Method for recovering valuable metals

Abstract

Provided in the present disclosure is a method for recovering valuable metals. The method comprises: a roasting procedure, involving roasting a material obtained by battery recycling and a lithium extraction agent under the protection of an inert gas, so as to obtain a roasted product; a first leaching separation procedure, involving subjecting the roasted product to first leaching separation, so as to obtain a lithium-containing leachate and leaching residues; a calcining procedure, involving mixing the leaching residues and a slag former, and then calcining same, so as to obtain a calcined product; and a matte smelting procedure, involving mixing the calcined product, nickel matte, pressure oxidation slag and a matte agent, and then subjecting the resulting mixture to matte smelting, so as to obtain nickel-containing matte. The recovery method provided in the embodiments of the present disclosure involves a simple process flow, has a high recovery rate and generates few pollutants, such that the recovery cost of valuable metals is low, and the recovery method is environmentally friendly.

Description

Methods for recovering valuable metals

[0001] Cross-references to related publications

[0002] This disclosure claims priority to Chinese Patent Publication No. 202310664670.2, filed on June 6, 2023, entitled “Method for Recovering Valuable Metals,” the entire contents of which are incorporated herein by reference. Technical Field

[0003] The present disclosure relates to the technical field of valuable metal recovery, and in particular to a valuable metal recovery method. Background Art

[0004] Valuable metals generally refer to metals with recycling value among the raw materials used in metal refining, such as nickel (Ni), cobalt (Co), manganese (Mn), and lithium (Li). These valuable metals can be used as manufacturing materials for batteries, giving batteries high energy density and high cycle performance, and are therefore widely used in electric vehicles, electronic equipment, aerospace, and other fields.

[0005] Currently, valuable metals in batteries can be recovered from nickel intermediates (such as nickel matte) or from batteries. However, the related art processes for recovering valuable metals are complex, have low recovery rates, and are prone to generating large amounts of pollutants, resulting in high recycling costs and environmental impact.

[0006] Summary of the Invention

[0007] The present disclosure provides a method for recovering valuable metals, which has a simple process flow, a high recovery rate, and is environmentally friendly.

[0008] The present disclosure provides a method for recovering valuable metals, the method comprising:

[0009] The roasting process is to roast the battery recycling material and the lithium extraction agent under a protective atmosphere to obtain a roasted product;

[0010] a first leaching and separation step of performing a first leaching and separation on the roasted product to obtain a lithium-containing leachate and a leaching residue;

[0011] a calcination step, mixing the leached slag and the slag-making agent and then calcining them to obtain a calcined product;

[0012] In the matte making and smelting process, the calcined product, nickel matte, oxygen pressed slag and matte making agent are mixed and then matte making and smelting is carried out to obtain nickel-containing matte.

[0013] The recovery method provided by the disclosed embodiments, through a roasting step and a first leaching and separation step, can significantly enrich valuable lithium metals in the battery recyclate in the leachate, thereby improving the lithium recovery rate. The leached residue obtained from the first leaching and separation step undergoes a calcination step for a slagging reaction, pre-removing some impurities, such as Ca and Mg, that enter the slagging agent. This helps shorten the matte smelting time. The calcined product, nickel matte, oxygen-pressed slag and matte-making agent obtained through the calcination process are mixed and then subjected to matte smelting, which can separate the valuable metal nickel in the calcined product, nickel matte and oxygen-pressed slag from other impurities, that is, valuable metals such as valuable metal nickel enter the sulfur phase to obtain matte containing nickel and other valuable metals. In this way, valuable metals such as nickel in the calcined product, nickel matte and oxygen-pressed slag can be recovered at the same time, simplifying the process of recovering valuable metals from battery recyclables, nickel matte and oxygen-pressed slag, and increasing the recovery rate of valuable metal nickel. In addition, the large amount of oxygen element contained in nickel matte, oxygen-pressed slag and calcined products can be used to reduce the amount of oxygen required for external introduction during the sulfur making process. This can further reduce the cost of recovering valuable metals while reducing the safety risks brought about by the large amount of oxygen introduced from the outside and the requirements for fire protection and recovery equipment in the factory. Moreover, impurities such as iron and manganese in the calcined products, nickel matte and oxygen-pressed slag will enter the slag phase, which can be directly used in fields such as construction as a by-product. This can reduce the amount of impurity removal in subsequent production lines, thereby helping to reduce the demand for impurity removal equipment and the floor space of the recycling plant, thereby further helping to reduce the recovery cost of valuable metals in battery recyclates, nickel matte and oxygen-pressed slag. In addition, in the above-mentioned processes, few pollutants are generated, thereby making the recovery method of valuable metals environmentally friendly. Therefore, the recovery method provided by the embodiment of the present disclosure has a simple process flow, a high recovery rate, and few pollutants generated, thereby making the recovery cost of valuable metals low and environmentally friendly.

[0014] In some embodiments of the present disclosure, during the roasting process, the lithium-extracting agent includes one or more of a reducing agent, a sulfate, or concentrated sulfuric acid.

[0015] In some embodiments of the present disclosure, the reducing agent includes one or more of activated carbon, hydrogen, lignite, anthracite, and carbon monoxide.

[0016] In some embodiments of the present disclosure, the sulfate includes one or more of ammonium sulfate and sodium sulfite.

[0017] In some embodiments of the present disclosure, the mass ratio of the battery recyclate to the lithium extraction agent is 20:(1-30).

[0018] In some embodiments of the present disclosure, the calcination temperature is 300° C.-1000° C., the heating rate is 2° C. / min-10° C. / min, and the calcination time is 0.5 h-5 h.

[0019] In some embodiments of the present disclosure, the first leaching and separation process includes:

[0020] a first leaching step of mixing the roasted product and a first leaching agent and performing a first leaching treatment to dissolve lithium in the roasted product into a solution to obtain a first leaching slurry containing lithium;

[0021] In the first separation step, the first leaching slurry is subjected to solid-liquid separation to obtain lithium-containing leaching solution and leaching residue.

[0022] In some embodiments of the present disclosure, in the first leaching process, the pH of the first leaching treatment is 3-7.

[0023] In some embodiments of the present disclosure, in the first leaching process, the temperature of the first leaching treatment is 30° C.-90° C., and the time of the first leaching treatment is 0.5 h-4 h.

[0024] In some embodiments of the present disclosure, during the calcination process, the mass ratio of the leached slag to the slag-forming agent is (2-20):1.

[0025] In some embodiments of the present disclosure, the calcination temperature is 600-1000° C., and the calcination time is 2 h-8 h.

[0026] In some embodiments of the present disclosure, the slag forming agent includes one or more of silicon dioxide, diatomaceous earth, bentonite, calcium oxide, and aluminum oxide.

[0027] In some embodiments of the present disclosure, in the matte smelting process, the mass ratio of the calcined product, nickel matte, oxygen pressed slag and matte forming agent is 10: (1-30): (1-25): (0.01-30).

[0028] In some embodiments of the present disclosure, the temperature of matte smelting is 1000° C.-1500° C., and the time of matte smelting is 1 hour-10 hours.

[0029] In some embodiments of the present disclosure, the matte-forming agent includes a sulfur-containing element and / or a sulfur-containing compound.

[0030] In some embodiments of the present disclosure, after the first leaching and separation process, the process further includes a drying process for drying the leached residue, nickel matte, and oxygen-pressed slag respectively.

[0031] In some embodiments of the present disclosure, the drying temperature is 100° C.-300° C., the drying time is 0.5 h-4 h, and air and / or oxygen are introduced during the drying process.

[0032] In some embodiments of the present disclosure, the method for recovering valuable metals further comprises:

[0033] a first slurrying step of mixing matte with a solvent and then performing a first slurrying treatment to obtain a first slurry liquid;

[0034] a second leaching and separation step, wherein the first slurry and the second leaching agent are mixed and then subjected to a second leaching treatment to dissolve the nickel in the first slurry into the solution to obtain a second leaching slurry containing nickel, and the second leaching slurry is subjected to a solid-liquid separation treatment to obtain a second filtrate and a second filter residue;

[0035] In the extraction step, the second filtrate is mixed with an extractant and subjected to extraction treatment to obtain an extract and a raffinate.

[0036] In some embodiments of the present disclosure, in the first slurrying process, the matte is ground, and the ground matte is mixed with a solvent at a solid-liquid ratio of 1 g: (2-15) mL to perform a first slurrying process to obtain a first slurry liquid.

[0037] In some embodiments of the present disclosure, the particle size of the ground matte is less than or equal to 48 μm.

[0038] In some embodiments of the present disclosure, the temperature of the first slurrying treatment is 30° C.-90° C., and the time of the first slurrying treatment is 0.5 h-5 h.

[0039] In some embodiments of the present disclosure, in the second leaching and separation process, the second leaching agent includes concentrated sulfuric acid, wherein the amount of concentrated sulfuric acid added is 1 to 1.6 times the theoretical amount required to completely leach out the nickel in the first slurry.

[0040] In some embodiments of the present disclosure, the temperature of the second leaching treatment is 50° C.-100° C., and the time of the second leaching treatment is 2 h-8 h.

[0041] In some embodiments of the present disclosure, the method for recovering valuable metals further comprises:

[0042] a third leaching and separation step, mixing the second filter residue and a third leaching agent, performing a third leaching treatment to obtain a third leaching slurry, and performing a solid-liquid separation treatment on the third leaching slurry to obtain a third filtrate and a third filter residue;

[0043] a second slurrying step of mixing the third filter residue with a solvent and then performing a second slurrying treatment to obtain a second slurry liquid;

[0044] In the fourth leaching and separation step, oxygen is introduced into the second slurry to perform oxygen pressure leaching treatment to obtain a fourth leaching slurry, and the fourth leaching slurry is subjected to solid-liquid separation treatment to obtain a fourth filtrate and a fourth filter residue.

