Method for producing precursor and method for producing positive electrode material

Abstract

According to the present invention, a reducing agent and a slag-forming agent having a mass ratio (CaO / SiO2) of 0.50 or less are 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, Co, and Mn; impurity elements copper and iron; and a quasi-valuable element Li. The oxide is reduced by heating the oxide-containing mixture to produce a slag that contains metal and the quasi-valuable element. 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 element.

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 (for example, Patent Document 1) in which the positive electrode material is heated together with a reducing agent and a slag-forming agent to reduce and generate the valuable elements.

[0005] Japanese Patent Publication No. 2021-95628, Japanese Patent Publication No. 2013-91826, Japanese Patent Publication No. 2012-224877

[0006] In the dry process, by reducing a composite oxide (LiNiO ２ , LiCoO ２ , LiMnO ２ ), slag is generated in addition to the metal containing valuable elements (Ni, Co, Mn). By the way, apart from valuable elements such as Ni, Co, and Mn, lithium (Li) may also be required. Hereinafter, Li is also referred to as a "semi-valuable element". However, in the dry process, a part of Li may volatilize, and in that case, it is necessary to separately recover Li, which is troublesome.

[0007] Further, 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 used 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 an oxide such as a positive electrode material of a used lithium-ion battery, and manufacturing a precursor and a positive electrode material containing the recovered valuable elements.

[0009] As a result of intensive studies, the present inventors have found that the above object can be achieved by adopting the following configuration, and have completed the present invention. That is, the present invention provides the following [1] to

[18] . [1] A method for manufacturing a precursor of a positive electrode material used in a lithium-ion battery, comprising: adding a reducing agent, CaO, and SiO ２ to an oxide containing at least one valuable element selected from the group consisting of nickel, cobalt, and manganese, impurity elements which are copper and iron, and a semi-valuable element which is lithium, and the mass ratio of CaO / SiO ２A method for producing a precursor, comprising: adding a slag-forming agent having a concentration of 0.50 or less to obtain a mixed oxide; heating the mixed oxide to reduce the oxide and obtain a slag containing a metal and the above-mentioned semi-valuable element; contacting the metal with an acid solution to obtain a leachate containing the above-mentioned valuable element and the above-mentioned impurity element; adding a sulfidating agent to the leachate to precipitate copper as copper sulfide; 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; obtaining the copper removal solution from which iron has been removed as a valuable element solution containing the above-mentioned valuable element; and introducing the valuable element solution, a complexing agent and an alkaline aqueous solution into a reaction vessel to obtain a precipitate containing the above-mentioned valuable element. [2] A method for producing a precursor according to [1], wherein the oxide is obtained from a waste lithium-ion battery. [3] A method for producing a precursor according to [1] or [2], wherein the amount of the reducing agent added is 2.0 equivalents or less. [4] A method for producing a precursor according to any one of [1] to [3] above, wherein the temperature at which the mixed oxide is heated is 1450°C or higher. [5] A method for producing a precursor according to any one of [1] to [4] above, wherein the metal obtained by heating the mixed oxide contains the valuable element and the impurity element. [6] A method for producing a precursor according to any one of [1] to [5] above, wherein the reducing agent is a carbon-containing substance containing carbon. [7] A method for producing a precursor according to any one of [1] to [6] above, wherein the reducing agent is at least one iron-containing substance selected from the group consisting of metallic iron and iron oxide. [8] A method for producing a precursor according to [7] above, wherein the iron oxide is ferrous oxide. [9] A method for producing a precursor according to [7] above, wherein the iron-containing substance is at least one selected from the group consisting of dust, scale, sludge and scrap.

[10] A method for producing a precursor according to any one of [1] to [9] above, wherein the metal is powdered and then brought into contact with the acid solution.

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

[10] above, 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.

[12] The method for producing a precursor according to

[11] above, wherein the oxidizing agent for the acid solution is hydrogen peroxide.

[13] A method for producing a precursor according to any one of [1] to

[12] 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.

[14] A method for producing a precursor according to any one of [1] to

[13] 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.

[15] The method for producing a precursor according to

[14] , wherein the temperature of the copper removal solution to which the oxidizing agent has been added is 10°C or higher.

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

[15] , 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.

[17] 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

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

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

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

[0010] The present invention provides a novel method for manufacturing precursors and cathode materials.

[0011] This is a flowchart illustrating an example of a method for recovering valuable elements. This is a flowchart illustrating an example of a process for manufacturing precursors and cathode materials. This is an Ellingham diagram (standard free energy change-temperature diagram). This is a potential-pH diagram (S-H) for Cu and Ni. ２ (O system). Potential-pH diagram of Fe and Ni (O ２ -H２ It is of the O type.

