JP2016037661A - Method for recovering valuable metal - Google Patents
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Abstract
Description
本発明は、湿式法と乾式法を組み合わせて廃電池から有価金属を回収する工程において、有価金属に残留する硫黄分を効果的に除去することができる回収方法に関する。 The present invention relates to a recovery method capable of effectively removing a sulfur content remaining in a valuable metal in a step of recovering the valuable metal from a waste battery by combining a wet method and a dry method.
リチウムイオン電池は、近年、小型化あるいは軽量化が進むモバイルコンピュータや携帯電話等の電子機器用の電池として注目され、その需要及び消費が飛躍的に増大している。
リチウムイオン電池は、現状においては、正極活物質として、コバルト酸リチウム（コバルト系）である場合が殆どであるが、希少元素であるコバルトが電池コストの大半を占めるため、大幅な低コスト化を目指して正極活物質として、ニッケル酸リチウム（ニッケル系）、マンガン酸リチウム（マンガン系）、リン酸鉄リチウム（リン酸鉄系）等が開発されている。
このため、市場で流通された使用済みのリチウムイオン電池（廃電池）に含まれる正極活物質から、コバルト、ニッケル、マンガン及びリチウム等の有価金属を回収することは、資源の有効利用の観点から極めて重要である。
廃電池からの有価金属の分別回収方法として、湿式法、乾式法、乾式法と湿式法との組合せ等種々の方式が提案されている。
In recent years, lithium ion batteries have attracted attention as batteries for electronic devices such as mobile computers and mobile phones, which are becoming smaller and lighter, and the demand and consumption thereof are dramatically increasing.
At present, lithium ion batteries are mostly lithium cobaltate (cobalt) as the positive electrode active material. However, since the rare element cobalt occupies most of the battery cost, the cost is greatly reduced. Aiming at positive electrode active materials, lithium nickelate (nickel), lithium manganate (manganese), lithium iron phosphate (iron phosphate) and the like have been developed.
For this reason, recovering valuable metals such as cobalt, nickel, manganese and lithium from the positive electrode active material contained in used lithium ion batteries (waste batteries) distributed in the market is from the viewpoint of effective use of resources. Very important.
Various methods such as a wet method, a dry method, and a combination of a dry method and a wet method have been proposed as a method for separating and recovering valuable metals from waste batteries.
乾式法としては、例えば、使用済みリチウム２次電池からの有価金属の回収方法として、電池構成材料の有機材料を除去する予備焙焼処理、予備焙焼処理後の焙焼物を粉砕処理して粉砕物を得、該粉砕物を篩分け処理して、篩下としてニッケル及びコバルトの１種以上とアルミニウムを含む１次濃縮物を得る工程、該１次濃縮物をカルシウム化合物と混合し、次に熔融し、前記粉砕物中のアルミニウム成分と前記カルシウム化合物から生成されるスラグを除去して、ニッケルおよびコバルトの１種以上の２次濃縮物を回収する方法が提案されている（例えば、特許文献１参照）。 As a dry method, for example, as a method for recovering valuable metals from used lithium secondary batteries, pre-roasting treatment for removing organic materials of battery constituent materials, and pulverizing and processing the pre-baked roasted material Obtaining a product, sieving the pulverized product to obtain a primary concentrate containing one or more of nickel and cobalt and aluminum as a sieve, and mixing the primary concentrate with a calcium compound; A method for recovering one or more secondary concentrates of nickel and cobalt by melting and removing slag produced from the aluminum component and the calcium compound in the pulverized product has been proposed (for example, Patent Documents). 1).
一方、電池構成材料として銅及び炭素を含む場合、廃電池を予備焙焼後、粉砕処理した粉砕物を、篩分け処理して、篩下としてニッケル及びコバルト等と銅及び炭素を含む1次濃縮物から、先ず、銅及び炭素を除いておくことが、以後に有価金属を回収する際の高純度化に効果的である。
そこで、銅及び炭素を除くため、硫酸を用いて酸溶解処理をし、硫酸に溶解しにくい銅および溶解しない炭素をろ過残渣として除くことが考えられる。
On the other hand, when copper and carbon are included as battery constituent materials, after the pre-baking of the waste battery, the pulverized pulverized product is subjected to sieving treatment, and the primary concentration containing nickel and cobalt, etc., and copper and carbon as the sieving First, removing copper and carbon from the product is effective for high purity when recovering valuable metals thereafter.
Therefore, in order to remove copper and carbon, it is conceivable that acid dissolution treatment is performed using sulfuric acid, and copper that is difficult to dissolve in sulfuric acid and carbon that does not dissolve are removed as filtration residues.
しかしながら、有価金属が溶解されたろ液中の硫酸由来の残留硫黄によって、以下の問題が発生することが知得された。
すなわち、以後の回収工程において、熔融過程で生成する残留硫黄由来の低融点組成スラグが熔融装置の耐火物、例えば不定形耐火材等を侵食し、消耗材コストを押し上げる要因となることが判明した。また、残留硫黄により有価金属の純度価値が低下する。
このように残留する硫黄分は、有価金属の純度価値を下げ、耐火材に悪影響を及ぼすので、回収方法の工程中で、硫酸を酸溶解処理工程に用いた際には、後工程で硫酸由来の硫黄分を除去することが望ましい。
However, it has been found that the following problems occur due to residual sulfur derived from sulfuric acid in the filtrate in which valuable metals are dissolved.
In other words, in the subsequent recovery process, it was found that the low-melting-point composition slag derived from residual sulfur generated in the melting process corrodes the refractory of the melting apparatus, for example, the irregular refractory material, and increases the cost of the consumables. . Moreover, the purity value of valuable metals is lowered by residual sulfur.
Since the residual sulfur content lowers the purity value of valuable metals and adversely affects the refractory material, when sulfuric acid is used in the acid dissolution treatment process during the recovery process, it is derived from sulfuric acid in the subsequent process. It is desirable to remove the sulfur content.