[0045] In some embodiments of the present disclosure, in the third leaching and separation step, the third leaching agent includes concentrated sulfuric acid, wherein the amount of concentrated sulfuric acid added is 1 to 1.3 times the theoretical amount required to completely leach out the nickel in the second filter residue.

[0046] In some embodiments of the present disclosure, in the third leaching and separation process, the temperature of the third leaching treatment is 40° C.-100° C., and the time of the third leaching treatment is 0.5 h-4 h;

[0047] In some embodiments of the present disclosure, in the second slurrying step, the third filter residue and the solvent have a solid-liquid ratio of 1 g: (3-15) mL.

[0048] In some embodiments of the present disclosure, in the fourth leaching and separation step, the oxygen partial pressure of the oxygen pressure leaching is 0.6 MPa-1.2 MPa.

[0049] In some embodiments of the present disclosure, in the fourth leaching and separation process, the temperature of the oxygen pressure leaching is 150° C.-220° C., and the time of the oxygen pressure leaching is 2 h-8 h. BRIEF DESCRIPTION OF THE DRAWINGS

[0050] In order to more clearly illustrate the specific embodiments of the present disclosure or the technical solutions in the prior art, the drawings required for use in the specific embodiments or the description of the prior art will be briefly introduced below. Obviously, the drawings described below are some embodiments of the present disclosure. For those skilled in the art, other drawings can be obtained based on these drawings without any creative work.

[0051] FIG1 is a flow chart of a method for recovering valuable metals according to some embodiments of the present disclosure;

[0052] FIG2 is a flow chart of a valuable metal recovery method provided in some embodiments of the present disclosure;

[0053] FIG3 is a flow chart of a method for recovering valuable metals according to some embodiments of the present disclosure;

[0054] FIG4 is a flow chart of a valuable metal recovery method provided in some embodiments of the present disclosure;

[0055] FIG5 is a flow chart of a valuable metal recovery method provided by some embodiments of the present disclosure. DETAILED DESCRIPTION

[0056] " scope " disclosed in the present disclosure is limited in the form of lower limit and upper limit, and given range is limited by selecting a lower limit and an upper limit, and selected lower limit and upper limit define the boundary of special scope.The scope that this mode limits can be to include end value or not include end value, and can be combined arbitrarily, and promptly any lower limit can form a scope with any upper limit combination.For example, if the scope of 60-120 and 80-110 is listed for specific parameter, it is understood that the scope of 60-110 and 80-120 is also expected.In addition, if the minimum range value 1 and 2 listed, and if the maximum range value 3,4 and 5 listed, then the following scope can all be expected: 1-3, 1-4, 1-5, 2-3, 2-4 and 2-5.In the present disclosure, unless otherwise specified, numerical range " ab " represents the abbreviation of any real number combination between a and b, and wherein a and b are all real numbers. For example, a numerical range of "0-5" indicates that all real numbers between "0-5" are listed herein, and "0-5" is merely an abbreviation for combinations of these values. Furthermore, when a parameter is expressed as an integer ≥ 2, this is equivalent to disclosing that the parameter is, for example, an integer of 2, 3, 4, 5, 6, 7, 8, 9, 10, 11, and 12.

[0057] Unless otherwise specified, all embodiments and optional embodiments of the present disclosure can be combined with each other to form new technical solutions.

[0058] Unless otherwise specified, all technical features and optional technical features disclosed herein can be combined with each other to form a new technical solution.

[0059] Unless otherwise specified, all steps of the present disclosure may be performed sequentially or randomly, preferably sequentially. For example, the method includes steps (a) and (b), which means that the method may include steps (a) and (b) performed sequentially, or may include steps (b) and (a) performed sequentially. For example, the method may further include step (c), which means that step (c) may be added to the method in any order, for example, the method may include steps (a), (b) and (c), or may include steps (a), (c) and (b), or may include steps (c), (a) and (b), etc.

[0060] Unless otherwise specified, the terms "include" and "comprising" mentioned in this disclosure may be open-ended or closed-ended. For example, the terms "include" and "comprising" may mean that other components not listed may also be included or that only the listed components are included.

[0061] Unless otherwise specified, the term "or" is inclusive in this disclosure. For example, the phrase "A or B" means "A, B, or both A and B." More specifically, the condition "A or B" is satisfied if any of the following conditions are met: A is true (or exists) and B is false (or does not exist); A is false (or does not exist) and B is true (or exists); or both A and B are true (or exist).

[0062] In the present disclosure, battery recyclate refers to a mixture of major battery components such as positive electrodes recovered from batteries, wherein the batteries may be nickel-sulfur secondary batteries, nickel-metal hydride secondary batteries, lithium-ion secondary batteries, etc. In some examples, the positive electrode in a lithium-ion secondary battery includes a positive electrode current collector and a positive electrode active material. The positive electrode current collector may be a metal foil such as aluminum, copper, or nickel, and the positive electrode active material may include valuable metals such as nickel (Ni), cobalt (Co), manganese (Mn), and lithium (Li). In some specific examples, battery recyclate may be a mixture containing valuable metals such as Ni, Co, Mn, and Li, and carbon powder, obtained through processes such as disassembly, crushing, screening, pyrolysis, and sorting.

[0063] Lithium extraction agent refers to a compound that can convert lithium metal in battery recycling into a compound that is easily soluble in water or a weakly acidic solution environment.

[0064] Nickel matte refers to a mixture of nickel and other metal sulfides, such as Ni3S2. In some examples, nickel matte can be an intermediate product in the nickel smelting process, such as low-grade nickel matte and high-grade nickel matte. The nickel content in low-grade nickel matte is generally below 40%, while the nickel content in high-grade nickel matte is generally above 60%.

[0065] Oxygen-pressed slag refers to slag containing elements such as nickel, iron, and oxygen. In some examples, oxygen-pressed slag can be produced through an oxygen-pressing process, primarily composed of iron oxide, with a combined nickel and cobalt content of approximately 2% to 15%. The oxygen-pressing process typically utilizes oxygen or air (oxygen in air) as an oxidant, which is thoroughly mixed with the slurry under stirring to oxidize low-valent minerals such as reduced sulfide minerals, leaching the target metal into solution and efficiently extracting the target metal from the ore.

[0066] Referring to FIG1 , the present disclosure provides a method for recovering valuable metals, the method comprising:

[0067] S100, a roasting process, roasting the battery recycling material and the lithium extraction agent under a protective atmosphere to obtain a roasted product.

[0068] S200, a first leaching and separation step, performing a first leaching and separation on the roasted product to obtain a lithium-containing leachate and a leaching residue.

[0069] S400, a calcination step, mixing the leached slag and the slag-forming agent and calcining them to obtain a calcined product;

[0070] S500, matte making and smelting process, mixing the calcined product, nickel matte, oxygen pressed slag and matte making agent and then performing matte making and smelting to obtain nickel-containing matte.

[0071] The recovery method provided by the disclosed embodiments, through a roasting step and a first leaching and separation step, can significantly enrich valuable lithium metals in the battery recyclate in the leachate, thereby improving the lithium recovery rate. The leached residue obtained from the first leaching and separation step undergoes a calcination step for a slagging reaction, pre-removing some impurities, such as Ca and Mg, that enter the slagging agent. This helps shorten the matte smelting time. The calcined product, nickel matte, oxygen-pressed slag and matte-making agent obtained through the calcination process are mixed and then subjected to matte-making smelting, which can separate the valuable metal nickel in the calcined product, nickel matte and oxygen-pressed slag from other impurities, that is, the valuable metal nickel enters the sulfur phase to obtain nickel-containing matte. In this way, valuable metals such as nickel in the calcined product, nickel matte and oxygen-pressed slag can be recovered at the same time, simplifying the process of recovering valuable metals from battery recyclables, nickel matte and oxygen-pressed slag, and increasing the recovery rate of valuable metal nickel. In addition, the large amount of oxygen element contained in nickel matte, oxygen-pressed slag and calcined products can be used to reduce the amount of oxygen required for external introduction during the sulfur-making process. This can further reduce the cost of recovering valuable metals while reducing the safety risks brought about by the large amount of oxygen introduced from the outside and the requirements for fire protection and recovery equipment in the factory. Moreover, impurities such as iron and manganese in the calcined products, nickel matte and oxygen-pressed slag will enter the slag phase, which can be directly used in fields such as construction as a by-product. This can reduce the amount of impurity removal in subsequent production lines, thereby helping to reduce the demand for impurity removal equipment and the floor space of the recycling plant, thereby further helping to reduce the recovery cost of valuable metals in battery recyclates, nickel matte and oxygen-pressed slag. In addition, in the above-mentioned processes, few pollutants are generated, thereby making the recovery method of valuable metals environmentally friendly. Therefore, the recovery method provided by the embodiment of the present disclosure has a simple process flow, a high recovery rate, and few pollutants generated, thereby making the recovery cost of valuable metals low and environmentally friendly.

