JP7790463B2 - 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

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).
At this time, it may be required to minimize the reduction of Mn (to leave Mn in the slag without transferring it to the metal) and to selectively transfer Ni and Co to the metal and recover them.

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 selectively recover Ni and Co.

After extensive research, the inventors discovered that the above objectives can be achieved by adopting the following configuration, and thus completed the present invention.

That is, the present invention provides the following [1] to [18].
[1] A method for recovering valuable elements, comprising the steps of: adding a reducing agent to an oxide containing valuable elements, which are at least one selected from the group consisting of nickel and cobalt, and manganese, and impurity elements, copper and iron, to obtain a mixed oxide; heating the mixed oxide to reduce the oxide to obtain a metal; contacting the metal with an acid solution to obtain a leachate containing the valuable element and the impurity element; adding a sulfiding agent to the leachate to precipitate copper as copper sulfide, 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.
However, the reducing agent contains a carbon-containing substance and an iron-containing substance, the iron-containing substance is at least one selected from the group consisting of metallic iron and iron oxide, and the added amounts of the carbon-containing substance and the iron-containing substance are 1.0 equivalent or more and 1.6 equivalents or less in total.
[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 oxide further contains lithium.
[4] The method for recovering a valuable element according to any one of [1] to [3] above, wherein the manganese content in the oxide is 3.0 mass% or more and 12.0 mass% or less.
[5] The method for recovering valuable elements according to any one of [1] to [4] above, wherein the amount of the carbon-containing substance added is 1.0 equivalent.
[6] The method for recovering valuable elements according to any one of [1] to [5] above, wherein a slag-forming agent containing CaO and SiO ₂ is further added to the oxide when obtaining the mixed oxide.
[7] The method for recovering valuable elements according to the above [6], wherein the mass ratio of CaO to SiO ₂ (CaO/SiO ₂ ) contained in the slag forming agent is 0.50 or less.
[8] The method for recovering valuable elements according to any one of [1] to [7] above, wherein the mixed oxide is heated to a temperature of 1450°C or higher.
[9] The method for recovering valuable elements according to any one of [1] to [8] above, wherein the iron oxide is ferrous oxide.
[10] The method for recovering valuable elements according to any one of [1] to [9] above, wherein the iron-containing material is at least one selected from the group consisting of dust, scale, sludge, and scrap.
[11] The method for recovering a valuable element according to any one of [1] to [10] above, wherein the metal obtained by heating the mixed oxide contains the valuable element and the impurity element.
[12] The method for recovering a valuable element according to any one of [1] to [11] above, wherein the metal is powdered and then brought into contact with the acid solution.
[13] The method for recovering valuable elements according to any one of [1] to [12] 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.
[14] The method for recovering valuable elements according to [13] above, wherein the oxidizing agent for the acid solution is hydrogen peroxide.
[15] The method for recovering valuable elements according to any one of [1] to [14] 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.
[16] The method for recovering a valuable element according to any one of [1] to [15] 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.
[17] The method for recovering valuable elements according to [16] above, wherein the temperature of the copper removal solution to which the oxidizing agent has been added is 10°C or higher.
[18] A method for producing a metal containing at least one element selected from the group consisting of nickel and cobalt, and iron, using the method for recovering valuable elements according to any one of [1] to [11] above.

The present invention provides a method for recovering valuable elements that can remove impurity elements and selectively recover Ni and Co.

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, a reducing agent (described later) is first added to oxides (Ni, Co, Mn, Cu, Fe) to obtain a mixed oxide. At this time, a slag former (described later) may also be added.
The resulting mixed oxide is then heated to reduce the oxides to obtain metals (Ni, Co, Cu, Fe) and slag (Mn), which are then separated as appropriate.
Before the wet treatment, the metals obtained are preferably powdered to obtain metal powders (Ni, Co, Cu, Fe).
In the wet treatment, metals (metal powders) are first brought into contact with an acid solution to obtain a leachate (Ni, Co, 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) to obtain a copper-removed solution (Ni, Co, 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).

In this way, impurity elements (Cu, Fe) can be removed from the oxide, while the valuable elements Ni and Co can be selectively recovered, distinguished from Mn, which is also a valuable element.
According to this recovery method, 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.

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 valuable elements, such as at least one selected from the group consisting of nickel (Ni) and cobalt (Co) and manganese (Mn), and impurity elements, such as copper (Cu) and iron (Fe), specifically, for example, a cathode material for used lithium-ion batteries. This cathode material (oxide) may further contain a quasi-valent element, such as lithium (Li).
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.

