Method for producing precursor and method for producing positive electrode material

Abstract

According to the present invention, a reducing agent is added to an oxide to produce an oxide-containing mixture, the oxide comprising: at least one valuable element selected from the group consisting of Ni and Co; a valuable element Mn; and impurity elements copper and iron. The oxide is reduced by heating the oxide-containing mixture to produce metal. The metal is brought into contact with an acid solution to produce a leachate that contains a valuable element and an impurity element. A sulfiding agent is added to the leachate to precipitate copper in the form of copper sulfide, thereby producing a copper-depleted solution. An oxidizing agent is added to the copper-depleted solution to precipitate iron in the form of iron hydroxide, thereby producing a valuable element solution that contains a valuable element. The valuable element solution, a complexing agent, and an alkaline aqueous solution are introduced into a liquid in a reaction tank to produce a precipitate that contains a valuable elements. The reducing agent comprises at least one agent selected from the group consisting of a C-based reducing agent, an Si-based reducing agent, and an Al-based reducing agent. The reducing agent is added in an amount R that satisfies the specific formula shown in Formula (1).

Description

Method for manufacturing precursor and method for manufacturing cathode material

[0001] This invention relates to a method for producing a precursor and a method for producing a cathode material.

[0002] In recent years, the demand for lithium-ion batteries has been rapidly increasing due to the spread of electric vehicles. In particular, the CO2 emissions of recent years ２ From the perspective of reducing emissions, the demand for electric vehicles that do not use fossil fuels is expected to expand further in the future, and consequently, the demand for lithium-ion batteries is also expected to increase further.

[0003] Generally, the positive electrode material of a lithium-ion battery consists of an oxide (composite oxide) containing nickel (Ni), cobalt (Co), manganese (Mn), etc. A specific example of this composite oxide is LiNiO. ２ LiCoO ２ LiMnO ２ These are some examples. Metal elements such as Ni, Co, and Mn are not abundant globally. Therefore, recovering these metal elements (valuable elements) from the cathode material of waste lithium-ion batteries is extremely beneficial from the standpoint of effective resource utilization. Here, "waste lithium-ion batteries" refers to waste lithium-ion batteries (used products); defective lithium-ion batteries (those generated during the manufacturing process of lithium-ion batteries); etc.

[0004] Lithium-ion batteries are composed of a combination of components such as positive electrode material, negative electrode material, and separator, and also include an electrolyte. Therefore, when recovering valuable elements from the positive electrode material of a waste lithium-ion battery, pretreatment such as removal of the electrolyte, pulverization, and crushing is performed prior to recovery. After such pretreatment, the positive electrode material is separated from the waste lithium-ion battery, and then the valuable elements are recovered from the separated positive electrode material. One example of a treatment for recovering valuable elements is a dry treatment in which the positive electrode material is heated together with a reducing agent to reduce and generate the valuable elements (for example, Patent Document 1).

[0005] Japanese Patent Publication No. 2021-95628

[0006] In dry processing, composite oxide (LiNiO ２, LiCoO ２ , LiMnO ２ ), by reducing, in addition to the metal containing valuable elements (Ni, Co, Mn), slag is produced. At this time, it is required not to reduce Mn as much as possible (to leave Mn in the slag without transferring it to the metal), and to selectively transfer Ni and Co to the metal and recover them in some cases.

[0007] Also, the metal obtained by the dry process may contain impurity elements in addition to valuable elements such as Ni and Co. Examples of the impurity elements include copper (Cu) and iron (Fe) derived from waste lithium-ion batteries. When the metal obtained by the dry process is reused as a positive electrode material for a lithium-ion battery, if this metal contains impurity elements (Cu and Fe), it may deteriorate the battery performance. Therefore, it is desirable to remove the impurity elements as much as possible.

[0008] The present invention has been made in view of the above points, and an object thereof is to provide a novel method for manufacturing a positive electrode material and a precursor thereof used in a lithium-ion battery. More specifically, an object is to provide a novel method for recovering valuable elements such as Ni while removing impurity elements from oxides such as the positive electrode material of a waste lithium-ion battery, and manufacturing a precursor and a positive electrode material containing the recovered valuable elements.

[0009] The inventors, after diligent study, have found that the above objective can be achieved by adopting the following configuration, and have completed the present invention. That is, the present invention provides the following [1] to

[13] . [1] A method for producing a precursor of a positive electrode material used in a lithium-ion battery, comprising: adding a reducing agent to an oxide containing a valuable element, which is at least one selected from the group consisting of nickel and cobalt, and manganese, and impurity elements, which are copper and iron, to obtain a mixed oxide; reducing the oxide by heating the mixed oxide to obtain a metal; contacting the metal with an acid solution to obtain a leachate containing the valuable element and the impurity element; adding a sulfiding agent to the leachate to precipitate copper as copper sulfide; and obtaining the leachate from which copper has been removed. A method for producing a precursor, comprising: obtaining a copper removal solution; adding an oxidizing agent to the copper removal solution to precipitate iron as iron hydroxide; obtaining the copper removal solution from which iron has been removed as a valuable element solution containing the valuable element; introducing the valuable element solution, a complexing agent, and an alkaline aqueous solution into a reaction vessel to obtain a precipitate containing the valuable element; wherein the reducing agent contains at least one selected from the group consisting of a carbon-containing C-type reducing agent, a silicon-containing Si-type reducing agent, and an aluminum-containing Al-type reducing agent, and the amount R of the reducing agent added satisfies the following formula (1). (0.30a - 0.15b + 0.60) × [Ni] ≤ R ≤ (0.33a - 0.17b + 0.67) × [Ni] + (0.33a - 0.17b + 0.67) × [Co] + (0.23a - 0.12b + 0.47) × [Mn] ... (1) However, in the above formula (1), R: Amount of the reducing agent added (unit: moles) [Ni]: Nickel content of the oxide (unit: moles) [Co]: Cobalt content of the oxide (unit: moles) [Mn]: Manganese content of the oxide (unit: moles) a: Molar ratio of the content of the C-based reducing agent in the reducing agent (unit: moles) to the total amount of the reducing agent added (unit: moles) b: The molar ratio of the content of the Si-based reducing agent in the reducing agent (in molar parts) to the total amount of the reducing agent added (in molar parts). [2] A method for producing the precursor described in [1] above, wherein the oxide is obtained from a waste lithium-ion battery.[3] The method for producing a precursor according to [1] or [2], wherein the temperature at which the mixed oxide is heated is 1400°C or higher. [4] The method for producing a precursor according to any one of [1] to [3], wherein the metal obtained by heating the mixed oxide contains the valuable element and the impurity element. [5] The method for producing a precursor according to any one of [1] to [4], wherein the metal is powdered and then brought into contact with the acid solution. [6] The method for producing a precursor according to any one of [1] to [5], wherein the acid solution contains an acid and an oxidizing agent for the acid solution, and the content of the oxidizing agent for the acid solution is 0.5% by volume or more relative to the acid. [7] The method for producing a precursor according to [6], wherein the oxidizing agent for the acid solution is hydrogen peroxide. [8] The method for producing a precursor according to any one of [1] to [7] above, wherein the amount of sulfiding agent added is 1.0 equivalent or more relative to the copper contained in the leachate, and when precipitation of the copper sulfide, the pH of the leachate to which the sulfiding agent has been added is 3.0 or less. [9] The method for producing a precursor according to any one of [1] to [8] above, wherein the oxidizing agent is at least one oxidizing agent A selected from the group consisting of air and ozone, or at least one oxidizing agent B selected from the group consisting of hydrogen peroxide, hypochlorous acid and potassium permanganate, the amount of oxidizing agent A added is 0.1 vvm or more relative to the copper removal solution, the amount of oxidizing agent B added is 0.005 volume% or more relative to the copper removal solution, and when precipitation of the iron hydroxide, the pH of the copper removal solution to which the oxidizing agent has been added is 3.0 or more and 7.0 or less.

[10] The method for producing a precursor according to [9] above, wherein the temperature of the copper removal solution to which the oxidizing agent has been added is 10°C or higher.

[11] The method for producing a precursor according to any one of [1] to

[10] , wherein the alkaline aqueous solution is an aqueous sodium hydroxide solution and the complexing agent is at least one ammonium source selected from the group consisting of ammonia and ammonium salts.

[12] A method for producing a positive electrode material for use in a lithium-ion battery, comprising mixing a precursor obtained by the method for producing a precursor according to any one of [1] to

[11] with a lithium-containing compound, and calcining the resulting mixture to obtain a calcined product containing the valuable element and lithium.

[13] The method for producing a positive electrode material according to

[12] above, wherein the lithium-containing compound is at least one selected from the group consisting of lithium hydroxide and lithium carbonate.

[0010] According to the present invention, a novel method for producing a precursor and a positive electrode material can be provided.

[0011] It is a flowchart showing an example of a method for recovering valuable elements. It is a flowchart showing an example of the process for producing a precursor and a positive electrode material. Potential-pH diagrams of Cu and Ni (S-H ２ O system). Potential-pH diagrams of Fe and Ni (O ２ -H ２ O system).

[0012] [Method for Producing Precursor and Positive Electrode Material] Hereinafter, a method for producing a positive electrode material used in a lithium-ion battery (method for producing a positive electrode material) will be described. The following description also serves as an explanation of a method for producing a precursor used in the positive electrode material (method for producing a precursor). First, as shown in FIG. 1A, valuable elements are recovered from an oxide such as the positive electrode material of a used lithium-ion battery. That is, by subjecting the oxide to dry treatment and wet treatment, a valuable element solution containing valuable elements is obtained. Thereafter, as shown in FIG. 1B, a precursor is produced using the obtained valuable element solution, and then a positive electrode material is produced.

[0013] Figure 1A is a flowchart illustrating an example of a method for recovering valuable elements. Based on Figure 1A, the method for recovering valuable elements will be outlined. In the dry treatment, first, a reducing agent described later is added to an oxide (Ni, Co, Mn, Cu, Fe) in the amount described later to obtain a mixed oxide. Next, the obtained mixed oxide is heated to reduce the oxide and obtain metal (Ni, Co, Cu, Fe) and slag (Mn). The two are separated as appropriate. Before the wet treatment, it is preferable to powder the obtained metal to obtain metal powder (Ni, Co, Cu, Fe). In the wet treatment, first, the metal (metal powder) is brought into contact with an acid solution to obtain a leachate (Ni, Co, Cu, Fe) and leachate residue. The two are separated as appropriate. Next, a sulfiding agent is added to the obtained leachate to precipitate copper sulfide (Cu) and obtain a copper removal solution (Ni, Co, Fe). The two are separated as appropriate. Subsequently, an oxidizing agent is added to the copper removal solution to precipitate iron hydroxide (Fe) and obtain a solution of valuable elements (Ni, Co).

