JP7790464B2 - Method for recovering valuable elements and method for producing metals - Google Patents

Description

The present invention relates to a method for recovering valuable elements and a method for producing metals.

In recent years, the demand for lithium-ion batteries has been increasing rapidly due to the widespread use of electric vehicles.
In particular, in light of the recent trend toward reducing _CO2 emissions, demand for electric vehicles that do not use fossil fuels is expected to expand further in the future, and the accompanying demand for lithium-ion batteries is also expected to increase further in the future.

Generally, the positive electrode 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 LiNiO ₂ , LiCoO ₂ , and LiMnO ₂ .
Metal elements such as Ni, Co, and Mn are not abundant worldwide.
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 utilization of resources.
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.

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 (see, for example, Patent Document 1).

Japanese Patent Application Laid-Open No. 2021-95628 JP 2013-91826 A JP 2012-224877 A

In the dry process, the composite oxides (LiNiO ₂ , LiCoO ₂ , LiMnO ₂ ) are reduced to produce slag in addition to metals containing valuable elements (Ni, Co, Mn).
In addition to valuable elements such as Ni, Co, and Mn, lithium (Li) may also be required. Hereinafter, Li may also be referred to as a "quasi-valuable element."
However, in the dry treatment, a part of Li may volatilize, and in that case, it is necessary to separately recover the Li, which is troublesome.

Furthermore, metals obtained by dry processing may contain impurity elements in addition to valuable elements such as Ni and Co. Examples of impurity elements include copper (Cu) and iron (Fe) derived from waste lithium-ion batteries.
When the metal obtained by dry processing is reused as a cathode material for a lithium-ion battery, if the metal contains impurity elements (Cu and Fe), the battery performance may be deteriorated. Therefore, it is desirable to remove the impurity elements as much as possible.

The present invention was made in consideration of the above points, and aims to provide a method for recovering valuable elements that can remove impurity elements and recover not only valuable elements but also lithium.

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 [16].
[1] A method for recovering valuable elements, 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, copper and iron as impurity elements, and lithium as a quasi-valuable element, to obtain a mixed oxide; heating the mixed oxide to reduce the oxide to obtain a metal and a slag containing the quasi-valuable 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-removed solution to precipitate iron as iron hydroxide, and obtaining the copper-removed solution from which iron has been removed as a valuable element solution containing the valuable element.
[2] The method for recovering valuable elements according to [1] above, wherein the oxide is obtained from waste lithium-ion batteries.
[3] The method for recovering valuable elements according to [1] or [2] above, wherein the amount of the reducing agent added is 2.0 equivalents or less.
[4] The method for recovering valuable elements according to any one of [1] to [3] above, wherein the mixed oxide is heated to a temperature of 1450°C or higher.
[5] The method for recovering a valuable element 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 recovering valuable elements according to any one of [1] to [5] above, wherein the reducing agent is a carbon-containing substance containing carbon.
[7] The method for recovering valuable elements according to any one of [1] to [6] above, wherein the reducing agent is at least one iron-containing substance selected from the group consisting of metallic iron and iron oxide.
[8] The method for recovering valuable elements according to [7] above, wherein the iron oxide is ferrous oxide.
[9] The method for recovering valuable elements 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 recovering a valuable element according to any one of [1] to [9] above, wherein the metal is powdered and then brought into contact with the acid solution.
[11] The method for recovering valuable elements 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 recovering valuable elements according to [11] above, wherein the oxidizing agent for the acid solution is hydrogen peroxide.
[13] The method for recovering valuable elements 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 recovering a valuable element 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 recovering valuable elements 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] A method for producing a metal containing at least one element selected from the group consisting of nickel and cobalt, using the method for recovering valuable elements according to any one of [1] to [9] above.

The present invention provides a method for recovering valuable elements that can remove impurity elements and recover not only valuable elements but also lithium.

1 is a flowchart showing an example of a method for recovering valuable elements. This is an Ellingham diagram (standard free energy change vs. temperature diagram). 1 shows potential-pH diagrams of Cu and Ni (S—H ₂ O system). 1 shows potential-pH diagrams of Fe and Ni (O ₂ —H ₂ O system).