先に述べたように、残留硫黄濃度が高いコバルト等の有価金属を回収すると、その残留硫黄の一部が有価金属へ混入し、有価金属の純度価値を下げ、さらに、熔融工程で生成する残留硫黄由来の低融点組成スラグが、熔融装置の耐火物、例えば不定形耐火材等を侵食し、消耗材コストを押し上げる要因となることが判明した。
本発明は、上記課題を解決するためになされたものであり、残留硫黄分の除去を酸化焙焼と水洗浄のみで行い、低コストで残留硫黄分を除去することにより、有価金属の純度価値を高め、かつ、熔融工程で生成される残留硫黄由来の低融点組成スラグが、熔融装置の耐火物、例えば不定形耐火材等を侵食することによる耐火材の損傷を防ぐことができるコバルト等の有価金属の回収方法を提供することにある。
As mentioned earlier, when valuable metals such as cobalt with a high residual sulfur concentration are recovered, a part of the residual sulfur is mixed into the valuable metals, reducing the purity value of the valuable metals, and the residual produced in the melting process. It has been found that the low melting point composition slag derived from sulfur erodes the refractory of the melting apparatus, for example, an indeterminate refractory material, and increases the cost of consumables.
The present invention has been made in order to solve the above-mentioned problems. The purity value of a valuable metal is obtained by removing residual sulfur with only oxidation roasting and washing with water and removing residual sulfur at low cost. And low melting point composition slag derived from residual sulfur produced in the melting process can prevent damage to the refractory material due to erosion of the refractory material of the melting apparatus, such as the amorphous refractory material, etc. It is to provide a method for recovering valuable metals.
本発明者らは、廃電池からの有価金属の１次濃縮物を硫酸で溶解する酸溶解処理工程を経て酸溶解液から水酸化物として回収した有価金属を、水洗処理する前に酸化焙焼する低硫黄化処理を施すことにより、水洗処理による単独処理よりも有価金属中の残留硫黄分が減少されること、並びに乾式法を用いて有価金属を回収する際に、残留硫黄由来の低融点組成スラグ生成による耐火物の損傷、例えば不定形耐火材等の損傷を防ぐことができることを見出し、本発明を完成するに至った。
本発明は以下の〔１〕〜〔６〕を提供する。
〔１〕コバルトまたはニッケルの少なくとも１種の有価金属を含む廃電池または工程屑からの前記有価金属の回収方法であって、
（１）予備焙焼処理、粉砕処理及び篩分け処理を経て、前記有価金属の１次濃縮物を得る１次濃縮工程、
（２）前記１次濃縮物を硫酸で溶解処理し、該溶解液を２次濃縮物として得る２次濃縮工程、
（３）前記２次濃縮物を水酸化処理後、低硫黄化処理し、前記有価金属の３次濃縮物を得る３次濃縮工程、及び
（４）前記３次濃縮物を熔融し、前記有価金属を回収する４次濃縮工程、
を含むことを特徴とする有価金属の回収方法。
〔２〕前記３次濃縮工程における２次濃縮物の水酸化処理は、前記２次濃縮物にアルカリ金属の水溶液を添加することを特徴とする前記〔１〕に記載の有価金属の回収方法。
〔３〕前記３次濃縮工程の低硫黄化処理は、酸化焙焼処理、及び水洗処理を含むことを特徴とする前記〔１〕又は〔２〕に記載の有価金属の回収方法。
〔４〕前記酸化焙焼処理は、前記３次濃縮工程における水酸化処理後の沈殿物を焙焼することを特徴とする前記〔３〕に記載の有価金属の回収方法。
〔５〕前記酸化焙焼処理は、１００〜８５０℃で、大気雰囲気もしくは酸素雰囲気下で行われることを特徴とする前記〔３〕又は〔４〕に記載の有価金属の回収方法。
〔６〕コバルトまたはニッケルの少なくとも１種を含む有価金属の硫酸溶解液からの前記有価金属の回収方法であって、
（１）前記硫酸溶解液を水酸化処理する工程、
（２）前記水酸化処理後の沈殿物を酸化焙焼処理する工程、
（３）水洗処理する工程、
を含むことを特徴とする有価金属の回収方法。
The present inventors oxidized and roasted valuable metals recovered as hydroxides from an acid-dissolved solution through an acid-dissolving treatment step of dissolving a primary concentrate of valuable metals from waste batteries with sulfuric acid before washing with water. By applying a low sulfur treatment, residual sulfur content in valuable metals is reduced compared to single treatment by washing with water, and when recovering valuable metals using the dry method, a low melting point derived from residual sulfur. It has been found that damage to refractories due to composition slag generation, such as damage to irregular refractory materials, can be prevented, and the present invention has been completed.
The present invention provides the following [1] to [6].
[1] A method for recovering the valuable metal from waste batteries or process waste containing at least one valuable metal of cobalt or nickel,
(1) A primary concentration step for obtaining a primary concentrate of the valuable metal through preliminary roasting treatment, pulverization treatment and sieving treatment,
(2) a secondary concentration step of dissolving the primary concentrate with sulfuric acid to obtain the solution as a secondary concentrate;
(3) A tertiary concentration step for obtaining a tertiary concentrate of the valuable metal by subjecting the secondary concentrate to a low sulfur treatment after hydroxylation, and (4) melting the tertiary concentrate, A fourth concentration step for recovering the metal,
A method for recovering valuable metals, comprising:
[2] The method for recovering a valuable metal as described in [1] above, wherein the hydroxylation treatment of the secondary concentrate in the tertiary concentration step comprises adding an aqueous solution of an alkali metal to the secondary concentrate.
[3] The method for recovering a valuable metal according to [1] or [2], wherein the sulfur reduction treatment in the tertiary concentration step includes an oxidation roasting treatment and a water washing treatment.
[4] The method for recovering a valuable metal according to [3], wherein the oxidation roasting treatment roasts the precipitate after the hydroxylation treatment in the tertiary concentration step.