[0072] In an embodiment of the present disclosure, the lithium extraction agent used in the roasting process can convert the lithium metal in the battery recyclate into a compound that is easily soluble in water or a weakly acidic solution environment. In some embodiments, the lithium extraction agent may include one or more of a reducing agent, a sulfate or concentrated sulfuric acid, wherein the reducing agent mainly undergoes a reduction reaction during the roasting process to form carbonate or bicarbonate, and the carbonate and bicarbonate can form soluble lithium carbonate or lithium bicarbonate with the lithium ions in the battery recyclate, thereby facilitating the leaching and recovery of lithium. Exemplarily, the reducing agent may include one or more of activated carbon, hydrogen, lignite, anthracite, and carbon monoxide, wherein the activated carbon may include coconut shell powdered activated carbon and / or wood activated carbon. The above-mentioned reducing agents are cost-effective and can help reduce the cost of lithium extraction while also helping to convert lithium ions in the battery recyclate into soluble lithium carbonate or lithium bicarbonate to improve the recovery rate of lithium.

[0073] Concentrated sulfuric acid and sulfates primarily react with lithium ions in the battery recyclate during the roasting process to form soluble lithium sulfate, thereby facilitating lithium leaching and recovery. For example, sulfates may include one or more of ammonium sulfate and sodium sulfite.

[0074] In this embodiment, the mass ratio of the lithium extraction agent to the battery recyclate is within a suitable range, which can also help improve the lithium recovery rate. In some embodiments, the mass ratio of the lithium extraction agent to the battery recyclate is 20:(1-30). When the mass ratio of the lithium extraction agent to the battery recyclate is within the above suitable range, it can increase the utilization rate of the lithium extraction agent and improve the lithium recovery rate.

[0075] In other embodiments, the mass ratio of the lithium extraction agent to the battery recyclate can also be 20:(3-28), 20:(5-24), 20:(8-21), etc.

[0076] Furthermore, in the calcination step of S100, the calcination parameters are within an appropriate range to facilitate the thermodynamic reaction of lithium ions to form soluble salts. In some embodiments of the present disclosure, the calcination temperature is 300°C to 1000°C, the heating rate is 2°C / min to 10°C / min, and the calcination time is 0.5h to 5h.

[0077] Furthermore, calcination under a protective atmosphere can facilitate oxidation of carbon in the battery recyclate, thereby helping to form soluble lithium carbonate or lithium bicarbonate with lithium ions, thereby improving the lithium recovery rate. In this embodiment, the gas used in the protective atmosphere can be any protective gas known in the art, such as an inert gas or nitrogen, wherein the inert gas can include helium (He), neon (Ne), argon (Ar), etc.

[0078] Referring to FIG. 2 , in some embodiments, the first leaching and separation step of S200 includes:

[0079] S210, a first leaching step of mixing the roasted product with a first leaching agent and performing a first leaching treatment to dissolve lithium in the roasted product into a solution to form a first leaching slurry;

[0080] S220, a first separation step, performing solid-liquid separation on the first leaching slurry to obtain lithium-containing leaching solution and leaching residue.

[0081] In the above-described embodiments, the first leaching step (S210) uses a solvent to mix and slurry the calcined product, providing a favorable leaching environment. The addition of a first leaching agent dissolves the lithium in the calcined product into a solution, forming a first leaching slurry. The first leaching slurry is then subjected to a solid-liquid separation process (S220) to produce a lithium-containing leachate and leaching residue.

[0082] In some embodiments, the solvent used in the first leaching step of S210 can be water, an acidic washing solution generated during the valuable metal recovery process, or a mixed solvent of water and the acidic washing solution. The acidic washing solution as a solvent can facilitate the reuse of waste liquid and achieve a pre-leaching effect, thereby reducing the amount of the first leaching agent added.

[0083] In some embodiments, in the first leaching step of S210, the liquid-to-solid ratio of the solvent to the calcined product is (1-10) mL:1 g. When the liquid-to-solid ratio of the solvent to the calcined product is within this range, it is beneficial for lithium leaching while also ensuring that the solvent content in the leached residue obtained by subsequent separation is within a reasonable range, thereby facilitating subsequent reactions of the leached residue.

[0084] In this embodiment, the liquid-to-solid ratio generally refers to the ratio of the volume of the solvent to the mass of the calcined product.

[0085] In some embodiments, in the first leaching step of S210 , the pH of the first leaching treatment is between 3 and 7. When the pH of the first leaching treatment is within the above range, lithium can be easily ionized and exist in the aqueous solution, thereby reducing the recovery cost of lithium, thereby helping to reduce the recovery cost of valuable metals.

[0086] In the above embodiments, the addition of the first leaching agent can make the pH of the first leaching treatment within the range of 3-7, and the first leaching agent can be any compound that can make the pH of the first leaching treatment within the above range, such as concentrated sulfuric acid.

[0087] Furthermore, in the first leaching step of S210 , the temperature of the first leaching treatment is between 30° C. and 90° C., and the duration of the first leaching treatment is between 0.5 and 4 hours. When the temperature of the first leaching treatment is within this range, it is beneficial for the occurrence of leaching reaction kinetics. When the duration of the first leaching treatment is within this range, it is helpful to improve the leaching reaction effect.

[0088] In the first separation step of S220 , the solid-liquid separation treatment method can be any method for separating solids and liquids known in the art, for example, the solid-liquid separation treatment can be performed by centrifugation, tilting, filtration, and the like.

[0089] Furthermore, the leached slag obtained through the solid-liquid separation treatment is mixed with a slag-forming agent and then calcined to initiate a slag-forming reaction. This allows some impurities (e.g., Ca and Mg) in the leached slag to enter the slag-forming agent, thereby pre-removing some of the impurities in the leached slag. When the leached slag with reduced impurity content is used in matte smelting, the time required for matte smelting is shortened, thereby improving matte smelting efficiency. Furthermore, the calcination step can remove some carbon contained in the leached slag, thereby reducing its impact on the subsequent sulfur-forming reaction.

[0090] In some embodiments, in the calcination process of S400, the mass ratio of the leached slag to the slag-forming agent is (2-20): 1. When the mass ratio of the leached slag to the slag-forming agent is within the above range, it can help improve the slag-forming reaction efficiency to increase the nickel recovery rate while further reducing the recovery cost of valuable metals.

[0091] In some embodiments, the slag-forming agent comprises one or more of silica, diatomaceous earth, bentonite, calcium oxide, and aluminum oxide. It is understood that the slag-forming agent may be any one of silica, diatomaceous earth, bentonite, calcium oxide, and aluminum oxide, or a mixture of any two of these. When the slag-forming agent comprises a mixture, the components of the mixture synergize with each other to selectively promote the incorporation of impurities in the leached residue into the slag phase, thereby reducing impurities in the sulfur phase and improving the recovery rate of valuable metals.

[0092] In some specific embodiments, the calcination process can cause the magnesium and part of the iron in the leached slag to react with the slag-forming agent to make them enter the slag phase, wherein the following slag-forming reactions may occur: 10Fe2O3+FeS=7Fe3O4+SO2; 3Fe3O4+FeS+5SiO2=5(2FeO·SiO2)+SO2; FeO+SiO2=2FeO·SiO2; CaO+SiO2=CaO·SiO2; MgO+SiO2=MgO·SiO2; 2Mn3O4+S=6MnO+SO2; MnO+SiO2=MnO·SiO2.

[0093] In this embodiment, the matte smelting process S500 mixes the calcined product, nickel matte, oxygen-pressed slag, and sulfur-forming agent and then performs matte smelting. This allows impurities in the calcined product, nickel matte, and oxygen-pressed slag to enter the slag phase, and also allows nickel in the calcined product, nickel matte, and oxygen-pressed slag to enter the sulfur phase, thereby simultaneously achieving the recovery of valuable metal nickel in battery recyclate, nickel matte, and oxygen-pressed slag, and also having a high recovery rate of valuable metals. In addition, the oxygen element contained in the oxygen-pressed slag reduces the amount of oxygen introduced during the matte smelting process. This can fully utilize the oxygen-pressed slag to further reduce the cost of valuable metal recovery while also reducing the safety risks brought about by the introduction of a large amount of oxygen from the outside and the requirements for fire protection and recovery equipment in the factory.

[0094] In some embodiments, the matte-forming agent includes elemental sulfur and / or sulfur-containing compounds. The rational combination of elemental sulfur and sulfur-containing compounds with the calcined product, nickel matte, and oxygen-pressed slag can enrich the valuable metallic nickel in the calcined product, nickel matte, and oxygen-pressed slag in the sulfur phase through matte-forming smelting, thereby helping to improve the recovery rate of valuable metals in the calcined product, nickel matte, and oxygen-pressed slag.

[0095] In the above embodiments, the sulfur-containing compound may be one or more of sulfur and sulfide ore.

[0096] In some embodiments, in the matte smelting process of S400, the mass ratio of the calcined product, nickel matte, oxygen-pressed slag, and matte-forming agent is 10:(1-30):(1-25):(0.01-30). When the mass ratio of the calcined product, nickel matte, oxygen-pressed slag, and matte-forming agent is within the above range, the valuable metal nickel in the calcined product, nickel matte, and oxygen-pressed slag can be absorbed into the sulfur phase to form nickel-containing matte, thereby further improving the recovery rate of the valuable metal.

[0097] In some embodiments, the matte smelting temperature is between 1000°C and 1500°C, and the matte smelting time is between 1 hour and 10 hours. When the matte smelting temperature and time are within the above ranges, it is not only beneficial for the thermodynamic chemical reaction to proceed, but also helps to improve the efficiency of the reaction. The reactions that may occur during matte smelting are as follows:

[0098] Decomposition reactions of high-valent sulfides: Fe7S8=7FeS+1 / 2S2; 2CuFeS2=Cu2S+2FeS+1 / 2S2; (Ni,Fe)9S8=(x / 3)Ni3S2+(9-x)FeS+(x / 6-1 / 2)S2; (Ni,Fe)9S8=(x / 3)Ni3S2+(9-x)FeS+(x / 6-1 / 2)S2;

[0099] Oxidation reactions of low-valent sulfides: 2FeS+3O2=2FeO+2SO2; Ni3S2+7 / 2O2=3NiO+2SO2; 3FeS+3NiO=Ni3S2+3FeO+1 / 2S2; 3NiS=Ni3S2+S; Co3S2+7 / 2O2=3CoO+2SO2; 3FeS+3CoO=Co3S2+3FeO+1 / 2S2; 3CoS=Co3S2+S;

[0100] The chemical reactions occurring in recycled batteries are mainly as follows: NiO+S=NiS+1 / 2O2; CoO+S=CoS+1 / 2O2; Ni+S=NiS; Co+S=CoS; MnO+SiO2=MnO·SiO2.

