leachate, further reducing the content of impurity ions in the valuable metal solution. At the same time, it can also avoid waste of hydrogen peroxide caused by too much actual addition amount of hydrogen peroxide.

[0093] It should be noted that the theoretical dosage of hydrogen peroxide is related to the content of ferrous ions in the leaching solution. The content of ferrous ions in the leaching solution can be detected by titration, and the theoretical dosage of hydrogen peroxide can be calculated accordingly.

[0094] Specifically, the chemical reaction equation between ferrous ions in the leaching solution and hydrogen peroxide is as follows:

[0095] 2Fe 2+ +H2O2+2H + =2Fe 3+ +2H2O

[0096] In the above reaction equation, the molar ratio of ferrous ions to hydrogen peroxide is 2:1. If the content of ferrous ions in the leaching solution detected by titration is C mol, then the theoretical dosage of hydrogen peroxide D1 = (0.5 × C) mol.

[0097] Exemplarily, the actual addition amount of hydrogen peroxide is D2 = D1 × 1.5;

[0098] That is: the actual addition amount of hydrogen peroxide is D2 = (0.75 × C) mol.

[0099] It should be noted that the hydrogen peroxide used in the present invention is industrial hydrogen peroxide with a concentration of 27.5% - 35%.

[0100] Examples and Comparative Examples

[0101] Example 1:

[0102] The waste batteries used in this example are ternary lithium-ion batteries. The method for recovering valuable metals from waste ternary lithium-ion batteries includes the following steps:

[0103] S1: Concentrated sulfuric acid (98 wt%) is added to the positive electrode powder of the ternary lithium-ion battery for sulfation roasting to obtain the roasted material. The addition amount of concentrated sulfuric acid is 1.04 times the theoretical amount, the roasting temperature is 700 °C, and the roasting time is 3 h.

[0104] S2: The roasted material is dissolved in water and subjected to solid-liquid separation to obtain lithium-extracted slag and lithium-rich solution. The roasted material is dissolved in pure water according to a liquid-solid ratio of 6 mL:1 g.

[0105] S3: An extractant solution is added to the lithium-extracted slag for leaching of valuable metals, and the leaching solution is obtained after solid-liquid separation. The extractant solution is a mixed solution of ammonium sulfate and sodium sulfite.

[0106] S4: The leaching solution is heated to 50 °C, and hydrogen peroxide is added to the leaching solution to convert Fe 2+ in the leaching solution into Fe3 + , remove the filter residue to obtain a valuable metal solution, wherein the addition amount of hydrogen peroxide is 1.5 times the theoretical amount.

[0107] In this embodiment, the liquid-solid ratio of the leaching agent solution to the lithium-extracted slag is 10 ml / g, the concentration of ammonium sulfate in the leaching agent solution is 3 mol / L, the concentration of sodium sulfite in the leaching agent solution is 0.8 mol / L, the leaching temperature is 150 °C, and the leaching time is 2 h.

[0108] Examples 2-14 and Comparative Examples 1-10:

[0109] Examples 2-14 and Comparative Examples 1-10 are all achieved by adjusting the concentration of ammonium sulfate in the leaching agent solution, the concentration of sodium sulfite in the leaching agent solution, the liquid-solid ratio of the leaching agent solution to the lithium-extracted slag, the leaching time, and the leaching temperature on the basis of Example 1. The specific adjustment measures and detailed data are shown in Table 1.

[0110] Table 1:

[0111] Testing method:

[0112] 1. Determination of the impurity concentration in the leaching solution:

[0113] Carry out sulfation roasting respectively according to the steps of the above examples. After dissolving the roasted material, carry out solid-liquid separation to obtain the lithium-extracted slag, and then carry out leaching of valuable metals. After solid-liquid separation, obtain the leaching solution and the leaching residue.

[0114] Use ICP to detect the concentration of aluminum (unit: ppm), the concentration of phosphorus (unit: ppm), and the concentration of iron (unit: ppm) in the leaching solution respectively.

[0115] After testing, in the leaching solution obtained from the above examples, the concentration of aluminum is lower than 1 ppm, the concentration of phosphorus is lower than 50 ppm, and the concentration of iron is lower than 50 ppm. The concentration of impurity ions in the obtained leaching solution is much lower than that of the acid leaching process in the prior art.