[5] The method for recovering a valuable metal according to [3] or [4], wherein the oxidation roasting treatment is performed at 100 to 850 ° C. in an air atmosphere or an oxygen atmosphere.
[6] A method for recovering the valuable metal from a sulfuric acid solution of the valuable metal containing at least one of cobalt and nickel,
(1) A step of hydroxylating the sulfuric acid solution,
(2) a step of oxidizing and roasting the precipitate after the hydroxylation treatment,
(3) A step of washing with water,
A method for recovering valuable metals, comprising:
本発明の有価金属の回収方法によれば、酸溶解処理後に回収した有価金属を水洗処理する前に酸化焙焼を施すことにより、有価金属中の残留硫黄分を減少させることができ、結果的に有価金属への硫黄分の混入の低減及び低融点組成スラグによる熔融装置の耐火物、例えば不定形耐火材等の損傷を防ぐことができる。さらに、酸化焙焼と水洗のみで硫黄分を除去でき、薬品を使用しないため低コストでコバルト等の有価金属を回収できる。 According to the method for recovering valuable metals of the present invention, the residual sulfur content in the valuable metals can be reduced by subjecting the valuable metals recovered after the acid dissolution treatment to oxidative roasting before washing with water. In addition, it is possible to reduce the contamination of valuable metals with sulfur and to prevent damage to the refractory material of the melting apparatus such as an indeterminate refractory material due to low melting point composition slag. Furthermore, sulfur can be removed only by oxidation roasting and washing, and since no chemical is used, valuable metals such as cobalt can be recovered at low cost.
以下、本発明を詳細に説明する。
なお、本発明において、「有価金属」とは、コバルト（Ｃｏ）またはニッケル（Ｎｉ）の少なくとも１種を含むものとし、「廃電池」には、使用済み電池および電池製造工程の工程屑を含むものとする。
また、「電池」には、リチウムイオン電池の他に、ニッケル水素電池、ニッケルカドミウム電池等が含まれる。
本発明の廃電池からの有価金属の回収方法は、下記の（１）〜（４）の濃縮工程を含み、特に（３）の３次濃縮工程において、低硫黄化処理された状態で３次濃縮物を得る点に特徴を有している。
Hereinafter, the present invention will be described in detail.
In the present invention, “valuable metal” includes at least one of cobalt (Co) and nickel (Ni), and “waste battery” includes used batteries and process waste from battery manufacturing processes. .
In addition to the lithium ion battery, the “battery” includes a nickel metal hydride battery, a nickel cadmium battery, and the like.
The method for recovering valuable metals from a waste battery of the present invention includes the following concentration steps (1) to (4), and in particular, in the tertiary concentration step (3), the tertiary is obtained in a state of low sulfur treatment. It is characterized in that a concentrate is obtained.
（１）１次濃縮工程
有価金属回収対象の廃電池を開孔処理した後、電池構成材料の有機材料を除去する予備焙焼処理、予備焙焼処理後の焙焼物の粉砕処理及び篩分け処理を経て、有価金属の１次濃縮物を得る工程である。本発明の1次濃縮工程には、以下の（ｉ）〜（ｉｖ）の処理工程が含まれる。
（i）開孔処理：廃電池は密閉系であり、内部に電解液等を有しており、このまま次工程の予備焙焼処理を行うと爆発の恐れがあり危険である。このため廃電池の爆発防止を目的として、何らかの方法でガス抜きのための開孔処理を施すことが望ましい。具体的な開孔処理方法としては、例えば、針状の刃先で廃電池を物理的に開孔する方法が挙げられる。
(ii)予備焙焼処理：電池構成材料である、セパレーター、電解液等の有機材料を分解、燃焼または揮発させて除去する目的で行われる。開孔処理した廃電池をコバルトやニッケルの還元が起こらない温度である４００℃以下の温度で予備焙焼して、廃電池中のポリエチレンやＮ−メチルピロリドンなど、熱分解する有機物質をガス化燃焼し系外へ除去する。
予備焙焼温度は、電池構成材料である有機材料の分解、燃焼を有効に行う観点から、３５０℃〜４００℃がより好ましい。また、予備焙焼を行う際の雰囲気は酸化性であることが好ましい。また、予備焙焼時間をあまり長くすると、Ｃｕが粉砕され易くなり、１次濃縮物へのＣｕの混入が増加する虞があることから、予備焙焼時間は３時間以上６時間以下とすることが適している。
(iii)粉砕処理Ａ：予備焙焼処理後の焙焼物を粉砕して粉砕物を得る目的で行われる。予備焙焼処理により、非常に粉砕され易くなっているコバルトに富む活物質と、Ａｌ箔、Ｃｕ箔、外装缶（Ｆｅ、Ａｌ）等のコバルトが少なく、かつ比較的粉砕され難い部分とを篩分けＡで分離し易くするために行う。粉砕は破砕機を含む破砕設備を用いて行うことができ、破砕機については特に制限はなく、いかなるものも使用できる。
（ｉｖ）篩分けＡ：粉砕処理Ａで得られた粉砕物を、篩分けする。篩目の開きは２.００ｍｍ以下、０．６０ｍｍ以上が好ましく、１．１８ｍｍ以下、０．８５ｍｍ以上が特に好ましい。篩目が２．００ｍｍより大きいと、有価金属以外のベースメタル（Ｆｅ、Ｃｕ、Ｚｎ、Ａｌ等）の混入が増え、後工程の処理が煩雑となる。一方、篩目が０．６０ｍｍより小さいと、ベースメタルの混入は減るが、１次濃縮物の収率を低下させるため好ましくない。
例えば篩目開き１.００ｍｍの篩を用いて篩分けし、１．００ｍｍ以下のコバルトに富む１次濃縮物を篩下として得る。１．００ｍｍより大きいＡｌ箔やＣｕ箔、外装缶は篩上物Ａとして分離・回収される。
(1) Primary concentration step After the waste battery subject to valuable metal recovery is opened, pre-roasting treatment to remove the organic material of the battery constituent material, pulverization treatment and sieving treatment of the roasted material after the pre-roasting treatment To obtain a primary concentrate of valuable metals. The primary concentration step of the present invention includes the following processing steps (i) to (iv).