[Method for recovering valuable elements]
In the method for recovering valuable elements of this embodiment (also referred to as "the present recovery method" for convenience), the positive electrode material (oxide) of a waste lithium ion battery is subjected to dry treatment and wet treatment.

FIG. 1 is a flowchart showing an example of a method for recovering valuable elements.
The recovery method will be explained briefly with reference to FIG.
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 a mixed oxide.
The resulting mixed oxide is then heated to reduce the oxides to obtain metals (Ni, Co, Mn, Cu, Fe) and slag (Li), which are optionally separated.
Before the wet treatment, the metals obtained are preferably powdered to obtain metal powders (Ni, Co, Mn, Cu, Fe).
In the wet treatment, 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), thereby obtaining 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) to obtain a valuable element solution (Ni, Co, Mn).

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.
According to this recovery method, valuable elements can be easily recovered from the cathode material (oxide) of waste lithium-ion batteries with a purity high enough to be reused as raw materials for lithium-ion batteries.

Next, the recovery method will be described in more detail.
The following description also serves as a description of the metal manufacturing method of this embodiment.

<Reducing target (oxide)>
The object to be reduced in this recovery method 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, the cathode material of a waste lithium-ion battery. The valuable element may be at least one element selected from the group consisting of Ni and Co.
The cathode material (oxide) is obtained from waste lithium-ion batteries by carrying out pre-processing such as removing the electrolyte, crushing, pulverizing, and sorting.

Addition of reducing agent and slag former (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.

<<Findings Obtained by the Inventors>>
The positive electrode material of a lithium ion battery is generally made of an oxide (composite oxide) such as LiNiO ₂ , LiCoO ₂ , or LiMnO ₂ .
Considering the dry process thermodynamically, for example, _LiNiO2 and _LiCoO2 decompose at high temperatures to produce NiO and CoO, respectively, as follows:
2LiNiO ₂ →Li ₂ O+2NiO+1/2O ₂
2LiCoO ₂ →Li ₂ O+2CoO+1/2O ₂

The standard free energy changes (ΔG ⁰ ) in the decomposition reactions of NiO and CoO are shown below, respectively.
NiO→Ni+1/2O ₂ :ΔG ⁰ =234900-84.68T[J]
CoO→Co+1/2O ₂ :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.

When valuable elements are recovered as metals from composite oxides by dry treatment, a slag-forming agent containing calcium oxide (CaO) and silicon dioxide (SiO ₂ ) may be used in addition to the reducing agent.
The mass ratio (CaO/SiO ₂ ) of CaO to SiO ₂ 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/SiO ₂ ) of 0.50 or less, it is possible to obtain a produced metal containing valuable elements such as Ni and Co, as well as a produced slag containing a large amount of Li, and 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.

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.

If the reducing power of the reducing agent is insufficient and reduction is insufficient, a part 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.

Examples of carbon-containing substances include solid carbon-containing substances such as graphite, coke, and solid hydrocarbons; and gaseous carbon-containing substances such as carbon monoxide (CO) and hydrocarbon gases (e.g., propane gas).
When a carbon-containing substance is used as the reducing agent, gases such as CO, CO ₂ and H ₂ O are generated after reduction, which is preferable in that the amount of slag generated does not increase.

The iron-containing substance is at least one selected from the group consisting of metallic iron (Fe) and iron oxide. Iron-containing substances as reducing agents are described below.

FIG. 2 is an Ellingham diagram (standard free energy change vs. 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/Fe ₃ O ₄ 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 reactions are expected:
NiO+Fe→Ni+FeO: ΔG ⁰ =-29530-19.95T[J]
CoO+Fe→Co+FeO: ΔG ⁰ =-28950-6.82T[J]

In the Ellingham diagram (FIG. 2), the higher the position, the easier it is to metallize.
When Si or Al is used as a reducing agent, Mn also tends to be metallized.
Therefore, it is expected that by using Fe (or FeO) as a reducing agent, only Ni and Co will be metallized without metallizing Mn.

Metallic iron (Fe) may be, for example, scrap or nuggets of iron used in steel mills.