[0101] Through the above reaction, the valuable metal nickel in the calcined product, nickel matte and oxygen-pressed slag can enter the sulfur phase to form matte, while manganese enters the slag phase, which can further remove impurities in the nickel phase, thereby further improving the recovery rate of valuable metal nickel.

[0102] In the above embodiments, the battery may be a ternary battery, that is, the battery contains valuable metals such as Ni, Co, and Mn. The valuable metals such as Ni and Co can be recycled through the above recycling method.

[0103] Referring to FIG. 3 , in some embodiments, after the first leaching and separation step in S200 , the method for recovering valuable metals further includes:

[0104] S300, drying process, drying the leached slag, nickel matte and oxygen pressed slag respectively.

[0105] In the above embodiment, the drying process of drying the leached residue, nickel matte and oxygen-pressed slag can remove part of the moisture in the leached residue, nickel matte and oxygen-pressed slag.

[0106] In some embodiments, the drying temperature is 100° C. to 300° C., and the drying time is 0.5 to 4 hours. The above drying parameters can further reduce the water content in the leached slag, nickel matte, and oxygen-pressed slag, thereby reducing the adverse effects of water on chemical reactions in the subsequent matte smelting process, such as the water vapor generated carrying away part of the sulfiding agent.

[0107] In the above embodiments, introducing air and / or oxygen during the drying process can help accelerate the drying of the leached slag, nickel matte and oxygen-pressed slag.

[0108] As shown in FIG4 , after the matte smelting process, the nickel-containing matte obtained can be sold directly as a commodity, or it can be further processed to separate nickel from other metal elements to prepare raw materials for battery production. Therefore, the present disclosure also provides a method for further processing the nickel-containing matte, including:

[0109] S600, a first slurrying step, mixing the matte with a solvent and performing a first slurrying treatment to obtain a first slurry liquid.

[0110] S700, a second leaching and separation step, wherein the first slurry and the second leaching agent are mixed and then subjected to a second leaching treatment to dissolve the nickel in the first slurry into the solution to obtain a second leaching slurry containing nickel, and the second leaching slurry is subjected to a solid-liquid separation treatment to obtain a second filtrate and a second filter residue;

[0111] S800, mixing the second filtrate and the extractant and performing extraction treatment to obtain an extract and a raffinate.

[0112] In the aforementioned embodiments, the first slurrying step uniformly disperses the matte in the solvent to form a first slurry, facilitating rapid completion of the leaching reaction in the subsequent second leaching and separation step. The second leaching and separation step leaches nickel and, after solid-liquid separation, separates a second filtrate containing nickel from a second filter residue containing other impurities. The second filtrate is then mixed with an extractant and subjected to extraction, extracting some impurities into the extract while concentrating nickel in the raffinate, thereby separating nickel from other metallic impurities.

[0113] In some embodiments, in the first slurrying step of S600 , water may be used as a solvent, which can help reduce the recovery cost of valuable metals.

[0114] In some embodiments, matte is ground, and the ground matte is mixed with a solvent at a solid-liquid ratio of 1 g: (2-15) mL to form a first slurry to obtain a first slurry. The grinding of matte can reduce the particle size of the matte, thereby facilitating its uniform dispersion in the solvent. Furthermore, the grinding process can also provide the matte particles with a larger reaction area, thereby facilitating the subsequent matte leaching reaction. Furthermore, mixing the ground matte with the solvent at the aforementioned solid-liquid ratio can further improve the dispersion of the matte, thereby facilitating nickel leaching.

[0115] In this embodiment, the solid-liquid ratio generally refers to the ratio of the mass of the ground matte to the volume of the solvent.

[0116] In some embodiments, the particle size of the ground matte is less than or equal to 48 μm. This ensures that the matte has an appropriate particle size, further improving its dispersion. Furthermore, a particle size within the aforementioned range provides a suitable reaction area, facilitating nickel leaching.

[0117] In the above embodiment, the ground matte can be screened through a sieve with an appropriate mesh size to ensure that the ground matte has a desired particle size. Large particles that do not pass through the sieve can be further ground to achieve the desired particle size. For example, a sieve with a mesh size of 300 or greater can be used to screen out matte with a particle size of 48 μm or less.

[0118] In some embodiments, the temperature of the first slurrying treatment is 30° C. to 90° C., and the duration of the first slurrying treatment is 0.5 to 5 hours. When the temperature and duration of the first slurrying treatment are within these ranges, it is beneficial to promote thermal motion of molecules and help evenly disperse the ground matte into the solvent, thereby shortening the subsequent leaching reaction time.

[0119] In some embodiments, in the second leaching and separation step of S700, the second leaching agent may include concentrated sulfuric acid, with the amount added being 1 to 1.6 times the theoretical amount required to completely leach the nickel from the first slurry. By properly combining the leaching agents, the nickel leaching rate can be improved while reducing leaching costs.

[0120] In some embodiments, the second leaching treatment temperature is 50-100° C., and the second leaching treatment time is 2-8 hours. When the temperature and time of the second leaching treatment are within the above ranges, it is beneficial to the thermal motion of molecules and helps to improve the leaching efficiency.

[0121] In addition, in some embodiments, in the extraction process of S800, the extractant may include one or more of di(2-ethylhexyl) phosphate (P204), 2-ethylhexyl phosphoric acid (P507), and di(2,4,4-trimethylpentyl)phosphinic acid (C272). The specific type of the extractant can be selected according to the impurity situation of the solution.

[0122] In the above embodiments, after the second filtrate is subjected to extraction treatment, the obtained raffinate includes nickel sulfate, and the extract can also be used to extract impurities from the extractant through the stripping solution (a mixed sulfate system formed by impurities such as Fe, Zn, Ca, Mg, and Cd), and the extractant can be recovered after separation, thereby further reducing the recovery cost of valuable metals.

[0123] In addition, the second filter residue produced by the second leaching and separation step in S700 can be further processed to further recover part of the valuable metal nickel contained in the filter residue and to form byproducts such as oxygen-pressed slag. Referring to FIG5 , in some embodiments, the method for recovering valuable metals further includes:

[0124] S900, a third leaching and separation step, mixing the second filter residue and a third leaching agent, performing a third leaching treatment to obtain a third leaching slurry, and performing a solid-liquid separation treatment on the third leaching slurry to obtain a third filtrate and a third filter residue;

[0125] S1000, a second slurrying step, mixing the third filter residue and the solvent and then performing a second slurrying treatment to obtain a second slurry liquid;

[0126] S2000, a fourth leaching and separation step, introducing oxygen into the second slurry to perform oxygen pressure leaching treatment to obtain a fourth leaching slurry, and performing solid-liquid separation treatment on the fourth leaching slurry to obtain a fourth filtrate and a fourth filter residue.

[0127] In the above embodiment, nickel in the second filter residue can be further recovered to further improve the recovery rate of valuable metals.

[0128] In some embodiments, in the third leaching and separation step of S900, the third leaching agent includes concentrated sulfuric acid, which is added in an amount that is 1 to 1.3 times the theoretical amount required to completely leach the nickel from the second filter residue. By properly combining leaching agents, the nickel leaching rate can be improved while reducing leaching costs.

[0129] In some embodiments, the temperature of the third leaching treatment is 40° C.-100° C., and the time of the third leaching treatment is 0.5 h-4 h.

[0130] In some embodiments, in the second slurrying step of S1000, the solvent may be water, and the third filter residue and solvent are mixed at a solid-liquid ratio of 1 g: (3-15) ml. This facilitates dispersion of the filter residue, facilitates subsequent leaching reactions, and reduces the possibility of water system imbalance in the system.

[0131] In some embodiments, in the second slurrying step of S1000, the temperature of the second slurrying treatment is 30° C. to 90° C., and the duration of the second slurrying treatment is 0.5 h to 5 h. When the temperature and duration of the second slurrying treatment are within the above ranges, it is beneficial to the thermal motion of molecules and helps to evenly disperse the filter residue in the third solvent, which can help shorten the time of the subsequent leaching reaction.

[0132] In some embodiments, the oxygen partial pressure of oxygen pressure leaching is 0.6 MPa-1.2 MPa. When the oxygen partial pressure of oxygen pressure leaching is within the above range, it can facilitate the oxygen pressure leaching reaction while reducing the safety risks caused by oxygen.

[0133] In some embodiments, the oxygen pressure leaching temperature is 150° C. to 220° C., and the oxygen pressure leaching time is 2 hours to 8 hours. This can help remove the alloy phase in the material and reduce the safety risk caused by the reaction between the alloy and concentrated sulfuric acid during the oxygen pressure reaction.