[0116] In the acid leaching process in the prior art, impurity ions are introduced into the leaching solution system, resulting in relatively high levels of phosphorus, aluminum, and ferric iron. The subsequent impurity removal process is cumbersome. In the present invention, ammonium sulfate and sodium sulfite are used as the leaching agent solution. During leaching, ammonium sulfate only forms complexes with nickel, cobalt, and manganese and enters the leaching solution system, while phosphorus, aluminum, and ferric iron basically do not enter the leaching solution system and all enter the leaching residue. Therefore, the concentrations of impurity ions of phosphorus, aluminum, and ferric iron in the obtained leaching solution are significantly lower than those in the acid leaching process in the prior art, completely solving the problem of excessive impurity content in the existing acid leaching process, simplifying the subsequent impurity removal process, and shortening the recovery process of nickel, cobalt, and manganese.

[0117] 2. Determination of the leaching rates of nickel, cobalt, and manganese:

[0118] Perform sulfation roasting according to the steps of the above-mentioned examples and comparative examples respectively. After dissolving the roasted material, perform solid-liquid separation to obtain the lithium extraction residue. Then, add the leaching agent solution to the lithium extraction residue for leaching of valuable metals. After solid-liquid separation, obtain the leaching solution and the leaching residue.

[0119] Weigh the mass M of the lithium extraction residue a , and weigh the mass M of the leaching residue b ;

[0120] Weigh a part of the lithium extraction residue and use an inductively coupled plasma optical emission spectrometer (ICP-OES) to detect the nickel content ω a1 , cobalt content ω a2 , and manganese content ω a3 ;

[0121] Weigh a part of the leaching residue and use an inductively coupled plasma optical emission spectrometer (ICP-OES) to detect the nickel content ω b1 , cobalt content ω b2 , and manganese content ω b3 ;

[0122] Calculate the leaching rate Q1 of nickel according to the following formula:

[0123]

[0124] Calculate the leaching rate Q2 of cobalt according to the following formula:

[0125]

[0126] Calculate the leaching rate Q3 of manganese according to the following formula:

[0127]

[0128] Testing results:

[0129] The detection results of the above Examples 1 to 14 and Comparative Examples 1 to 10 are shown in Table 2.

[0130] Table 2: Detection data of examples and comparative examples

[0131] It can be seen from the data in Table 2 that:

[0132] 1. Comparing Examples 1 to 3 with Comparative Examples 1 to 2, the leaching rates of nickel, cobalt, and manganese in Examples 1 to 3 are significantly higher than those in Comparative Example 1. Although the leaching rates of nickel and cobalt in Comparative Example 2 are slightly higher than those in Examples 1 to 3, with the increase in the concentration of ammonium sulfate, the leaching rates of nickel and cobalt do not increase significantly, and too high a concentration of ammonium sulfate will affect the leaching rate of manganese, and the increase in the concentration of ammonium sulfate will also lead to an increase in cost. Therefore, when the concentration of ammonium sulfate in the leaching agent solution is 2 mol / L - 4 mol / L, the leaching rates of nickel, cobalt, and manganese are relatively high and the recovery cost is relatively low.

[0133] Also, since the leaching rates of nickel, cobalt, and manganese in Examples 1 and 3 are higher than those in Example 2, it can be seen that when the concentration of ammonium sulfate in the leaching agent solution is 3 mol / L - 4 mol / L, the leaching rates of nickel, cobalt, and manganese are relatively high.

[0134] 2. Comparing Examples 1, 4 to 5 with Comparative Examples 3 to 4, the leaching rates of nickel, cobalt, and manganese in Examples 1, 4 to 5 are significantly higher than those in Comparative Example 3. Although the leaching rates of nickel, cobalt, and manganese in Comparative Example 4 are slightly higher than those in Examples 1, 4 to 5, with the increase in the liquid-solid ratio of the leaching agent solution to the lithium-extracted slag, the leaching rates of nickel, cobalt, and manganese do not increase significantly, and too high a liquid-solid ratio will lead to an increase in the consumption of the leaching agent solution, and thus an increase in cost. It can be seen that when the liquid-solid ratio of the leaching agent solution to the lithium-extracted slag is 8 ml / g - 12 ml / g, the leaching rates of nickel, cobalt, and manganese are relatively high and the recovery cost is relatively low.