Iron oxides are generally classified into three types: ferrous oxide (FeO), also known as wustite; triiron tetroxide (Fe ₃ O ₄ ), also known as magnetite; and ferric oxide (Fe ₂ O ₃ ), 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 as by-products 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.

(Amount of reducing agent added)
The amount of reducing agent required to reduce the oxide to be reduced is referred to as 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) + Fe ₃ O ₄
2Al+3(NiO, CoO, MnO) → 3(Ni, Co, Mn)+Al ₂ O ₃
Si+2(NiO, CoO, MnO) → 2(Ni, Co, Mn)+SiO ₂
C+2 (NiO, CoO, MnO) → 2 (Ni, Co, Mn) + CO ₂
C ₃ H ₈ + 10 (NiO, CoO, MnO) → 10 (Ni, Co, Mn) + 3 CO ₂ + 4 H ₂ O

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).

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.

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 poor reduction.
However, in the present recovery method, Fe can be removed by wet treatment, so the amount of reducing agent added can be increased (for example, 1.6 equivalents or more), making it easier to suppress insufficient reduction.

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.
Therefore, the amount of the reducing agent added is preferably 2.0 equivalents or less, more preferably 1.8 equivalents or less.

Slag former
In this recovery method, a slag-forming agent containing calcium oxide (CaO) and silicon dioxide (SiO ₂ ) is used.
The content (total content) of CaO and SiO ₂ in the slag former is preferably 90% by mass or more, more preferably 95% by mass or more, even more preferably 98% by mass or more, and particularly preferably 100% by mass.

(Mass ratio (CaO/SiO ₂ ))
In this recovery method, a low basicity slag former having a low mass ratio of CaO to SiO ₂ (CaO/SiO ₂ ) is used.
The mass ratio of the slag former (CaO/SiO ₂ ) is 0.50 or less, preferably 0.48 or less, more preferably 0.46 or less, even more preferably 0.44 or less, particularly preferably 0.42 or less, and most preferably 0.35 or less.
On the other hand, the lower limit is not particularly limited, and the mass ratio of the slag former (CaO/SiO ₂ ) 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.

(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.

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.

《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 suppress insufficient reduction.
Although there is no particular upper limit, the heating temperature is preferably 1800°C or less, and more preferably 1700°C or less.

<Heated atmosphere>
Suitable atmospheres for heating the mixed oxide (heating atmosphere) include, for example, inert atmospheres such as nitrogen gas (N ₂ ) atmosphere and argon gas (Ar) atmosphere; and reducing atmospheres such as carbon monoxide gas (CO) atmosphere.

<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.

《Product (metal)》
Metals are produced by reducing the oxide (cathode material) to be reduced, that is, valuable elements (Ni, Co, Mn) contained in the oxide are recovered as metals.

The metal obtained by reducing the oxide (also referred to as 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.

Product (slag)
By reducing the oxide (cathode material), slag is produced in addition to the metal.
As described above, the slag obtained in this recovery method (also referred to as "produced slag") contains a large amount of Li.
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 included in the produced metal.
When reducing an oxide containing Mn, by using an iron-containing substance as a reducing agent, the Mn/MnO equilibrium is more base than the Fe/FeO equilibrium and the FeO/ _{Fe3O4 equilibrium} , and therefore it is possible to prevent Mn from being mixed into the product metal obtained 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.

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

<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.

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).

The method for powdering the resulting metal is not particularly limited as long as it keeps the particle size of the resulting metal powder within an appropriate range. Examples include methods using a grinding device such as a jaw crusher or a vibration mill; atomization; etc.

<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 contacting the metal 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.

《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 saturation 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.

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

(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 viewpoint of realizing the "battery-to-battery" concept, in which used lithium-ion batteries are recycled and reused as raw materials for lithium-ion batteries, it is preferable to use sulfuric acid as the acid, because valuable elements can 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.

((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.
There is no particular upper limit, but 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.

(Oxidizing agent for acid solution)
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, ozone, etc. 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.