[0012] [Method for Manufacturing Precursors and Cathode Materials] The following describes a method for manufacturing cathode materials used in lithium-ion batteries (method for manufacturing cathode materials). The following description also describes a method for manufacturing precursors used in cathode materials (method for manufacturing precursors). First, as shown in Figure 1A, valuable elements are recovered from oxides such as cathode materials of waste lithium-ion batteries. That is, a solution of valuable elements containing valuable elements is obtained by subjecting the oxides to dry treatment and wet treatment. Then, as shown in Figure 1B, a precursor is manufactured using the obtained solution of valuable elements, and then the cathode material is manufactured.

[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 process, first, a reducing agent and a slag-forming agent, described later, are added to oxides (Ni, Co, Mn, Cu, Fe, Li) to obtain a mixed oxide. Next, the obtained mixed oxide is heated to reduce the oxide and obtain metals (Ni, Co, Mn, Cu, Fe) and slag (Li). The two are separated as appropriate. Before the wet process, it is preferable to powder the obtained metal to obtain metal powder (Ni, Co, Mn, Cu, Fe). In the wet process, first, the metal (metal powder) is brought into contact with an acid solution to obtain leachate (Ni, Co, Mn, 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-removing solution (Ni, Co, Mn, Fe). The two solutions are separated as appropriate. Then, an oxidizing agent is added to the copper-removing solution to precipitate iron hydroxide (Fe) and obtain a valuable element solution (Ni, Co, Mn).

[0014] In this way, valuable elements (Ni, Co, Mn) can be recovered from oxides while removing impurity elements (Cu, Fe), and even semi-valuable elements (Li) can be recovered. 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. Note that the following explanation also serves as an explanation of the metal manufacturing methods.

[0016] <Target for Reduction (Oxide)> The target for reduction is an oxide containing at least one valuable element selected from the group consisting of nickel (Ni), cobalt (Co), and manganese (Mn), impurity elements such as copper (Cu) and iron (Fe), and a semi-valuable element such as lithium (Li). Specifically, for example, it is the positive electrode material of a waste lithium-ion battery. The valuable element may be at least one element selected from the group consisting of Ni and Co. 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 and slag-forming agent (obtaining mixed oxide)> First, a reducing agent and a slag-forming agent are added to the oxide to be reduced to obtain a mixed oxide, which is a mixture of the oxide, the reducing agent, and the slag-forming agent.

[0018] [Knowledge obtained by the inventors] The cathode material of a lithium-ion battery is generally LiNiO ２ LiCoO ２ LiMnO ２ It consists of oxides (complex oxides) such as the above. If we consider the dry process thermodynamically, for example, LiNiO ２ and LiCoO ２ At high temperatures, it decomposes as follows, producing NiO and CoO, respectively: 2LiNiO ２ →Li ２ O+2NiO+1 / 2O ２ 2LiCoO ２ →Li ２ O+2CoO+1 / 2O ２

[0019] Standard free energy change (ΔG) in the decomposition reactions of NiO and CoO ０ The following are the respective reactions: NiO → Ni + 1 / 2O ２ ΔG ０ =234900-84.68T[J] CoO→Co+1 / 2O ２ ΔG ０= 235480 - 71.55T [J] Substances having a free energy change value lower than these standard free energy change values ​​at any high temperature can be used as reducing agents.

[0020] When recovering valuable elements as metals from complex oxides by dry treatment, calcium oxide (CaO) and silicon dioxide (SiO₂) are used in addition to the reducing agent. ２ A slag-forming agent containing ) may be used. CaO and SiO contained in the slag-forming agent ２ The mass ratio of (CaO / SiO ２ ) is also called basicity, and conventionally, a value of 0.66 to 2.00 has been considered preferable (Patent Documents 2-3). However, the present inventors have found that the mass ratio (CaO / SiO ２ By using a low-basicity slag-forming agent with a basicity of 0.50 or less, it was found that a generated slag containing valuable elements such as Ni and Co, as well as a large amount of Li, can be obtained, and Li can be recovered simply and efficiently. The method for further recovering Li from the generated slag is not particularly limited, and various methods can be cited, such as recovering it in the form of lithium carbonate by wet treatment.

[0021] Reducing agents include, for example, aluminum-containing substances such as metallic aluminum (Al); silicon-containing substances such as metallic silicon (Si) and FeSi; carbon-containing substances; iron-containing substances; and so on.

[0022] If the reducing agent's reducing power is insufficient, resulting in incomplete reduction, a portion of the cathode material may separate as slag in its oxide form, reducing the content of valuable elements in the metal obtained by reducing the cathode material (reducing the reduction rate of valuable elements). For example, aluminum-containing materials, silicon-containing materials, and carbon-containing materials have strong reducing power, which can help avoid such incomplete reduction and is expected to yield a high reduction rate.