[0014] In this way, impurity elements (Cu, Fe) can be removed from the oxide while valuable elements Ni and Co can be selectively recovered, distinguishing them from Mn, which is also a valuable element. Valuable elements can be easily recovered from the cathode material (oxide) of waste lithium-ion batteries with a purity high enough to be reused as raw materials for lithium-ion batteries.

[0015] Next, we will explain in more detail the methods for recovering valuable elements.

[0016] <Reduction Target (Oxide)> The reduction target is an oxide containing at least one element obtained from the group consisting of nickel (Ni) and cobalt (Co), as well as manganese (Mn), and impurity elements such as copper (Cu) and iron (Fe). Specifically, for example, it is the positive electrode material of a waste lithium-ion battery. The positive electrode material (oxide) is obtained by subjecting the waste lithium-ion battery to pretreatment such as removal of the electrolyte, crushing, pulverization, and sorting.

[0017] <Addition of reducing agent (obtaining mixed oxide)> First, a reducing agent is added to the oxide to be reduced to obtain a mixed oxide, which is a mixture of the oxide and the reducing agent.

[0018] Reducing Agents: In the dry treatment, at least one reducing agent selected from the group consisting of carbon-containing C-based reducing agents, silicon-containing Si-based reducing agents, and aluminum-containing Al-based reducing agents is used. Examples of C-based reducing agents include graphite, coal, and coke. Examples of Si-based reducing agents include ferrosilicon, metallic silicon, and silicon sludge. Examples of Al-based reducing agents include metallic aluminum, aluminum sludge, and aluminum dross.

[0019] 《Amount of Reducing Agent Added R》 As described later, by heating a mixed oxide, which is a mixture of an oxide and a reducing agent, the oxide is reduced to obtain a metal (product metal) and slag (product slag). At this time, the amount of reducing agent added is R that satisfies the following formula (1). This suppresses the reduction of Mn, and a product metal with a low Mn content is obtained. In addition, a high reduction rate can be obtained for Ni and Co (especially Ni). For details, please refer to Test A described later.

[0020] (0.30a - 0.15b + 0.60) × [Ni] ≤ R ≤ (0.33a - 0.17b + 0.67) × [Ni] + (0.33a - 0.17b + 0.67) × [Co] + (0.23a - 0.12b + 0.47) × [Mn] ... (1) However, in the above formula (1), R: Amount of reducing agent added (unit: moles) [Ni]: Nickel content of oxide (unit: moles) [Co]: Cobalt content of oxide (unit: moles) [Mn]: Manganese content of oxide (unit: moles) a: Molar ratio of the content of C-based reducing agent in the reducing agent (unit: moles) to the total amount of reducing agent added (unit: moles) b: Molar ratio of the content of the Si-based reducing agent in the reducing agent (unit: moles) to the total amount of reducing agent added (unit: moles)

[0021] When determining the amount of reducing agent to add, first, the content of NiO, CoO, and MnO in the oxide to be reduced is determined. Specifically, the content of Ni, Co, and Mn in the oxide to be reduced is measured and considered to be the content of NiO, CoO, and MnO, respectively. The content of Ni, Co, and Mn is measured using an energy-dispersive X-ray spectrometer (EDX). X-ray fluorescence (XRF) analysis may also be used.

[0022] <Heating of mixed oxides (acquisition of metals)> Next, the mixed oxide (a mixture of oxide and reducing agent) is heated. This reduces the oxide. Note that, in addition to the reducing agent, CaO and SiO are also used during heating. ２ Fluxes such as the following may be added. In other words, the mixed oxide may further contain flux. The equipment used for heating the mixed oxide is not particularly limited and includes conventionally known equipment such as electric furnaces, resistance furnaces, high-frequency melting furnaces, low-frequency melting furnaces, rotary kilns, vertical furnaces, and steelmaking furnaces.

[0023] <Heating Temperature> The heating temperature when heating the mixed oxide is preferably above the melting point of the oxide (cathode material) containing nickel oxide, cobalt oxide, and manganese oxide. Considering the heating efficiency of the electric furnace, the heating temperature is preferably 1400°C or higher, more preferably 1450°C or higher, and even more preferably 1450°C or higher.

[0024] Incidentally, the cathode material of waste lithium-ion batteries may contain lithium. For example, lithium carbonate, when heated within the above temperature range, exceeds its melting point and liquefies, making it difficult to handle. However, the lithium contained in the cathode material is an oxide, and its melting point is about 500°C higher than that of lithium carbonate, so it transitions to the generated slag while maintaining its form. Therefore, the handling problems caused by liquefaction mentioned above do not occur.

[0025] The higher the heating temperature, the faster the reduction reaction rate. Therefore, increasing the heating temperature can shorten the time required to reduce the oxide. However, if the heating temperature is too high, it may cause the volatilization of Ni and Co. For this reason, the heating temperature is preferably 1600°C or lower, and more preferably 1550°C or lower.

[0026] 《Heating Atmosphere》 When heating mixed oxides, the atmosphere (heating atmosphere) can be, for example, nitrogen gas (N ２ Suitable atmospheres include inert atmospheres such as argon gas (Ar) atmospheres and reducing atmospheres such as carbon monoxide gas (CO) atmospheres.

[0027] 《Heating Time》 The heating time for the mixed oxide is preferably 1 hour or more, more preferably 2 hours or more, and even more preferably 3 hours or more, because it helps to suppress poor reduction (insufficient reduction reaction). There is no particular upper limit, but the heating time is preferably 6 hours or less, and more preferably 5 hours or less.

[0028] 《Products (Metal and Slag)》 By reducing the oxide (cathode material) that is the target of reduction, metal is produced. That is, the valuable elements Ni and Co contained in the oxide are recovered as metal. In addition, some of the Mn contained in the oxide may also be recovered as this metal.

[0029] The metal obtained by the reduction of an oxide (also called the "product metal") is an alloy containing at least one of the valuable elements (Ni, Co) and impurity elements (Cu, Fe). The product metal may contain only one of the valuable elements (Ni, Co).

[0030] By reducing the oxide (cathode material), in addition to the metal, slag is also generated. Slag (also called "generated slag") is formed, for example, when an Al-based reducing agent is used as the reducing agent. ２ O ３It contains oxides such as those listed above. In addition, the generated slag contains Mn, a valuable element not present in the generated metal, in the form of an oxide (e.g., MnO). Separating Mn from the generated metal by wet processing is computationally intensive. Retaining Mn in the generated slag by suppressing its incorporation into the generated metal is beneficial because it reduces this computational burden.

[0031] <Separation of Metal and Slag> Regarding the generated metal and slag obtained by the reduction of oxides, it is preferable to separate them before pulverizing the generated metal, as will be described later. The separation method is not particularly limited, and known methods can be used.

[0032] <Powdering of Metals (Obtaining Metal Powder)> Next, it is preferable to powder the resulting metal to obtain metal powder. In the wet process, as will be described later, first, leaching is carried out on the resulting metal using an acid solution. At this time, if the resulting metal is in the state obtained by the reduction of oxides, the leaching efficiency may be insufficient. For this reason, it is preferable to powder the resulting metal before carrying out leaching with an acid solution.

[0033] The smaller the particle size of the metal powder, the better the leaching efficiency. However, if the particle size of the metal powder is too small, handling may deteriorate and the risk of explosive reactions may increase. Therefore, these points should also be taken into consideration, and the particle size of the metal powder should be kept within an appropriate range. Specifically, for example, the particle size of the metal powder is preferably 250 to 6000 μm, and more preferably 300 to 5000 μm. The particle size is the median diameter (particle size at 50% cumulative value) on a volume basis in the particle size distribution determined by laser diffraction and scattering (the same applies hereinafter).

[0034] The method for pulverizing the generated metal is not particularly limited as long as the particle size of the resulting metal powder can be kept within an appropriate range. Examples include methods using grinding equipment such as jaw crushers and vibrating mills; atomization; and so on.

[0035] <Contact of metal with acid solution (obtaining leachate)> Next, the metal (metal powder) is brought into contact with the acid solution to leach out valuable elements (Ni, Co) and impurity elements (Cu, Fe). That is, a leachate containing valuable elements and impurity elements is obtained. The metal from which the valuable elements and impurity elements have leached out becomes a residue (leachate residue). There are no particular limitations on the method of contacting the metal with the acid solution, but examples include immersing the metal in the acid solution; spraying the metal with the acid solution; etc.

[0036] 《Solid-Liquid Ratio (Metal / Acid Solution)》 If the amount of acid solution in contact with the metal is too small (the amount of metal is too large relative to the amount of acid solution), some of the metal elements, such as valuable elements that have dissolved in the acid solution, may reach their saturation solubility and precipitate, resulting in insufficient leaching. For this reason, the ratio of the mass of the solid metal (in units: g) to the volume of the liquid acid solution (in units: mL) (also called the "solid-liquid ratio (metal / acid solution)") is preferably 1 / 5 or less, more preferably 1 / 7 or less, and even more preferably 1 / 10 or less. When the solid-liquid ratio (metal / acid solution) is 1 / 10, for example, 1 g of metal is immersed in 10 mL of acid solution. On the other hand, the solid-liquid ratio (metal / acid solution) is preferably 1 / 50 or more, more preferably 1 / 35 or more, and even more preferably 1 / 20 or more.

[0037] Acidic solutions: Acidic solutions that come into contact with metals must contain at least an acid.

[0038] (Acid) Examples of acids used in the acid solution include inorganic acids such as sulfuric acid, hydrochloric acid, and nitric acid. These may be used individually or in combination of two or more. From the viewpoint of realizing "battery-to-battery" recycling, where waste lithium-ion batteries are recycled and reused as raw materials for lithium-ion batteries, it is preferable to use sulfuric acid as the acid. This is because valuable elements can be obtained in the form of sulfates, which are easily used as cathode materials for lithium-ion batteries. Chlorides may also be included in the sulfuric acid and used as the acid.