((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 vol% or more, more preferably 1.0 vol% or more, even more preferably 3.0 vol% or more, even more preferably 5.0 vol% or more, particularly preferably 6.0 vol% or more, and most preferably 6.9 vol% or more, relative to the acid (e.g., sulfuric acid).
On the other hand, the content of the oxidizing agent for the acid solution (e.g., hydrogen peroxide) in the acid solution is preferably 15.0 vol % or less, more preferably 13.0 vol % or less, and even more preferably 10.0 vol % or less, relative to the acid (e.g., sulfuric acid).

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.

<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.

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.

Figure 3 shows the potential-pH diagram for copper (Cu) and nickel (Ni) (S-H ₂ O system). Figure 3 shows the region on the potential-pH diagram for the copper (Cu)-sulfur (S)-water (H ₂ O) system where precipitates of oxides (hydroxides) or sulfides of copper (Cu) and nickel (Ni) are formed, taking into account their solubility.
Cobalt precipitates in the same manner as nickel, so cobalt is not shown in FIG.
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 a leachate from which copper (Cu) has been removed.

Sulfurizing agents
Examples of sulfiding agents to be added to the leaching solution include sulfur (S), hydrogen sulfide (H ₂ S), sodium hydrogen sulfide (NaSH), and sodium sulfide (Na ₂ S). 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.

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

(Amount of sulfurizing agent added)
From the viewpoint of sufficiently removing copper contained in the leaching solution, 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 leaching solution.
On the other hand, if an excessive amount of sulfurizing agent is added, the amount of sulfides (precipitates) of valuable elements (such as Ni and Co) increases, which may reduce the amount of the valuable elements desired to remain in the resulting copper removal solution. From this perspective, the amount of sulfurizing 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 leaching solution.

For example, when using 1.0 equivalent of sodium hydrogen sulfide (NaSH) as the 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.

《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 desired 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 leachate. The pH adjuster is not particularly limited, and examples thereof include sulfuric acid and sodium hydroxide.

Sulfurization Time
The time (sulfiding time) for sulfiding the copper contained in the leachate by reacting it with the sulfiding 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.

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.

<Addition of oxidizing agent (obtaining valuable element solution)>
Next, an oxidizing agent is added to the copper-removing 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-removing solution from which iron (Fe) has been selectively removed is obtained as a valuable element solution containing valuable elements (Ni, Co, Mn).

Figure 4 shows the potential-pH diagram for iron (Fe) and nickel (Ni) (O ₂ -H ₂ O system). Figure 4 shows the region where precipitates of oxides (hydroxides) of iron (Fe) and nickel ( _{Ni) form, taking into account their solubility, on the potential-pH diagram for the iron (Fe)-oxygen (O 2 )} -water (H 2 O) system.
Cobalt precipitates in the same manner as nickel, so cobalt is not shown in FIG.
As shown in Fig. 4, iron (Fe) selectively precipitates in a region where the pH is 3.0 or higher and 7.0 or lower, and where the oxidation-reduction potential is high. Although not shown in Fig. 4, in this region, iron precipitates as iron (III) oxide hydroxide (FeO(OH)). Utilizing this, the copper-removal solution is made acidic to neutral and oxidizing, so that the iron (Fe) contained in the copper-removal solution precipitates as iron (III) oxide hydroxide and is selectively removed. In other words, a valuable element solution is obtained, which is a copper-removal solution from which iron (Fe) has been removed.

Oxidizing agent
Examples of the oxidizing agent 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.

(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.

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.

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.

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.

Oxidation 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 (oxidation pH) is too low, precipitation may be difficult to form. Therefore, the oxidation pH is preferably 3.0 or higher, more preferably 3.7 or higher, even more preferably 4.0 or higher, and particularly preferably 4.5 or higher.
On the other hand, if the oxidation pH is too high, 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. The pH adjuster is not particularly limited, but examples thereof include sulfuric acid and sodium hydroxide.

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.

<<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 (Fe ₂ O ₃ ) and iron (III) oxide hydroxide (FeO(OH)), and the oxidation aid is preferably in the form of powder.
The mechanism by which the oxidation aid increases the reaction rate is catalytic action. Specifically, because the oxidation aid tends to become negatively charged in the aqueous solution (copper removal solution), it adsorbs Fe ²⁺ ions and weakens the bond with e ^- inside the Fe ^{2+ ions} . This is thought to lower the activation energy of the reaction Fe ²⁺ → Fe ³⁺ + e ^- (Fe oxidation reaction), accelerating the reaction.