[0023] Examples of carbon-containing substances include solid carbon-containing substances such as graphite, coke, and solid hydrocarbons; and gaseous carbon-containing substances such as carbon monoxide (CO) and hydrocarbon gases (e.g., propane gas). When a carbon-containing substance is used as a reducing agent, CO is produced after reduction.２ H ２ This is preferable because it does not generate gases such as oxygen, thus avoiding an increase in the amount of slag produced.

[0024] The iron-containing substance is at least one selected from the group consisting of metallic iron (Fe) and iron oxide. The iron-containing substance as a reducing agent is described below.

[0025] Figure 2 is the Ellingham diagram (standard free energy change-temperature diagram). Referring to the standard free energy change and the Ellingham diagram (Figure 2) described above, the Fe / FeO equilibrium is less virtuous than the Ni / NiO equilibrium and the Co / CoO equilibrium, suggesting the possibility of reduction by Fe. Also, FeO / Fe ３ O ４ The equilibrium is less noble than the Ni / NiO equilibrium but more noble than the Co / CoO equilibrium. Therefore, it is expected that Ni will be included in the resulting metal and Co in the resulting slag. Specifically, the following reaction is expected: NiO + Fe → Ni + FeO: ΔG ０ =-29530-19.95T[J] CoO+Fe→Co+FeO:ΔG ０ =-28950-6.82T[J]

[0026] In the Ellingham diagram (Figure 2), the higher up a compound is, the more easily it is metallized. When Si or Al is used as a reducing agent, Mn is also easily metallized. Therefore, by using Fe (or FeO) as a reducing agent, it is possible to metallize only Ni and Co while preventing Mn from being metallized.

[0027] For metallic iron (Fe), for example, scrap or granular iron used in steel mills may be used.

[0028] Iron oxides are generally ferrous oxide (FeO), also called wustite, and triiron tetroxide (FeO), also called magnetite. ３ O ４ ) and ferric oxide (Fe), also known as hematite ２ O ３Iron oxide is classified into three types. Of these, magnetite and hematite have higher standard free energy changes than wustite at the same temperature, and may not readily induce reduction reactions. For this reason, ferrous oxide (wustite) is preferred as iron oxide because it readily induces reduction reactions. Iron oxide may also be at least one of the dust, scale, and sludge (hereinafter referred to as "dust" for convenience) that are produced as by-products in the steelmaking process. Using dust as iron oxide is preferable from the viewpoint of effectively utilizing by-products of the steelmaking process and utilizing an inexpensive iron source.

[0029] (Amount of reducing agent added) The amount of reducing agent required to reduce the oxide to be reduced is called 1.0 equivalent. For example, the reducing agent may be metallic iron (Fe), ferrous oxide (FeO), metallic aluminum (Al), metallic silicon (Si), coke (C), and propane (C). ３ H ８ In the case of ), the reduction using 1.0 equivalent of reducing agent is expressed as follows: Fe + (NiO, CoO, MnO) → (Ni, Co, Mn) + FeO 3FeO + (NiO, CoO, MnO) → (Ni, Co, Mn) + Fe ３ O ４ 2Al+3(NiO, CoO, MnO) → 3(Ni, Co, Mn)+Al ２ O ３ Si+2(NiO, CoO, MnO) → 2(Ni, Co, Mn)+SiO ２ C+2(NiO, CoO, MnO) → 2(Ni, Co, Mn)+CO ２ C ３ H ８ +10 (NiO, CoO, MnO) → 10 (Ni, Co, Mn) +3CO ２ +4H ２ O

[0030] 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).

[0031] The amount of reducing agent added is preferably 1.0 equivalent or more, more preferably 1.2 equivalents or more, even more preferably 1.4 equivalents or more, and particularly preferably 1.6 equivalents or more, because it helps to suppress insufficient reduction.

[0032] Incidentally, when a large amount of reducing agent is added to an oxide to avoid poor reduction, the amount of elements other than valuable elements (e.g., Fe) in the resulting metal tends to increase. For example, when using an iron-containing substance as a reducing agent, in order to suppress the Fe content in the resulting metal (e.g., to 1% by mass or less), it is necessary to reduce the amount of reducing agent added (e.g., to 0.8 equivalents or less), and in this case, it is difficult to suppress poor reduction. However, since Fe can be removed by wet treatment, the amount of reducing agent added can be increased (e.g., to 1.6 equivalents or more), making it easier to suppress poor reduction.

[0033] Therefore, there is no particular upper limit to the amount of reducing agent to be added. However, from the viewpoint of easily removing Fe by wet treatment, it is preferable that the amount of Fe contained in the generated metal be small. Also, the effect saturates even if an excess of reducing agent is added. For this reason, the amount of reducing agent to be added is preferably 2.0 equivalents or less, and more preferably 1.8 equivalents or less.

