Precursor manufacturing method and cathode material manufacturing method

Abstract

A reducing agent and a slag-forming agent with a mass ratio (CaO / SiO2) of 0.50 or less are added to an oxide containing 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 to obtain a mixed oxide. The mixed oxide is reduced by heating to obtain a slag containing the metal and the quasi-valuable element. The metal is contacted with an acid solution to obtain a leachate containing the valuable element and the impurity element. A sulfiding agent is added to the leachate to precipitate copper as copper sulfide, thereby obtaining a copper-removed solution. An oxidizing agent is added to the copper-removed solution to precipitate iron as iron hydroxide, thereby obtaining a valuable element solution containing the valuable element. The valuable element solution, a complexing agent, and an alkaline aqueous solution are introduced into a reaction vessel liquid to obtain a precipitate containing the valuable element.

Description

[Technical Field]

[0001] The present invention relates to a method for producing a precursor and a method for producing a cathode material. [Background technology]

[0002] In recent years, the demand for lithium-ion batteries has been increasing rapidly due to the widespread use of electric vehicles. In particular, given the current trend toward reducing CO2 emissions, demand for electric vehicles that do not use fossil fuels is expected to continue to expand in the future, and the resulting demand for lithium-ion batteries is also expected to increase further.

[0003] Generally, the cathode material of a lithium-ion battery is made of an oxide (composite oxide) containing nickel (Ni), cobalt (Co), manganese (Mn), etc. Specific examples of this composite oxide include LiNiO2, LiCoO2, and LiMnO2. Metal elements such as Ni, Co, and Mn cannot be said to be abundant even on a global scale. Therefore, recovering these metal elements (valuable elements) from the positive electrode material of waste lithium-ion batteries is extremely beneficial from the perspective 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] A lithium ion battery is composed of a combination of components such as a positive electrode material, a negative electrode material, and a separator, and also contains an electrolyte solution. Therefore, when recovering valuable elements from the positive electrode material of used lithium-ion batteries, preliminary treatment such as removal of the electrolyte, crushing, and crushing is carried out prior to recovery. After such pre-treatment, the positive electrode material is separated from the waste lithium-ion batteries, and then valuable elements are recovered from the separated positive electrode material. Examples of treatments for recovering valuable elements include dry treatments in which the positive electrode material is heated together with a reducing agent and a slag-forming agent to reduce and generate valuable elements (for example, Patent Document 1). [Prior art documents] [Patent documents]

[0005] [Patent Document 1] Patent Publication No. 2021-95628 [Patent Document 2] Japanese Patent Application Laid-Open No. 2013-91826 [Patent Document 3] Japanese Patent Application Laid-Open No. 2012-224877 Summary of the Invention [Problem to be solved by the invention]

[0006] In dry processing, the reduction of complex oxides (LiNiO2, LiCoO2, LiMnO2) produces metals containing valuable elements (Ni, Co, Mn) as well as slag. In addition to valuable elements such as Ni, Co, and Mn, lithium (Li) is also sometimes required. Hereinafter, Li will also be referred to as a "quasi-valuable element." However, in dry processing, some of the Li may volatilize, and in that case, it is necessary to recover the Li separately, which is cumbersome.

[0007] Furthermore, metals obtained by dry processing may contain impurity elements in addition to valuable elements such as Ni and Co. Impurity elements include copper (Cu) and iron (Fe) derived from waste lithium-ion batteries. When metals obtained by dry processing are reused as cathode materials for lithium-ion batteries, impurity elements (Cu and Fe) contained in the metals can deteriorate battery performance, so it is desirable to remove as many impurity elements as possible.

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

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

[18] . [1] A method for producing a precursor of a cathode material for use in a lithium-ion battery, comprising: adding a reducing agent and a slag-forming agent having a CaO to SiO2 mass ratio (CaO / SiO2) of 0.50 or less to an oxide containing at least one valuable element selected from the group consisting of nickel, cobalt, and manganese, impurity elements such as copper and iron, and a quasi-valuable element such as lithium, to obtain a mixed oxide; heating the mixed oxide to reduce the oxide to obtain a slag containing a metal and the quasi-valuable element; and a method for producing a precursor, comprising the steps of: contacting the metal with an acid solution to obtain a leachate containing the valuable element and the impurity element; adding a sulfiding agent to the leachate to precipitate copper as copper sulfide; obtaining the leachate from which copper has been removed as a copper-removed solution; adding an oxidizing agent to the copper-removed solution to precipitate iron as iron hydroxide; obtaining the copper-removed solution from which iron has been removed as a valuable element solution containing the valuable element; and introducing the valuable element solution, a complexing agent, and an alkaline aqueous solution into a reaction bath liquid to obtain a precipitate containing the valuable element. [2] The method for producing the precursor according to [1] above, wherein the oxide is obtained from waste lithium ion batteries. [3] The method for producing a precursor according to [1] or [2] above, wherein the amount of the reducing agent added is 2.0 equivalents or less. [4] The method for producing a precursor according to any one of the above [1] to [3], wherein the mixed oxide is heated to a temperature of 1450° C. or higher. [5] The 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] The method for producing a precursor according to any one of the above [1] to [5], wherein the reducing agent is a carbon-containing substance. [7] The method for producing a precursor according to any one of the above [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. [8] The method for producing a precursor according to [7] above, wherein the iron oxide is ferrous oxide. [9] The method for producing a precursor according to [7] above, wherein the iron-containing material is at least one selected from the group consisting of dust, scale, sludge, and scrap.

[10] The method for producing a precursor according to any one of the above [1] to [9], 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] The method for producing a precursor according to any one of [1] to

[12] above, wherein the amount of the sulfurizing agent added is 1.0 equivalent or more relative to the copper contained in the leachate, and the pH of the leachate to which the sulfurizing agent has been added is adjusted to 3.0 or less when precipitating the copper sulfide.

[14] The 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 the oxidizing agent A added is 0.1 vvm or more relative to the copper-removal solution, the amount of the oxidizing agent B added is 0.005 vol% or more relative to the copper-removal solution, and the pH of the copper-removal solution to which the oxidizing agent has been added is adjusted to 3.0 or more and 7.0 or less when precipitating the iron hydroxide.

[15] The method for producing a precursor according to

[14] above, 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] above, wherein the alkaline aqueous solution is a sodium hydroxide aqueous 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 cathode 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] above 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] above, wherein the lithium-containing compound is at least one selected from the group consisting of lithium hydroxide and lithium carbonate. [Effects of the Invention]

[0010] According to the present invention, a novel method for producing a precursor and a cathode material can be provided. [Brief explanation of the drawings]

[0011] [Figure 1A] 1 is a flowchart showing an example of a method for recovering valuable elements. [Figure 1B] 1 is a flowchart showing an example of a flow for producing a precursor and a cathode material. [Figure 2]This is an Ellingham diagram (standard free energy change-temperature diagram). [Figure 3] Potential-pH diagrams of Cu and Ni (S-H2O system). [Figure 4] Potential-pH diagrams of Fe and Ni (O2-H2O system). DETAILED DESCRIPTION OF THE INVENTION

[0012] [Precursor and cathode material manufacturing method] Hereinafter, a method for producing a cathode material used in a lithium ion battery (a method for producing a cathode material) will be described. The following description also includes a method for producing a precursor used in the cathode material (a method for producing a precursor). First, as shown in Figure 1A, valuable elements are recovered from oxides such as cathode materials of used lithium-ion batteries. That is, the oxides are subjected to dry and wet treatments to obtain a valuable element solution containing the valuable elements. Thereafter, as shown in FIG. 1B, the obtained valuable element solution is used to produce a precursor, and then a cathode material is produced.