(I) Opening treatment: The waste battery is a closed system, and has an electrolyte or the like inside. If the pre-baking treatment in the next step is performed as it is, there is a risk of explosion, which is dangerous. For this reason, for the purpose of preventing the explosion of the waste battery, it is desirable to perform an opening process for degassing by some method. As a specific hole treatment method, for example, there is a method of physically opening a waste battery with a needle-like blade edge.
(ii) Pre-roasting treatment: It is performed for the purpose of decomposing, burning or volatilizing and removing organic materials such as separators and electrolytes, which are battery constituent materials. The waste battery that has been subjected to pore opening treatment is pre-roasted at a temperature of 400 ° C. or less, at which cobalt or nickel is not reduced, and gasified organic substances such as polyethylene and N-methylpyrrolidone in the waste battery are gasified. Burn and remove out of the system.
The preliminary roasting temperature is more preferably 350 ° C. to 400 ° C. from the viewpoint of effectively decomposing and burning the organic material that is the battery constituent material. Moreover, it is preferable that the atmosphere at the time of preliminary roasting is oxidizing. Also, if the pre-roasting time is too long, Cu is likely to be pulverized and there is a risk that Cu contamination into the primary concentrate may increase, so the pre-roasting time should be 3 hours or more and 6 hours or less. Is suitable.
(iii) Grinding treatment A: It is carried out for the purpose of obtaining a pulverized product by pulverizing the baked product after the preliminary roasting treatment. The active material rich in cobalt, which is very easily pulverized by the pre-roasting process, and the portion of Al foil, Cu foil, outer can (Fe, Al), etc. that are low in cobalt and relatively difficult to pulverize. This is done in order to facilitate separation in division A. The pulverization can be performed using a crushing facility including a crusher, and the crusher is not particularly limited, and any one can be used.
(Iv) Sieving A: The pulverized product obtained by the pulverizing treatment A is sieved. The opening of the sieve is preferably 2.00 mm or less and 0.60 mm or more, particularly preferably 1.18 mm or less and 0.85 mm or more. When the mesh size is larger than 2.00 mm, the base metal (Fe, Cu, Zn, Al, etc.) other than valuable metals is mixed, and the subsequent process becomes complicated. On the other hand, if the mesh size is smaller than 0.60 mm, the contamination of the base metal is reduced, but the yield of the primary concentrate is lowered, which is not preferable.
For example, sieving is performed using a sieve having a sieve opening of 1.00 mm, and a cobalt-rich primary concentrate having a size of 1.00 mm or less is obtained as a sieve. Al foil, Cu foil, and outer can larger than 1.00 mm are separated and collected as sieve top A.
（２）２次濃縮工程
前記１次濃縮物を硫酸で溶解処理して、ろ過し、Ｃｕ及びＣをろ過残渣として回収する一方、ろ液を２次濃縮物として得る工程であり、以下の（ｖｉ）〜（ｖii）の処理が含まれる。
（ｖｉ）酸溶解処理：１次濃縮物に硫酸を加え、ｐＨを０．６〜１．０の範囲に調整する。硫酸に溶解しにくいＣｕ及び硫酸に溶解しないＣを次の除銅処理工程で分離するために行われる。ｐＨを０．６未満にすると、Ｃｕが溶解しやすくなり、溶解液であるろ液（２次濃縮物）へのＣｕの混入量が増える虞がある。そこで、ｐＨを０．６以上とし、２次濃縮物へのＣｕの混入を抑制することが好ましい。
（vii）除銅処理：前工程の酸溶解処理物をろ過し、Ｃｕ及びＣをろ過残渣として分離、回収する。
(2) Secondary concentration step The primary concentrate is dissolved in sulfuric acid, filtered, and Cu and C are recovered as a filtration residue, while the filtrate is obtained as a secondary concentrate. The processes of vi) to (vii) are included.
(Vi) Acid dissolution treatment: sulfuric acid is added to the primary concentrate to adjust the pH to a range of 0.6 to 1.0. This is performed in order to separate Cu, which is difficult to dissolve in sulfuric acid, and C, which does not dissolve in sulfuric acid, in the next copper removal process. When the pH is less than 0.6, Cu is easily dissolved, and there is a possibility that the amount of Cu mixed into the filtrate (secondary concentrate), which is a solution, increases. Therefore, it is preferable to set the pH to 0.6 or more and suppress the mixing of Cu into the secondary concentrate.
(Vii) Copper removal treatment: The acid-dissolved treatment product in the previous step is filtered, and Cu and C are separated and collected as a filtration residue.
（３）３次濃縮工程
前記２次濃縮物を水酸化処理して、有価金属の水酸化物を低硫黄化した３次濃縮物として得る工程であり、以下の（ｉｘ）及び（ｘ）の処理が含まれる。
（ｉｘ）水酸化処理：前記２次濃縮物にアルカリ金属の水溶液を添加し、有価金属を水酸化物として沈殿させ、回収することを目的とする。
（ｘ）低硫黄化処理：前記水酸化処理工程で回収した水酸化物中の残留硫黄分を低減させる目的で、回収した水酸化物（水酸化処理後の沈殿物）を、１００℃〜８５０℃の範囲、好ましくは４００℃〜６００℃の範囲で、大気雰囲気もしくは酸素雰囲気下で焙焼（酸化焙焼）する。下限値は、水酸化物中に包含された芒硝が、水の蒸発により放出され易くなる温度（１００℃）とした。すなわち、芒硝は流動性に劣るので、単に水洗処理しただけでは十分に除去できない。特に、水酸化物の孔隙部に存在する芒硝は除去され難い。そこで本発明においては、前記水酸化物を加熱することに伴い発生する水蒸気を利用し、前記水酸化物の周囲や孔隙部から芒硝を追い出すことにより、通常の水洗だけでは硫黄成分が残留してしまう課題を解決した。また、８５０℃より高くなると芒硝が溶解し、その後の水洗処理による芒硝除去が困難になるため、上限値は８５０℃とした。また、焙焼温度を４００〜６００℃とすることで、焙焼時間の短縮化を図ることができ、効率的に３次濃縮物を得ることが可能となる。
(3) Tertiary concentration step This is a step of subjecting the secondary concentrate to a hydroxylation treatment to obtain a valuable metal hydroxide as a low-sulfurized tertiary concentrate, wherein the following (ix) and (x) Processing is included.