[0034] 《Slag-forming agents》 Calcium oxide (CaO) and silicon dioxide (SiO₂ ２ A slag-forming agent containing ) is used. CaO and SiO in the slag-forming agent ２ The content (total content) is preferably 90% by mass or more, more preferably 95% by mass or more, even more preferably 98% by mass or more, and particularly preferably 100% by mass.

[0035] (Mass ratio (CaO / SiO ２CaO and SiO ２ The mass ratio of (CaO / SiO ２ Use a low-basicity slag-forming agent. Mass ratio of slag-forming agent (CaO / SiO ２ The mass ratio of the slag-forming agent (CaO / SiO) is 0.50 or less, preferably 0.48 or less, more preferably 0.46 or less, even more preferably 0.44 or less, particularly preferably 0.42 or less, and most preferably 0.35 or less. On the other hand, the lower limit is not particularly limited, and the mass ratio of the slag-forming agent (CaO / SiO) is also acceptable. ２ For example, the ratio is 0.15 or higher, preferably 0.20 or higher, more preferably 0.25 or higher, and even more preferably 0.30 or higher.

[0036] (Amount of slag-forming agent added) The amount of slag-forming agent added is not particularly limited, but the mass ratio of the slag-forming agent to the oxide to be reduced (slag-forming agent / oxide) is preferably 0.40 to 1.00, more preferably 0.45 to 0.85, and even more preferably 0.50 to 0.80.

[0037] <Heating the mixed oxide (obtaining metal and slag)> Next, the mixed oxide (a mixture of oxide, reducing agent, and slag-forming agent) is heated. This reduces the oxide. The equipment used for heating the mixed oxide is not particularly limited and includes conventionally known equipment such as arc furnaces, submerged arc furnaces, resistance furnaces, high-frequency melting furnaces, low-frequency melting furnaces, rotary kilns, vertical furnaces, and steelmaking furnaces.

[0038] <Heating Temperature> The heating temperature when heating the mixed oxide is preferably 1300°C or higher, more preferably 1350°C or higher, even more preferably 1400°C or higher, and particularly preferably 1450°C or higher, because it helps to suppress poor reduction. There is no particular upper limit, but the heating temperature is preferably 1800°C or lower, and more preferably 1700°C or lower.

[0039] 《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 a argon (Ar) atmosphere and a reducing atmosphere such as a carbon monoxide (CO) atmosphere.

[0040] 《Heating Time》 The heating time for the mixed oxide (heating time) 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. There is no particular upper limit, but the heating time is preferably 6 hours or less, and more preferably 5 hours or less.

[0041] 《Products (Metals)》 By reducing the oxide (cathode material) that is the target of reduction, metals are produced. In other words, the valuable elements (Ni, Co, Mn) contained in the oxide are recovered as metals.

[0042] The metal obtained by the reduction of oxides (also called the "product metal") is an alloy containing valuable elements (Ni, Co, Mn) and impurity elements (Cu, Fe). The product metal may contain only one of the valuable elements (Ni, Co, Mn). The product metal may contain, for example, at least one element selected from the group consisting of nickel (Ni) and cobalt (Co). The product metal may contain one valuable element in proportion greater than the proportion of other valuable elements.

[0043] 《Products (Slag)》 By reducing the oxide (cathode material), slag is produced in addition to the metal. The slag produced (also called "generated slag") contains a large amount of Li, as mentioned above. When an iron-containing substance is used as a reducing agent, the generated slag contains FeO, etc. In addition, the generated slag may also contain oxides of valuable elements not included in the generated metal (e.g., MnO). When reducing an oxide containing Mn, by using an iron-containing substance as a reducing agent, the Mn / MnO equilibrium changes to an Fe / FeO equilibrium and FeO / Fe ３ O ４ Because it is less efficient than equilibrium, it can suppress the incorporation of Mn into the resulting metal obtained by reduction. Separating Mn from the resulting metal by wet processing is computationally intensive. Suppressing the incorporation of Mn into the resulting metal and retaining Mn in the resulting slag is beneficial because it reduces this computational burden.

[0044] <Separation of Metal and Slag> It is preferable to separate the generated metal and slag obtained by the reduction of oxides before pulverizing the generated metal (described later). The separation method is not particularly limited, and known methods can be used.

[0045] <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.

[0046] 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).

[0047] 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.

[0048] <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, Mn) 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 (leaching 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.

[0049] 《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.

[0050] Acidic solutions: Acidic solutions that come into contact with metals must contain at least an acid.

[0051] (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.

[0052] ((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.

[0053] (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.

[0054] ((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).

[0055] <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.

[0056] <Separation of leachate and leachate residue> It is preferable to separate the leachate and leachate residue before adding the sulfurizing agent (described later). The separation method is not particularly limited, and known solid-liquid separation methods can be used.

[0057] <Addition of sulfiding agent (obtaining copper removal solution)> Next, a sulfiding agent is added to the leachate containing valuable elements (Ni, Co, Mn) 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.