[0013] FIG. 1A is a flowchart showing an example of a method for recovering valuable elements. A method for recovering valuable elements will be explained briefly with reference to FIG. 1A. In the dry treatment, first, a reducing agent and a slag former, which will be described later, are added to oxides (Ni, Co, Mn, Cu, Fe, Li) to obtain mixed oxides. The resulting mixed oxide is then heated to reduce the oxides, yielding metals (Ni, Co, Mn, Cu, Fe) and slag (Li), which are then separated as appropriate. Before the wet treatment, the metals obtained are preferably powdered to obtain metal powders (Ni, Co, Mn, Cu, Fe). In hydroprocessing, metals (metal powders) are first brought into contact with an acid solution to obtain a leachate (Ni, Co, Mn, Cu, Fe) and a leach residue, which are then separated as appropriate. Next, a sulfiding agent is added to the resulting leachate to precipitate copper sulfide (Cu), yielding a copper-removed solution (Ni, Co, Mn, Fe). If necessary, the two are separated. An oxidizing agent is then added to the copper removal solution to precipitate iron hydroxide (Fe) and obtain a solution of valuable elements (Ni, Co, Mn).

[0014] In this way, valuable elements (Ni, Co, Mn) can be recovered from the oxide while removing impurity elements (Cu, Fe), and further, semi-valuable elements (Li) can also be recovered. Valuable elements can be easily recovered from the cathode material (oxide) of used lithium-ion batteries with a purity high enough to be reused as raw materials for lithium-ion batteries.

[0015] Next, the method for recovering valuable elements will be described in more detail. The following description also includes a description of a method for manufacturing metal.

[0016] <Reducing target (oxide)> The object to be reduced 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 quasi-valuable element such as lithium (Li), and specifically, for example, a cathode material for a waste lithium-ion battery. The valuable element may be at least one element selected from the group consisting of Ni and Co. The cathode material (oxide) is obtained from waste lithium-ion batteries by undergoing pre-processing such as removing the electrolyte, crushing, pulverizing, 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 an 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] <<Findings Obtained by the Inventors>> The positive electrode material of lithium-ion batteries is generally made of oxides (composite oxides) such as LiNiO2, LiCoO2, and LiMnO2. Considering the dry processing thermodynamically, for example, LiNiO2 and LiCoO2 decompose at high temperatures to produce NiO and CoO, respectively, as follows: 2LiNiO2 → Li2O + 2NiO + 1 / 2O2 2LiCoO2 → Li2O + 2CoO + 1 / 2O2

[0019] The standard free energy change (ΔG 0 ) are shown below, respectively. NiO→Ni+1 / 2O2:ΔG 0 =234900-84.68T[J] CoO → Co + 1 / 2O2: ΔG 0 =235480-71.55T[J] Any substance having a free energy change value lower than these standard free energy change values ​​at any elevated temperature can be used as a reducing agent.

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

[0021] Reducing Agent Examples of reducing agents include aluminum-containing substances such as metallic aluminum (Al); silicon-containing substances such as metallic silicon (Si) and FeSi; carbon-containing substances containing carbon; and iron-containing substances.

[0022] If the reducing power of the reducing agent is insufficient and reduction is insufficient, a portion of the cathode material may be separated as slag in the form of oxide, and the content of valuable elements in the metal obtained by reducing the cathode material may decrease (the reduction rate of valuable elements may decrease). For example, aluminum-containing substances, silicon-containing substances, and carbon-containing substances have strong reducing power, so it is expected that such poor reduction can be avoided and a high reduction rate can be obtained.

[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 (for example, propane gas). When a carbon-containing substance is used as a reducing agent, gases such as CO, CO2, and H2O are generated after reduction, which is preferable in that the amount of slag generated does not increase.

[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 will be described below.

[0025] Figure 2 is an Ellingham diagram (standard free energy change-temperature diagram). Referring to the standard free energy change and the Ellingham diagram (FIG. 2) described above, the Fe / FeO equilibrium is less noble than the Ni / NiO equilibrium and the Co / CoO equilibrium, suggesting the possibility of reduction by Fe. Also, the FeO / Fe3O4 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 contained in the produced metal and Co will be contained in the produced 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 addition, the higher up on the Ellingham diagram (Figure 2), the easier it is to metallize. When Si or Al is used as a reducing agent, Mn also easily becomes metallized. Therefore, by using Fe (or FeO) as a reducing agent, it is expected that only Ni and Co will be metallized without metallizing Mn.

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

[0028] Iron oxide is generally classified into three types: ferrous oxide (FeO), also known as wustite; iron tetraoxide (Fe3O4), also known as magnetite; and ferric oxide (Fe2O3), also known as hematite. Of these, magnetite and hematite have a higher standard free energy change than that of wustite at the same temperature, and may be less likely to cause a reduction reaction. For this reason, ferrous oxide (wustite) is preferred as the iron oxide because it is more likely to cause a reduction reaction. The iron oxide may be at least one of dust, scale, and sludge (hereinafter referred to as "dusts" for convenience) that are generated secondarily in the iron-making process. The use of dusts as iron oxide is preferable from the viewpoint of effectively utilizing by-products of the iron manufacturing process and from the viewpoint of 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, when the reducing agent is metallic iron (Fe), ferrous oxide (FeO), metallic aluminum (Al), metallic silicon (Si), coke (C), and propane (C3H8), reduction with 1.0 equivalent of reducing agent is shown as follows, respectively: Fe+(NiO,CoO,MnO)→(Ni,Co,Mn)+FeO 3FeO+(NiO,CoO,MnO)→(Ni,Co,Mn)+Fe3O4 2Al+3(NiO,CoO,MnO)→3(Ni,Co,Mn)+Al2O3 Si+2(NiO,CoO,MnO)→2(Ni,Co,Mn)+SiO2 C+2(NiO,CoO,MnO)→2(Ni,Co,Mn)+CO2 C3H8+10(NiO,CoO,MnO)→10(Ni,Co,Mn)+3CO2+4H2O

[0030] When determining the amount of reducing agent to be added, first, the contents of NiO, CoO, and MnO in the oxide to be reduced are determined. Specifically, the contents of Ni, Co, and Mn in the object to be reduced (oxide) are measured and regarded as the contents of NiO, CoO, and MnO in the object to be reduced (oxide), respectively. The contents of Ni, Co and Mn are measured using an energy dispersive X-ray analyzer (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 this makes it easier to prevent insufficient reduction.

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

[0033] Therefore, there is no particular upper limit to the amount of reducing agent added. However, from the viewpoint of facilitating the removal of Fe by wet treatment, it is preferable that the amount of Fe contained in the produced metal is small. Also, even if an excessive amount of reducing agent is added, the effect becomes saturated. For this reason, the amount of the reducing agent added is preferably 2.0 equivalents or less, more preferably 1.8 equivalents or less.

[0034] Slag former A slag former containing calcium oxide (CaO) and silicon dioxide (SiO2) is used. The content (total content) of CaO and SiO2 in the slag former is preferably 90 mass % or more, more preferably 95 mass % or more, even more preferably 98 mass % or more, and particularly preferably 100 mass %.

[0035] (Mass ratio (CaO / SiO2)) A low-basicity slag former with a low mass ratio of CaO to SiO2 (CaO / SiO2) is used. The mass ratio (CaO / SiO2) of the slag former 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 (CaO / SiO2) of the slag former is, for example, 0.15 or more, preferably 0.20 or more, more preferably 0.25 or more, and even more preferably 0.30 or more.

[0036] (Amount of slag former added) The amount of the slag former to be added is not particularly limited, but the mass ratio of the slag former to the oxide to be reduced (slag former / 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 of mixed oxides (obtaining metal and slag) Next, the mixed oxide (a mixture of oxide, reducing agent, and slag-forming agent) is heated, whereby the oxide is reduced. The equipment used for heating the mixed oxide is not particularly limited, and examples thereof include conventionally known equipment such as an arc furnace, a submerged arc furnace, a resistance furnace, a high-frequency melting furnace, a low-frequency melting furnace, a rotary kiln, a shaft furnace, and a steelmaking furnace.

[0038] 《Heating temperature》 The temperature (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 this makes it easier to prevent insufficient reduction. Although there is no particular upper limit, the heating temperature is preferably 1800°C or lower, and more preferably 1700°C or lower.