In general, in view of the composition range of oxides (composite oxides) used as positive electrode materials for lithium ion batteries, the Mn content in the oxide is preferably 3.0 mass % or more, and more preferably 5.0 mass % or more.
Similarly, the Mn content in the oxide is preferably 12.0 mass % or less, and more preferably 10.0 mass % or less.

The Mn content in the oxide is determined using ICP (inductively coupled plasma) atomic emission spectroscopy. After confirming that measurement results similar to those obtained using ICP atomic emission spectroscopy are obtained, X-ray fluorescence (XRF) elemental analysis may also be used for simpler determination.

Addition of reducing agent (obtaining mixed oxides)
First, a reducing agent is added to an oxide to be reduced to obtain a mixed oxide, which is a mixture of the oxide and the reducing agent. At this time, a slag-forming agent, which will be described later, may be added.

<<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, Mn is generally inevitably reduced and converted to metal.
However, since it is difficult to separate Mn that has been transferred to the metal by subsequent hydroprocessing, it is desirable to minimize the reduction of Mn (leaving Mn in the slag).

Conventionally, substances with strong reducing power, such as Al-containing and Si-containing substances, have been used as reducing agents to avoid insufficient reduction. If the reducing power of the reducing agent is insufficient and insufficient reduction occurs, a portion of the cathode material is separated as slag in the form of oxide, and the content of valuable elements in the metal obtained by reducing the cathode material is reduced.
However, since elements located higher in the Ellingham diagram (FIG. 2) described below are more likely to metallize, when Si or Al is used as a reducing agent, Mn is also likely to metallize, and this does not satisfy the need to minimize the reduction of Mn.
Therefore, the present inventors investigated substances that have a lower reducing power than Al-containing and Si-containing substances as potential new reducing agents, and found that metallic iron (Fe) or iron oxide is effective.

The standard free energy change (ΔG ⁰ ) of the decomposition reaction of iron oxide is as follows:
FeO=Fe+1/2O ₂ :ΔG ⁰ =264430-64.73T[J]
Fe ₃ O ₄ =3FeO+1/2O ₂ :ΔG ⁰ =302370-108.15T[J]

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 can be recovered as a metal and Co can remain in the slag. Specifically, the following reactions are expected:
NiO+Fe→Ni+FeO: ΔG0=-29530-19.95T[J]
CoO+Fe→Co+FeO: ΔG0=-28950-6.82T[J]

As mentioned above, elements higher up on the Ellingham diagram (Figure 2) are more likely to metallize, so by using Fe (or FeO) as a reducing agent, it is expected that only Ni and Co will metallize, without metallizing Mn.

However, further investigation by the inventors revealed that when Fe (or FeO) is used as a reducing agent, the metallization of Mn can be avoided in the metal (product metal) obtained by reducing the cathode material (oxide), but the proportion of Fe may increase and the proportions of the valuable elements Ni and Co may decrease.

Therefore, the present inventors have considered using not only an iron-containing substance but also a carbon-containing substance as a reducing agent.
In the Ellingham diagram (FIG. 2), carbon (C) is located below Mn at temperatures above 1400° C., and therefore may be more likely to reduce Mn.
However, the present inventors have discovered that by using an appropriate amount of an iron-containing substance and a carbon-containing substance as a reducing agent, the reduction of Mn is suppressed and Mn remains in the slag (produced slag), while Ni and Co are obtained as the produced metals at a high reduction rate and the proportion of Fe in the produced metals can be reduced, i.e., Ni and Co can be selectively transferred to the produced metals (see Examples below).

Reducing Agent
The reducing agent used in this recovery method contains a carbon-containing substance and an iron-containing substance, that is, a carbon-containing substance and an iron-containing substance are used in combination as the reducing agent.
The content (total content) of the carbon-containing substance and the iron-containing substance in the reducing agent is preferably 90 mass % or more, more preferably 95 mass % or more, even more preferably 98 mass % or more, and particularly preferably 100 mass %.

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 material is at least one selected from the group consisting of metallic iron (Fe) and iron oxide. 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, if the reducing agents are metallic iron (Fe), ferrous oxide (FeO), coke (C), and propane (C ₃ H ₈ ), reduction with 1.0 equivalent of reducing agent is shown below, respectively:
Fe+ (NiO, CoO, MnO) → (Ni, Co, Mn) + FeO
3FeO+ (NiO, CoO, MnO) → (Ni, Co, Mn) + Fe ₃ O ₄
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

In calculating the equivalent of a carbon-containing substance, the fixed carbon content and the carbon content and hydrogen content in the volatile content contained in the carbon-containing substance are taken into consideration as components that contribute to reduction.
For example, when the carbon-containing substance is coke, the amount of coke added multiplied by the carbon content in the coke (unit: mass %) is calculated.
Furthermore, when the carbon-containing substance is propane gas, the following is calculated: (amount of propane added (unit: Nm ³ )/22.4)×(12×3+8).