[0058] Figure 3 shows the potential-pH diagram (S-H) of copper (Cu) and nickel (Ni). ２ It is an O-system. In Figure 3, copper (Cu) - sulfur (S) - water (H) ２ Figure 3 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 3. As shown in Figure 3, 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 3, in this region, copper precipitates as copper(II) sulfide (CuS). By utilizing this, the copper (Cu) contained in the leachate is precipitated as copper(II) sulfide and selectively removed by making the leachate low pH and reducing. That is, a copper removal solution is obtained, which is an leachate from which copper (Cu) has been removed.

[0059] 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.

[0060] 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.

[0061] (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.

[0062] For example, when producing copper(II) sulfide (CuS) using 1.0 equivalent of sodium hydrogen sulfide (NaSH) as a sulfiding agent, 1 mol of sodium hydrogen sulfide (NaSH) is used for every 1 mol of copper (Cu) contained in the leachate.

[0063] 《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.

[0064] <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.

[0065] <Separation of copper sulfide and copper removal solution> It is preferable to separate the copper sulfide and the copper removal solution before adding the oxidizing agent (described later). The separation method is not particularly limited, and known solid-liquid separation methods can be used.

[0066] <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, Mn) 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, Mn).

[0067] Figure 4 shows the potential-pH diagram (O) of iron (Fe) and nickel (Ni). ２ -H ２ It is an O system. In Figure 4, iron (Fe) - oxygen (O ２ ) - Water (H ２ Figure 4 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 4. As shown in Figure 4, 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 4, in this region, iron precipitates as iron(III) oxide hydroxide (FeO(OH)). By utilizing this, by making the copper removal solution acidic to neutral and oxidizing, the iron (Fe) contained in the copper removal solution is precipitated as iron(III) oxide hydroxide and selectively removed. That is, a valuable element solution, which is a copper removal solution from which iron (Fe) has been removed, is obtained.

[0068] <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.

[0069] (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.

[0070] 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.

[0071] 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.

[0072] <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.

[0073] 《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.

[0074] <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.

[0075] 《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.

[0076] (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.

[0077] (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.

[0078] <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.

[0079] <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.

[0080] 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.

[0081] 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.

[0082] 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.

[0083] 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.

[0084] 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.

[0085] 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.

[0086] 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.

[0087] As the alkaline aqueous solution, an aqueous solution of sodium hydroxide (NaOH) is preferred.

[0088] 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.

[0089] 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.

[0090] 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.

[0091] 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.

[0092] As described above, for example, a precursor of a positive electrode material can be obtained by subjecting a valuable element precipitate to, for example, water washing and drying. In the precursor, the molar ratio of nickel content (Ni / (Ni + Co + Mn)) to the total content of valuable elements (Ni, Co, Mn) is preferably 0.3 or more, more preferably 0.4 or more. Also, this molar ratio (Ni / (Ni + Co + Mn)) is preferably 1.0 or less, more preferably 0.8 or less.

[0093] The tap density of the precursor is preferably 0.8 g / cm ３ or more, more preferably 1.0 g / cm ３ or more, still more preferably 1.2 g / cm ３ or more. Also, the tap density of the precursor may be 1.8 g / cm ３ or less, and may be 1.5 g / cm ３ or less. The tap density is determined by putting a sample into a container with a volume of 100 cm ３ which is left standing, tapping with a tapping device until the volume of the sample no longer decreases, and then dividing the mass of the sample (unit: g) by the volume of the sample (unit: cm ３ )(hereinafter the same).

[0094] The particle size D １０ of the precursor is preferably 3.0 μm or more, more preferably 4.0 μm or more. Also, the particle size D １０ of the precursor is preferably 10.0 μm or less, more preferably 8.0 μm or less. The particle size D ５０ of the precursor is preferably 8.0 μm or more, more preferably 9.0 μm or more. Also, the particle size D ５０ of the precursor is preferably 16.0 μm or less, more preferably 14.0 μm or less. The particle size D ９０ of the precursor is preferably 12.0 μm or more, more preferably 14.0 μm or more. Also, the particle size D ９０ of the precursor is preferably 24.0 μm or less, more preferably 22.0 μm or less. The particle size D １０ , the particle size D ５０ and the particle size D ９０ are respectively the particle sizes at which the cumulative frequency of the particle size distribution determined by the laser diffraction / scattering method is 10%, 50% and 90% in volume percentage (hereinafter the same).

[0095] <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.

[0096] 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).

[0097] The following describes each step in the manufacturing method of the cathode material in more detail.

[0098] 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.

[0099] Suitable lithium-containing compounds include, for example, at least one selected from the group consisting of lithium hydroxide and lithium carbonate.

[0100] 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.

[0101] 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.

[0102] 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.

[0103] 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.