[0039] <Heated atmosphere> Suitable atmospheres for heating the mixed oxide (heating atmosphere) include, for example, inert atmospheres such as nitrogen gas (N2) atmosphere and argon gas (Ar) atmosphere; reducing atmospheres such as carbon monoxide gas (CO) atmosphere; and the like.

[0040] <Heating time> The time for heating 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 this makes it easier to prevent insufficient reduction. Although there is no particular upper limit, the heating time is preferably 6 hours or less, and more preferably 5 hours or less.

[0041] 《Product (metal)》 Metals are produced by reducing the oxides (cathode materials) to be reduced. In other words, valuable elements (Ni, Co, Mn) contained in the oxides are recovered as metals.

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

[0043] Product (Slag) By reducing the oxide (cathode material), slag is produced in addition to the metal. The resulting slag (also referred to as "resulting slag") contains a large amount of Li, as described above. When an iron-containing substance is used as a reducing agent, the resulting slag contains FeO and the like. In addition, the produced slag may also contain oxides of valuable elements (for example, MnO) that were not contained in the produced metal. When reducing oxides containing Mn, by using an iron-containing substance as a reducing agent, the Mn / MnO equilibrium is less noble than the Fe / FeO equilibrium and the FeO / Fe3O4 equilibrium, and therefore it is possible to prevent Mn from being mixed into the metal produced by reduction. However, separating Mn from the produced metal by hydrometallurgy is a heavy burden, so it is beneficial to prevent Mn from being mixed into the produced metal and keep Mn in the produced slag, as this burden can be reduced.

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

[0045] <Metal powderization (obtaining metal powder)> Next, it is preferable to powder the resulting metal to obtain a metal powder. In the wet treatment, as described below, the produced metal is first subjected to leaching using an acid solution. At this time, if the produced metal remains in the state obtained by oxide reduction, the leaching efficiency may be insufficient. Therefore, it is preferable to powder the produced metal before performing leaching using 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, it may become difficult to handle or may increase the risk of an explosive reaction. Therefore, taking these points into consideration, the particle size of the metal powder is 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 volume-based median diameter (particle size at 50% cumulative value) in the particle size distribution determined by a laser diffraction / scattering method (the same applies hereinafter).

[0047] The method for powdering the produced metal is not particularly limited as long as it can keep the particle size of the resulting metal powder within an appropriate range, and examples include methods using a grinding device such as a jaw crusher or a vibration mill; atomization; and the like.

[0048] <Contact of metal with acid solution (obtaining leachate)> Next, the metal (metal powder) is brought into contact with an acid solution to leach out the valuable elements (Ni, Co, Mn) and impurity elements (Cu, Fe), thereby obtaining a leachate containing the valuable elements and impurity elements. The metal from which the valuable elements and impurity elements have been leached becomes a residue (leaching residue). The method for bringing the metal into contact with the acid solution is not particularly limited, but examples thereof include a method of immersing the metal in the acid solution, and a method of spraying the acid solution onto the metal.

[0049] 《Solid-liquid ratio (metal / acid liquid)》 If the amount of acid solution brought into contact with the metal is too small (the amount of metal is too large compared to the amount of acid solution), some of the metal elements, such as valuable elements, that have dissolved in the acid solution may reach saturated solubility and precipitate, resulting in an insufficient leaching rate. Therefore, the ratio of the mass (unit: g) of the solid metal to the volume (unit: mL) of the liquid acid solution (also referred to as 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 liquid) is preferably 1 / 50 or more, more preferably 1 / 35 or more, and even more preferably 1 / 20 or more.

[0050] 《Acid liquid》 The acid solution to be brought into contact with the metal contains at least an acid.

[0051] (acid) The acid used in the acid solution includes inorganic acids such as sulfuric acid, hydrochloric acid, and nitric acid, and these may be used alone or in combination of two or more kinds. From the perspective of realizing the "battery-to-battery" concept of recycling used lithium-ion batteries and using them again as raw materials for lithium-ion batteries, it is preferable to use sulfuric acid as the acid because it allows valuable elements to be obtained in the form of sulfates, which are easily usable as cathode materials for lithium-ion batteries. Sulfuric acid containing chloride may also be used as the acid.

[0052] ((acid concentration)) The concentration (acid concentration) of the acid (e.g., sulfuric acid) used in the acid solution is preferably 0.1 mol / L or more, more preferably 0.5 mol / L or more, and even more preferably 1.0 mol / L or more, because this can increase the leaching rate. Although there is no particular upper limit, the acid concentration is preferably 8.0 mol / L or less, more preferably 6.0 mol / L or less, even more preferably 4.0 mol / L or less, and particularly preferably 3.0 mol / L or less.

[0053] (Oxidizing agent for acid solutions) The inventors have found through their investigations that even when the solid-liquid ratio (metal / acid solution) and acid concentration are within the above-mentioned ranges, 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. Among these, hydrogen peroxide and ozone are preferred because the use of hypochlorous acid and potassium permanganate may require complicated post-treatment for chlorine, potassium, manganese, etc.

[0054] ((Oxidizing agent content for acid solution)) From the viewpoint of carrying out sufficient leaching, the content of the oxidizing agent for the acid solution (e.g., hydrogen peroxide) in the acid solution is preferably 0.5% by volume or more, more preferably 1.0% by volume or more, even more preferably 3.0% by volume or more, even more preferably 5.0% by volume or more, particularly preferably 6.0% by volume or more, and most preferably 6.9% by 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 (for example, hydrogen peroxide) in the acid solution is preferably 15.0% by volume or less, more preferably 13.0% by volume or less, and even more preferably 10.0% by volume or less, relative to the acid (for example, sulfuric acid).

[0055] Contact Time The time for which the metal is brought into contact with the acid solution (contact time) 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, more preferably 1.5 hours or less.

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

[0057] Adding sulfurizing 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, the leachate from which copper (Cu) has been selectively removed is obtained as a copper-removed solution.

[0058] Figure 3 shows the potential-pH diagram for copper (Cu) and nickel (Ni) (S-H2O system). Figure 3 shows the region on the potential-pH diagram for the copper (Cu)-sulfur (S)-water (H2O) system where precipitation of copper (Cu) and nickel (Ni) oxides (hydroxides) or sulfides occurs, taking into account their solubility. Cobalt precipitates in the same manner as nickel, so cobalt is not shown in Figure 3. As shown in Figure 3, copper (Cu) is selectively precipitated in the region where the pH is 3.0 or less and the oxidation-reduction potential is low. Although not shown in Figure 3, in this region, copper precipitates as copper (II) sulfide (CuS). Utilizing this, by making the leachate low pH and reducing, the copper (Cu) contained in the leachate is precipitated as copper (II) sulfide and selectively removed. In other words, a copper-removed solution is obtained, which is the leachate from which copper (Cu) has been removed.

[0059] Sulfurizing agents Examples of sulfurizing agents to be added to the leaching solution include sulfur (S), hydrogen sulfide (HS), sodium hydrogen sulfide (NaSH), and sodium sulfide (NaS). These may be used alone or in combination of two or more. Of these, sulfur, sodium hydrogen sulfide, and sodium sulfide, which can be handled as solids or solutions, are preferable to hydrogen sulfide, which is a toxic gas, from the viewpoint of ease of handling. However, in all cases, hydrogen sulfide gas may be generated by the sulfurization reaction, so care must be taken when carrying out the process.

[0060] The temperature of the leachate to which the sulfurizing agent has been added (sulfurization temperature) is not particularly limited, and may be, for example, room temperature.

[0061] (Amount of sulfurizing 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 copper (Cu) contained in the leachate. On the other hand, adding an excessive amount of sulfiding agent may result in an increase in the amount of sulfides (precipitates) of valuable elements (Ni, Co, etc.), potentially reducing the amount of the valuable elements desired to remain in the resulting copper removal solution. From this perspective, 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 copper (Cu) contained in the leachate.