[0135] Also, since the leaching rates of nickel, cobalt, and manganese in Examples 1 and 5 are significantly higher than those in Example 4, it can be seen that when the liquid-solid ratio of the leaching agent solution to the lithium-extracted slag is 10 ml / g - 12 ml / g, the leaching rates of nickel, cobalt, and manganese are relatively high.

[0136] 3. Compare Example 1, Examples 6 to 7 with Comparative Examples 5 to 6. The leaching rates of nickel, cobalt, and manganese in Example 1, Examples 6 to 7 are significantly higher than those in Comparative Example 5. Although the leaching rates of nickel, cobalt, and manganese in Comparative Example 6 are slightly higher than those in Example 1, Examples 6 to 7, with the further increase of temperature, the leaching rates of nickel, cobalt, and manganese in Comparative Example 6 do not increase significantly. Moreover, the increase of leaching temperature will cause problems of energy waste and equipment safety. Therefore, considering comprehensively the leaching rates of nickel, cobalt, and manganese and the safety issues of equipment caused by leaching temperature, when the leaching temperature is 140°C - 180°C, the leaching temperature is relatively optimal.

[0137] Also, since the leaching rates of nickel, cobalt, and manganese in Example 1 and Example 7 are significantly higher than those in Example 6, therefore, when the leaching temperature is 150°C - 180°C, the leaching rates of nickel, cobalt, and manganese are relatively optimal.

[0138] 4. Compare Example 1, Examples 8 to 9 with Comparative Example 7 and Comparative Example 8. The leaching rates of nickel-cobalt-manganese in Example 1, Examples 8 to 9 are significantly higher than those in Comparative Example 7. With the further extension of leaching time, the leaching rates of nickel, cobalt, and manganese in Comparative Example 8 do not increase significantly. However, too long leaching time will lead to too low leaching efficiency and affect the recovery efficiency of valuable leaching. Therefore, when the leaching time is 1.5 h - 2.5 h, the leaching rates of nickel, cobalt, and manganese are relatively high, and the leaching efficiency of valuable metals is also relatively high.

[0139] Also, since the leaching rates of nickel, cobalt, and manganese in Example 1 and Example 9 are higher than those in Example 8, therefore, when the leaching time is 2 h - 2.5 h, the leaching rates and leaching efficiency of nickel, cobalt, and manganese are both high.

[0140] 5. Compare Example 1, Examples 10 to 11 with Comparative Example 9 to Comparative Example 10. The leaching rates of nickel, cobalt, and manganese in Example 1, Examples 10 to 11 are significantly higher than those in Comparative Example 9. Although the leaching rates of nickel and cobalt in Comparative Example 10 are higher than those in Example 1, Examples 10 and 11, the leaching rate of manganese in Comparative Example 10 is much lower than those in Example 1, Examples 10 and 11. Therefore, it can be seen that when the concentration of sodium sulfite in the leaching agent solution is 0.6 mol / L - 1.0 mol / L, the leaching rates of nickel, cobalt, and manganese can be relatively high.

[0141] Also, since the leaching rates of nickel, cobalt, and manganese in Example 1 are higher than those in Example 10, although the difference in the leaching rate of manganese between Example 11 and Example 10 is not significant, the leaching rates of nickel and cobalt in Example 11 are higher than those in Example 10. Therefore, it can be seen that when the concentration of sodium sulfite in the leaching agent solution is 0.8 mol / L - 1.0 mol / L, the leaching rates of nickel, cobalt, and manganese are all relatively high.

[0142] 6. Compare Examples 12 to 14 with Comparative Example 10. The leaching rates of nickel and cobalt in Examples 12 to 14 are not much different from those in Comparative Example 10. However, the leaching rate of manganese in Examples 12 to 14 is much lower than that in Comparative Example 10. From this, it can be seen that when recovering nickel and cobalt, when the concentration of sodium sulfite in the leaching agent solution is 1.5 mol / L - 2 mol / L, the leaching rates of nickel and cobalt can be relatively high, and the leaching rate of manganese can be relatively low.

[0143] Also, since the leaching rates of nickel and cobalt in Examples 12 and 13 are slightly higher than those in Example 14, and the leaching rates of manganese in Examples 12 and 13 are already relatively low. As the concentration of sodium sulfite in the leaching agent solution is further increased, the leaching rate of manganese does not decrease significantly, while the increase in the concentration of sodium sulfite will increase the recovery cost. From this, it can be seen that when recovering nickel and cobalt, when the concentration of sodium sulfite in the leaching agent solution is 1.5 mol / L - 1.8 mol / L, the leaching rates of nickel and cobalt can be relatively high, and the leaching rate of manganese can be relatively low.