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 total amount of the carbon-containing substance and iron-containing substance added as reducing agents is 1.0 equivalent or more and 1.6 equivalents or less.
As a result, as described above, reduction of Mn is suppressed and Mn remains in the produced slag, while Ni and Co are obtained as the produced metals at a high reduction rate, and furthermore, the proportion of Fe in the produced metals can be reduced.

The total amount of the carbon-containing substance and the iron-containing substance added is 1.0 equivalent or more, which is the stoichiometric composition, preferably 1.2 equivalents or more, and more preferably 1.3 equivalents or more.
On the other hand, the total amount added is 1.6 equivalents or less, and preferably 1.4 equivalents or less.

Of the carbon-containing substances and iron-containing substances used as reducing agents, the amount of only the carbon-containing substance such as coke (C) added is preferably, for example, 1.0 equivalent.
This minimizes insufficient reduction, and as described above, a higher reduction rate for Ni and Co can be obtained.

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.
However, in the present recovery method, Fe can be removed by wet treatment, so the amount of reducing agent added can be increased, making it easier to prevent insufficient reduction.

Slag former
In the present recovery method, it is preferable to use a slag-forming agent containing calcium oxide (CaO) and silicon dioxide (SiO ₂ ).
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 ))
Regarding the slag former, the mass ratio of CaO to _SiO2 (CaO/ _SiO2 ) is also called basicity. The mass ratio of the slag former (CaO/ _SiO2 ) is, for example, 2.00 or less, preferably 1.80 or less, and more preferably 1.60 or less.

However, it is preferable that the slag former has a low basicity, since this can further suppress the reduction rate of Mn. Specifically, the mass ratio of the slag former (CaO/SiO ₂ ) is preferably 1.50 or less, more preferably 1.00 or less, even more preferably 0.50 or less, and particularly preferably 0.35 or less.
Furthermore, when the oxide (cathode material) contains lithium (Li), by using a slag former with such a low basicity, it is possible to obtain a slag containing a large amount of Li in addition to the produced metal containing Ni and Co. This allows lithium to 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.

The lower limit of the mass ratio of the slag former (CaO/SiO ₂ ) is not particularly limited and is, for example, 0.15, preferably 0.20, more preferably 0.25, and even more preferably 0.30.

(Mass ratio {(CaO+Li ₂ O)/SiO ₂ })
The slag former may further contain lithium oxide (Li ₂ O) in addition to CaO and SiO ₂ .
The slag former is preferably prepared taking into consideration the amount of lithium.
Specifically, the mass ratio of the slag former {(CaO+Li ₂ O)/SiO 2 } is preferably 0.05 or more, more preferably 0.10 or more, and even more preferably _0.15 or more, because this makes it easier to maintain the reduction smelting ability (it makes it easier to prevent a decrease in the reduction reaction rate) and because it makes it easier to fix Li ₂ O in the slag (it makes it easier to recover Li in a later process).
On the other hand, because it is easy to prevent an increase in slag volume and to fix Li ₂ O in the slag, the mass ratio of the slag former {(CaO+Li ₂ O)/SiO ₂ } is preferably 2.50 or less, more preferably 2.00 or less, and even more preferably 1.50 or less.

(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, the valuable elements Ni and Co 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) and impurity elements (Cu, Fe).
The resulting metal may contain only one of the valuable elements (Ni, Co).
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), and iron (Fe).
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), in addition to the metal, slag (also referred to as "produced slag") is obtained. The produced slag contains Mn, a valuable element not contained in the produced metal, in the form of an oxide (MnO).
When reducing an oxide (cathode material) 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.

By using an iron-containing material as a reducing agent, the resulting slag contains FeO.
Furthermore, as described above, when the oxide (cathode material) contains lithium (Li), the generated slag also contains Li.

<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) 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) 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 the valuable elements (Ni, Co) 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 the valuable elements (Ni, Co).

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 Mn content in the positive electrode material was 11.3 mass%. 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 a reducing agent, powder of coke (C) was 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 former, a slag former containing CaO, SiO ₂ and Li ₂ O was prepared. A plurality of types of slag formers 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.
Table 2 below shows the type and amount (unit: equivalent) of the reducing agent used, as well as the mass ratio (CaO/SiO ₂ ) and mass ratio {(CaO+Li ₂ O)/SiO ₂ } of the slag forming agent used.