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

[0063] 《Sulfide pH》 When a sulfiding agent is added to the leachate to precipitate copper sulfides, if the pH of the leachate to which the sulfiding agent has been added (sulfiding pH) is high, the amount of sulfides (precipitates) of valuable elements that are to be left in the copper removal solution may increase. Therefore, the sulfiding 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 sulfurization pH is adjusted, for example, by adding a pH adjuster to the leaching solution. The pH adjuster is not particularly limited, and examples thereof include sulfuric acid and sodium hydroxide.

[0064] Sulfurization Time The time (sulfurization time) for sulfurizing the copper contained in the leachate by reacting it with the sulfurizing agent 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 sulfurization 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 from 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 valuable element solution)> 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, the copper-removal solution from which iron (Fe) has been selectively removed is obtained as a valuable element solution containing valuable elements (Ni, Co, Mn).

[0067] Figure 4 shows the potential-pH diagram for iron (Fe) and nickel (Ni) (O2-H2O system). Figure 4 shows the region where precipitates of oxides (hydroxides) of iron (Fe) and nickel (Ni) form on the potential-pH diagram for the iron (Fe)-oxygen (O2)-water (H2O) system, taking into account their solubility. Cobalt precipitates in the same manner as nickel, so cobalt is not shown in Figure 4. As shown in Figure 4, iron (Fe) is selectively precipitated in the region where the pH is between 3.0 and 7.0 and where the redox potential is high. Although not shown in Figure 4, in this region, iron precipitates as iron (III) oxide hydroxide (FeO(OH)). By utilizing this, the copper-removal solution is made acidic to neutral and oxidizing, so that the iron (Fe) contained in the copper-removal solution is precipitated as iron (III) oxide hydroxide and selectively removed. In other words, a valuable element solution is obtained, which is a copper-removal solution from which iron (Fe) has been removed.

[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 the use of 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) added as a gas 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" is a unit that expresses the volume ratio of gas blown into the liquid per minute. For example, if the amount of oxidant A added is 2 vvm, then 2 L of oxidant A is blown into 1 L of copper removal solution per minute.

[0071] For the same reason, 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 conducted research and found that the use of the above-mentioned oxidizing agents alone may not be sufficient to oxidize the iron contained in the copper removal solution. Therefore, from the viewpoint of promoting oxidation, it is preferable to increase 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, more preferably 80°C or lower.

[0073] Oxidative pH When an oxidizing agent is added to a copper removal solution to precipitate iron hydroxide, if the pH of the copper removal solution to which the oxidizing agent has been added (oxidizing pH) is too low, precipitation may be difficult to form. Therefore, the oxidizing 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, there is a concern that the coprecipitation of valuable elements (Ni, Co, etc.) will increase, and the amount of the valuable elements desired to remain in the obtained valuable element solution will decrease. Therefore, the oxidation pH is preferably 7.0 or less, more preferably 6.0 or less, and even more preferably 5.0 or less. The oxidation pH is adjusted, for example, by adding a pH adjuster to the copper removal solution, which is not particularly limited and includes sulfuric acid, sodium hydroxide, etc.

[0074] Oxidation Time The time (oxidation time) for reacting the iron contained in the copper removal solution with the oxidizing agent 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] <<oxidation promoter>> In order to increase the reaction rate for producing the precipitate of iron hydroxide, an oxidizing agent may be used in combination with the above-mentioned oxidizing agent. The oxidation aid may be, for example, at least one selected from the group consisting of ferric oxide (Fe2O3) and iron (III) oxide hydroxide (FeO(OH)), and the oxidation aid is preferably in the form of powder. The principle behind the increase in reaction rate due to the oxidation aid is catalytic action. Specifically, the oxidation aid tends to be negatively charged in the aqueous solution (copper removal solution), so Fe 2+ Adsorbs ions and Fe 2+ Internal e -This weakens the bond with Fe 2+ →Fe 3+ +e - It is thought that the activation energy of this reaction (Fe oxidation reaction) decreases, accelerating the reaction.

[0076] (Amount of auxiliary oxidation agent added) It is believed that the larger the amount of oxidation aid added, the larger the reaction surface area and the faster the rate of the Fe oxidation reaction. Therefore, the amount of oxidation aid 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 the amount of the oxidation aid is too large, there is a concern that the coprecipitation of valuable elements (such as Ni and Co) will increase. Therefore, the amount of the oxidation aid 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 oxidation aid) If the particle size of the oxidation promoter is too small, the reaction surface area becomes too large, which may increase the coprecipitation of valuable elements (Ni, Co, etc.). Therefore, the particle size of the oxidation promoter is preferably 0.1 μm or more, more preferably 0.3 μm or more, and even more preferably 0.5 μm or more. On the other hand, if the particle size of the oxidation promoter is too large, the reaction surface area becomes too small and the desired effect may not be obtained. Therefore, the particle size of the oxidation promoter is preferably 3.0 μm or less, more preferably 2.0 μm or less, and even more preferably 1.0 μm or less.

[0078] <Separation of iron hydroxide and valuable element solution> It is preferable to separate the iron hydroxide and the valuable element solution from each other. The separation method is not particularly limited, and known solid-liquid separation methods can be used. The valuable elements in the valuable element solution thus obtained can be used, for example, as a positive electrode material for lithium ion batteries.

[0079] Obtaining Valuable Element Precipitates FIG. 1B is a flow chart showing an example of a process for producing a precursor. Next, the valuable element solution obtained by the above-mentioned valuable element recovery method, a complexing agent, and an alkaline aqueous solution are introduced (dropped) into the reaction vessel liquid, and a precipitate containing the valuable element (valuable element precipitate) is obtained by the so-called coprecipitation method, as shown in FIG. 1B. Specifically, the valuable element precipitate is at least one selected from the group consisting of a composite hydroxide containing a valuable element and a composite oxide containing a valuable element. By using the coprecipitation method, the valuable elements (Ni, Co, Mn) can be uniformly dispersed at the atomic level. The resulting precipitate of valuable elements is filtered out from the reaction vessel liquid, washed with water and dried as necessary, to obtain a precursor of the positive electrode material.

[0080] When obtaining a valuable element precipitate, a raw material aqueous solution may be used instead of the valuable element solution. The raw material aqueous solution is prepared by adding at least one selected from the group consisting of a nickel source, a cobalt source, and a manganese source to the valuable element solution. In the raw material aqueous solution, the molar ratio of the nickel content to the cobalt content to the manganese content (Ni / Co / Mn) is preferably 1 / 1 / 1, 5 / 2 / 3, 6 / 2 / 2, or 8 / 1 / 1.

[0081] The nickel source is, for example, a nickel salt such as nickel sulfate, nickel carbonate, nickel nitrate, nickel acetate, or nickel chloride, with nickel sulfate (NiSO4) being preferred. The cobalt source is, for example, a cobalt salt such as cobalt sulfate, cobalt carbonate, cobalt nitrate, cobalt acetate, cobalt chloride, etc., with cobalt sulfate (CoSO4) being preferred. The manganese source is, for example, a manganese salt such as manganese sulfate, manganese carbonate, manganese nitrate, manganese acetate, manganese chloride, etc., with manganese sulfate (MnSO4) being preferred. The nickel source, cobalt source, and manganese source are preferably used in the form of an aqueous solution, and the concentrations (contents) of the nickel source, cobalt source, and manganese source in each aqueous solution are 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. 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 aqueous solution) is preferably 1.0 mL / min or more, more preferably 2.5 mL / min or more. The dropping rate of the valuable element solution (raw aqueous solution) is preferably 7.0 mL / min or less, and more preferably 5.5 mL / min or less.

[0084] The complexing agent may be, for example, at least one ammonium source selected from the group consisting of ammonia (NH3) and ammonium salts. Examples of ammonium salts include ammonium sulfate, ammonium chloride, ammonium nitrate, and ammonium carbonate. The complexing agent that is the ammonium source is preferably ammonia. 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 below.

[0085] The dropwise addition rate of the complexing agent is preferably 0.1 mL / min or more, more preferably 0.3 mL / min or more, and is preferably 1.0 mL / min or less, more preferably 0.8 mL / min or less.