[0039] ((Acid Concentration)) The concentration of the acid (e.g., sulfuric acid) used in the acid solution (acid concentration) is preferably 0.1 mol / L or higher, more preferably 0.5 mol / L or higher, and even more preferably 1.0 mol / L or higher, because it can increase the rate of leaching. There is no particular upper limit, but the acid concentration is preferably 8.0 mol / L or lower, more preferably 6.0 mol / L or lower, even more preferably 4.0 mol / L or lower, and particularly preferably 3.0 mol / L or lower.

[0040] (Oxidizing agent for acid solution) The inventors have found that even when the solid-liquid ratio (metal / acid solution) and acid concentration are within the above-mentioned range, leaching may be insufficient. For this reason, it is preferable to add an oxidizing agent (oxidizing agent for acid solution) to the acid solution as a leaching accelerator. Examples of oxidizing agents for acid solutions include hydrogen peroxide, hypochlorous acid, potassium permanganate, and ozone. Of these, hydrogen peroxide and ozone are preferred because the use of hypochlorous acid and potassium permanganate may require complicated post-treatment of chlorine, potassium, manganese, etc.

[0041] ((Content of oxidizing agent for acid solution)) From the viewpoint of ensuring sufficient leaching, the content of the oxidizing agent for the acid solution (e.g., hydrogen peroxide) in the acid solution is preferably 0.5 volume% or more, more preferably 1.0 volume% or more, even more preferably 3.0 volume% or more, even more preferably 5.0 volume% or more, particularly preferably 6.0 volume% or more, and most preferably 6.9 volume% or more, relative to the acid (e.g., sulfuric acid). On the other hand, the content of the oxidizing agent for the acid solution (e.g., hydrogen peroxide) in the acid solution is preferably 15.0 volume% or less, more preferably 13.0 volume% or less, and even more preferably 10.0 volume% or less, relative to the acid (e.g., sulfuric acid).

[0042] <Contact Time> The contact time between the metal and the acid solution is preferably 0.5 hours or more, more preferably 0.8 hours or more, and even more preferably 1.0 hour or more, in order to ensure sufficient leaching. On the other hand, from the viewpoint of productivity, the contact time is preferably 3.0 hours or less, and more preferably 1.5 hours or less.

[0043] <Separation of leachate and leachate residue> As described later, it is preferable to separate the leachate and leachate residue before adding the sulfurizing agent to the leachate. The separation method is not particularly limited, and known solid-liquid separation methods can be employed.

[0044] <Addition of sulfiding agent (obtaining copper removal solution)> Next, a sulfiding agent is added to the leachate containing valuable elements (Ni, Co) and impurity elements (Cu, Fe) to precipitate the impurity element copper (Cu) as copper sulfide. In this way, a leachate from which copper (Cu) has been selectively removed is obtained as a copper removal solution.

[0045] Figure 2 shows the potential-pH diagram (S-H) of copper (Cu) and nickel (Ni). ２ It is an O-system. In Figure 2, copper (Cu) - sulfur (S) - water (H) ２ Figure 2 shows the region on the potential-pH diagram of system O) where precipitates of copper (Cu) and nickel (Ni) oxides (hydroxides) or sulfides form, taking solubility into account. Note that cobalt precipitates in a similar manner to nickel, so its illustration is omitted in Figure 2. As shown in Figure 2, in the region where the pH is 3.0 or less and the oxidation-reduction potential is low, copper (Cu) precipitates selectively. Although not shown in Figure 2, in this region, copper precipitates as copper(II) sulfide (CuS). By utilizing this, the copper (Cu) contained in the leachate is selectively removed by making the leachate low pH and reducing. In other words, a copper removal solution is obtained, which is an leachate from which copper (Cu) has been removed.

[0046] Sulfidating agents: Sulfur (S), hydrogen sulfide (H) are used as sulfidating agents added to the leachate. ２ S), sodium hydrogen sulfide (NaSH), sodium sulfide (Na ２ Examples include S), and these may be used individually or in combination of two or more. Of these, from the viewpoint of ease of handling, sulfur, sodium hydrogen sulfide, and sodium sulfide, which can be handled as solids or solutions, are preferred over hydrogen sulfide, which is a toxic gas. However, in all cases, hydrogen sulfide gas may be generated by the sulfurization reaction, so caution is required when carrying out the experiment.

[0047] The temperature of the leachate to which the sulfiding agent has been added (sulfidation temperature) is not particularly limited and may be, for example, at room temperature.

[0048] (Amount of sulfiding agent added) From the viewpoint of sufficiently removing copper contained in the leachate, the amount of sulfiding agent added is preferably 1.0 equivalent or more, more preferably 1.5 equivalents or more, and even more preferably 2.0 equivalents or more, relative to the amount of copper (Cu) contained in the leachate. On the other hand, if an excessive amount of sulfiding agent is added, the amount of sulfides (precipitates) of valuable elements (Ni, Co, etc.) will increase, and the amount of valuable elements that we want to remain in the resulting copper removal solution may decrease. From this viewpoint, the amount of sulfiding agent added is preferably 3.0 equivalents or less, more preferably 2.5 equivalents or less, and even more preferably 2.0 equivalents or less, relative to the amount of copper (Cu) contained in the leachate.

[0049] For example, when producing copper(II) sulfide (CuS) using 1.0 equivalent of sodium hydrogen sulfide (NaSH) as a sulfiding agent, 1 mole of sodium hydrogen sulfide (NaSH) is used for every 1 mole of copper (Cu) contained in the leachate.

[0050] 《Sulfidation pH》 When adding a sulfiding agent to the leachate to precipitate copper sulfide, if the pH of the leachate with the added sulfiding agent (sulfidation pH) is high, there is a possibility that the amount of valuable element sulfides (precipitates) that we want to retain in the copper removal solution will increase. For this reason, the sulfidation pH is preferably 3.0 or less, more preferably 2.0 or less, even more preferably 1.0 or less, and particularly preferably 0 (zero). The sulfidation pH can be adjusted, for example, by adding a pH adjusting agent to the leachate. The pH adjusting agent is not particularly limited, and examples include sulfuric acid and sodium hydroxide.

[0051] <Sulfidation Time> The time required for the copper contained in the leachate to react with the sulfidating agent to sulfide (sulfidation time) is preferably 0.1 hours or more, more preferably 0.2 hours or more, and even more preferably 0.3 hours or more. On the other hand, from the viewpoint of productivity, the sulfidation time is preferably 3.0 hours or less, more preferably 2.0 hours or less, and even more preferably 1.0 hour or less.

[0052] <Separation of copper sulfide and copper removal solution> As described later, it is preferable to separate the copper sulfide and the copper removal solution before adding the oxidizing agent to the copper removal solution. The separation method is not particularly limited, and known solid-liquid separation methods can be used.

[0053] <Addition of oxidizing agent (obtaining a solution of valuable elements)> Next, an oxidizing agent is added to the copper removal solution containing valuable elements (Ni, Co) and iron (Fe) to precipitate the impurity element iron (Fe) as iron hydroxide. In this way, a copper removal solution from which iron (Fe) has been selectively removed is obtained as a solution of valuable elements containing valuable elements (Ni, Co).

[0054] Figure 3 shows the potential-pH diagram (O) of iron (Fe) and nickel (Ni). ２ -H ２ It is an O system. In Figure 3, iron (Fe) - oxygen (O ２ ) - Water (H ２ Figure 3 shows the region on the potential-pH diagram of system O) where precipitates of iron (Fe) and nickel (Ni) oxides (hydroxides) form, taking solubility into account. Note that cobalt precipitates in a similar manner to nickel, so its illustration is omitted in Figure 3. As shown in Figure 3, in the region where the pH is between 3.0 and 7.0 and the oxidation-reduction potential is high, iron (Fe) precipitates selectively. Although not shown in Figure 3, in this region, iron precipitates as iron(III) oxide hydroxide (FeO(OH)). By utilizing this, the copper removal solution is made acidic to neutral and oxidizing, causing the iron (Fe) contained in the copper removal solution to precipitate as iron(III) oxide hydroxide and be selectively removed. That is, a valuable element solution, which is a copper removal solution from which iron (Fe) has been removed, is obtained.

[0055] <Oxidizing Agent> Examples of oxidizing agents to be added to the copper removal solution include at least one oxidizing agent A selected from the group consisting of air and ozone; and at least one oxidizing agent B selected from the group consisting of hydrogen peroxide, hypochlorous acid, and potassium permanganate. Of these, air, hydrogen peroxide, and ozone are preferred because using hypochlorous acid and potassium permanganate may require complicated post-treatment for chlorine, potassium, manganese, etc.

[0056] (Amount of oxidizing agent added) From the viewpoint of sufficiently oxidizing the iron contained in the copper removal solution, the amount of oxidizing agent A (air, ozone), which is a gas, added is preferably 0.1 vvm or more, more preferably 0.3 vvm or more, and even more preferably 0.5 vvm or more, relative to the copper removal solution. On the other hand, the amount of oxidizing agent A added is preferably 5.0 vvm or less, more preferably 4.0 vvm or less, and even more preferably 3.0 vvm or less, relative to the copper removal solution.

[0057] The unit "vvm" represents the volume ratio of gas injected per minute relative to the liquid. For example, if the amount of oxidizing agent A added is 2vvm, then 2 liters of oxidizing agent A are injected per minute into 1 liter of copper removal solution.

[0058] For similar reasons, the amount of oxidizing agent B (hydrogen peroxide, hypochlorous acid, potassium permanganate) added is preferably 0.005% by volume or more, more preferably 0.015% by volume or more, even more preferably 0.050% by volume or more, and particularly preferably 0.100% by volume or more, relative to the copper removal solution. On the other hand, the amount of oxidizing agent B added is preferably 1.500% by volume or less, more preferably 1.000% by volume or less, even more preferably 0.500% by volume or less, and particularly preferably 0.300% by volume or less, relative to the copper removal solution.

[0059] <Oxidation Temperature> The inventors have found that using only the oxidizing agent described above may result in insufficient oxidation of the iron contained in the copper removal solution. Therefore, from the viewpoint of promoting oxidation, it is preferable to raise the temperature (oxidation temperature) of the copper removal solution to which the oxidizing agent has been added. Specifically, the oxidation temperature is preferably 10°C or higher, more preferably 30°C or higher, and even more preferably 50°C or higher. On the other hand, the oxidation temperature is preferably 90°C or lower, and more preferably 80°C or lower.