[0134] In some specific embodiments, the method for recovering valuable metals may include:

[0135] In the roasting process, the battery recycling material and the lithium extraction agent are mixed evenly and then roasted under protective gas to obtain a roasted product;

[0136] a first leaching step of mixing the roasted product and a first leaching agent and performing a first leaching treatment to dissolve lithium in the roasted product into a solution to form a first leaching slurry;

[0137] a first separation step of performing solid-liquid separation on the first leaching slurry to obtain a lithium-containing leachate and a leach residue, wherein the lithium-containing leachate can be used as a lithium source for producing a positive electrode active material in a lithium-ion secondary battery;

[0138] Drying process: drying the leached residue, nickel matte and oxygen pressed slag separately;

[0139] Calcination process: In the calcination process, the leached slag and the slag-making agent are mixed and calcined to obtain a calcined product;

[0140] Matte smelting process: In the matte smelting process, the calcined product, nickel matte, oxygen-pressed slag and matte-making agent are mixed and then smelted to obtain nickel-containing matte, which can be used as a material for producing ternary precursors;

[0141] In the first slurrying step, the matte is mixed with a solvent and then subjected to a first slurrying treatment to obtain a first slurry liquid.

[0142] a second leaching and separation step, wherein the first slurry and the second leaching agent are mixed and then subjected to a second leaching treatment to dissolve the nickel in the first slurry into the solution to obtain a second leaching slurry containing nickel, and the second leaching slurry is subjected to a solid-liquid separation treatment to obtain a second filtrate and a second filter residue;

[0143] An extraction step, wherein the second filtrate is mixed with an extractant and subjected to an extraction treatment to obtain an extract and a raffinate containing nickel, wherein the raffinate can be used as a ternary precursor material;

[0144] a third leaching and separation step, mixing the second filter residue and a third leaching agent and then performing a third leaching treatment to obtain a third leaching slurry, and performing a solid-liquid separation treatment on the third leaching slurry to obtain a third filtrate and a third filter residue, wherein the third filtrate can be recycled, for example, can be added to the leaching and separation step again;

[0145] a second slurrying step of mixing the third filter residue with a solvent and then performing a second slurrying treatment to obtain a second slurry liquid;

[0146] In the fourth leaching and separation step, oxygen is introduced into the second slurry to perform oxygen pressure leaching treatment to obtain a fourth leaching slurry, and the fourth leaching slurry is subjected to solid-liquid separation treatment to obtain a fourth filtrate containing nickel and a fourth filter residue. The fourth filtrate can be used as a ternary precursor material, and the fourth filter residue can be directly used in building materials.

[0147] The following examples describe the present disclosure in more detail. These examples are intended to be illustrative only, as various modifications and variations within the scope of the present disclosure will be apparent to those skilled in the art. Unless otherwise stated, all parts, percentages, and ratios reported in the following examples are by mass, and all reagents and starting materials used in the examples are commercially available or synthesized according to conventional methods, as well as the instruments used in the examples are commercially available.

[0148] Example 1

[0149] This embodiment provides a method for recovering valuable metals, comprising the following steps:

[0150] In the calcination process, 200 g of battery recycled material (the mass fractions of each component in the battery recycled material are shown in Table 1) and 10 g of coconut shell powdered activated carbon were mixed evenly and placed in a crucible. The mixture was calcined at 400 °C for 0.5 h under inert gas protection to obtain a calcined product at a heating rate of 2 °C / min.

[0151] In the first leaching and separation step, the calcined product is added to pure water at a liquid-solid ratio of 5 ml:1 g, concentrated sulfuric acid is added to adjust the initial solution pH to 4±0.1, and the reaction is stirred at 40°C for 2 hours. During the process, the pH is monitored in real time and acid is added in time to ensure that the system pH is maintained at 4±0.1 for the first leaching treatment. After the leaching is completed, a lithium-containing leachate and a leach residue are obtained by filtration. Among them, the Li leaching rate in the leachate is 96.18%, the Ni leaching rate is 4.98%, the Co leaching rate is 0.44%, and the Mn leaching rate is 0.5%.

[0152] In the drying process, the leached residue, low nickel matte and oxygen pressed residue are dried separately. Air is introduced during the drying process. The drying temperature is 100°C, the drying time is 4 hours, and the moisture content of the dried material is 0.5%.

[0153] In the calcination process, the dried leaching residue is mixed with a slagging agent in a mass ratio of 8:1 and then calcined. The slagging agent is silica and bentonite. The calcination temperature is 850°C and the calcination time is 6 hours.

[0154] In the matte smelting process, the calcined product, nickel matte, oxygen-pressed slag and sulfur are mixed in a mass ratio of 10:20:2:10 and then matte smelting is carried out. The matte smelting temperature is 1450°C and the matte smelting time is 1.5 hours to obtain matte containing nickel and cobalt. Among them, the composition of low nickel matte and oxygen-pressed slag is shown in Table 1. It is calculated that the Ni recovery rate is 99.23% and the Co recovery rate is 99.85%.

[0155] The method for recovering valuable metals provided in this embodiment also includes:

[0156] In the first slurrying step, the obtained matte is ball-milled and sieved through 300 mesh. The sieved matte is mixed with water at a solid-liquid ratio of 1g:9mL for the first slurrying treatment to obtain a first slurry liquid, wherein the temperature of the first slurrying treatment is 30°C and the time of the first slurrying treatment is 2h.

[0157] In the second leaching and separation step, concentrated sulfuric acid is added to the first slurry to perform a second leaching treatment. The amount of sulfuric acid added is 1.1 times the theoretical amount. The temperature of the second leaching treatment is 70°C and the time of the second leaching treatment is 4 hours. After the second leaching treatment, the product is filtered to obtain a second filtrate and a second filter residue.

[0158] In the extraction process, the second filtrate is mixed with a P507 extractant and then subjected to extraction treatment to obtain an extract and a raffinate, wherein the raffinate is a nickel sulfate and cobalt sulfate solution.

[0159] In the third leaching and separation step, concentrated sulfuric acid is added to the second filter residue for a third leaching treatment, the amount of sulfuric acid added is 1.1 times the theoretical amount, the temperature of the third leaching treatment is 80° C., and the time of the third leaching treatment is 2 hours. After the third leaching treatment, the product is filtered to obtain a third filtrate and a third filter residue, wherein the third filtrate contains nickel sulfate and cobalt sulfate.

[0160] In the second slurrying step, the third filter residue and water are subjected to a second slurrying treatment at a solid-liquid ratio of 1 g:10 mL to obtain a second slurry liquid.

[0161] In the fourth leaching and separation step, oxygen is introduced into the second slurry for oxygen pressure leaching. The oxygen pressure leaching process is performed at a temperature of 185°C, an oxygen partial pressure of 0.8 MPa, and a duration of 6 hours. After the oxygen pressure leaching process, pressure filtration is performed. The resulting fourth filtrate contains nickel sulfate and cobalt sulfate, and the fourth filter residue contains metals such as calcium, which can be directly sold as a construction material byproduct.

[0162] See Table 1 for the ingredients of the raw materials in the examples of this disclosure.

[0163] Table 1 Composition of raw materials

[0164] It should be noted that the components of battery recyclate, low-grade nickel matte and oxygen-pressed slag used in the embodiments of the present disclosure are merely exemplary and should not be understood as a limitation on the idea of ​​the present disclosure, namely, recycling valuable metals such as lithium, nickel and cobalt in battery recyclate, low-grade nickel matte and oxygen-pressed slag.

[0165] Example 2

[0166] This embodiment is similar to the first embodiment in that it uses recycled battery materials, low-grade nickel matte, and oxygen-pressed slag. The difference lies in other process parameters. Specifically, the following steps are included:

[0167] In the roasting process, 200 g of battery recycled material and 30 g of wood activated carbon were mixed evenly and placed in a crucible. The mixture was roasted at 800 °C for 2 h under inert gas protection to obtain a roasted product at a heating rate of 10 °C / min.

[0168] In the first leaching and separation step, the calcined product was added to pure water at a liquid-solid ratio of 2 ml:1 g, concentrated sulfuric acid was added to adjust the initial solution pH to 6.5±0.1, and the mixture was stirred at 60°C for 2 hours. During the process, the pH was monitored in real time and acid was added in time to ensure that the system pH was maintained at 6.5±0.1 for the first leaching treatment. After the leaching was completed, a lithium-containing leachate and a leach residue were obtained by filtration. Among them, the Li leaching rate in the leachate was 97.36%, the Ni leaching rate was 3.49%, the Co leaching rate was 0.72%, and the Mn leaching rate was 0.81%.

[0169] In the drying process, the leached residue, low nickel matte and oxygen pressed residue are dried separately. Air is introduced during the drying process. The drying temperature is 150°C and the drying time is 2 hours. The moisture content of the dried material is 0.25%.

[0170] In the calcination process, the dried leaching residue is mixed with a slagging agent in a mass ratio of 10:1 and then calcined. The slagging agent is silicon dioxide, bentonite, and calcium oxide. The calcination temperature is 800°C and the calcination time is 2 hours.

[0171] In the matte smelting process, the calcined product, low-grade nickel matte, oxygen-pressed slag and liquid sulfur are mixed in a mass ratio of 2:3:2:4 and then smelted. The matte smelting temperature is 1300°C and the matte smelting time is 2 hours to obtain matte containing nickel and cobalt. The Ni recovery rate is calculated to be 99.12% and the Co recovery rate is 99.29%.

[0172] The method for recovering valuable metals provided in this embodiment also includes:

[0173] In the first slurrying step, the obtained matte is ball-milled and sieved through 300 mesh. The sieved matte is mixed with water at a solid-liquid ratio of 1g:6mL for the first slurrying treatment to obtain a first slurry liquid. The temperature of the first slurrying treatment is 60°C, and the time of the first slurrying treatment is 1.5h.