[0086] The molar ratio (NH4 / (Ni+Co+Mn)) of the content of the ammonium source (complexing agent) in terms of ammonium to the total content of the valuable elements (Ni, Co, Mn) in the raw material aqueous solution is preferably greater than 0, more preferably equal to or greater than 2, and even more preferably equal to or greater than 4. Furthermore, this molar ratio (NH4 / (Ni+Co+Mn)) is preferably 20 or less, more preferably 15 or less, and even more preferably 12 or less.

[0087] The alkaline aqueous solution is preferably an aqueous sodium hydroxide (NaOH) solution.

[0088] The reaction tank liquid is the liquid contained in the reaction tank, and is, for example, an aqueous solution prepared by adding an alkaline aqueous solution such as an aqueous sodium hydroxide solution to pure water. The pH of the reaction bath liquid is preferably 9.0 or higher, more preferably 9.5 or higher, and is preferably 12.0 or lower, more preferably 11.5 or lower. The temperature of the reaction vessel liquid is preferably 30° C. or higher, more preferably 35° C. or higher. The temperature of the reaction vessel liquid is preferably 60° C. or lower, 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 raw material aqueous solution, and a homogeneous precipitate of valuable elements may not be obtained. In these cases, the precursor obtained by drying the precipitate of valuable elements may not be uniformly spherical, making it difficult to achieve 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. The stirring speed may be changed during the reaction. For example, the stirring speed may be started at a slow speed and then changed to a faster, more suitable stirring speed as the amount of liquid in the reaction vessel increases.

[0090] During the dropwise addition of the valuable element solution (raw material aqueous solution) and the complexing agent, it is preferable to control the pH of the reaction bath solution within the above range by dropping an alkaline aqueous solution into the reaction bath solution.

[0091] The obtained precipitate of valuable elements is preferably filtered out 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, and 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, and is preferably 15 hours or less, more preferably 12 hours or less.

[0092] As described above, for example, the valuable element precipitate is washed with water and dried to obtain a precursor of the positive electrode material. In the precursor, the molar ratio of nickel content to the total content of valuable elements (Ni, Co, Mn) (Ni / (Ni+Co+Mn)) is preferably 0.3 or more, more preferably 0.4 or more, and 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 was 0.8 g / cm 3 More than 1.0 g / cm is preferable. 3 More preferably, 1.2 g / cm 3 The tap density of the precursor is more preferably 1.8 g / cm. 3 It may be less than 1.5 g / cm 3 It may be the following: The tap density is the volume of the solution left standing at 100 cm 3 The sample is placed in a container and tapped with a tapping device until the volume of the sample no longer decreases. Then, the mass of the sample (unit: g) is calculated based on the volume of the sample (unit: cm 3 ) to calculate the value (same below).

[0094] Precursor particle size D 10 is preferably 3.0 μm or more, more preferably 4.0 μm or more. 10 is preferably 10.0 μm or less, more preferably 8.0 μm or less. Precursor particle size D 50 is preferably 8.0 μm or more, more preferably 9.0 μm or more. 50is preferably 16.0 μm or less, more preferably 14.0 μm or less. Precursor particle size D 90 is preferably 12.0 μm or more, more preferably 14.0 μm or more. 90 is preferably 24.0 μm or less, more preferably 22.0 μm or less. Particle size D 10 , particle size D 50 and particle size D 90 are 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% by volume (the same applies below).

[0095] <Acquisition of fired products> FIG. 1B is a flowchart showing an example of a flow for producing a cathode material. Next, the obtained precursor is mixed with a lithium-containing compound, and the resulting mixture is fired to obtain a fired product containing a valuable element and lithium (a composite oxide containing a valuable element and lithium), as shown in FIG. The resulting fired product is then crushed appropriately to obtain a cathode material for use in a lithium ion battery. The cathode material is also called a cathode active material.

[0096] The resulting positive electrode 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] Each step in the method for producing the cathode material will be described in more detail below.

[0098] First, the precursor and the lithium-containing compound are mixed to obtain a mixture. In this case, the molar ratio (Li / (Ni+Co+Mn)) of the lithium-equivalent content of the lithium-containing compound to the sum of the nickel-equivalent content of the precursor, the cobalt-equivalent content of the precursor, and the manganese-equivalent content of the precursor is preferably greater than 1.03, more preferably 1.04 or greater, and is preferably less than 1.10, more preferably 1.08 or less.

[0099] The lithium-containing compound is preferably at least one selected from the group consisting of lithium hydroxide and lithium carbonate.

[0100] When the resulting positive electrode material contains the above-mentioned element A, the mixture may further contain a compound containing element A (hereinafter also referred to as an "A-containing compound"). Examples of A-containing compounds include, but are not limited to, hydroxides, oxides, chlorides, and salts (eg, sulfates, carbonates, nitrates, etc.) of element A. The amount of the A-containing compound to be mixed is adjusted appropriately depending on the desired composition.

[0101] Next, the mixture obtained by mixing is fired to obtain a fired product. In this case, it is preferable to calcinate the mixture and then calcinate it. The calcination temperature for the calcination is preferably 400° C. or higher, more preferably 500° C. or higher, and is preferably 700° C. or lower, more preferably 680° C. or lower. The firing temperature in the main firing is preferably 800° C. or higher, more preferably 900° C. or higher, and is preferably 1000° C. or lower, more preferably 980° C. or lower.

[0102] The atmosphere for the calcination may be an oxidizing atmosphere (for example, air) or a non-oxidizing atmosphere. The non-oxidizing atmosphere may be, for example, an atmosphere with an oxygen concentration of 10% by volume or less, and a specific example of such an atmosphere is a nitrogen atmosphere. The atmosphere for the main firing may be an oxidizing atmosphere (for example, air) or a non-oxidizing atmosphere.

[0103] The calcination time is preferably 2 hours or more, more preferably 3 hours or more, and is preferably 48 hours or less, more preferably 12 hours or less. The firing time of the main firing is preferably 1 hour or more, more preferably 2 hours or more, and even more preferably 3 hours or more, and is preferably 30 hours or less, more preferably 15 hours or less, and even more preferably 8 hours or less.

[0104] The fired product may be washed with water. By washing with water, excess lithium that has not penetrated into the interior is washed away. After washing with water, the fired product is dried as appropriate. The fired product may be further fired at 200° C. or higher and 800° C. or lower, or crushed. In this way, a positive electrode material for use in a lithium ion battery is obtained.

[0105] The tap density of the cathode material is 1.0 g / cm 3 More than 1.5g / cm is preferable. 3 More preferably, the tap density of the positive electrode material is 3.5 g / cm. 3 It may be 3.0 g / cm or less. 3 It may be the following:

[0106] Positive electrode material particle size D 10 is preferably 3.0 μm or more, more preferably 4.0 μm or more. 10 is preferably 10.0 μm or less, more preferably 8.0 μm or less. Positive electrode particle size D 50 is preferably 8.0 μm or more, more preferably 9.0 μm or more. 50 is preferably 16.0 μm or less, more preferably 14.0 μm or less. Positive electrode material particle size D 90 is preferably 12.0 μm or more, more preferably 14.0 μm or more. 90is preferably 24.0 μm or less, more preferably 22.0 μm or less.

[0107] A lithium ion battery generally comprises a positive electrode, a negative electrode, and an ion-conductive medium (e.g., an electrolyte such as a non-aqueous electrolyte) interposed between the positive electrode and the negative electrode to conduct lithium ions, and may further comprise a separator. The cathode material thus obtained is used to produce a cathode by a known method, and the produced cathode is used to produce a lithium ion battery, which has excellent discharge capacity and cycle characteristics. [Example]

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

[0109] [Test A] <Preparing the cathode material> A cathode material for a used lithium-ion battery was prepared. Specifically, the used lithium-ion batteries were subjected to pre-treatments such as disassembly, discharge, and removal of the electrolyte, and the positive electrode material was separated. The composition ratio (molar ratio) of nickel (Ni), cobalt (Co), and manganese (Mn) in the positive electrode material is shown in Table 1 below. The positive electrode material also contained copper (Cu) and iron (Fe) as impurity elements. The positive electrode material also contained lithium (Li) as a quasi-valent element.