[0060] 《Oxidation pH》 When adding an oxidizing agent to a copper removal solution to precipitate iron hydroxide, if the pH of the copper removal solution with the oxidizing agent added (oxidation pH) is too low, precipitation may be difficult. For this reason, the oxidation pH is preferably 3.0 or higher, more preferably 3.7 or higher, even more preferably 4.0 or higher, and particularly preferably 4.5 or higher. On the other hand, if the oxidation pH is too high, the coprecipitation of valuable elements (such as Ni and Co) increases, and there is a concern that the amount of valuable elements to be retained in the resulting valuable element solution will decrease. For this reason, the oxidation pH is preferably 7.0 or lower, more preferably 6.0 or lower, and even more preferably 5.0 or lower. The oxidation pH can be adjusted, for example, by adding a pH adjusting agent to the copper removal solution. Examples of pH adjusting agents are not particularly limited, but include sulfuric acid and sodium hydroxide.

[0061] <Oxidation Time> The time required for the iron contained in the copper removal solution to react with the oxidizing agent (oxidation time) is preferably 0.3 hours or more, more preferably 0.5 hours or more, and even more preferably 1.0 hour or more. On the other hand, from the viewpoint of productivity, the oxidation time is preferably 3.0 hours or less, more preferably 2.0 hours or less, and even more preferably 1.5 hours or less.

[0062] 《Oxidizing Agents》 From the viewpoint of improving the reaction rate for forming iron hydroxide precipitates, oxidizing agents may be used in combination with the oxidizing agent described above. Examples of oxidizing agents include ferric oxide (Fe ２ O ３ At least one selected from the group consisting of ) and iron(III) oxide hydroxide (FeO(OH)), and the form of the oxidizing agent is preferably in powder form. The principle by which the reaction rate is improved by the oxidizing agent is catalytic action. Specifically, since the oxidizing agent tends to become negatively charged in aqueous solution (copper removal solution), Fe ２＋ It adsorbs ions, Fe ２＋ internal e － This weakens the connection with Fe ２＋ →Fe ３＋ +e － It is thought that the activation energy of this reaction (Fe oxidation reaction) is lowered, and the reaction is accelerated.

[0063] (Amount of oxidizing agent added) It is thought that the more oxidizing agent added, the larger the reaction surface area and the faster the Fe oxidation reaction. For this reason, the amount of oxidizing agent added is preferably 0.1 g / L or more, more preferably 0.5 g / L or more, and even more preferably 1.0 g / L or more, relative to the copper removal solution. On the other hand, if too much oxidizing agent is added, there is a concern that the coprecipitation of valuable elements (Ni, Co, etc.) will increase. For this reason, the amount of oxidizing agent added is preferably 40.0 g / L or less, more preferably 10.0 g / L or less, and even more preferably 5.0 g / L or less, relative to the copper removal solution.

[0064] (Particle size of oxidizing agent) If the particle size of the oxidizing agent is too small, the reaction surface area will be excessive, raising concerns about increased coprecipitation of valuable elements (Ni, Co, etc.). For this reason, the particle size of the oxidizing agent is preferably 0.1 μm or larger, more preferably 0.3 μm or larger, and even more preferably 0.5 μm or larger. On the other hand, if the particle size of the oxidizing agent is too large, the reaction surface area will be insufficient, and the desired effect may not be obtained. For this reason, the particle size of the oxidizing agent is preferably 3.0 μm or smaller, more preferably 2.0 μm or smaller, and even more preferably 1.0 μm or smaller.

[0065] <Separation of iron hydroxide and valuable element solution> It is preferable to separate the iron hydroxide and the valuable element solution. The separation method is not particularly limited, and known solid-liquid separation methods can be used. The valuable elements in the valuable element solution obtained in this way can be used, for example, as a positive electrode material for lithium-ion batteries.

[0066] <Obtaining Valuable Element Precipitate> Figure 1B is a flowchart showing an example of the process for producing the precursor. Next, the valuable element solution obtained by the valuable element recovery method described above, a complexing agent, and an alkaline aqueous solution are introduced (dropwise) into the reaction vessel liquid, and a precipitate containing valuable elements (valuable element precipitate) is obtained by the so-called coprecipitation method, as shown in Figure 1B. Specifically, the valuable element precipitate is, for example, at least one selected from the group consisting of a composite hydroxide containing valuable elements and a composite oxide containing valuable elements. By using the coprecipitation method, the valuable elements (Ni, Co, Mn) can be uniformly dispersed at the atomic level. The obtained valuable element precipitate is filtered from the reaction vessel liquid, and washed with water and dried as necessary to obtain a cathode material precursor.

[0067] When obtaining a precipitate of valuable elements, an aqueous solution of the raw materials may be used instead of a solution of valuable elements. The aqueous solution of the raw materials is prepared by adding at least one selected from the group consisting of a nickel source, a cobalt source, and a manganese source to a solution of valuable elements. In the aqueous solution of the raw materials, the molar ratio of nickel content, cobalt content, and manganese content (Ni / Co / Mn) is preferably 1 / 1 / 1, 5 / 2 / 3, 6 / 2 / 2, or 8 / 1 / 1.

[0068] Nickel sources include nickel salts such as nickel sulfate, nickel carbonate, nickel nitrate, nickel acetate, and nickel chloride, and nickel sulfate (NiSO4). ４ ) is preferred. The cobalt source is, for example, a cobalt salt such as cobalt sulfate, cobalt carbonate, cobalt nitrate, cobalt acetate, or cobalt chloride, and cobalt sulfate (CoSO4) ４ ) is preferred. Manganese sources are, for example, manganese salts such as manganese sulfate, manganese carbonate, manganese nitrate, manganese acetate, and manganese chloride, and manganese sulfate (MnSO4) ４ ) is preferred. The nickel source, cobalt source, and manganese source are each preferably used in the form of aqueous solutions. In each aqueous solution, the concentrations (contents) of the nickel source, cobalt source, and manganese source are each preferably adjusted to the molar ratios described above.

[0069] The pH of the raw material aqueous solution is preferably 8 or less, more preferably 7 or less, and even more preferably 6 or less. Furthermore, the pH of the raw material aqueous solution is, for example, 1 or more, and preferably 2 or more.

[0070] The dropping rate of the valuable element solution (raw material aqueous solution) is preferably 1.0 mL / min or more, and more preferably 2.5 mL / min or more. Furthermore, the dropping rate of the valuable element solution (raw material aqueous solution) is preferably 7.0 mL / min or less, and more preferably 5.5 mL / min or less.

[0071] Examples of complexing agents include ammonia (NH₃). ３ An ammonium source is selected from the group consisting of ammonium sulfate and ammonium salts. Examples of ammonium salts include ammonium sulfate, ammonium chloride, ammonium nitrate, and ammonium carbonate. Ammonia is preferred as the complexing agent that is the ammonium source. The ammonium source is preferably used in the form of an aqueous solution. In the aqueous solution, the concentration (content) of the ammonium source is preferably adjusted to the molar ratio described later.

[0072] The dropping rate of the complexing agent is preferably 0.1 mL / min or more, and more preferably 0.3 mL / min or more. Furthermore, the dropping rate of the complexing agent is preferably 1.0 mL / min or less, and more preferably 0.8 mL / min or less.

[0073] The molar ratio (NH₄) of the ammonium source (complexing agent) content relative to the total content of valuable elements (Ni, Co, Mn) in the raw material aqueous solution. ４ The molar ratio ( / (Ni + Co + Mn)) is preferably greater than 0, more preferably 2 or greater, and even more preferably 4 or greater. ４ The ratio of (Ni + Co + Mn) is preferably 20 or less, more preferably 15 or less, and even more preferably 12 or less.

[0074] As the alkaline aqueous solution, an aqueous solution of sodium hydroxide (NaOH) is preferred.

[0075] The reaction vessel liquid is the liquid inside the reaction vessel, and is an aqueous solution prepared, for example, by adding an alkaline aqueous solution such as an aqueous sodium hydroxide solution to pure water. The pH of the reaction vessel liquid is preferably 9.0 or higher, and more preferably 9.5 or higher. Furthermore, the pH of the reaction vessel liquid is preferably 12.0 or lower, and more preferably 11.5 or lower. The temperature of the reaction vessel liquid is preferably 30°C or higher, and more preferably 35°C or higher. Furthermore, the temperature of the reaction vessel liquid is preferably 60°C or lower, and more preferably 45°C or lower.

[0076] When obtaining the precipitate, it is preferable to stir the reaction vessel liquid using a stirring rod or the like. If the stirring speed (rotation speed of the stirring blade) is too slow, a homogeneous precipitate of valuable elements may not be obtained. On the other hand, if the stirring speed is too fast, the complexing agent (e.g., aqueous ammonia solution) may be scattered during stirring, destabilizing the reaction between the complexing agent and the aqueous raw material solution, and again, a homogeneous precipitate of valuable elements may not be obtained. In these cases, the precursor obtained by drying the precipitate of valuable elements does not become uniformly spheroidal, making it difficult to obtain a high tap density. Therefore, from the viewpoint of obtaining a high tap density, the stirring speed is, for example, 150 to 550 rpm, preferably 200 to 500 rpm, more preferably 250 to 450 rpm, and even more preferably 300 to 400 rpm. Note that the stirring speed may be changed during the process. For example, it may be possible to start with a slow stirring speed and change to a faster, more suitable stirring speed as the volume of liquid in the reaction vessel increases.

[0077] It is preferable to control the pH of the reaction vessel solution within the above range by adding an alkaline aqueous solution to the reaction vessel solution dropwise while adding the valuable element solution (aqueous solution of raw materials) and the complexing agent.

[0078] The obtained valuable element precipitate is preferably filtered from the reaction vessel liquid (solid-liquid separation), washed with water, and then dried. The drying temperature is preferably 90°C or higher, more preferably 95°C or higher. Furthermore, the drying temperature is preferably 120°C or lower, more preferably 110°C or lower. The drying time is preferably 5 hours or more, more preferably 8 hours or more. Furthermore, the drying time is preferably 15 hours or less, more preferably 12 hours or less.

[0079] As described above, for example, a precursor for the cathode material can be obtained by washing and drying the precipitate of valuable elements. In the precursor, the molar ratio of nickel content to the total content of valuable elements (Ni, Co, Mn) (Ni / (Ni+Co+Mn)) is preferably 0.3 or higher, and more preferably 0.4 or higher. Furthermore, this molar ratio (Ni / (Ni+Co+Mn)) is preferably 1.0 or lower, and more preferably 0.8 or lower.