(Amount of added oxidation promoter)
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) may 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.

(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 (such as Ni and Co). 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.

<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.

The present invention will be specifically explained below using examples. However, the present invention is not limited to the examples described below.

[Test A]
<Preparation of cathode material: Test A>
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 further contained copper (Cu) and iron (Fe) as impurity elements. The positive electrode material also contained lithium (Li) as a quasi-valent element.

<Preparation of reducing agent: Test A>
As reducing agents, powder of metallic aluminum (Al), powder of metallic silicon (Si), powder of coke (C), and propane gas (C ₃ H ₈ ) were prepared.
Furthermore, as reducing agents, powder of metallic iron (Fe) obtained by atomization and powder of ferrous oxide (FeO) were prepared.

<Preparation of slag former: Test A>
As the slag forming agent, a slag forming agent composed of CaO and SiO ₂ was prepared. A plurality of types of slag forming agents having different mass ratios of CaO to SiO ₂ (CaO/SiO ₂ ) were prepared.

Addition of reducing agent and slag-forming agent: Test A
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 cathode material, that is, the mass ratio of the slag former to the cathode material (slag former/cathode material) was about 0.67.
The type and amount (unit: equivalent) of the reducing agent used, and the mass ratio (CaO/SiO ₂ ) of the slag forming agent used are shown in Table 2 below.

Heating of Mixed Oxides: Test A
The resulting mixed oxide was heated for 3 hours in an Ar atmosphere at a heating temperature (unit: °C) as 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.

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

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 produced slag = 100 × (amount of Li contained in produced slag [kg]) / (amount of Li contained in oxide to be reduced [kg])

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

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 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.
Note that 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 at the time of generating FeO is slightly lower than the free energy at the time of generating CoO (see FIG. 2). However, since FeO generates 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.

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 ₂ ), 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.

Metal Powdering: Test A
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, metals 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.

<Contact between metal and acid solution: Test A>
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).
A metal (metal powder) having the composition shown in Table 3 below was contacted with the prepared acid solution 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 %, meaning that all of the elements were leached from the metal into the leachate.

<Addition of sulfurizing agent: Test A>
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 the mixture was 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 sulfiding agent was added (sulfidation pH) was adjusted to 0 (zero) using sulfuric acid and sodium hydroxide as pH adjusters.
In this way, 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). Thereafter, the copper sulfide was separated from the copper-removed solution, which was the leachate from which copper had been removed.
The content (unit: g/L) of each element in the copper removal solution was determined by ICP-AES (inductively coupled plasma atomic emission spectroscopy). 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 was very low, and therefore it can be seen that Cu was removed from the leachate with very high efficiency.

<Addition of Oxidizing Agent: Test A>
Next, the copper removal solution was first diluted with water. The contents (unit: g/L) of each element in the diluted copper removal solution are shown in Table 4 below. The reason for dilution was that in a preliminary experiment, 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 the mixture was stirred. The amount of the 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, iron (Fe) contained in the copper removal solution was oxidized by reacting with the oxidizing agent (oxidation time: 1.0 hour) and precipitated as iron hydroxide (iron (III) oxide hydroxide). Thereafter, the iron hydroxide was separated from the valuable element solution, which was the copper removal 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 (diluted) copper removal solution 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, and therefore it can be seen that Fe was removed from the copper removal solution with very high efficiency.

Furthermore, for each element, the final residual rate in the valuable element solution was calculated as the total residual rate (unit: mass%) from the residual rate a and the residual rate b. The results are shown in Table 4 below.
The results shown in Table 4 below show that by carrying out dry processing, then powdering the metal obtained by the dry processing, and then carrying out wet processing, valuable elements (Ni, Co, Mn) could be recovered with very high purity.

[Test B]
<Preparation of cathode material - Metal powder: Test B>
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.

<Contact between metal and acid solution: Test B>
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 within the scope of this example, 6.9% by volume is the most suitable amount of oxidizing agent (hydrogen peroxide) for the acid solution.

<Addition of sulfurizing agent: Test B>
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, thereby obtaining 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 residual rate and Co residual 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).