[0144] So far, the technical solutions of the present invention have been described in conjunction with the preferred embodiments shown in the drawings. However, it is easy for those skilled in the art to understand that the protection scope of the present invention is obviously not limited to these specific embodiments. Without departing from the principle of the present invention, those skilled in the art can make equivalent changes or substitutions to the relevant technical features, and the technical solutions after these changes or substitutions will all fall within the protection scope of the present invention.

Claims

1. A method for recovering valuable metals from waste lithium batteries, characterized in that, The method includes the following steps: S1: Sulfate roast the positive electrode powder of the waste lithium battery to obtain the roasted material; S2: Dissolve the roasted material in water and perform solid-liquid separation to obtain lithium extraction residue and lithium-rich solution; S3: Add a leaching agent solution to the lithium extraction residue to leach valuable metals to obtain a leachate, wherein the leaching agent solution includes at least an ammonium salt and a reducing agent; S4: Add an oxidizing agent to the leachate to form precipitates of impurity ions in the leachate, remove the filter residue, and obtain a valuable metal solution.

2. The method for recovering valuable metals from waste lithium batteries according to claim 1, characterized in that, In the above step S3, the leaching agent solution satisfies at least one of the following conditions: The ammonium salt includes at least one of ammonium sulfate, ammonium chloride, and ammonium nitrate; The reducing agent includes at least one of sodium sulfite, sodium thiosulfate, starch, and glucose; The liquid-solid ratio of the leaching agent solution to the lithium extraction residue is 8 ml / g - 12 ml / g; The concentration of the ammonium salt in the leaching agent solution is 2 mol / L - 4 mol / L; The concentration of the reducing agent in the leaching agent solution is 0.6 mol / L - 1.0 mol / L; The concentration of the reducing agent in the leaching agent solution is 1.5 mol / L - 2.0 mol / L.

3. The method for recovering valuable metals from waste lithium batteries according to claim 2, characterized in that, In the above step S3, the leaching agent solution satisfies at least one of the following conditions: The ammonium salt is ammonium sulfate; The reducing agent is sodium sulfite; The liquid-solid ratio of the leaching agent solution to the lithium extraction residue is 10 ml / g - 12 ml / g; The concentration of ammonium sulfate in the leaching agent solution is 3 mol / L - 4 mol / L; The concentration of sodium sulfite in the leaching agent solution is 0.8 mol / L - 1.0 mol / L; The concentration of sodium sulfite in the leaching agent solution is 1.5 mol / L - 1.8 mol / L.

4. The method for recovering valuable metals from waste lithium batteries according to claim 1, characterized in that, In the above step S3, during the leaching of valuable metals, the leaching temperature is 140°C - 180°C; and / or, the leaching time is 1.5 h - 2.5 h.

5. The method for recovering valuable metals from waste lithium batteries according to claim 4, characterized in that, In the above step S3, during the leaching of valuable metals, the leaching temperature is 150°C - 180°C; and / or, the leaching time is 2 h - 2.5 h.

6. The method for recovering valuable metals from waste lithium batteries according to claim 1, characterized in that, In the above step S1, the step of sulfate roasting the positive electrode powder of the lithium battery specifically includes: Add concentrated sulfuric acid to the positive electrode powder of the lithium battery for sulfate roasting; wherein, the addition amount of the concentrated sulfuric acid is 1.02 times - 1.06 times of the theoretical amount.

7. The method for recovering valuable metals from waste lithium batteries according to claim 1, characterized in that, The roasting temperature for sulfate roasting the positive electrode powder of the lithium battery is 550°C - 700°C, and / or, the roasting time is 2 h - 3 h.

8. The method for recovering valuable metals from waste lithium batteries according to claim 7, characterized in that, The roasting temperature for sulfate roasting the positive electrode powder of the lithium battery is 600°C - 700°C, and / or, the roasting time is 2.5 h - 3 h.

9. The method for recovering valuable metals from waste lithium batteries according to claim 1, characterized in that, In the above step S4, the oxidizing agent is hydrogen peroxide.

10. The method for recovering valuable metals from waste lithium batteries according to claim 9, characterized in that, In the above step S4, when heating the leachate to 40°C - 60°C, add the hydrogen peroxide to the leachate.

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