[0080] The tap density of the precursor is 0.8 g / cm³. ３ The above is preferable, 1.0 g / cm³ ３ The above is more preferable, specifically 1.2 g / cm³. ３ The above is even more preferable. Furthermore, the tap density of the precursor is 1.8 g / cm³. ３ It may also be less than 1.5 g / cm³. ３ The following may also be true: Tap density is per 100 cm³ of standing volume. ３ The sample is placed in the container, and using a tapping device, it is tapped until the volume of the sample no longer decreases. Then, the mass of the sample (in grams) is measured against the volume of the sample (in centimeters). ３ This can be calculated by dividing by (the same applies below).

[0081] Precursor particle size D １０ The particle size D of the precursor is preferably 3.0 μm or larger, and more preferably 4.0 μm or larger. １０ The particle size of the precursor is preferably 10.0 μm or less, and more preferably 8.0 μm or less. ５０ The particle size D of the precursor is preferably 8.0 μm or larger, and more preferably 9.0 μm or larger. ５０ The particle size of the precursor is preferably 16.0 μm or less, and more preferably 14.0 μm or less. ９０ The particle size D of the precursor is preferably 12.0 μm or larger, and more preferably 14.0 μm or larger. ９０ Particle size D is preferably 24.0 μm or less, and more preferably 22.0 μm or less. １０ , particle size D ５０ and particle size D ９０ These are the particle sizes at which the cumulative frequency of the particle size distribution determined by laser diffraction and scattering methods accounts for 10%, 50%, and 90% of the volume, respectively (the same applies hereafter).

[0082] <Obtaining the Calcined Product> Figure 1B is a flowchart showing an example of the process for manufacturing a positive electrode material. Next, the obtained precursor is mixed with a lithium-containing compound, and the resulting mixture is calcined. In this way, as shown in Figure 1B, a calcined product containing valuable elements and lithium (a composite oxide containing valuable elements and lithium) is obtained. By appropriately crushing the obtained calcined product, a positive electrode material for use in lithium-ion batteries is obtained. The positive electrode material is also called a positive electrode active material.

[0083] The resulting cathode material is a composite oxide containing valuable elements (Ni, Co, Mn) and lithium (Li), and may further contain at least one element A selected from the group consisting of aluminum (Al), silicon (Si), titanium (Ti), zirconium (Zr), calcium (Ca), potassium (K), barium (Ba), strontium (Sr), and sulfur (S).

[0084] The following describes each step in the manufacturing method of the cathode material in more detail.

[0085] First, the precursor and the lithium-containing compound are mixed to obtain a mixture. At this time, the molar ratio (Li / (Ni+Co+Mn)) of the lithium-constituted content of the lithium-containing compound to the sum of the nickel-constituted content of the precursor, the cobalt-constituted content of the precursor, and the manganese-constituted content of the precursor is preferably greater than 1.03, and more preferably 1.04 or greater. Furthermore, this molar ratio (Li / (Ni+Co+Mn)) is preferably less than 1.10, and more preferably 1.08 or less.

[0086] Suitable lithium-containing compounds include, for example, at least one selected from the group consisting of lithium hydroxide and lithium carbonate.

[0087] If the resulting cathode material contains element A as described above, a compound containing element A (hereinafter also referred to as "A-containing compound") may be further mixed into the mixture. Examples of A-containing compounds include, but are not limited to, hydroxides, oxides, chlorides, and salts of element A (e.g., sulfates, carbonates, nitrates, etc.). The amount of A-containing compound mixed can be adjusted as appropriate according to the desired composition.

[0088] Next, the mixture obtained by mixing is fired to obtain a fired product. At this time, it is preferable to pre-fire the mixture and then fire it. The firing temperature for pre-fire is preferably 400°C or higher, and more preferably 500°C or higher. Furthermore, the firing temperature for pre-fire is preferably 700°C or lower, and more preferably 680°C or lower. The firing temperature for the main firing is preferably 800°C or higher, and more preferably 900°C or higher. Furthermore, the firing temperature for the main firing is preferably 1000°C or lower, and more preferably 980°C or lower.

[0089] The atmosphere for calcination can be an oxidizing atmosphere (e.g., an atmospheric atmosphere) or a non-oxidizing atmosphere. An example of a non-oxidizing atmosphere is an atmosphere with an oxygen concentration of 10% by volume or less, and a specific example of this is a nitrogen atmosphere. The atmosphere for the final calcination can be an oxidizing atmosphere (e.g., an atmospheric atmosphere) or a non-oxidizing atmosphere.

[0090] The firing time for the preliminary firing is preferably 2 hours or more, and more preferably 3 hours or more. Furthermore, the firing time for the preliminary firing is preferably 48 hours or less, and more preferably 12 hours or less. The firing time for the main firing is preferably 1 hour or more, more preferably 2 hours or more, and even more preferably 3 hours or more. Furthermore, the firing time for the main firing is preferably 30 hours or less, more preferably 15 hours or less, and even more preferably 8 hours or less.

[0091] The calcined material may be washed with water. Washing with water removes any excess lithium that has not penetrated the interior. After washing, it is dried as appropriate. The calcined material may be further calcined at a temperature between 200°C and 800°C, or crushed. In this way, a positive electrode material for use in lithium-ion batteries is obtained.

[0092] The tap density of the positive electrode material is 1.0 g / cm³.３ The above is preferable, 1.5 g / cm³ ３ The above is more preferable. Furthermore, the tap density of the positive electrode material is 3.5 g / cm³. ３ The following is also acceptable: 3.0 g / cm³ ３ The following is also acceptable.

[0093] Particle size D of the positive electrode material １０ The particle size D of the positive electrode material is preferably 3.0 μm or larger, and more preferably 4.0 μm or larger. １０ The particle size D of the positive electrode material is preferably 10.0 μm or less, and more preferably 8.0 μm or less. ５０ The particle size D of the positive electrode material is preferably 8.0 μm or larger, and more preferably 9.0 μm or larger. ５０ The particle size D of the positive electrode material is preferably 16.0 μm or less, and more preferably 14.0 μm or less. ９０ The particle size D of the positive electrode material is preferably 12.0 μm or larger, and more preferably 14.0 μm or larger. ９０ The particle size is preferably 24.0 μm or less, and more preferably 22.0 μm or less.

[0094] A lithium-ion battery generally comprises a positive electrode, a negative electrode, and an ion-conducting medium (for example, an electrolyte such as a non-aqueous electrolyte) interposed between the positive and negative electrodes to conduct lithium ions, and may further include a separator. A positive electrode is manufactured using the above-mentioned positive electrode material by a known method, and a lithium-ion battery is manufactured using the manufactured positive electrode. The manufactured lithium-ion battery has excellent discharge capacity and cycle characteristics.

[0095] The present invention will be specifically described below with reference to examples. However, the present invention is not limited to the examples described below.

[0096] [Test A] In Test A, the amount R added of reducing agents (C-based reducing agent, Si-based reducing agent, and Al-based reducing agent) was investigated.

[0097] <Test A1: C-type reducing agent> NiO, CoO, and MnO were mixed to prepare oxides. The content of Ni, Co, and Mn in the prepared oxides (in molar parts) is shown in Table 1 below. As a reducing agent, a C-type reducing agent, which is coke, was prepared.

[0098] A mixed oxide was obtained by adding a C-based reducing agent to the prepared oxide in the amounts (in moles) shown in Table 1 below. The obtained mixed oxide was filled into a mold, and pressure was applied to the mold from the outside to obtain a molded body of the mixed oxide. Next, the molded body was heated for 3 hours in an Ar atmosphere using a heater (electric resistance furnace) maintained at a temperature of 1450°C to reduce the oxide and obtain a metal (produced metal) and a slag (produced slag), which were then separated.

[0099] The content (in mole percent) of each element in the generated metal was determined by X-ray fluorescence (XRF) analysis. The results are shown in Table 1 below. In addition, the Ni reduction rate (in percent) was determined from the Ni content of the oxide and the Ni content of the generated metal. The results are shown in Table 1 below.

[0100]

[0101] As shown in Table 1 above, samples 1 to 6 had a Mn content of 9.0 mol% or less in the generated metal and a Ni reduction rate of 85% or more. In contrast, sample 7 had a Ni reduction rate of 70%. Samples 8 to 9 had a Mn content of 28.5 mol% or more in the generated metal.

[0102] From the perspective of reducing the Mn content of the resulting metal, an upper limit of 2.70 molar parts can be derived for the amount R of reducing agent added. This upper limit can be expressed by the equation [Ni] + [Co] + α(C) × [Mn], and α(C) = 0.70 is obtained.

[0103] If the amount of reducing agent added R is too low, NiO will not be reduced, Ni will be included in the resulting slag, and the Ni reduction rate will decrease. Specifically, when the amount of reducing agent added R is 0.70 moles, the Ni reduction rate is 70% (No. 7). Here, when the amount of reducing agent added R is 0.90 moles or more, the Ni reduction rate is 85% or more (Nos. 1 to 6). From this, a lower limit of 0.90 moles can be derived for the amount of reducing agent added R. This lower limit can be expressed by the formula β(C) × [Ni], and β(C) = 0.90 is obtained.

[0104] To efficiently recover not only Ni but also Co, the amount of reducing agent R added should be 2.00 molar parts or more. In this case, the lower limit of the amount of reducing agent R corresponds to [Ni] + [Co].

[0105] <Test A2: Si-based reducing agent> Oxides were prepared in the same manner as in Test A1 described above. The content of Ni, Co, and Mn in the prepared oxide (in moles) is shown in Table 2 below. A Si-based reducing agent, which is metallic silicon, was prepared as a reducing agent. The Si-based reducing agent was added to the prepared oxide in the amount (in moles) shown in Table 2 below to obtain a mixed oxide. The obtained mixed oxide was molded and heated in the same manner as in Test A1 described above to obtain the generated metal and generated slag. Next, the content of each element (in mole %) and the Ni reduction rate (in %) in the generated metal were determined in the same manner as in Test A1 described above. The results are shown in Table 2 below.

[0106]

[0107] As shown in Table 2 above, samples No. 10 to 15 had a Mn content of 8.9 mol% or less in the generated metal and a Ni reduction rate of 86% or more. In contrast, sample No. 16 had a Ni reduction rate of 70%. Samples No. 17 to 18 had a Mn content of 27.6 mol% or more in the generated metal.

[0108] From the perspective of reducing the Mn content of the resulting metal, an upper limit of 1.35 molar parts can be derived for the amount R of reducing agent added. This upper limit can be expressed by the equation 0.50 × [Ni] + 0.50 × [Co] + α(Si) × [Mn], where α(Si) = 0.35.