[0174] In the second leaching and separation step, concentrated sulfuric acid is added to the first slurry to perform a second leaching treatment. The amount of sulfuric acid added is 1.2 times the theoretical amount. The temperature of the second leaching treatment is 70° C. and the time of the second leaching treatment is 6 hours. After the second leaching treatment, the product is filtered to obtain a second filtrate and a second filter residue.

[0175] In the extraction process, the second filtrate is mixed with a P507 extractant and then subjected to extraction treatment to obtain an extract and a raffinate, wherein the raffinate is a nickel sulfate and cobalt sulfate solution.

[0176] In the third leaching and separation step, concentrated sulfuric acid is added to the second filter residue for a third leaching treatment, the amount of sulfuric acid added is 1.2 times the theoretical amount, the temperature of the third leaching treatment is 60° C., and the time of the third leaching treatment is 2 hours. After the third leaching treatment, the product is filtered to obtain a third filtrate and a third filter residue, wherein the third filtrate contains nickel sulfate and cobalt sulfate.

[0177] In the second slurrying step, the third filter residue and water are subjected to a second slurrying treatment at a solid-liquid ratio of 1 g:8 mL to obtain a second slurry liquid.

[0178] In the fourth leaching and separation step, oxygen is introduced into the second slurry liquid for oxygen pressure leaching treatment, wherein the oxygen pressure leaching temperature is 185°C, the oxygen partial pressure is 0.85 MPa, and the oxygen pressure leaching time is 4 hours. After the oxygen pressure leaching is completed, pressure filtration is performed to obtain a fourth filtrate containing nickel sulfate and cobalt sulfate and a fourth filter residue containing metals such as Ca.

[0179] Example 3

[0180] This embodiment is similar to the first embodiment in that it uses recycled battery materials, low-grade nickel matte, and oxygen-pressed slag. The difference lies in other process parameters. Specifically, the following steps are included:

[0181] In the calcination process, 200 g of battery recycling, 180 g of ammonium sulfate and concentrated sulfuric acid were mixed evenly and placed into a crucible. The mixture was calcined at 750 ° C for 2 h under inert gas protection to obtain a calcined product at a heating rate of 5 ° C / min.

[0182] In the first leaching and separation step, the calcined product is added to pure water at a liquid-solid ratio of 2 ml:1 g, concentrated sulfuric acid is added to adjust the initial solution pH to 6.5±0.1, and the reaction is stirred at 60°C for 2 hours. During the process, the pH is monitored in real time and acid is added in time to ensure that the system pH is maintained at 6.5±0.1 for the first leaching treatment. After the leaching is completed, a lithium-containing leachate and a leach residue are obtained by filtration. Among them, the Li leaching rate in the leachate is 99.28%, the Ni leaching rate is 3.86%, the Co leaching rate is 1.2%, and the Mn leaching rate is 0.36%.

[0183] In the drying process, the leached residue, low nickel matte and oxygen pressed residue are dried separately. Air is introduced during the drying process. The drying temperature is 200°C, the drying time is 1.5 hours, and the moisture content of the dried material is 0.15%.

[0184] In the calcination process, the dried leached slag is mixed with a slag-forming agent in a mass ratio of 12:1 and then calcined. The slag-forming agent is a mixture of silicon dioxide and calcium oxide. The calcination temperature is 900°C and the calcination time is 3 hours.

[0185] In the matte smelting process, the calcined product, low-grade nickel matte, oxygen-pressed slag and liquid sulfur are mixed in a mass ratio of 10:2:2:5 and then smelted. The matte smelting temperature is 1250°C and the matte smelting time is 2 hours to obtain matte containing nickel and cobalt. The Ni recovery rate is calculated to be 99.14% and the Co recovery rate is 98.99%.

[0186] The method for recovering valuable metals provided in this embodiment also includes:

[0187] In the first slurrying step, the obtained matte is ball-milled and sieved through 300 mesh. The sieved matte is mixed with water at a solid-liquid ratio of 1g:7mL for the first slurrying treatment to obtain a first slurry liquid. The temperature of the first slurrying treatment is 70°C, and the time of the first slurrying treatment is 2h.

[0188] In the second leaching and separation step, concentrated sulfuric acid is added to the first slurry to perform a second leaching treatment. The amount of sulfuric acid added is 1.15 times the theoretical amount. The temperature of the second leaching treatment is 60° C. and the time of the second leaching treatment is 5 hours. After the second leaching treatment, the product is filtered to obtain a second filtrate and a second filter residue.

[0189] In the extraction process, the second filtrate is mixed with a P507 extractant and then subjected to extraction treatment to obtain an extract and a raffinate, wherein the raffinate is a nickel sulfate and cobalt sulfate solution.

[0190] In the third leaching and separation step, concentrated sulfuric acid is added to the second filter residue for a third leaching treatment, the amount of sulfuric acid added is 1.3 times the theoretical amount, the temperature of the third leaching treatment is 60° C., and the time of the third leaching treatment is 2 hours. After the third leaching treatment, the product is filtered to obtain a third filtrate and a third filter residue, wherein the third filtrate contains nickel sulfate and cobalt sulfate.

[0191] In the second slurrying step, the third filter residue and water are subjected to a second slurrying treatment at a solid-liquid ratio of 1 g:12 mL to obtain a second slurry liquid.

[0192] In the fourth leaching and separation step, oxygen is introduced into the second slurry liquid for oxygen pressure leaching treatment, wherein the oxygen pressure leaching temperature is 185°C, the oxygen partial pressure is 1 MPa, and the oxygen pressure leaching time is 3 hours. After the oxygen pressure leaching is completed, pressure filtration is performed to obtain a fourth filtrate containing nickel sulfate and cobalt sulfate and a fourth filter residue containing metals such as Ca.

[0193] Example 4

[0194] This embodiment is similar to the first embodiment in that it uses recycled battery materials, low-grade nickel matte, and oxygen-pressed slag. The difference lies in other process parameters. Specifically, the following steps are included:

[0195] In the roasting process, 200 g of battery recycled material and 300 g of concentrated sulfuric acid were mixed evenly and then loaded into a crucible. The mixture was roasted at 1000°C for 5 h under inert gas protection to obtain a roasted product at a heating rate of 8°C / min.

[0196] In the first leaching and separation step, the roasted product is added to pure water at a liquid-solid ratio of 10 ml:1 g, concentrated sulfuric acid is added to adjust the initial solution pH to 7±0.1, and the reaction is stirred at 90°C for 4 hours. During the process, the pH is monitored in real time and acid is added in time to ensure that the system pH is maintained at 7±0.1 for the first leaching treatment. After the leaching is completed, a lithium-containing leachate and a leach residue are obtained by filtration. Among them, the Li leaching rate in the leachate is 99.89%, the Ni leaching rate is 4.50%, the Co leaching rate is 2.1%, and the Mn leaching rate is 1.34%.

[0197] In the drying process, the leached residue, low nickel matte and oxygen pressed slag are dried separately. Oxygen is introduced during the drying process. The drying temperature is 300°C, the drying time is 0.5h, and the moisture content of the dried material is 0.05%.

[0198] In the calcination process, the dried leaching residue is mixed with a slagging agent in a mass ratio of 20:1 and then calcined. The slagging agent is a mixture of silicon dioxide, calcium oxide and aluminum oxide. The calcination temperature is 1000°C and the calcination time is 8 hours.

[0199] In the matte smelting process, the calcined product, low-grade nickel matte, oxygen-pressed slag and sulfide ore are mixed in a mass ratio of 10:10:7:6 and then smelted. The matte smelting temperature is 1500°C and the matte smelting time is 5 hours to obtain matte containing nickel and cobalt. The Ni recovery rate is calculated to be 98.22% and the Co recovery rate is 98.89%.

[0200] The method for recovering valuable metals provided in this embodiment also includes:

[0201] In the first slurrying step, the obtained matte is ball-milled and sieved through 300 mesh. The sieved matte is mixed with water at a solid-liquid ratio of 1g:15mL for the first slurrying treatment to obtain a first slurry liquid, wherein the temperature of the first slurrying treatment is 90°C and the time of the first slurrying treatment is 5h.

[0202] In the second leaching and separation step, concentrated sulfuric acid is added to the first slurry to perform a second leaching treatment. The amount of sulfuric acid added is 1.6 times the theoretical amount. The temperature of the second leaching treatment is 100° C. and the time of the second leaching treatment is 4 hours. After the second leaching treatment, the product is filtered to obtain a second filtrate and a second filter residue.

[0203] In the extraction process, the second filtrate is mixed with a P507 extractant and then subjected to extraction treatment to obtain an extract and a raffinate, wherein the raffinate is a nickel sulfate and cobalt sulfate solution.

[0204] In the third leaching and separation step, concentrated sulfuric acid is added to the second filter residue for a third leaching treatment, the amount of sulfuric acid added is 1.3 times the theoretical amount, the temperature of the third leaching treatment is 100° C., and the time of the third leaching treatment is 2 hours. After the third leaching treatment, the product is filtered to obtain a third filtrate and a third filter residue, wherein the third filtrate contains nickel sulfate and cobalt sulfate.

[0205] In the second slurrying step, the third filter residue and water are subjected to a second slurrying treatment according to a solid-liquid ratio of 1 g:15 mL to obtain a second slurry liquid.

[0206] In the fourth leaching and separation step, oxygen is introduced into the second slurry liquid for oxygen pressure leaching treatment, wherein the oxygen pressure leaching temperature is 220°C, the oxygen partial pressure is 1 MPa, and the oxygen pressure leaching time is 8 hours. After the oxygen pressure leaching is completed, pressure filtration is performed to obtain a fourth filtrate containing nickel sulfate and cobalt sulfate and a fourth filter residue containing metals such as Ca.