[0110] [Table 1]

[0111] <Preparing the reducing agent> As reducing agents, powder of metallic aluminum (Al), powder of metallic silicon (Si), powder of coke (C), and propane gas (C3H8) were prepared. Furthermore, as reducing agents, powder of metallic iron (Fe) obtained by atomization treatment and powder of ferrous oxide (FeO) were prepared.

[0112] <Preparing the slag former> A slag former consisting of CaO and SiO2 was prepared. Several types of slag formers with different mass ratios of CaO to SiO2 (CaO / SiO2) were prepared.

[0113] Addition of reducing agents and slag-forming agents The prepared cathode material was placed in a submerged arc furnace with a heat size of 150 kg, and a reducing agent and a slag-forming agent were added to obtain a mixed oxide. More specifically, 30 kg of the slag former was added to 45 kg of the positive electrode material, that is, the mass ratio of the slag former to the positive electrode material (slag former / positive electrode material) was about 0.67. The type and amount (unit: equivalent) of the reducing agent used, as well as the mass ratio (CaO / SiO2) of the slag forming agent used, are shown in Table 2 below.

[0114] <Heating mixed oxides> The resulting mixed oxide was heated for 3 hours in an Ar atmosphere. The heating temperature (unit: °C) is shown in Table 2 below. In this way, the cathode material was reduced to obtain a product metal and a product slag, which were then separated.

[0115] The reduction rate (unit: mass %) for each of the metal elements Ni, Co, and Mn 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 produced metal [kg]) / (amount of metal element contained in the oxide to be reduced [kg])

[0116] Furthermore, the residual rate (unit: mass %) of Li in the produced slag 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 the produced slag = 100 × (amount of Li contained in the produced slag [kg]) / (amount of Li contained in the oxide to be reduced [kg])

[0117] As shown in Table 2 below, Test Examples 2-1 to 2-8, which used a slag former with a mass ratio (CaO / SiO2) of 0.50 or less, had a higher residual Li rate in the produced slag than Comparative Test Examples 2-1 to 2-2, which used a slag former with a mass ratio (CaO / SiO2) of more than 0.50.

[0118] In all of Test Examples 2-1 to 2-8, the reduction rates of Ni and Co were high, while the reduction rate of Mn was able to be suppressed to 20% or less. In particular, in Test Examples 2-5 to 2-6, in which Fe or FeO was used as the reducing agent, the reduction of Mn could be further suppressed. In addition, the reduction rate of Co in Test Example 2-6 is lower than that in Test Example 2-5. This is thought to be because the free energy when FeO is produced is slightly lower than the free energy when CoO is produced (see Figure 2). However, since FeO produces Fe by reduction, its reduction potential is higher than that of FeO. Overall, it has a certain degree of reduction potential, and this result is thought to be reflected in the results.

[0119] Comparing Test Examples 2-1 to 2-6 with Test Examples 2-7 to 2-8, Test Examples 2-7 to 2-8, which had a lower mass ratio of the slag former (CaO / SiO), were able to further suppress the reduction of Mn compared to Test Examples 2-1 to 2-6 without causing a significant decrease in the reduction rates of Ni and Co.

[0120] [Table 2]

[0121] <Metal powdering> The compositions of the metals and slag produced by the reduction of the cathode material were determined. Of the metals produced by reduction of the positive electrode material, those having the compositions shown in Table 3 below were pulverized using a vibration mill to obtain metal powders. The particle size of the resulting metal powder was 1100 μm.

[0122] <Contact between metal and acidic solution> An acid solution was prepared by adding 7.0% by volume of hydrogen peroxide as an oxidizing agent to sulfuric acid (concentration: 2.0 mol / L). The prepared acid solution was contacted with a metal (metal powder) having the composition shown in Table 3 below at a solid-liquid ratio (metal / acid solution) of 1 / 10 (contact time: 1.0 hour). Specifically, the metal powder was immersed in the acid solution. In this way, a leachate and a leach residue were obtained, and the two were 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 to 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 %, and all of the elements were leached from the metal into the leachate.

[0123] [Table 3]

[0124] <Addition of sulfurizing agent> The content (unit: g / L) of each element in the obtained leachate is shown in Table 4 below. Sodium hydrogen sulfide (NaSH) was added as a sulfiding agent to the obtained leachate and stirred at room temperature (25°C). The amount of sulfiding agent (sodium hydrogen sulfide) added was 2.0 equivalents relative to the Cu contained in the leachate. The pH of the leachate to which the sulfurizing agent was added (sulfurization 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 sulfided by reacting with the sulfiding agent (sulfiding time: 20 minutes) and precipitated as copper sulfide (copper (II) sulfide).The copper sulfide was then separated from the copper-removed leachate, which was the copper-removed solution. The content (unit: g / L) of each element in the copper removal solution was determined by ICP-AES (inductively coupled plasma atomic emission spectroscopy), and the results are shown in Table 4 below. Furthermore, for each element, the ratio of the content in the copper removal solution to the content in the leaching solution was calculated 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, which indicates that Cu was removed from the leachate with very high efficiency.

[0125] <Addition of oxidizing agent> Next, the copper removal solution was first diluted with water. The contents (units: g / L) of each element in the diluted copper removal solution are 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 composition similar to that of 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% by volume relative to the diluted copper removal solution. The pH of the copper removal solution to which the oxidizing agent had been added (oxidation pH) was adjusted to 4.5 using sulfuric acid and sodium hydroxide as pH adjusters, and the temperature of the copper removal solution to which the oxidizing agent had been added (oxidation temperature) was set to and maintained at 70°C. In this way, the iron (Fe) contained in the copper-removed solution was oxidized by reacting with the oxidizing agent (oxidation time: 1.0 hour) and precipitated as iron hydroxide (iron (III) oxide hydroxide).The iron hydroxide was then separated from the valuable element solution, which was the copper-removed solution from which iron had been removed. The content (unit: g / L) of each element in the valuable element solution 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 calculated as the residual rate b (unit: mass %). The results are shown in Table 4 below. As shown in Table 4 below, the content of Fe in the valuable element solution is very low, which indicates 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 calculated as the total residual rate (unit: mass%) from residual rate a and residual rate b. The results are shown in Table 4 below. The results shown in Table 4 below show that by performing dry processing, then powdering the metals obtained by the dry processing, and then performing wet processing, valuable elements (Ni, Co, Mn) could be recovered with very high purity.

[0127] [Table 4]

[0128] [Test B] <Preparing the cathode material ~ Powdering the metal> The steps from preparation of the positive electrode material to powdering of the metal were the same as those in Test A, and therefore the explanation will be omitted.

[0129] <Contact between metal and acidic solution> A plurality of acid solutions were prepared by adding an oxidizing agent for the acid solution (hydrogen peroxide) to sulfuric acid (concentration: 2.0 mol / L) in the amount (unit: volume %) shown in Table 5 below. A leachate was obtained by contacting a metal (metal powder) with an acid solution in the same manner as in Test A described above, except that the amount of oxidizing agent for the acid solution was changed. Furthermore, the leaching rate (unit: mass %) of each element from the metal into the leachate was calculated in the same manner as in the above-mentioned Test A. The results are shown in Table 5 below. As shown in Table 5 below, the leaching rate increased as the amount of oxidizing agent for the acid solution increased. It was found that the amount of oxidizing agent for the acid solution (hydrogen peroxide) added should be 6.9% by volume or more in order to sufficiently leach valuable metals (Ni, Co, Mn). However, when the amount exceeds 6.9% by volume, the leaching rate reaches a plateau. Therefore, from the viewpoint of cost, it was found that the amount of oxidizing agent (hydrogen peroxide) for the acid solution in this example is preferably 6.9% by volume.