(Ix) Hydroxidation treatment: An object is to add an aqueous solution of an alkali metal to the secondary concentrate to precipitate valuable metals as hydroxides and collect them.
(X) Low sulfurization treatment: For the purpose of reducing the residual sulfur content in the hydroxide recovered in the hydroxylation step, the recovered hydroxide (precipitate after the hydroxylation treatment) is treated at 100 ° C to 850 ° C. Roasting (oxidation roasting) is performed in an air atmosphere or an oxygen atmosphere in the range of ° C, preferably in the range of 400 ° C to 600 ° C. The lower limit value was set to a temperature (100 ° C.) at which the sodium sulfate contained in the hydroxide was easily released by water evaporation. That is, since salt cake is inferior in fluidity, it cannot be removed sufficiently by simply washing with water. In particular, sodium sulfate present in the pores of hydroxide is difficult to remove. Therefore, in the present invention, by utilizing water vapor generated by heating the hydroxide and expelling the salt cake from the periphery of the hydroxide and the pores, the sulfur component remains only by ordinary water washing. Solved the problem. Further, when the temperature is higher than 850 ° C., the salt cake is dissolved, and it is difficult to remove the salt cake by subsequent water washing treatment. Therefore, the upper limit value is set to 850 ° C. In addition, by setting the roasting temperature to 400 to 600 ° C., it is possible to shorten the roasting time and to obtain a tertiary concentrate efficiently.
（４）４次濃縮工程
前記３次濃縮物を熔融し、スラグと有価金属とを回収する工程であり、以下の（ｘｉ）熔融処理が含まれる。
（ｘｉ）熔融処理：３次濃縮物に含まれる金属化合物の融点以上の温度で還元剤も併用しつつ、３次濃縮物を熔融処理し有価金属を回収する。
(4) Fourth concentration step This step is a step of melting the tertiary concentrate and recovering slag and valuable metals, and includes the following (xi) melting process.
(Xi) Melting treatment: The tertiary concentrate is melt treated at a temperature equal to or higher than the melting point of the metal compound contained in the tertiary concentrate to recover valuable metals.
本発明の廃電池からの有価金属の回収方法は、上記の工程のうち、特に（３）３次濃縮工程において、低硫黄化処理された状態の３次濃縮物を得る点に特徴を有しており、以下に、３次濃縮工程について詳しく説明する。
３次濃縮工程は、正極活物質の２次濃縮物を水酸化処理して、有価金属の水酸化物を３次濃縮物として得る工程である。前もって、１次濃縮物を硫酸で処理し、硫酸に溶解しにくいＣｕ及び硫酸に溶解しないＣが分離されたろ液（溶解液）が、２次濃縮物として供される。
この２次濃縮物にアルカリ金属の水溶液、例えば、水酸化ナトリウム（ＮａＯＨ）を添加し、有価金属としてコバルトを回収する場合は、水酸化物として水酸化コバルト（Ｃｏ（ＯＨ）2）を沈殿させる。アルカリ金属の水溶液としては、汎用性及び副生物の除去のし易さから、水酸化ナトリウム（ＮａＯＨ）が推奨される。
より具体的には、硫酸による２次濃縮物にＮａＯＨを添加して、ｐＨを１１まで上昇させ、コバルトの水酸化物（Ｃｏ（ＯＨ）2）として回収する。なお、ｐＨは、７．５以上、好ましくは１０から１１である。有価金属は、ｐＨ１１に至る過程で全て水酸化物として沈殿するため、これ以上ｐＨを上げる必要はない。
この処理過程で水酸化物を形成し得るものは全て水酸化物として回収される。
この場合、副生物として硫酸ナトリウム（Ｎａ2ＳＯ4、芒硝）が生成し、生成する前記水酸化物中に芒硝が包含されてしまう。
そこで、芒硝を除去するため、低硫黄化処理として、回収した水酸化物を、１００℃〜８５０℃の範囲、好ましくは４００℃〜６００℃の範囲で、大気雰囲気もしくは酸素雰囲気下で焙焼する酸化焙焼処理（図示省略）を施す。その後、焙焼物を水洗浄し、焙焼物中の芒硝を除去することで、低硫黄化が達成できる。
The method for recovering valuable metals from a waste battery according to the present invention is characterized in that, among the above-mentioned steps, in particular, in the (3) tertiary concentration step, a low-sulfurized tertiary concentrate is obtained. The tertiary concentration step will be described in detail below.
The tertiary concentration step is a step in which the secondary concentrate of the positive electrode active material is subjected to a hydroxylation treatment to obtain a valuable metal hydroxide as a tertiary concentrate. The primary concentrate is treated with sulfuric acid in advance, and a filtrate (dissolved solution) from which Cu that is difficult to dissolve in sulfuric acid and C that does not dissolve in sulfuric acid is separated is provided as a secondary concentrate.
When an aqueous solution of an alkali metal such as sodium hydroxide (NaOH) is added to the secondary concentrate and cobalt is recovered as a valuable metal, cobalt hydroxide (Co (OH) 2 ) is precipitated as a hydroxide. . As the alkali metal aqueous solution, sodium hydroxide (NaOH) is recommended because of its versatility and ease of removal of by-products.