[0207] Example 5

[0208] This embodiment is similar to the first embodiment in that it uses recycled battery materials, low-grade nickel matte, and oxygen-pressed slag. The difference lies in other process parameters. Specifically, the following steps are included:

[0209] In the roasting process, 200 g of battery recycled material and 200 g of concentrated sulfuric acid were mixed evenly and then loaded into a crucible. The mixture was roasted at 750°C for 2 h under inert gas protection to obtain a roasted product at a heating rate of 7°C / min.

[0210] In the first leaching and separation step, the calcined product is added to pure water at a liquid-solid ratio of 5 ml:1 g, concentrated sulfuric acid is added to adjust the initial solution pH to 6.5±0.1, and the reaction is stirred at 80°C for 2 hours. During the process, the pH is monitored in real time and acid is added in time to ensure that the system pH is maintained at 6.5±0.1 for the first leaching treatment. After the leaching is completed, a lithium-containing leachate and a leach residue are obtained by filtration. Among them, the Li leaching rate in the leachate is 99.78%, the Ni leaching rate is 2.50%, the Co leaching rate is 1.69%, and the Mn leaching rate is 0.65%.

[0211] In the drying process, the leached residue, low nickel matte and oxygen pressed residue are dried separately. Compressed air is introduced during the drying process. The drying temperature is 200°C, the drying time is 2 hours, and the moisture content of the dried material is 0.1%.

[0212] In the calcination process, the dried leaching residue is mixed with a slagging agent in a mass ratio of 5:1 and then calcined. The slagging agent is a mixture of silicon dioxide, calcium oxide and bentonite. The calcination temperature is 750°C and the calcination time is 3 hours.

[0213] In the matte smelting process, the calcined product, low-grade nickel matte, oxygen-pressed slag and sulfide ore are mixed in a mass ratio of 10:20:20:25 and then smelted. The matte smelting temperature is 1250°C and the matte smelting time is 3 hours to obtain matte containing nickel and cobalt. Calculation shows that the Ni recovery rate is 99.85% and the Co recovery rate is 99.56%.

[0214] The method for recovering valuable metals provided in this embodiment also includes:

[0215] In the first slurrying step, the obtained matte is ball-milled and sieved through 300 mesh. The sieved matte is mixed with water at a solid-liquid ratio of 1g:5mL for the first slurrying treatment to obtain a first slurry liquid, wherein the temperature of the first slurrying treatment is 90°C and the time of the first slurrying treatment is 2h.

[0216] In the second leaching and separation step, concentrated sulfuric acid is added to the first slurry to perform a second leaching treatment. The amount of sulfuric acid added is 1.2 times the theoretical amount. The temperature of the second leaching treatment is 80° C. and the time of the second leaching treatment is 8 hours. After the second leaching treatment, the product is filtered to obtain a second filtrate and a second filter residue.

[0217] In the extraction process, the second filtrate is mixed with a P507 extractant and then subjected to extraction treatment to obtain an extract and a raffinate, wherein the raffinate is a nickel sulfate and cobalt sulfate solution.

[0218] In the third leaching and separation step, concentrated sulfuric acid is added to the second filter residue for a third leaching treatment, the amount of sulfuric acid added is 1.2 times the theoretical amount, the temperature of the third leaching treatment is 80° C., and the time of the third leaching treatment is 2 hours. After the third leaching treatment, the product is filtered to obtain a third filtrate and a third filter residue, wherein the third filtrate contains nickel sulfate and cobalt sulfate.

[0219] In the second slurrying step, the third filter residue and water are subjected to a second slurrying treatment according to a solid-liquid ratio of 1 g:5 mL to obtain a second slurry liquid.

[0220] In the fourth leaching and separation step, oxygen is introduced into the second slurry liquid for oxygen pressure leaching treatment, wherein the temperature of the oxygen pressure leaching is 200°C, the oxygen partial pressure is 0.9 MPa, and the oxygen pressure leaching time is 8 hours. After the oxygen pressure leaching is completed, pressure filtration is performed to obtain a fourth filtrate containing nickel sulfate and cobalt sulfate and a fourth filter residue containing metals such as Ca.

[0221] Example 6

[0222] This embodiment is similar to the first embodiment in that it uses recycled battery materials, low-grade nickel matte, and oxygen-pressed slag. The difference lies in other process parameters. Specifically, the following steps are included:

[0223] In the roasting process, 200 g of battery recycled material and 300 g of concentrated sulfuric acid were mixed evenly and then loaded into a crucible. The mixture was roasted at 600 °C for 2 h under inert gas protection to obtain a roasted product at a heating rate of 6 °C / min.

[0224] In the first leaching and separation step, the roasted product is added to pure water at a liquid-solid ratio of 5 ml:1 g, concentrated sulfuric acid is added to adjust the pH of the initial solution to 5±0.1, and the reaction is stirred at 80°C for 1 hour. During the process, the pH is monitored in real time and acid is added in time to ensure that the system pH is maintained at 5±0.1 for the first leaching treatment. After the leaching is completed, a lithium-containing leachate and a leach residue are obtained by filtration. Among them, the Li leaching rate of the leachate is 99.18%, the Ni leaching rate is 3.50%, the Co leaching rate is 3.25%, and the Mn leaching rate is 4.78%.

[0225] In the drying process, the leached residue, low nickel matte and oxygen pressed slag are dried separately. Oxygen is introduced during the drying process. The drying temperature is 150°C and the drying time is 3 hours. The moisture content of the dried material is 0.13%.

[0226] In the calcination process, the dried leaching residue is mixed with a slagging agent in a mass ratio of 7:1 and then calcined. The slagging agent is a mixture of silica, alumina, calcium oxide and bentonite. The calcination temperature is 750°C and the calcination time is 3 hours.

[0227] In the matte smelting process, the calcined product, low-grade nickel matte, oxygen-pressed slag and liquid sulfur are mixed in a mass ratio of 10:30:25:30 and then smelted. The matte smelting temperature is 1250°C and the matte smelting time is 8 hours. Matte containing nickel and cobalt is obtained. The Ni recovery rate is calculated to be 98.23% and the Co recovery rate is 98.65%.

[0228] The method for recovering valuable metals provided in this embodiment also includes:

[0229] In the first slurrying step, the obtained matte is ball-milled and sieved through 300 mesh. The sieved matte is mixed with water at a solid-liquid ratio of 1g:10mL for the first slurrying treatment to obtain a first slurry liquid, wherein the temperature of the first slurrying treatment is 60°C and the time of the first slurrying treatment is 2h.

[0230] In the second leaching and separation step, concentrated sulfuric acid is added to the first slurry to perform a second leaching treatment. The amount of sulfuric acid added is 1.2 times the theoretical amount. The temperature of the second leaching treatment is 80° C. and the time of the second leaching treatment is 2 hours. After the second leaching treatment, the product is filtered to obtain a second filtrate and a second filter residue.

[0231] In the extraction process, the second filtrate is mixed with a P507 extractant and then subjected to extraction treatment to obtain an extract and a raffinate, wherein the raffinate is a nickel sulfate and cobalt sulfate solution.

[0232] In the third leaching and separation step, concentrated sulfuric acid is added to the second filter residue for a third leaching treatment, the amount of sulfuric acid added is 1.2 times the theoretical amount, the temperature of the third leaching treatment is 80° C., and the time of the third leaching treatment is 2 hours. After the third leaching treatment, the product is filtered to obtain a third filtrate and a third filter residue, wherein the third filtrate contains nickel sulfate and cobalt sulfate.

[0233] In the second slurrying step, the third filter residue and water are subjected to a second slurrying treatment at a solid-liquid ratio of 1 g:7 mL to obtain a second slurry liquid.

[0234] In the fourth leaching and separation step, oxygen is introduced into the second slurry liquid for oxygen pressure leaching treatment, wherein the temperature of the oxygen pressure leaching is 200°C, the oxygen partial pressure is 1.2 MPa, and the oxygen pressure leaching time is 8 hours. After the oxygen pressure leaching is completed, pressure filtration is performed to obtain a fourth filtrate containing nickel sulfate and cobalt sulfate and a fourth filter residue containing metals such as Ca.

[0235] Comparative Example 1

[0236] On the basis of Example 1, oxygen was additionally introduced only during the matte smelting process, with an oxygen flow rate of 50 Nm 3 / h, nickel yield was 98.02%, and Co yield was 98.67%.

[0237] Comparative Example 2

[0238] On the basis of Example 1, only the matte smelting process was changed, and oxygen was not added to press the slag during this process. The nickel yield was 97.28% and the Co yield was 97.99%.

[0239] Comparative Example 3

[0240] On the basis of Example 1, only the matte smelting process was changed. No oxygen was added to the slag and additional oxygen was introduced. The oxygen flow rate was 50 Nm 3 / h, nickel yield was 98.12%, and Co yield was 98.93%.

[0241] Test section

[0242] (1) Valuable metal leaching rate test

[0243] Take 5 mL of the leachate and dilute it with 45 mL of concentrated sulfuric acid to obtain the test sample solution;

[0244] 10 mL of sample solution was placed in an inductively coupled plasma emission spectrometer (ICP) for testing. The concentration of each valuable metal in the sample solution was obtained through testing. The concentration of the valuable metal in the leachate was calculated based on the concentration. The leaching rate of the valuable metal (such as Li, Ni, Co, Mn, etc.) was then calculated according to the following formula:

[0245] Leaching rate = (volume of leachate * concentration of valuable metals in the leachate) / (mass of battery recyclate * mass content of valuable metals in battery recyclate).