[0104] 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.

[0105] 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.

[0106] 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.

[0107] 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.

[0108] The present invention will be specifically described below with reference to examples. However, the present invention is not limited to the examples described below.

[0109] [Test A] <Preparation of Cathode Material> Cathode material 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 composition ratio (molar ratio) of nickel (Ni), cobalt (Co), and manganese (Mn) in the cathode material is shown in Table 1 below. The cathode material also contained copper (Cu) and iron (Fe) as impurity elements. Furthermore, the cathode material also contained lithium (Li) as a semi-valuable element.

[0110]

[0111] <Preparation of Reducing Agents> The reducing agents are: metallic aluminum (Al) powder, metallic silicon (Si) powder, coke (C) powder, and propane gas (C) ３ H ８ ) was prepared. Furthermore, as reducing agents, powder of metallic iron (Fe) obtained by atomization and powder of ferrous oxide (FeO) were prepared.

[0112] <Preparation of slag-forming agent> CaO and SiO ２ A slag-forming agent consisting of CaO and SiO was prepared. ２ The mass ratio of (CaO / SiO ２ Several types of slag-forming agents with different properties were prepared.

[0113] <Addition of Reducing Agent and Slag-Building Agent> The prepared cathode material was placed in a submerged arc furnace with a heat size of 150 kg, and a reducing agent and slag-building agent were added to obtain a mixed oxide. More specifically, 30 kg of slag-building agent was added for every 45 kg of cathode material. That is, the mass ratio of slag-building agent to cathode material (slag-building agent / cathode material) was approximately 0.67. The type and amount of reducing agent used (unit: equivalent), and the mass ratio of the slag-building agent used (CaO / SiO) are as follows: ２ The results are shown in Table 2 below.

[0114] <Heating of mixed oxide> The obtained mixed oxide was heated. The heating time was 3 hours, and the heating atmosphere was an Ar atmosphere. The heating temperature (unit: °C) is shown in Table 2 below. In this way, the cathode material was reduced to obtain the generated metal and generated slag, and the two were separated.

[0115] For each of the metal elements Ni, Co, and Mn, the reduction rate (in mass%) was calculated based on the following formula, and the unit was converted from "mass%" to "mol%". The results are shown in Table 2 below. Reduction rate = 100 × (Amount of metal element contained in the generated metal [kg]) / (Amount of metal element contained in the oxide to be reduced [kg])

[0116] Furthermore, the residual rate of Li in the generated slag (unit: mass%) was calculated based on the following formula, and the unit was converted from "mass%" to "mol%". The results are shown in Table 2 below. Residual rate in generated slag = 100 × (amount of Li contained in the generated slag [kg]) / (amount of Li contained in the oxide to be reduced [kg])

[0117] As shown in Table 2 below, the mass ratio (CaO / SiO ２ Test Examples 2-1 to 2-8, which used a slag-forming agent with a mass ratio (CaO / SiO2) of 0.50 or less, showed that the mass ratio (CaO / SiO2) was 0.50 or less. ２ Compared to comparative test examples 2-1 to 2-2, which used a slag-forming agent with a ratio of over 0.50, the residual rate of Li in the generated slag was higher.

[0118] In Test Examples 2-1 to 2-8, the reduction rates of Ni and Co were high, while the reduction rate of Mn was suppressed to 20% or less. In particular, in Test Examples 2-5 to 2-6, where Fe or FeO was used as the reducing agent, the reduction of Mn was suppressed even more effectively. Note that in Test Example 2-6, the reduction rate of Co was lower than in Test Example 2-5. This is thought to be because the free energy during FeO formation is slightly lower than the free energy during CoO formation (see Figure 2). However, since FeO produces Fe upon reduction, its reduction potential is higher than that of FeO, and overall, it is thought to have a certain degree of reduction potential, and this result is reflected in the results.

[0119] Comparing Test Examples 2-1 to 2-6 with Test Examples 2-7 to 2-8, the mass ratio of the slag-forming agent (CaO / SiO ２ In Test Examples 2-7 to 2-8, where the reduction rate was lower, the reduction of Mn was further suppressed without causing a significant decrease in the reduction rates of Ni and Co, compared to Test Examples 2-1 to 2-6.

[0120]

[0121] <Powdering of Metals> The composition of the metals and slag produced by the reduction of the cathode material was determined. Of the metals produced by the reduction of the cathode material, those having the composition shown in Table 3 below were powdered using a vibratory mill to obtain metal powder. The particle size of the obtained metal powder was 1100 μm.

[0122] <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 3 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 a 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 3 below. As shown in Table 3 below, the leaching rate for each element was 100 mass%, indicating that all elements were successfully leached from the metal into the leachate.