Heating of Mixed Oxides: Test A
The resulting mixed oxide was heated at 1600° C. for 3 hours in an Ar atmosphere.
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, Comparative Test Examples 2-1 and 2-2, in which only coke (C) was used as the reducing agent, had a high reduction rate of Mn.
In addition, in Comparative Test Examples 2-3 to 2-5, in which only metallic iron (Fe) or ferrous oxide (FeO) was used as the reducing agent, the reduction of Mn was suppressed, but the reduction of Ni and Co was insufficient.
In contrast, Test Examples 2-1 to 2-6, which used metallic iron (Fe) or ferrous oxide (FeO) in combination with coke (C) as the reducing agent, suppressed the reduction of Mn while achieving high reduction rates for Ni and Co. In other words, Ni and Co could be selectively recovered.

Furthermore, the proportion of Fe in the produced metal was lower in Test Examples 2-1 to 2-4, which used metallic iron (Fe) or ferrous oxide (FeO) in combination with coke (C), than in Comparative Test Examples 2-3 to 2-5, which used only metallic iron (Fe) or ferrous oxide (FeO) as the reducing agent.

Comparing Test Example 2-1 and Test Example 2-2, Test Example 2-2, in which the amount of coke (C), a carbon-containing substance, added was 1.0 equivalent, achieved a higher reduction rate for Ni and Co than Test Example 2-1, in which the amount was 0.4 equivalent.

Comparing Test Examples 2-2 to 2-4, Test Examples 2-3 to 2-4, which used a slag former with a CaO/SiO ₂ mass ratio of 0.50, were able to more effectively suppress the reduction of Mn than Test Example 2-2, which used a slag former with a CaO/ _{SiO 2} mass ratio of 1.50. They also increased the residual rate of Li in the resulting slag.

Comparing Test Examples 2-3 to 2-4 with Test Examples 2-5 to 2-6, Test Examples 2-5 to 2-6, 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 2-3 to 2-4 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) 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).
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 and 8-2 differed in the amounts of Fe and C added as reducing agents. In Comparative Test Example 8-1, only dry treatment was performed. In Comparative Test Example 8-2, only Fe was used as the reducing agent.

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 with Test Examples 8-1 to 8-2, Test Examples 8-1 to 8-2, which used a combination of Fe and C as the reducing agent, had a lower mass of Fe in the produced metal, a lower total mass of the reducing agent, sulfiding agent, and oxidizing agent, and a higher final Ni + Co recovery rate than Comparative Test Example 8-2, which used only Fe.

Comparing Test Example 8-1 and Test Example 8-2, Test Example 8-2, in which a larger amount of C was added as a reducing agent (a smaller amount of Fe was added), had a larger amount of Mn mass in the resulting metal than Test Example 8-1, in which a smaller amount of C was added (a larger amount of Fe was added), but the final Ni + Co recovery rate was higher.

Claims (18)

A reducing agent is added to an oxide containing at least one valuable element selected from the group consisting of nickel and cobalt, and manganese, and impurity elements, copper and iron, to obtain a mixed oxide;
heating the mixed oxide to reduce the oxide to obtain a metal;
contacting the metal with an acid solution to obtain a leachate containing the valuable element and the impurity element;
adding a sulfiding agent to the leachate to precipitate copper as copper sulfide, 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 reducing agent contains a carbon-containing substance and an iron-containing substance, and the iron-containing substance is at least one selected from the group consisting of metallic iron and iron oxide;
The total amount of the carbon-containing substance and the iron-containing substance added is 1.0 equivalent or more and 1.6 equivalents or less. 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 oxide further contains lithium. The method for recovering valuable elements described in claim 1, wherein the manganese content in the oxide is 3.0 mass% or more and 12.0 mass% or less. The method for recovering valuable elements described in claim 1, wherein the amount of carbon-containing substance added is 1.0 equivalent. The method for recovering valuable elements according to claim 1, wherein a slag-forming agent containing CaO and _SiO2 is further added to the oxide when obtaining the mixed oxide. 7. The method for recovering valuable elements according to claim 6, wherein the mass ratio of CaO to _SiO2 (CaO/ _SiO2 ) contained in the slag forming agent is 0.50 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 iron oxide is ferrous oxide. The method for recovering valuable elements described in claim 1, 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 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 described in any one of claims 1 to 11, 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 11, 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 13, 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 11, 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 11, 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 16, 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 11 to produce a metal containing at least one element selected from the group consisting of nickel and cobalt, and iron.