[0246] (2) Valuable metal recovery test

[0247] Using strong acid to leach valuable metals in the sulfur phase material to obtain a leachate;

[0248] Take 5 mL of the leachate and dilute it with 45 mL of concentrated sulfuric acid to obtain the test sample solution;

[0249] Take 10mL of sample solution and put it into ICP for testing. The concentration of each valuable metal in the sample solution is obtained through testing. The total amount of valuable metals in the leachate is calculated based on the concentration. The recovery rate of valuable metals (such as Ni and Co) is then calculated according to the following formula:

[0250] Mass fraction of Ni / Co in matte = (Ni / Co concentration of matte leachate * leachate volume) / mass of matte material (Note: matte leachate is obtained by completely dissolving matte material in concentrated sulfuric acid.)

[0251] Ni / Co recovery rate = (mass of matte material * mass fraction of Ni / Co in matte) / ((mass content of Ni / Co in nickel matte * mass of nickel matte) + (mass content of Ni / Co in oxygen-pressed slag * mass of oxygen-pressed slag) + (mass content of Ni / Co in battery recycling after lithium extraction * mass of battery recycling after lithium extraction)).

[0252] Table 2 Recovery rates of valuable metals in various examples and comparative examples

[0253] Referring to Table 2, it can be seen from Examples 1 to 6 that by adopting the recycling method provided by the present invention, the recovery rate of lithium in battery recyclate can be as high as over 92%, which is significantly improved compared to the traditional wet lithium extraction rate of only about 85%. At the same time, the recovery rates of nickel and cobalt are also over 92%, which can achieve the purpose of recovering valuable metals such as lithium, nickel and cobalt from battery recyclate, nickel matte and oxygen-pressed slag, and the recovery rate of valuable metals is high and the recovery cost is relatively low.

[0254] Comparing the experimental data of Comparative Example 1 with that of Example 1, although additional oxygen was introduced during the matte smelting process in Comparative Example 1, the recovery rates of nickel and cobalt were not improved. Therefore, it can be seen that there is no need to additionally introduce oxygen during the matte smelting process.

[0255] Comparing the experimental data of Comparative Example 2 with that of Example 1, the nickel and cobalt recovery rate is significantly reduced due to the fact that no oxygen slag is added during the matte smelting process in Comparative Example 2. Furthermore, comparing the experimental data of Comparative Example 3 with that of Example 1 shows that, in Comparative Example 3, no oxygen slag is added during the matte smelting process, and additional oxygen is introduced, yet the nickel and cobalt recovery rate in Comparative Example 3 is still lower than that in Example 1. This shows that, in the matte smelting process, the addition of oxygen slag eliminates the need for additional oxygen, thereby increasing the utilization value of the oxygen slag to reduce the cost of valuable metal recovery, improving the recovery rate of valuable metals nickel and cobalt, and reducing the safety risks associated with the introduction of large amounts of external oxygen, as well as the requirements for fire protection and recovery equipment in the plant.

[0256] The various technical features described above can be combined arbitrarily. Although not all possible combinations of these technical features are described, any combination of these technical features should be considered to be covered by this specification as long as such combination does not conflict.

[0257] Finally, it should be noted that the above embodiments are only used to illustrate the technical solutions of the present disclosure, rather than to limit them. Although the present disclosure has been described in detail with reference to the above embodiments, those skilled in the art should understand that they can still modify the technical solutions described in the above embodiments, or replace some or all of the technical features therein with equivalents. However, these modifications or replacements do not cause the essence of the corresponding technical solutions to deviate from the scope of the technical solutions of the embodiments of the present disclosure. Industrial Applicability

[0258] The recovery method provided by the present disclosure has a simple process flow, a high recovery rate, and generates few pollutants, thereby reducing the recovery cost of valuable metals and being environmentally friendly.

Claims

1. A method for recovering valuable metals, the method comprising: A roasting process is to roast the battery recycling material and the lithium extraction agent under a protective atmosphere to obtain a roasting product; A first leaching and separation step, performing a first leaching and separation on the roasted product to obtain a lithium-containing leaching solution and leaching residue; a calcination step, mixing the leached slag and the slag-making agent and then calcining to obtain a calcined product; In the matte making and smelting process, the calcined product, nickel matte, oxygen pressed slag and matte making agent are mixed and then matte making and smelting is performed to obtain nickel-containing matte.

2. The recycling method according to claim 1, wherein: In the roasting process, the lithium extracting agent includes one or more of a reducing agent, a sulfate or concentrated sulfuric acid; Optionally, the reducing agent includes one or more of activated carbon, hydrogen, lignite, anthracite, and carbon monoxide; Optionally, the sulfate includes one or more of ammonium sulfate and sodium sulfite; Optionally, the mass ratio of the battery recyclate to the lithium extraction agent is 20:(1-30); Optionally, the calcination temperature is 300°C-1000°C, the heating rate is 2°C / min-10°C / min, and the calcination time is 0.5h-5h.

3. The recovery method according to claim 1 or 2, wherein: In the first leaching and separation step, it includes: a first leaching step, mixing the roasted product and a first leaching agent to perform a first leaching treatment, so that the lithium in the roasted product is dissolved into a solution to obtain a first leaching slurry containing lithium; A first separation step, performing solid-liquid separation on the first leaching slurry to obtain the lithium-containing leaching solution and the leaching residue; Optionally, in the first leaching step, the pH of the first leaching treatment is 3-7; Optionally, in the first leaching process, the temperature of the first leaching treatment is 30°C-90°C, and the time of the first leaching treatment is 0.5h-4h.

4. The recovery method according to any one of claims 1 to 3, wherein: In the calcining process, the mass ratio of the leached slag to the slag-forming agent is (2-20):1; Optionally, the calcination temperature is 600°C-1000°C, and the calcination time is 2h-8h; Optionally, the slag-forming agent includes one or more of silicon dioxide, diatomaceous earth, bentonite, calcium oxide, and aluminum oxide.

5. The recovery method according to any one of claims 1 to 4, wherein: In the matte smelting process, the mass ratio of the calcined product, the nickel matte, the oxygen pressed slag and the matte making agent is 10:(1-30):(1-25):(0.01-30); Optionally, the matte-forming agent includes a sulfur-containing element and / or a sulfur-containing compound; Optionally, the temperature of the matte smelting is 1000°C-1500°C, and the time of the matte smelting is 1h-10h.

6. The recovery method according to any one of claims 1 to 5, wherein: After the first leaching and separation step, the method further includes: a drying step of drying the leached residue, nickel matte and oxygen-pressed residue respectively; Optionally, the drying temperature is 100° C.-300° C., the drying time is 0.5 h-4 h, and air and / or oxygen are introduced during the drying process.

7. The recovery method according to any one of claims 1 to 6, wherein: Also includes: A first slurrying step, mixing the matte with a solvent and then performing a first slurrying treatment to obtain a first slurry liquid; a second leaching and separation step, mixing the first slurry and the second leaching agent and performing a second leaching treatment to dissolve the nickel in the first slurry into the solution to obtain a second leaching slurry containing nickel, and performing a solid-liquid separation treatment on the second leaching slurry to obtain a second filtrate and a second filter residue; In the extraction step, the second filtrate is mixed with an extractant and then subjected to extraction treatment to obtain an extract and a raffinate.

8. The recycling method according to claim 7, wherein: In the first slurrying step, the matte is ground, and the ground matte is mixed with the solvent at a solid-liquid ratio of 1 g: (2-15) mL to perform the first slurrying treatment to obtain the first slurry liquid; Optionally, the particle size of the matte after grinding is less than or equal to 48 μm; Optionally, the temperature of the first slurry treatment is 30°C-90°C, and the time of the first slurry treatment is 0.5h-5h; Optionally, in the second leaching and separation step, the second leaching agent includes concentrated sulfuric acid, wherein the amount of the concentrated sulfuric acid added is 1 to 1.6 times the theoretical amount required to completely leach the nickel in the first slurry; Optionally, the temperature of the second leaching treatment is 50°C-100°C, and the time of the second leaching treatment is 2h-8h.

9. The recycling method according to claim 7, wherein: Also includes: a third leaching and separation step, mixing the second filter residue and a third leaching agent, performing a third leaching treatment to obtain a third leaching slurry, and performing a solid-liquid separation treatment on the third leaching slurry to obtain a third filtrate and a third filter residue; a second slurrying step, mixing the third filter residue and the solvent and then performing a second slurrying treatment to obtain a second slurry liquid; The fourth leaching and separation step is to introduce oxygen into the second slurry to perform oxygen pressure leaching treatment to obtain a fourth leaching slurry, and to perform solid-liquid separation treatment on the fourth leaching slurry to obtain a fourth filtrate and a fourth filter residue.

10. The recycling method according to claim 9, wherein: In the third leaching and separation step, the third leaching agent includes concentrated sulfuric acid, wherein the amount of concentrated sulfuric acid added is 1 to 1.3 times the theoretical amount required to completely leach nickel from the second filter residue; Optionally, in the third leaching and separation step, the temperature of the third leaching treatment is 40° C.-100° C., and the time of the third leaching treatment is 0.5 h-4 h; Optionally, in the second slurrying step, the third filter residue and the solvent are in a solid-liquid ratio of 1 g: (3-15) ml; Optionally, in the fourth leaching and separation step, the oxygen partial pressure of the oxygen pressure leaching is 0.6MPa-1.2MPa; Optionally, in the fourth leaching and separation step, the temperature of the oxygen pressure leaching is 150° C.-220° C., and the time of the oxygen pressure leaching is 2 h-8 h.

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