[0109] If the amount of reducing agent added, R, is too low, NiO will not be reduced, and Ni will be included in the resulting slag, reducing the Ni reduction rate. Specifically, when the amount of reducing agent added, R, is 0.35 moles, the Ni reduction rate is 70% (No. 16). Here, when the amount of reducing agent added, R, is 0.45 moles or more, the Ni reduction rate is 86% or more (Nos. 10-15). From this, a lower limit of 0.45 moles can be derived for the amount of reducing agent added, R. This lower limit can be expressed by the formula β(Si) × [Ni], and β(Si) = 0.45 is obtained.

[0110] Furthermore, if the amount of reducing agent R added is 1.25 molar parts or more, Si will be mixed into the resulting metal, but this will not hinder the subsequent separation and recovery of Ni and Co in the wet treatment.

[0111] To efficiently recover not only Ni but also Co, the amount of reducing agent R added should be at least 1.00 molar. In this case, the lower limit of the amount of reducing agent R corresponds to 0.50 × [Ni] + 0.50 × [Co].

[0112] <Test A3: Al-based reducing agent> Oxides were prepared in the same manner as in Test A1 described above. The content of Ni, Co, and Mn in the prepared oxides (in moles) is shown in Table 3 below. An Al-based reducing agent, which is metallic aluminum, was prepared as the reducing agent.

[0113] A mixed oxide was obtained by adding an Al-based reducing agent to the prepared oxide in the amounts (in moles) shown in Table 3 below. The obtained mixed oxide was molded and heated in the same manner as in Test A1 described above to obtain the generated metal and generated slag. Then, in the same manner as in Test A1 described above, the content of each element (in moles %) and the Ni reduction rate (in %) in the generated metal were determined. The results are shown in Table 3 below.

[0114]

[0115] As shown in Table 3 above, samples No. 19 to 24 had a Mn content of 8.8 mol% or less in the generated metal and a Ni reduction rate of 85% or more. In contrast, sample No. 25 had a Ni reduction rate of 70%. Samples No. 26 to 27 had a Mn content of 27.1 mol% or more in the generated metal.

[0116] From the perspective of reducing the Mn content of the resulting metal, an upper limit of 1.80 moles can be derived for the amount R of reducing agent added. This upper limit can be expressed by the equation 0.67 × [Ni] + 0.67 × [Co] + α(Al) × [Mn], and α(Al) = 0.47 is obtained.

[0117] If the amount of reducing agent added, R, is too low, NiO will not be reduced, and Ni will be included in the resulting slag, reducing the Ni reduction rate. Specifically, when the amount of reducing agent added, R, is 0.47 moles, the Ni reduction rate is 70% (No. 25). Here, when the amount of reducing agent added, R, is 0.60 moles or more, the Ni reduction rate is 85% or more (Nos. 19-24). From this, a lower limit of 0.60 moles can be derived for the amount of reducing agent added, R. This lower limit can be expressed by the formula β(Al) × [Ni], and β(Al) = 0.60 is obtained.

[0118] To efficiently recover not only Ni but also Co, the amount of reducing agent R added should be at least 1.34 molar parts. In this case, the lower limit of the amount of reducing agent R corresponds to 0.67 × [Ni] + 0.67 × [Co].

[0119] Furthermore, if the amount of reducing agent R added is 1.67 molar parts or more, Al will be mixed into the generated metal, but this will not hinder the subsequent separation and recovery of Ni and Co in the wet treatment.

[0120] The appropriate range for the amount R of reducing agent added, as described above, also holds true when the reducing agent is a mixture of C-based reducing agents, Si-based reducing agents, and Al-based reducing agents. That is, the amount R of reducing agent added should satisfy the following formula (1): (0.30a - 0.15b + 0.60) × [Ni] ≤ R ≤ (0.33a - 0.17b + 0.67) × [Ni] + (0.33a - 0.17b + 0.67) × [Co] + (0.23a - 0.12b + 0.47) × [Mn] ... (1)

[0121] In the above formula (1), a is the molar ratio of the content of C-based reducing agent in the reducing agent (in molar parts) to the total amount of reducing agent added (in molar parts) (C-based reducing agent / total reducing agent). Also, b is the molar ratio of the content of Si-based reducing agent in the reducing agent (in molar parts) to the total amount of reducing agent added (in molar parts) (Si-based reducing agent / total reducing agent).

[0122] By adopting this additive amount R, a metal with a low Mn content can be obtained. Furthermore, by reducing the amount of Ni contained in the generated slag, a high Ni reduction rate can be obtained.

[0123] <Test A4: C-based reducing agent, Si-based reducing agent, and Al-based reducing agent> Oxides were prepared in the same manner as in Test A1 described above. The content of Ni, Co, and Mn in the prepared oxides (in moles) is shown in Table 4 below. Mixed oxides were obtained by adding at least one of the C-based reducing agent, Si-based reducing agent, and Al-based reducing agent to the prepared oxide in the amounts (in moles) shown in Table 4 below. Graphite was used as the C-based reducing agent, metallic silicon as the Si-based reducing agent, and metallic aluminum as the Al-based reducing agent.

[0124] The resulting mixed oxide was subjected to molding and heating in the same manner as in Test A1 described above, thereby reducing the cathode material (oxide) to obtain the generated metal and generated slag. Next, the content (in mole %) of each element in the generated metal and the reduction rate (in percent) of each element were determined in the same manner as in Test A1 described above. The results are shown in Table 4 below.

[0125]

[0126] As described above, when using a carbon-based reducing agent, the amount of reducing agent added (in moles) R is set to satisfy 0.90 × [Ni] ≤ R ≤ [Ni] + [Co] + 0.70 × [Mn]. Specifically, an amount of R that satisfies 0.216 ≤ R ≤ 0.376 is set. As shown in Table 4 above, for Nos. 28 to 33, where the amount of added R satisfies the above range, the Ni reduction rate was 79% or higher, and the Mn content of the resulting metal was 4.6 mol% or less. In contrast, for No. 34, where the amount of added R was below the above range, the Ni reduction rate was 69%, which was lower than Nos. 28 to 33. Furthermore, for Nos. 35 to 36, where the amount of added R was above the above range, the Mn content of the resulting metal was 17.0% or higher, which was higher than Nos. 28 to 33.

[0127] As described above, when using a Si-based reducing agent, the amount of reducing agent added (in moles) R is set to satisfy 0.45 × [Ni] ≤ R ≤ 0.5 × [Ni] + 0.5 × [Co] + 0.35 × [Mn]. Specifically, an amount of R that satisfies 0.108 ≤ R ≤ 0.188 is set. As shown in Table 4 above, for Nos. 37 to 42, where the amount of added R satisfies the above range, the Ni reduction rate was 76% or higher, and the Mn content of the resulting metal was 4.6 mol% or less. In contrast, for No. 43, where the amount of added R is below the above range, the Ni reduction rate was 70%, which was lower than Nos. 37 to 42. Also, for Nos. 44 to 45, where the amount of added R is above the above range, the Mn content of the resulting metal was 16.2 mol% or higher, which was higher than Nos. 37 to 42.

[0128] As described above, when using an Al-based reducing agent, the amount of reducing agent added (in moles) R is set to satisfy 0.60 × [Ni] ≤ R ≤ 0.67 × [Ni] + 0.67 × [Co] + 0.47 × [Mn]. Specifically, an amount of R that satisfies 0.144 ≤ R ≤ 0.252 is set. As shown in Table 4 above, for Nos. 46 to 51, where the amount of added R satisfies the above range, the Ni reduction rate was 77% or higher, and the Mn content of the resulting metal was 4.4 mol% or less. In contrast, for No. 52, where the amount of added R is below the above range, the Ni reduction rate was 60%, which was lower than Nos. 46 to 51. Also, for Nos. 53 to 54, where the amount of added R is above the above range, the Mn content of the resulting metal was 12.0 mol or higher, which was higher than Nos. 46 to 51.

[0129] Furthermore, as shown in Table 4 above, for samples No. 55-58, where the reducing agent was a mixture of a C-based reducing agent, a Si-based reducing agent, and an Al-based reducing agent, and the amount R added satisfied formula (1), the resulting metal had a low Mn content and a high Ni reduction rate. In contrast, for samples No. 59-60, where the amount R added did not satisfy formula (1), the resulting metal had a higher Mn content than No. 55-58, or a lower Ni reduction rate than No. 55-58. A similar trend was observed in the comparison results between No. 61-64 and No. 65-66, No. 67-70 and No. 71-72, and No. 73-76 and No. 77-78.

[0130] [Test B] <Preparation of Cathode Material> Cathode material (oxide) from a waste lithium-ion battery was prepared. Specifically, the waste lithium-ion battery was subjected to pretreatment such as decomposition, discharge, and removal of the electrolyte to separate the cathode material. The cathode material used was obtained by crushing and grinding into a powder. The content of Ni, Co, and Mn in the cathode material (oxide) (unit: moles) was 0.24 moles, 0.08 moles, and 0.08 moles, respectively, the same as the oxide used in Test A4 described above. In addition to Ni, Co, and Mn, the cathode material also contained Cu and Fe as impurity elements.

[0131] <Addition of Reducing Agent> A mixed oxide was obtained by adding at least one reducing agent selected from the group consisting of C-based reducing agents, Si-based reducing agents, and Al-based reducing agents to the prepared cathode material (oxide) in the same amount R (unit: moles) as in Test A4 Nos. 28-30, 37-39, 46-48, 55-56, 61-62, 67-68, and 73-74 described above. Graphite was used as the C-based reducing agent, metallic silicon as the Si-based reducing agent, and metallic aluminum as the Al-based reducing agent.

[0132] <Heating of mixed oxides> The obtained mixed oxides were molded and heated in the same manner as in Test A1 described above to reduce the cathode material (oxide) and obtain the generated metal and generated slag. Next, the content (in mole %) of each element in the generated metal and the reduction rate (in %) of each element were determined in the same manner as in Test A1 described above. The results were all the same as those for Test A4 Nos. 28-30, 37-39, 46-48, 55-56, 61-62, 67-68 and 73-74 described above.

[0133] <Powdering of Metals> From the metals (produced metals) obtained by the reduction of the cathode material, metals having the composition shown in Table 5 below were powdered using a vibratory mill to obtain metal powder. The particle size of the obtained metal powder was 1100 μm.