[0130] [Table 5]

[0131] <Addition of sulfurizing agent> Sodium hydrogen sulfide (NaSH) was added as a sulfiding agent to the leachate shown in Table 4 in the amount (unit: equivalent) shown in Table 6 below, and the mixture was stirred. At this time, the sulfiding pH was adjusted to the value shown in Table 6 below. Copper contained in the leachate was precipitated as copper sulfide in the same manner as in Test A described above, except that the amount of sulfiding agent added and the sulfiding pH were changed, to obtain a copper-removed solution. Furthermore, the Cu content (unit: mg / L), Ni residual rate (unit: mass %), and Co residual rate (unit: mass %) in the obtained copper removal solution were determined in the same manner as in the above-mentioned Test A. The results are shown in Table 6 below. As shown in Table 6 below, in order to sufficiently remove copper, the amount of sulfurizing agent added is preferably 2.0 equivalents or more relative to copper, but it was found that the Ni retention rate and Co retention rate decrease as the amount of sulfurizing agent added increases. Furthermore, as shown in Table 6 below, as the sulfurization pH increased, there was a tendency for copper removal to become insufficient and for the Ni and Co residual rates to decrease. From the above, it was found that within the range of this example, the amount of sulfurizing agent added was preferably 2.0 equivalents and the sulfurization pH was preferably 0 (zero).

[0132] [Table 6]

[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 as an oxidizing agent to the diluted copper removal solution in the amount (volume %) shown in Table 7 below, followed by stirring. At this time, the oxidation pH and oxidation temperature (°C) were adjusted to the values ​​shown in Table 7 below. The iron contained in the copper removal solution was precipitated as iron hydroxide to obtain a valuable element solution in the same manner as in Test A described above, except that the amount of oxidizing agent added, the oxidation pH, and the oxidation temperature were changed. In both Test A and Test B, no auxiliary oxidation agent was used. Furthermore, the Fe content (unit: mg / L), Ni residual rate (unit: mass %) and Co residual rate (unit: mass %) in the obtained valuable element solution were determined in the same manner as in the above-mentioned Test A. 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 more efficiently removed, but the Ni and Co residual rates tended to decrease. From the above, it was found that within the scope of this example, the amount of oxidizing agent (hydrogen peroxide) added was preferably 0.030% by volume, and the oxidation pH was preferably in the range of 4.5 to 5.0.

[0134] [Table 7]

[0135] [Test C] A valuable element solution was obtained by carrying out dry and wet treatments (i.e., from preparation of the cathode material to addition of the oxidizing agent) in accordance with Test A. For convenience, the mass of the cathode material (oxide) used is set to 100 kg, as shown in Table 8 below. Test Examples 8-1 to 8-3 differ from each other in the type and amount of reducing agent added. In Comparative Test Example 8-1, only dry treatment was carried out. In Comparative Test Example 8-2, a slag former with a mass ratio (CaO / SiO2) of 1.50 was used.

[0136] The Ni+Co recovery rate (unit: mass%) was calculated from the Ni+Co reduction rate (unit: mass%) of the metal produced by the dry treatment and the Ni+Co residual rate (unit: mass%) of the valuable element solution obtained by the wet treatment. The results are shown in Table 8 below. When only dry treatment was carried out, the Ni+Co reduction rate of the produced metal was shown in Table 8 below as the Ni+Co recovery rate.

[0137] Furthermore, for the valuable element solution obtained by the wet treatment, the proportion of Ni+Co in the metal elements was calculated as the final Ni+Co purity (unit: mass%). The results are shown in Table 8 below. However, in Comparative Test Example 8-1 in which only dry treatment was carried out, the Ni+Co purity of the metal produced by dry treatment is shown in Table 8 below as the final Ni+Co purity.

[0138] Comparing Comparative Test Example 8-1 and Test Example 8-1, the dry treatment conditions were the same for both, but Test Example 8-1, in which wet treatment was performed, ultimately yielded valuable elements (Ni + Co) with a higher purity than Comparative Test Example 8-1, in which wet treatment was not performed.

[0139] Comparing Comparative Test Example 8-2, which used Al as a reducing agent, with Test Example 8-1, Test Example 8-1, in which the mass ratio of the slag-forming agent (CaO / SiO2) was 0.50, had a higher residual Li rate in the generated slag than Comparative Test Example 8-2, in which the mass ratio was 1.50, and more Li was recovered.

[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 Li residual rate in the generated slag than Test Example 8-1, which used Al as the reducing agent, and was able to recover more Li.

[0141] Comparing Test Example 8-2 and Test Example 8-3, which used Fe as the reducing agent, Test Example 8-3, in which the amount of reducing agent added was 1.6 equivalents, had a lower Ni+Co purity of the produced metal than Test Example 8-2, in which the amount of reducing agent added was 1.4 equivalents. That is, by increasing the amount of Fe added as a reducing agent in the dry treatment, the amount of Fe contained in the produced metal increased, and the Ni+Co purity of the produced metal decreased. However, ultimately, Test Example 8-3 had a higher Ni+Co recovery rate and an equally high Ni+Co purity compared to Test Example 8-2. In this way, by carrying out wet treatment after dry treatment, high purity of valuable elements can be obtained in the end, and the amount of Fe added in the dry treatment can be increased to sufficiently reduce the valuable elements (Ni + Co). As a result, the final Ni + Co recovery rate including wet treatment can also be high.

[0142] [Table 8]

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

[0144] Precursor 1 Nickel sulfate (NiSO4) as a nickel source, cobalt sulfate (CoSO4) as a cobalt source, and manganese sulfate (MnSO4) as a manganese source were added to the valuable element solution of Test Example 8-1 listed in Table 8. In this way, a raw material aqueous solution was prepared having a valuable element (Ni, Co, Mn) content of 1.25 mol / L and a Ni / Co / Mn (molar ratio) of 6 / 2 / 2. 0.35 L of pure water, an aqueous sodium hydroxide solution, and an aqueous ammonia solution were added to the reaction vessel to prepare a reaction vessel liquid with a pH of 11.0. A precipitate (valuable element precipitate) was obtained by adding dropwise the raw material aqueous solution, an aqueous ammonia solution (concentration: 28% by mass) as a complexing agent, and an aqueous sodium hydroxide solution (concentration: 48% by mass) as an alkaline aqueous solution to the reaction vessel liquid. More specifically, the raw material aqueous solution was added dropwise to the reaction vessel liquid 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, an alkaline aqueous solution was also added dropwise to control the pH of the reaction vessel liquid to 11.0. During this process, the reaction vessel liquid was stirred with a stirring blade, while the temperature of the reaction vessel liquid was controlled at 40°C. The stirring speed (rotational speed of the stirring blade) was initially 200 rpm, but was increased to 350 rpm as the amount of liquid in the reaction vessel increased. The resulting precipitate was then filtered and washed with water, and then dried in a dryer at 100° C. for 10 hours, thereby obtaining Precursor 1.

[0145] Precursor 2 Precursor 2 was obtained in the same manner as Precursor 1, except that the composition of the raw material aqueous solution was Ni / Co / Mn (molar ratio) = 5 / 2 / 3.

[0146] Precursor 3 Precursor 3 was obtained in the same manner as Precursor 1, except that the composition of the raw material aqueous solution was Ni / Co / Mn (molar ratio) = 1 / 1 / 1.

[0147] Precursor 4 Precursor 4 was obtained in the same manner as Precursor 1, except that the dropwise addition rate of the complexing agent was changed to 0.6 mL / min.

[0148] Precursor 5 Instead of using the valuable element solution listed in Table 8, 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. Otherwise, Precursor 5 was obtained in the same manner as Precursor 1.

[0149] Precursor 6 Precursor 6 was obtained in the same manner as Precursor 1, except that the stirring speed was kept at 200 rpm from the beginning to the end of stirring (it was not increased to 350 rpm).

[0150] Precursor 7 Precursor 7 was obtained in the same manner as Precursor 1, except that the stirring speed was kept at 500 rpm from the beginning to the end of stirring.

[0151] <Characteristics of precursor> The tap density and particle size (D 10 , D 50 and D 90 The results are shown in Table 9 below.

[0152] [Table 9]

[0153] As shown in Table 9 above, precursors 1 to 5 had higher tap densities than precursor 6 with a stirring speed of 200 rpm and precursor 7 with a stirring speed of 500 rpm.