[0123]

[0124] <Addition of Sulfidating Agent> The content of each element in the obtained leachate (unit: g / L) is shown in Table 4 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 4 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 4 below. As shown in Table 4 below, the Cu content in the copper removal solution is very low, indicating that Cu was removed from the leachate with very high efficiency.

[0125] <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 4 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 4 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 4 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 4 below. As shown in Table 4 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.

[0126] 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 4 below. From the results shown in Table 4 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, Mn) could be recovered with very high purity.

[0127]

[0128] [Test B] <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 A, so the explanation will be omitted.

[0129] <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 5 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 A 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 A described above. The results are shown in Table 5 below. As shown in Table 5 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, Mn). 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.

[0130]

[0131] <Addition of Sulfidating Agent> Sodium hydrogen sulfide (NaSH) was added as a sulfidating agent to the leachate shown in Table 4 above, in the amount (unit: equivalent) shown in Table 6 below, and the mixture was stirred. At this time, the sulfidation pH was adjusted to the value shown in Table 6 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 A 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 A described above. The results are shown in Table 6 below. As shown in Table 6 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 6 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) are preferable.

[0132]

[0133] <Addition of Oxidizing Agent> First, the copper removal solution obtained in Test Example 6-4 in Table 6 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 7 below and stirred. At this time, the oxidation pH and oxidation temperature (unit: °C) were adjusted to the values ​​shown in Table 7 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 A 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 A above. The results are shown in Table 7 below. As shown in Table 7 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 7 below, as the oxidation pH increased, iron was removed more efficiently, while the residual rates of Ni and Co 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.

[0134]

[0135] [Test C] In accordance with Test A, dry and wet treatments (i.e., from the preparation of the cathode material to the addition of the oxidizing agent) were carried out to obtain a solution of valuable elements. For convenience, as shown in Table 8 below, the mass of the cathode material (oxide) used was assumed to be 100 kg. Test Examples 8-1 to 8-3 differ in the type and amount of reducing agent added. In comparative test example 8-1, only the dry treatment was performed. In comparative test example 8-2, the mass ratio (CaO / SiO ２ A slag-forming agent with a ratio of 1.50 was used.

[0136] The Ni+Co recovery rate (unit: mass%) was determined from the Ni+Co reduction rate (unit: mass%) of the generated metal obtained by dry treatment and the Ni+Co residue rate (unit: mass%) of the valuable element solution obtained by wet treatment. The results are shown in Table 8 below. In the case where only dry treatment was performed, the Ni+Co reduction rate of the generated metal was used as the Ni+Co recovery rate and is listed in Table 8 below.

[0137] Furthermore, for the valuable element solutions obtained by wet treatment, the Ni + Co ratio in the metal elements was determined as the final Ni + Co purity (unit: mass%). The results are shown in Table 8 below. However, in comparative test example 8-1, where only dry treatment was performed, the Ni + Co purity of the generated metal obtained by dry treatment was recorded in Table 8 below as the final Ni + Co purity.

[0138] Comparing comparative test example 8-1 with test example 8-1, although the dry treatment conditions were the same, test example 8-1, which underwent wet treatment, ultimately yielded a higher purity of valuable elements (Ni + Co) compared to comparative test example 8-1, which did not undergo wet treatment.

[0139] When comparing comparative test example 8-2, which uses Al as a reducing agent, with test example 8-1, the mass ratio of the slag-forming agent (CaO / SiO ２ In Test Example 8-1, where the ratio was 0.50, the residual rate of Li in the generated slag was higher and more Li was recovered compared to Comparative Test Example 8-2, where the ratio was 1.50.

[0140] Comparing Test Example 8-1 and Test Example 8-2, Test Example 8-2, which used Fe as the reducing agent, had a higher residual rate of Li in the generated slag and allowed for the recovery of more Li than Test Example 8-1, which used Al as the reducing agent.

[0141] Comparing Test Example 8-2 and Test Example 8-3, both using Fe as a reducing agent, Test Example 8-3, with an added amount of 1.6 equivalents of the reducing agent, had a lower Ni+Co purity in the resulting metal than Test Example 8-2, which used 1.4 equivalents. In other words, increasing the amount of Fe added as a reducing agent during the dry treatment increased the amount of Fe in the resulting metal, thus decreasing the Ni+Co purity of the resulting metal. However, ultimately, Test Example 8-3 had a higher Ni+Co recovery rate and a similarly high Ni+Co purity compared to Test Example 8-2. Thus, by performing a wet treatment after the dry treatment, a high purity of valuable elements can ultimately be obtained. Therefore, by increasing the amount of Fe added during the dry treatment, the valuable elements (Ni+Co) can be sufficiently reduced. As a result, a high final Ni+Co recovery rate can be obtained by combining the dry and wet treatments.

[0142]

[0143] [Test D] <Preparation of Precursors> Precursors 1 to 5 were prepared as follows.

[0144] Precursor 1: To the valuable element solution of Test Example 8-1 described in Table 8 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.