More specifically, NaOH is added to the secondary concentrate with sulfuric acid to raise the pH to 11 and recover as cobalt hydroxide (Co (OH) 2 ). The pH is 7.5 or more, preferably 10 to 11. Since valuable metals are all precipitated as hydroxides in the process of reaching pH 11, it is not necessary to raise the pH any more.
Anything that can form hydroxide during this process is recovered as hydroxide.
In this case, sodium sulfate (Na 2 SO 4 , sodium sulfate) is generated as a by-product, and sodium sulfate is included in the generated hydroxide.
Therefore, in order to remove mirabilite, the recovered hydroxide is roasted in the air atmosphere or oxygen atmosphere in the range of 100 ° C. to 850 ° C., preferably in the range of 400 ° C. to 600 ° C., as a low sulfur treatment. An oxidation roasting process (not shown) is performed. Thereafter, the roasted product is washed with water, and the sodium sulfate in the roasted product is removed, thereby reducing the sulfur content.
また、本発明の有価金属の回収方法の発明は、有価金属の硫酸溶解液からの回収方法と捉えることができる。すなわち、コバルトまたはニッケルの少なくとも１種を含む有価金属の硫酸溶解液からの前記有価金属の回収方法であって、（１）前記硫酸溶解液を水酸化処理する工程、（２）前記水酸化処理後の沈殿物を酸化焙焼処理する工程、（３）水洗処理する工程、を含むことを特徴とする有価金属の回収方法であり、前記（１）〜（３）の各処理工程は、既に述べた処理工程と同じなので、説明を割愛する。 The invention of the method for recovering valuable metals of the present invention can be regarded as a method for recovering valuable metals from a sulfuric acid solution. That is, a method for recovering the valuable metal from a sulfuric acid solution of a valuable metal containing at least one of cobalt or nickel, (1) a step of hydroxylating the sulfuric acid solution, (2) the hydroxylation treatment A method for recovering valuable metals, comprising: a step of subjecting a subsequent precipitate to oxidative roasting; and (3) a step of washing with water. Each of the treatment steps (1) to (3) has already been performed. Since it is the same as the processing process described, explanation is omitted.
本発明の有価金属の回収方法において、予備焙焼処理や、酸化焙焼処理において発生する粉塵や排ガス等は、従来公知の排ガス処理装置を用いて無害化処理される。 In the valuable metal recovery method of the present invention, dust, exhaust gas and the like generated in the preliminary roasting treatment and the oxidation roasting treatment are detoxified using a conventionally known exhaust gas treatment apparatus.
また、本発明の有価金属の回収方法においては、さらに展開した態様（以下、「展開態様（Ｉ）」という。）として、４次濃縮工程の（ｘｉ）熔融処理に至るまでの過程で、分離・回収された電池構成材料（Ｆｅ、Ａｌ、Ｃ）を還元剤として使用する有価金属の回収方法を提案することができる。
すなわち、本発明の展開態様（Ｉ）では、前記４次濃縮工程の（ｘｉ）熔融処理において、３次濃縮物を熔融処理するに際して、還元剤として、本発明の一次濃縮工程の（ｖ）Ｃｕ濃縮処理で得られる、Ｆｅ・Ａｌ殻、及び２次濃縮工程の（ｖiii）Ｃｕ・Ｃの分離処理において分離されたＣを還元剤として投入し、利用することができる（図２参照。）。
In the method for recovering valuable metals according to the present invention, as a further developed mode (hereinafter referred to as “deployed mode (I)”), separation is performed in the process up to (xi) melting treatment in the quaternary concentration step. A method for recovering valuable metals using the recovered battery constituent materials (Fe, Al, C) as a reducing agent can be proposed.
That is, in the development mode (I) of the present invention, in the (xi) melting process of the quaternary concentration process, when the tertiary concentrate is melted, (v) Cu of the primary concentration process of the present invention is used as a reducing agent. The Fe · Al shell obtained by the concentration treatment and the C separated in the separation treatment of (viii) Cu · C in the secondary concentration step can be introduced and used as a reducing agent (see FIG. 2).
以下、図２のフローチャートを用いて説明する。
図２の一次濃縮工程の「（ｖ）Ｃｕ濃縮処理」によって、篩上物Ａに含まれるＣｕに富むＣｕ濃縮物とＦｅ、Ａｌに富むＦｅ・Ａｌ殻とを分離して回収することを目的として、篩上物Ａを４００℃〜６００℃、より好ましくは５００℃〜６００℃で、大気雰囲気もしくは酸素雰囲気下で焙焼する。冷却後、その酸化焙焼物を粉砕する。粉砕には、周知の衝撃、摩擦、せん断などの粉砕手段を単独で又は組み合わせて利用する粉砕装置を適宜使用することができる。例えば、ボールミル等で粉砕し、次いで粉砕物を、篩分けする。篩目の開きは２.００ｍｍ以下、０．６０ｍｍ以上が好ましく、１．１８ｍｍ以下、０．８５ｍｍ以上が特に好ましい。篩目が２．００ｍｍより大きいと、前記Ｃｕ濃縮物と前記Ｆｅ・Ａｌ殻の分離度が低下する。
例えば、篩目の開きは１.００ｍｍの篩で篩分けし、１．００ｍｍ以下のＣｕに富むＣｕ濃縮物を篩下として得る一方、１．００ｍｍよりも大きいＡｌ箔や外装缶等は、Ｆｅ・Ａｌ殻として分離、回収される。
一方、２次濃縮工程の（viii）Ｃｕ・Ｃ分離処理によって、前工程の酸溶解処理物のろ過残渣に含まれるＣｕ及びＣの混合物を硫酸で溶解する。この処理により、Ｃｕは溶解し、Ｃは溶解しないので濾別分離が可能となる。
硫酸を用いる場合は、例えば、Ｃｕ及びＣの混合物に７１％硫酸を加え、ｐＨを０.６以下にしてろ過し、Ｃｕをろ液（溶解液）、Ｃは残渣とする（viii）Ｃｕ・Ｃ分離処理により、Ｃが回収される。
これらの回収されたＦｅ・Ａｌ殻及びＣが、還元剤として４次濃縮工程の（ｘｉ）熔融工程に投入される。
このように、本発明において、展開態様(I)を採用して、有価金属の熔融に至るまでの過程で分離・回収される電池材料（Ｆｅ、Ａｌ、Ｃ）を熔融時の還元剤として使用すれば、系外からの還元剤投入というコスト要因を低減できると共に、電池材料の有効利用が可能となる。
This will be described below with reference to the flowchart of FIG.