<Addition of oxidizing agent: Test B>
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 (unit: vol %) shown in Table 7 below, followed by stirring. At this time, the oxidation pH and oxidation temperature (unit: °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 range 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.

[Test C]
A valuable element solution was obtained by carrying out dry treatment and wet treatment (i.e., from preparation of the cathode material to addition of the oxidizing agent) in accordance with Test A. As shown in Table 8 below, for convenience, the mass of the cathode material (oxide) used is set to 100 kg.
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/SiO ₂ ) of 1.50 was used.

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.

Furthermore, the proportion of Ni and Co in the metal elements of the valuable element solution obtained by the wet treatment was calculated as the final Ni and 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.

Comparing Comparative Test Example 8-1 and Test Example 8-1, both had the same dry processing conditions, but Test Example 8-1, which underwent wet processing, ultimately yielded valuable elements (Ni + Co) with a higher purity than Comparative Test Example 8-1, which did not undergo wet processing.

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/SiO ₂ ) 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.

Comparing Test Example 8-1 and Test Example 8-2, Test Example 8-2, which used Fe as the reducing agent, had a higher residual Li rate in the resulting slag than Test Example 8-1, which used Al as the reducing agent, and was able to recover even more Li.

Comparing Test Example 8-2 and Test Example 8-3, in which Fe was used 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, in the dry treatment, the amount of Fe added as a reducing agent was increased, so that 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, a high purity of valuable elements can be finally obtained, and the amount of Fe added in the dry treatment can be increased to sufficiently reduce the valuable elements (Ni + Co), resulting in a high final recovery rate of Ni + Co, including the wet treatment.

Claims (16)

A mixed oxide is obtained by adding a reducing agent and a slag former having a mass ratio of CaO to _SiO2 (CaO/SiO2 ₎ of 0.50 or less to an oxide containing at least one valence element selected from the group consisting of nickel, cobalt, and manganese, impurity elements being copper and iron, and a semi-valence element being lithium,
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;
a method for recovering a valuable element, comprising adding an oxidizing agent to the copper-removing solution to precipitate iron as iron hydroxide, and obtaining the copper-removing solution from which iron has been removed as a valuable element solution containing the valuable element. The method for recovering valuable elements described in claim 1, wherein the oxide is obtained from waste lithium-ion batteries. The method for recovering valuable elements described in claim 1, wherein the amount of the reducing agent added is 2.0 equivalents or less. The method for recovering valuable elements described in claim 1, wherein the mixed oxide is heated to a temperature of 1450°C or higher. The method for recovering valuable elements described in claim 1, wherein the metal obtained by heating the mixed oxide contains the valuable element and the impurity element. The method for recovering valuable elements according to claim 1, wherein the reducing agent is a carbon-containing substance. The method for recovering valuable elements described in claim 1, wherein the reducing agent is at least one iron-containing substance selected from the group consisting of metallic iron and iron oxide. The method for recovering valuable elements described in claim 7, wherein the iron oxide is ferrous oxide. The method for recovering valuable elements described in claim 7, wherein the iron-containing material is at least one selected from the group consisting of dust, scale, sludge, and scrap. The method for recovering valuable elements described in any one of claims 1 to 9, wherein the metal is powdered before being brought into contact with the acid solution. The acid solution contains an acid and an oxidizing agent for the acid solution,
The method for recovering valuable elements according to any one of claims 1 to 9, wherein the content of the oxidizing agent for the acid solution is 0.5% by volume or more relative to the acid. The method for recovering valuable elements described in claim 11, wherein the oxidizing agent for the acid solution is hydrogen peroxide. 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 recovering valuable elements according to any one of claims 1 to 9, wherein the pH of the leachate to which the sulfiding agent has been added is adjusted to 3.0 or less when precipitating the copper sulfide. 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 recovering a valuable element according to any one of claims 1 to 9, 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. The method for recovering valuable elements described in claim 14, wherein the temperature of the copper removal solution to which the oxidizing agent has been added is 10°C or higher. A metal production method that uses the valuable element recovery method described in any one of claims 1 to 9 to produce a metal containing at least one element selected from the group consisting of nickel and cobalt.