[0145] 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.

[0146] 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.

[0147] 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.

[0148] <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.

[0149] 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).

[0150] 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.

[0151] <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 following was calculated. The results are shown in Table 9 below.

[0152]

[0153] As shown in Table 9 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.

[0154] <Manufacturing of Cathode Materials> Cathode materials 1 to 10 were manufactured using the obtained precursors 1 to 5 as follows.

[0155] 《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.

[0156] 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.

[0157] 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.

[0158] 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.

[0159] <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 10 below.

[0160] 《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.

[0161] 《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

[0162] 《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.

[0163]

[0164] As shown in Table 10 above, cathode materials 1 to 4, which use 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 use 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

A method for producing a precursor for a positive electrode material used in lithium-ion batteries, An oxide containing at least one valuable element selected from the group consisting of nickel, cobalt, and manganese, impurity elements being copper and iron, and a semi-valuable element being lithium, is mixed with a reducing agent, CaO, and SiO ２ The mass ratio of (CaO / SiO ２ A slag-forming agent having a value of 0.50 or less is added to obtain a mixed oxide. By heating the mixed oxide, the oxide is reduced to obtain a metal and a slag containing the aforementioned quasi-valuable element. The metal is brought into contact with an acid solution to obtain a leachate containing the valuable element and the impurity element. A sulfiding agent is added to the aforementioned leachate to precipitate copper as copper sulfide, and the leachate from which copper has been removed is obtained as a copper removal solution. An oxidizing agent is added to the copper removal solution to precipitate iron as iron hydroxide, and the copper removal solution from which iron has been removed is obtained as a valuable element solution containing the valuable element. A method for producing a precursor, comprising introducing the above-mentioned valuable element solution, a complexing agent, and an alkaline aqueous solution into a reaction vessel to obtain a precipitate containing the above-mentioned valuable element. 　 The method for producing the precursor according to claim 1, wherein the oxide is obtained from a waste lithium-ion battery. 　 A method for producing a precursor according to claim 1 or 2, wherein the amount of the reducing agent added is 2.0 equivalents or less. 　 A method for producing a precursor according to any one of claims 1 to 3, wherein the temperature at which the mixed oxide is heated is 1450°C or higher. 　 A method for producing a precursor according to any one of claims 1 to 4, wherein the metal obtained by heating the mixed oxide contains the valuable element and the impurity element. 　 A method for producing a precursor according to any one of claims 1 to 5, wherein the reducing agent is a carbon-containing substance containing carbon. 　 The method for producing a precursor according to any one of claims 1 to 6, wherein the reducing agent is at least one iron-containing substance selected from the group consisting of metallic iron and iron oxide. 　 The method for producing the precursor according to claim 7, wherein the iron oxide is ferrous oxide. 　 The method for producing a precursor according to claim 7, wherein the iron-containing substance is at least one selected from the group consisting of dust, scale, sludge, and scrap. 　 A method for producing a precursor according to any one of claims 1 to 9, wherein the metal is powdered and then brought into contact with the acid solution. 　 The aforementioned acid solution contains an acid and an oxidizing agent for the acid solution. A method for producing a precursor according to any one of claims 1 to 10, wherein the content of the oxidizing agent for the acid solution is 0.5% by volume or more relative to the acid. 　 The method for producing a precursor according to claim 11, wherein the oxidizing agent for the acid solution is hydrogen peroxide. 　 The amount of sulfidizing agent added is 1.0 equivalent or more relative to the copper contained in the leachate. A method for producing a precursor according to any one of claims 1 to 12, wherein when precipitating the copper sulfide, the pH of the leachate to which the sulfiding agent has been added is set to 3.0 or less. 　 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% by volume or more relative to the copper removal solution. A method for producing a precursor according to any one of claims 1 to 13, wherein when precipitating the iron hydroxide, the pH of the copper removal solution to which the oxidizing agent has been added is set to 3.0 or more and 7.0 or less. 　 The method for producing a precursor according to claim 14, wherein the temperature of the copper removal solution to which the oxidizing agent has been added is 10°C or higher. 　 The aforementioned alkaline aqueous solution is an aqueous solution of sodium hydroxide. The method for producing a precursor according to any one of claims 1 to 15, wherein the complexing agent is at least one ammonium source selected from the group consisting of ammonia and ammonium salts. 　 A method for manufacturing a positive electrode material used in lithium-ion batteries, A method for producing a positive electrode material, comprising mixing a precursor obtained by the method for producing a precursor described in any one of claims 1 to 16 with a lithium-containing compound, and calcining the resulting mixture to obtain a calcined product containing the valuable element and lithium. 　 The method for producing a cathode material according to claim 17, wherein the lithium-containing compound is at least one selected from the group consisting of lithium hydroxide and lithium carbonate.

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