The purpose is to separate and recover the Cu-rich Cu concentrate and the Fe / Al-rich Fe / Al shell contained in the sieve top A by “(v) Cu concentration process” in the primary concentration step of FIG. As described above, the sieved product A is roasted at 400 ° C. to 600 ° C., more preferably 500 ° C. to 600 ° C. in an air atmosphere or an oxygen atmosphere. After cooling, the oxidized roast is pulverized. For the pulverization, a pulverization apparatus using a known pulverization means such as impact, friction, shear or the like alone or in combination can be appropriately used. For example, the mixture is pulverized with a ball mill or the like, and then the pulverized product is sieved. The opening of the sieve is preferably 2.00 mm or less and 0.60 mm or more, particularly preferably 1.18 mm or less and 0.85 mm or more. If the sieve mesh is larger than 2.00 mm, the degree of separation between the Cu concentrate and the Fe · Al shell decreases.
For example, the opening of the sieve is sieved with a 1.00 mm sieve to obtain Cu concentrate rich in Cu of 1.00 mm or less as the sieve, while Al foil and outer cans larger than 1.00 mm are Fe -Separated and recovered as Al shell.
On the other hand, the mixture of Cu and C contained in the filtration residue of the acid-dissolved treatment product in the previous step is dissolved with sulfuric acid by (viii) Cu / C separation treatment in the secondary concentration step. By this treatment, Cu is dissolved and C is not dissolved, so that separation by filtration becomes possible.
In the case of using sulfuric acid, for example, 71% sulfuric acid is added to a mixture of Cu and C, and the pH is adjusted to 0.6 or less, and Cu is filtrated (dissolved solution), and C is a residue (viii) Cu · C is recovered by the C separation process.
These recovered Fe · Al shells and C are supplied as a reducing agent to the (xi) melting step of the fourth concentration step.
Thus, in the present invention, the battery material (Fe, Al, C) separated and recovered in the process up to melting of valuable metals is used as the reducing agent at the time of melting by adopting the development mode (I). By doing so, it is possible to reduce the cost factor of introducing the reducing agent from outside the system and to effectively use the battery material.
なお、上記の展開態様（Ｉ）の発明は、以下の背景からなされたものである。
リチウムイオン電池等の廃電池から有価金属を回収する方法が提案されている。しかしながら、高価なレアメタルであるニッケル及びコバルト以外の電池材料を回収し、資源として別途有効利用するという観点では記載されていない。
一方、電池構成材料の開発は、日進月歩であり、今後益々ベースメタルの比重が増してくることが予測される。このような状況において、これからの循環型リサイクル社会を維持・発展させていくためには、ベースメタルを含めた資源も個別に回収し、有効利用するという視点が不可欠であると考えられる。
The invention of the development mode (I) is made from the following background.
A method for recovering valuable metals from waste batteries such as lithium ion batteries has been proposed. However, it is not described in terms of recovering battery materials other than expensive rare metals such as nickel and cobalt and separately effectively using them as resources.
On the other hand, the development of battery constituent materials is progressing steadily, and the specific gravity of base metal is expected to increase more and more in the future. In such a situation, in order to maintain and develop a recycling-based recycling society in the future, it is indispensable to collect and effectively use resources including base metals individually.
本展開態様（Ｉ）の発明は、上述の通り、ＣｏやＮｉといった廃電池由来の有価金属とともに、廃電池からベースメタルを採算性よく回収し有効活用するという課題を解決するためになされたものである。すなわち、熔融に至るまでの過程でＦｅ、Ａｌ、Ｃを回収し、有価金属を回収するための熔融時に還元剤として用いることにより、系外から還元剤を取り込むことなくリサイクルを行うことができる。また、余剰のＦｅ、Ａｌ、Ｃは別途還元剤等の目的で使用可能となる。 As described above, the invention of the present development mode (I) was made to solve the problem of recovering and effectively utilizing base metal from waste batteries together with valuable metals derived from waste batteries such as Co and Ni. It is. That is, by collecting Fe, Al, and C in the process until melting and using them as a reducing agent during melting for recovering valuable metals, recycling can be performed without taking in the reducing agent from outside the system. Further, surplus Fe, Al, and C can be separately used for the purpose of a reducing agent or the like.
さらに、本発明の有価金属の回収方法は、さらに展開された態様として、銅（Ｃｕ）の回収方法（以下、「展開態様（II）」という。）を提案することができる。
すなわち、リチウムイオン電池等の廃電池からＣｕ濃縮物を簡単に回収する方法を提供する。リチウムイオン電池等の廃電池からコバルト等の有価金属を回収する際、「粉砕処理」、「篩分け」を行う。この処理工程により、コバルト等の有価金属を多く含む活物質は篩下として分離、回収される。一方、Ｆｅ、Ａｌ、Ｃｕが多く含まれる外装缶や極板等は篩上として分離、回収される。
Furthermore, the method for recovering valuable metals of the present invention can propose a method for recovering copper (Cu) (hereinafter referred to as “development mode (II)”) as a further developed mode.
That is, a method for easily recovering a Cu concentrate from a waste battery such as a lithium ion battery is provided. When recovering valuable metals such as cobalt from waste batteries such as lithium ion batteries, "grinding" and "sieving" are performed. By this processing step, an active material containing a large amount of valuable metals such as cobalt is separated and recovered as a sieve. On the other hand, outer cans, electrode plates, and the like containing a large amount of Fe, Al, and Cu are separated and collected as a sieve.