[0154] <Production of cathode materials> Using the obtained precursors 1 to 5, cathode materials 1 to 10 were produced as follows.

[0155] <Cathode material 1> Precursor 1 was mixed with lithium hydroxide as a lithium-containing compound to obtain a mixture. The molar ratio (Li / (Ni+Co+Mn)) during mixing was 1.075. The obtained mixture was fired to obtain a fired product. More specifically, the mixture was pre-fired at 650°C for 8 hours in an air atmosphere, and then fired at 950°C for 3 hours in an air atmosphere. The fired product was roughly crushed using a mortar. The fired product was not washed with water. In this way, cathode material 1 was obtained.

[0156] <Cathode Material 2 to Cathode Material 5> Cathode materials 2 to 5 were obtained in the same manner as for cathode material 1, except that precursors 2 to 5 were used, respectively.

[0157] <Cathode material 6> Cathode material 6 was obtained in the same manner as cathode material 1, except that precursor 1 and lithium carbonate as a lithium-containing compound were mixed to obtain a mixture.

[0158] <Cathode Material 7 to Cathode Material 10> Cathode materials 7 to 10 were obtained in the same manner as for cathode material 6, except that precursor 2 to cathode material 5 were used, respectively.

[0159] <Characteristics and evaluation of cathode materials> The tap density and particle size (D 10 , D 50and D 90 ) was determined. Furthermore, the charge capacity, discharge 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: Charge and discharge capacity N-methyl-2-pyrrolidone was added to the positive electrode material (90% by mass), acetylene black (5% by mass), and polyvinylidene fluoride (5% by mass), and the mixture was kneaded to obtain a mixture. 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 then roll-pressed to a density of 3.1 to 3.3 g / cm. 3 A pressure was applied so that the thickness was within the range of 14 mm. A disk having a diameter of 14 mm was punched out from the pressed laminate. The punched disk was dried in a vacuum at 150°C for 10 hours. The disk after vacuum drying was used as a positive electrode. A lithium metal sheet was used as the negative electrode, and a porous polyethylene film (thickness: 16 μm, manufactured by Staryuan Materials Co., Ltd.) was used as the separator. A non-aqueous electrolyte solution was obtained by dissolving 1 mol of LiPF6 in 1 L of a mixed solution of ethylene carbonate (EC) and dimethyl carbonate (DMC) in a volume ratio (EC / DMC) of 1 / 1. Using these positive electrodes, negative electrodes, separators, and nonaqueous electrolytes, a lithium-ion battery (test cell) for evaluation was fabricated in an argon-filled glove box. The fabricated test cell was charged and discharged at 25°C. Specifically, the cell was first charged at a constant current of 0.05C. When the voltage reached 4.3V, the charge was switched to constant voltage charging. When the charge current dropped to 0.01C, the charge was terminated. The cell was then discharged at a constant current of 0.05C until the voltage reached 2.75V. The charge capacity (unit: mAh / g) and discharge capacity (unit: mAh / g) were then determined.

[0161] Test 2 (Cycle Test): Cycle Characteristics First, a 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 carboxymethyl cellulose (1% by mass), and the mixture was kneaded to obtain a mixture. The 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 then roll-pressed to a density of 1.3 to 1.5 g / cm. 3 A disk was punched out from the pressed laminate to serve as a negative electrode. A lithium-ion battery (test cell) for evaluation was fabricated in the same manner as in Test 1 above, except for the negative electrode. Using the fabricated test cell, charge and discharge were repeated 500 times (500 cycles) at 60°C with a current of 1.0 C and a voltage range of 2.75 to 4.2 V. The cycle characteristics (unit: %) were calculated from the obtained discharge capacity (unit: mAh / g) using the following formula. Cycle characteristics = (discharge capacity at the 500th cycle / discharge capacity at the 1st cycle) x 100

[0162] Test 3: Discharge capacity after cycle test The positive electrode was removed from the test cell after the cycle test. A lithium ion battery (test cell) for evaluation was fabricated in the same manner as in Test 1 above, except for the removed positive electrode. Using the fabricated test cell, charge and discharge were performed in the same manner as in Test 1 above, and the discharge capacity (unit: mAh / g) was determined.

[0163] [Table 10]

[0164] As shown in Table 10 above, cathode materials 1 to 4, which used precursors 1 to 4 obtained using a valuable element solution, were equivalent in charge capacity, discharge capacity, cycle characteristics, and discharge capacity after cycle testing to cathode material 5, which did not use a valuable element solution. Furthermore, cathode materials 6 to 9, which used precursors 1 to 4 obtained using a valuable element solution, were equivalent in charge capacity, discharge capacity, cycle characteristics, and discharge capacity after cycle testing to cathode material 10, which did not use a valuable element solution.

Claims

1. 1. A method for producing a precursor of a cathode material for use in a lithium ion battery, comprising: The present invention relates to an oxide containing at least one valuable element selected from the group consisting of nickel, cobalt, and manganese, impurity elements of copper and iron, and a semi-valuable element of lithium, and a reducing agent, CaO, and SiO ２ Mass ratio (CaO / SiO ２ ) is 0.50 or less to obtain a mixed oxide, heating the mixed oxide to reduce the oxide and obtain a metal and a slag containing the semi-valent element; contacting the metal with an acid solution to obtain a leachate containing the valuable element and the impurity element; adding a sulfiding agent to the leachate to precipitate copper as copper sulfide, thereby obtaining the leachate from which copper has been removed as a copper-removed solution; adding an oxidizing agent to the copper-removing solution to precipitate iron as iron hydroxide, thereby obtaining the copper-removing solution from which iron has been removed as a valuable element solution containing the valuable element; The method for producing a precursor includes introducing the valuable element solution, a complexing agent, and an alkaline aqueous solution into a reaction bath liquid to obtain a precipitate containing the valuable element.

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

3. The method for producing a precursor according to claim 1 , wherein the amount of the reducing agent added is 2.0 equivalents or less.

4. The method for producing a precursor according to claim 1, wherein the mixed oxide is heated to a temperature of 1450°C or higher.

5. The method for producing a precursor according to claim 1 , wherein the metal obtained by heating the mixed oxide contains the valuable element and the impurity element.

6. The method for producing a precursor according to claim 1 , wherein the reducing agent is a carbon-containing substance.

7. The method for producing a precursor according to claim 1 , wherein the reducing agent is at least one iron-containing substance selected from the group consisting of metallic iron and iron oxide.

8. The method for producing a precursor according to claim 7 , wherein the iron oxide is ferrous oxide.

9. The method for producing a precursor according to claim 7, wherein the iron-containing material is at least one selected from the group consisting of dust, scale, sludge, and scrap.

10. The method for producing a precursor according to claim 1 , wherein the metal is powdered before being brought into contact with the acid solution.

11. The acid solution contains an acid and an oxidizing agent for the acid solution, The method for producing a precursor according to claim 1 , wherein 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 claim 11, wherein the oxidizing agent for the acid solution is hydrogen peroxide.

13. the amount of the sulfiding agent added is 1.0 equivalent or more relative to the copper contained in the leaching solution; The method for producing a precursor according to claim 1 , wherein the pH of the leaching solution to which the sulfiding agent has been added is adjusted to 3.0 or less when precipitating the copper sulfide.

14. 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 the oxidizing agent A added is 0.1 vvm or more relative to the copper removal solution; the amount of the oxidizing agent B added is 0.005% by volume or more relative to the copper removal solution; The method for producing a precursor according to claim 1 , wherein the pH of the copper removal solution to which the oxidizing agent has been added is adjusted to 3.0 or more and 7.0 or less when precipitating the iron hydroxide.

15. 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.

16. the alkaline aqueous solution is a sodium hydroxide aqueous solution, 2. The method for producing a precursor according to claim 1, wherein 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 cathode material for use in a lithium ion battery, comprising: A method for producing a cathode material, comprising mixing a precursor obtained by the method for producing a precursor according to 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.

18. The method for producing a positive electrode 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.

Images