[0134] <Contact between metal and acid solution> An acid solution was prepared by adding 7.0 volume% hydrogen peroxide as an oxidizing agent to sulfuric acid (concentration: 2.0 mol / L). A metal (metal powder) having the composition shown in Table 5 below was brought into contact with the prepared acid solution at a solid-liquid ratio of 1 / 10 (metal / acid solution) (contact time: 1.0 hour). Specifically, the metal powder was immersed in the acid solution. In this way, a leachate and leachate residue were obtained and separated. The concentration of each element in the leachate was determined using XRF (X-ray fluorescence) analysis, and the leaching rate (unit: mass%) of each element from the metal into the leachate was calculated. The results are shown in Table 5 below. As shown in Table 5 below, the leaching rate for each element was 100 mass%, indicating that all elements were successfully leached from the metal into the leachate.

[0135]

[0136] <Addition of Sulfidating Agent> The content of each element in the obtained leachate (unit: g / L) is shown in Table 6 below. Sodium hydrogen sulfide (NaSH) was added to the obtained leachate as a sulfidating agent and stirred at room temperature (25°C). The amount of sulfidating agent (sodium hydrogen sulfide) added was 2.0 equivalents relative to the Cu contained in the leachate. The pH of the leachate with the added sulfidating agent (sulfidation pH) was adjusted to 0 (zero) using sulfuric acid and sodium hydroxide as pH adjusters. In this way, the copper (Cu) contained in the leachate was reacted with the sulfidating agent and sulfidized (sulfidation time: 20 minutes), and precipitated as copper sulfide (copper(II) sulfide). Subsequently, the copper sulfide and the copper removal solution, which is the leachate from which copper has been removed, were separated. The content of each element in the copper removal solution (unit: g / L) was determined by ICP-AES (inductively coupled plasma atomic emission spectrometry). The results are shown in Table 6 below. Furthermore, for each element, the ratio of its content in the copper removal solution to its content in the leachate was determined as the residual rate a (unit: mass%). The results are shown in Table 6 below. As shown in Table 6 below, the Cu content in the copper removal solution is very low, indicating that Cu was removed from the leachate with very high efficiency.

[0137] <Addition of Oxidizing Agent> Next, the copper removal solution was diluted with water. The content of each element in the diluted copper removal solution (unit: g / L) is shown in Table 6 below. The reason for dilution was that in preliminary experiments, when an oxidizing agent was added to a model solution having a similar composition to the copper removal solution shown in Table 6 below, excessive precipitation occurred, making stirring impossible in some cases. Hydrogen peroxide was added as an oxidizing agent to the diluted copper removal solution and stirred. The amount of oxidizing agent (hydrogen peroxide) added was 0.020 volume% relative to the diluted copper removal solution. The pH (oxidation pH) of the copper removal solution with the added oxidizing agent was adjusted to 4.5 using sulfuric acid and sodium hydroxide as pH adjusters. The temperature (oxidation temperature) of the copper removal solution with the added oxidizing agent was set to 70°C and maintained at that temperature. In this way, the iron (Fe) contained in the copper removal solution was oxidized by reacting with the oxidizing agent (oxidation time: 1.0 hour) and precipitated as iron hydroxide (iron(III) oxide). Subsequently, the iron hydroxide and the valuable element solution, which was the copper removal solution from which iron had been removed, were separated. The content of each element in the valuable element solution (unit: g / L) was determined by ICP-AES. The results are shown in Table 6 below. Furthermore, for each element, the ratio of the content in the valuable element solution to the content in the copper removal solution (diluted) was determined as the residual rate b (unit: mass%). The results are shown in Table 6 below. As shown in Table 6 below, the Fe content in the valuable element solution is very low, indicating that Fe was removed from the copper removal solution with very high efficiency.

[0138] Furthermore, for each element, the final residual rate in the valuable element solution was determined as the overall residual rate (unit: mass%) from residual rates a and b. The results are shown in Table 6 below. From the results shown in Table 6 below, it can be seen that by performing a dry treatment, then pulverizing the metal obtained by the dry treatment, and then performing a wet treatment, the valuable elements (Ni, Co) could be recovered with very high purity.

[0139]

[0140] [Test C] <Preparation of positive electrode material to metal pulverization> The preparation of the positive electrode material and the pulverization of the metal are the same as in Test B, so the explanation will be omitted.

[0141] <Contact between metal and acid solution> Multiple acid solutions were prepared by adding the oxidizing agent for the acid solution (hydrogen peroxide) in the amounts (unit: volume %) shown in Table 7 below to sulfuric acid (concentration: 2.0 mol / L). Except for changing the amount of oxidizing agent added, the metal (metal powder) was brought into contact with the acid solution in the same manner as in Test B described above to obtain a leachate. Furthermore, the leaching rate (unit: mass %) of each element from the metal into the leachate was calculated in the same manner as in Test B described above. The results are shown in Table 7 below. As shown in Table 7 below, the leaching rate increased as the amount of oxidizing agent added for the acid solution increased. It was found that an amount of oxidizing agent for the acid solution (hydrogen peroxide) of 6.9 volume % or more is suitable for sufficiently leaching valuable metals (Ni, Co). However, when the amount added exceeds 6.9 volume %, the leaching rate plateaus. Therefore, from a cost perspective, it was found that within the scope of this embodiment, 6.9% by volume is a suitable amount of oxidizing agent (hydrogen peroxide) for the acid solution.

[0142]

[0143] <Addition of Sulfidating Agent> Sodium hydrogen sulfide (NaSH) was added as a sulfidating agent to the leachate shown in Table 6 above, in the amount (unit: equivalent) shown in Table 8 below, and the mixture was stirred. At this time, the sulfidation pH was adjusted to the value shown in Table 8 below. Except for changing the amount of sulfidating agent and the sulfidation pH, the copper contained in the leachate was precipitated as copper sulfide in the same manner as in Test B described above to obtain a copper removal solution. Furthermore, the Cu content (unit: mg / L), Ni residue rate (unit: mass%), and Co residue rate (unit: mass%) in the obtained copper removal solution were determined in the same manner as in Test B described above. The results are shown in Table 8 below. As shown in Table 8 below, in order to sufficiently remove copper, it is preferable to add 2.0 equivalents or more of sulfidating agent relative to the copper, but it was found that the Ni residue rate and Co residue rate decrease as the amount of sulfidating agent added increases. Furthermore, as shown in Table 8 below, as the sulfidation pH increased, copper removal became insufficient, and the residual rates of Ni and Co tended to decrease. From the above, it was found that within the scope of this embodiment, an amount of sulfidating agent added of 2.0 equivalents and a sulfidation pH of 0 (zero) were preferable.

[0144]

[0145] <Addition of Oxidizing Agent> First, the copper removal solution obtained in Test Example 8-4 in Table 8 above was diluted five times. Next, hydrogen peroxide was added to the diluted copper removal solution as an oxidizing agent in the amount (unit: volume %) shown in Table 9 below and stirred. At this time, the oxidation pH and oxidation temperature (unit: °C) were adjusted to the values ​​shown in Table 9 below. Except for changing the amount of oxidizing agent added, the oxidation pH and oxidation temperature, the iron contained in the copper removal solution was precipitated as iron hydroxide in the same manner as in Test B described above to obtain a valuable element solution. In both Test A and Test B, no oxidizing aid was used. Furthermore, the Fe content (unit: mg / L), Ni residue rate (unit: mass %), and Co residue rate (unit: mass %) in the obtained valuable element solution were determined in the same manner as in Test B described above. The results are shown in Table 9 below. As shown in Table 9 below, it was found that in order to sufficiently remove iron, it is preferable to set the oxidation pH to 6.0, or to set the oxidation pH to 4.5 or higher and add an oxidizing agent. Furthermore, as shown in Table 9 below, as the oxidation pH increased, iron was removed more efficiently, while the Ni and Co residue rates tended to decrease. From the above, it was found that within the scope of this embodiment, the amount of oxidizing agent (hydrogen peroxide) added was preferably 0.030 volume%, and the oxidation pH was preferably in the range of 4.5 to 5.0.

[0146]

[0147] [Test D] <Preparation of Precursors> Precursors 1 to 5 were prepared as follows.

[0148] Precursor 1: To the valuable element solution of Test Example 9-4 described in Table 9 above, nickel sulfate (NiSO4) is used as a nickel source. ４ ) and cobalt sulfate (CoSO4) as a cobalt source. ４ ) and manganese sulfate (MnSO4) as a source of manganese ４) was added. In this way, a raw material aqueous solution was prepared in which the content of valuable elements (Ni, Co, Mn) was 1.25 mol / L and the Ni / Co / Mn (molar ratio) was 6 / 2 / 2. 0.35 L of pure water, sodium hydroxide aqueous solution, and ammonia aqueous solution were added to the reaction vessel to prepare a reaction vessel solution with a pH of 11.0. The raw material aqueous solution, ammonia aqueous solution as a complexing agent (concentration: 28 mass%), and sodium hydroxide aqueous solution as an alkaline aqueous solution (concentration: 48 mass%) were added dropwise to the reaction vessel solution to obtain a precipitate (precipitate of valuable elements). More specifically, the raw material aqueous solution was added dropwise to the reaction vessel solution at a rate of 4.0 mL / min while the complexing agent was added dropwise at a rate of 0.8 mL / min. During the addition of the raw material aqueous solution and the complexing agent, the alkaline aqueous solution was also added dropwise to control the pH of the reaction vessel solution to 11.0. During this time, the reaction vessel liquid was stirred with a stirring blade, and the temperature of the reaction vessel liquid was controlled to 40°C. The stirring speed (rotation speed of the stirring blade) was initially set to 200 rpm, but was increased to 350 rpm as the volume of liquid in the reaction vessel increased. Next, the obtained precipitate was filtered and washed with water, and then dried in a dryer at 100°C for 10 hours. Thus, precursor 1 was obtained.

[0149] Precursor 2 was obtained in the same manner as Precursor 1, except that the composition of the raw material aqueous solution was changed to Ni / Co / Mn (molar ratio) = 5 / 2 / 3.

[0150] Precursor 3 was obtained in the same manner as Precursor 1, except that the composition of the raw material aqueous solution was changed to Ni / Co / Mn (molar ratio) = 1 / 1 / 1.

[0151] Precursor 4 was obtained in the same manner as Precursor 1, except that the dropping rate of the complexing agent was changed to 0.6 mL / min.

[0152] <Precursor 5> Without using the valuable element solutions listed in Table 8 above, a raw material aqueous solution with a Ni / Co / Mn (molar ratio) of 6 / 2 / 2 was prepared using nickel sulfate, cobalt sulfate, and manganese sulfate as reagents. Otherwise, Precursor 5 was obtained in the same manner as Precursor 1.

[0153] Precursor 6 was obtained in the same manner as Precursor 1, except that the stirring speed was kept at 200 rpm from the beginning until the end of stirring (it was not increased to 350 rpm).

[0154] Precursor 7 was obtained in the same manner as Precursor 1, except that the stirring speed was set to 500 rpm from the beginning until the end of stirring.

[0155] <Characteristics of Precursors> For each of the obtained precursors 1 to 5, the tap density and particle size (D) were determined by the method described above. １０ , D ５０ and D ９０ The results were calculated and are shown in Table 10 below.

[0156]

[0157] As shown in Table 10 above, precursors 1 to 5 showed higher tap densities compared to precursor 6, which was stirred at a stirring speed of 200 rpm, and precursor 7, which was stirred at a stirring speed of 500 rpm.

[0158] <Manufacturing of Cathode Materials> Cathode materials 1 to 10 were manufactured using the obtained precursors 1 to 5 as follows.

[0159] 《Cathode Material 1》 A mixture was obtained by mixing precursor 1 with lithium hydroxide as a lithium-containing compound. The molar ratio (Li / (Ni+Co+Mn)) during mixing was 1.075. The obtained mixture was calcined to obtain a calcined product. More specifically, the mixture was calcined at 650°C for 8 hours in an air atmosphere, and then calcined at 950°C for 3 hours in an air atmosphere. The obtained calcined product was roughly crushed using a mortar. The calcined product was not washed with water. Thus, cathode material 1 was obtained.

[0160] Cathode materials 2 to 5 were obtained in the same manner as for cathode material 1, except that precursor materials 2 to 5 were used for each.

[0161] Cathode material 6 was obtained in the same manner as cathode material 1, except that a mixture was obtained by mixing precursor 1 with lithium carbonate as a lithium-containing compound.

[0162] Cathode materials 7 to 10 were obtained in the same manner as for cathode material 6, except that precursors 2 to cathode material 5 were used for each.

[0163] <Characteristics and Evaluation of Cathode Materials> For the obtained cathode materials 1 to 10, the tap density and particle size (D) were determined by the method described above. １０ , D ５０ and D ９０ The following was determined. Furthermore, the charging capacity, discharging capacity, cycle characteristics, and discharge capacity after the cycle test were determined as follows. The results are shown in Table 11 below.

[0164] 《Test 1: Charging and Discharging Capacity》 A mixture was obtained by adding N-methyl-2-pyrrolidone to positive electrode material (90% by mass), acetylene black (5% by mass), and polyvinylidene fluoride (5% by mass) and kneading. The obtained mixture was applied to an aluminum current collector to form a coating film. The laminate of the coating film and the aluminum current collector was pressed using a roll press to achieve a density of 3.1 to 3.3 g / cm³. ３ The laminate was pressurized to the specified range. A 14 mm diameter disc was punched out from the pressurized laminate. The punched disc was vacuum-dried at 150°C for 10 hours. The vacuum-dried disc was used as the positive electrode. A lithium metal sheet was used as the negative electrode. A porous polyethylene membrane (thickness: 16 μm, manufactured by Seigen Materials Co., Ltd.) was used as the separator. 1 mole of LiPF was added to 1 L of a mixed solution of ethylene carbonate (EC) and dimethyl carbonate (DMC) with a volume ratio (EC / DMC) of 1 / 1. ６ A non-aqueous electrolyte was obtained by dissolving the material. Using these positive electrode, negative electrode, separator, and non-aqueous electrolyte, a lithium-ion battery (test cell) for evaluation was fabricated in a glove box purged with argon. The fabricated test cell was charged and discharged at 25°C. Specifically, first, it was charged with a constant current of 0.05C, and when the voltage reached 4.3V, it was switched to constant voltage charging, and charging was terminated when the charging current dropped to 0.01C. Next, it was discharged with a constant current of 0.05C until the voltage reached 2.75V. In this way, the charging capacity (unit: mAh / g) and discharging capacity (unit: mAh / g) were determined.

[0165] 《Test 2 (Cycle Test): Cycle Characteristics》 First, the negative electrode was prepared. Specifically, pure water was added to artificial graphite (96.5% by mass), acetylene black (0.5% by mass), styrene-butadiene rubber (2% by mass), and carboxymethylcellulose (1% by mass) and kneaded to obtain a mixture. The obtained mixture was applied to a copper current collector to form a coating film. The laminate of the coating film and the copper current collector was pressed using a roll press to a density of 1.3 to 1.5 g / cm³. ３ The laminate was pressurized to maintain a certain range. A disc punched out from the pressurized laminate was used as the negative electrode. Except for the negative electrode, a lithium-ion battery (test cell) for evaluation was fabricated in the same manner as in Test 1 above. Using the fabricated test cell, 500 charge-discharge cycles (500 cycles) were repeated at 60°C with a current of 1.0C in the voltage range of 2.75 to 4.2V. From the obtained discharge capacity (unit: mAh / g), the cycle characteristics (unit: %) were determined using the following formula: Cycle characteristics = (Discharge capacity of the 500th cycle / Discharge capacity of the 1st cycle) × 100

[0166] 《Test 3: Discharge Capacity After Cycle Testing》 The positive electrode was removed from the test cell after the cycle test. Except for the removed positive electrode, a lithium-ion battery (test cell) for evaluation was prepared in the same manner as in Test 1 above. Using the prepared test cell, the discharge capacity (unit: mAh / g) was determined by charging and discharging in the same manner as in Test 1 above.

[0167]

[0168] As shown in Table 11 above, cathode materials 1 to 4, which used precursors 1 to 4 obtained using a valuable element solution, exhibited equivalent charging capacity, discharging capacity, cycle characteristics, and discharge capacity after cycle testing compared to cathode material 5, which did not use a valuable element solution. Furthermore, cathode materials 6 to 9, which used precursors 1 to 4 obtained using a valuable element solution, exhibited equivalent charging capacity, discharging capacity, cycle characteristics, and discharge capacity after cycle testing compared to cathode material 10, which did not use a valuable element solution.

Claims

1. A method for producing a precursor for a positive electrode material used in a lithium-ion battery, comprising: adding a reducing agent to an oxide containing at least one valuable element selected from the group consisting of nickel and cobalt, and manganese, and impurity elements consisting of copper and iron to obtain a mixed oxide; reducing the oxide by heating the mixed oxide to obtain a metal; contacting the metal with an acid solution to obtain a leachate containing the valuable element and the impurity element; adding a sulfiding agent to the leachate to precipitate copper as copper sulfide, and obtaining the leachate from which copper has been removed as a copper removal solution; adding an oxidizing agent to the copper removal solution to precipitate iron as iron hydroxide, and obtaining the copper removal solution from which iron has been removed as a valuable element solution containing the valuable element; introducing the valuable element solution, a complexing agent, and an alkaline aqueous solution into a reaction vessel to obtain a precipitate containing the valuable element; and the reducing agent containing at least one selected from the group consisting of a carbon-containing C-based reducing agent, a silicon-containing Si-based reducing agent, and an aluminum-containing Al-based reducing agent. A method for producing a precursor, wherein the amount R of the reducing agent added satisfies the following formula (1): (0.30a - 0.15b + 0.60) × [Ni] ≤ R ≤ (0.33a - 0.17b + 0.67) × [Ni] + (0.33a - 0.17b + 0.67) × [Co] + (0.23a - 0.12b + 0.47) × [Mn] ... (1) where, R: Amount of the reducing agent added (unit: moles) [Ni]: Nickel content of the oxide (unit: moles) [Co]: Cobalt content of the oxide (unit: moles) [Mn]: Manganese content of the oxide (unit: moles) a: Molar ratio of the content of the C-based reducing agent in the reducing agent (unit: moles) to the total amount of the reducing agent added (unit: moles) b: This is the molar ratio of the content of the Si-based reducing agent in the reducing agent (in molar parts) to the total amount of the reducing agent added (in molar parts).

2. The method for producing the precursor according to claim 1, wherein the oxide is obtained from a waste lithium-ion battery.

3. The method for producing a precursor according to claim 1 or 2, wherein the temperature at which the mixed oxide is heated is 1400°C or higher.

4. A method for producing a precursor according to any one of claims 1 to 3, wherein the metal obtained by heating the mixed oxide contains the valuable element and the impurity element.

5. A method for producing a precursor according to any one of claims 1 to 4, wherein the metal is powdered and then brought into contact with the acid solution.

6. The method for producing a precursor according to any one of claims 1 to 5, wherein the acid solution contains an acid and an oxidizing agent for the acid solution, and the content of the oxidizing agent for the acid solution is 0.5% by volume or more relative to the acid.

7. The method for producing a precursor according to claim 6, wherein the oxidizing agent for the acid solution is hydrogen peroxide.

8. A method for producing a precursor according to any one of claims 1 to 7, wherein the amount of sulfiding agent added is 1.0 equivalent or more relative to the copper contained in the leachate, and when precipitating the copper sulfide, the pH of the leachate to which the sulfiding agent has been added is 3.0 or less.

9. A method for producing a precursor according to any one of claims 1 to 8, wherein the oxidizing agent is at least one oxidizing agent A selected from the group consisting of air and ozone, or at least one oxidizing agent B selected from the group consisting of hydrogen peroxide, hypochlorous acid and potassium permanganate, the amount of oxidizing agent A added is 0.1 vvm or more relative to the copper removal solution, the amount of oxidizing agent B added is 0.005 volume% or more relative to the copper removal solution, and when precipitation of iron hydroxide, the pH of the copper removal solution to which the oxidizing agent has been added is 3.0 or more and 7.0 or less.

10. The method for producing a precursor according to claim 9, wherein the temperature of the copper removal solution to which the oxidizing agent has been added is 10°C or higher.

11. The method for producing a precursor according to any one of claims 1 to 10, wherein the alkaline aqueous solution is an aqueous solution of sodium hydroxide, and the complexing agent is at least one ammonium source selected from the group consisting of ammonia and ammonium salts.

12. A method for producing a positive electrode material for use in a lithium-ion battery, comprising: mixing a precursor obtained by a method for producing a precursor described in any one of claims 1 to 11 with a lithium-containing compound; and calcining the resulting mixture to obtain a calcined product containing the valuable element and lithium.

13. The method for producing a cathode material according to claim 12, wherein the lithium-containing compound is at least one selected from the group consisting of lithium hydroxide and lithium carbonate.

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