本発明の展開態様(II)は、その篩上品を、「焙焼処理」「粉砕処理」「篩分け」することで、主にＣｕが粉状となり、Ｃｕ濃縮物を篩下成分として回収できることを見出し完成に至った。
すなわち、本発明の展開態様(II)は、廃電池からのＣｕ濃縮物回収方法であって、前記廃電池の予備焙焼処理と、粉砕処理と、篩分けにより１次濃縮物と篩上物に分離する工程と、前記篩上物の酸化焙焼と、粉砕処理と、篩分けによりＣｕ濃縮物とそれ以外のベースメタルに分離する工程とを備え、前記Ｃｕ濃縮物を粉状で回収することを特徴とするＣｕ濃縮物回収方法、を提供できる。
The development mode (II) of the present invention is that the sieved product is “roasted”, “pulverized”, and “sieved”, so that Cu is mainly powdered and the Cu concentrate can be recovered as an under-sieving component. The headline was completed.
That is, the development mode (II) of the present invention is a method for recovering Cu concentrate from a waste battery, wherein the primary concentrate and sieved material are obtained by preliminary roasting treatment, pulverization treatment and sieving of the waste battery. And the step of separating the copper concentrate and the other base metal by sieving, and collecting the Cu concentrate in powder form. A Cu concentrate recovery method characterized by the above can be provided.
以下、本展開態様(II)について図２を用いて説明する。図２の一次濃縮工程において、（ii）予備焙焼処理を経た焙焼物を、(iii)粉砕処理Ａで、予備焙焼処理により非常に粉砕され易くなっているコバルトに富む活物質と、Ａｌ箔、Ｃｕ箔、外装缶（Ｆｅ・Ａｌ）等のコバルトが少なく、かつ比較的粉砕され難い部分とを篩分けＡで分離し易くするために粉砕される。
粉砕は破砕機を含む破砕設備を用いて行うことができ、破砕機については特に制限はなく、いかなるものも使用できる。
（ｉｖ）篩分けＡ処理で、粉砕処理Ａで得られた粉砕物を篩分けする。篩目の開きは２.００ｍｍ以下、０．６０ｍｍ以上が好ましく、１．１８ｍｍ以下、０．８５ｍｍ以上が特に好ましい。篩目が２．００ｍｍより大きいと、篩下への有価金属以外のベースメタルの混入が増える。一方、篩目が０．６０ｍｍより小さいと、篩上に有価金属の混入が増えるため好ましくない。例えば篩目開き１.００ｍｍの篩を用いて篩分けし、１.００ｍｍ以下のコバルトに富む１次濃縮物を篩下として得る。１．００ｍｍより大きいＡｌ箔やＣｕ箔、外装缶（Ｆｅ・Ａｌ）等は篩上物Ａとして分離・回収される。
篩上物Ａとして分離・回収されたＡｌ箔やＣｕ箔、外装缶（Ｆｅ・Ａｌ）等は、篩上物Ａに含まれるＣｕに富むＣｕ濃縮物と、Ｆｅ、Ａｌに富むＦｅ・Ａｌ殻とを分離して回収することを目的として行われる（ｖ）Ｃｕ濃縮処理に供される。
該（ｖ）Ｃｕ濃縮処理は、篩上物Ａを４００℃〜６００℃の範囲、好ましくは５００℃〜６００℃の範囲で、大気雰囲気もしくは酸素雰囲気下で焙焼する。
冷却後、その酸化焙焼物を粉砕する。粉砕には、周知の衝撃、摩擦、せん断などの粉砕手段を単独で又は組み合わせて利用する粉砕装置を適宜使用することができる。例えば、ボールミル等で粉砕し、次いで粉砕物を、篩分けする。篩目の開きは２.００ｍｍ以下、０．６０ｍｍ以上が好ましく、１．１８ｍｍ以下、０．８５ｍｍ以上が特に好ましい。篩目が２．００ｍｍより大きいと、不純物（Ａｌ等）の混入率が増える。
より具体的には、例えば篩目開き１．００ｍｍの篩を用いて篩分けし、１．００ｍｍ以下のＣｕに富むＣｕ濃縮物を篩下として回収する。一方、１．００ｍｍよりも大きいＡｌ箔や外装缶等は、Ｆｅ・Ａｌ殻として分離、回収される。
本発明の展開態様(II)の発明では、粉砕処理Ａによる篩上品を、「焙焼処理」、「粉砕処理」、「篩分け」することで、硬く脆い性状の酸化銅成分のみが粉状となり、Ｃｕ濃縮物を篩下成分として回収することができる。
Hereinafter, this development mode (II) will be described with reference to FIG. In the primary concentration step of FIG. 2, (ii) the roasted product that has undergone the preliminary roasting treatment is converted into (iii) an active material rich in cobalt that is easily pulverized by the preliminary roasting treatment in pulverization treatment A, and Al It is pulverized in order to make it easy to separate it by sieving A from a portion of cobalt, such as foil, Cu foil, and outer can (Fe · Al), which is low in cobalt and relatively difficult to pulverize.
The pulverization can be performed using a crushing facility including a crusher, and the crusher is not particularly limited, and any one can be used.
(Iv) In the sieving A process, the pulverized product obtained in the pulverizing process A is sieved. The opening of the sieve is preferably 2.00 mm or less and 0.60 mm or more, particularly preferably 1.18 mm or less and 0.85 mm or more. If the mesh size is larger than 2.00 mm, the base metal other than the valuable metal is mixed under the sieve. On the other hand, if the sieve mesh is smaller than 0.60 mm, it is not preferable because valuable metals are mixed on the sieve. For example, sieving is performed using a sieve having a sieve opening of 1.00 mm, and a primary concentrate rich in cobalt of 1.00 mm or less is obtained as a sieve. Al foil, Cu foil, outer can (Fe · Al), etc. larger than 1.00 mm are separated and collected as sieved product A.
Al foil, Cu foil, outer can (Fe · Al), etc. separated and recovered as sieve top A are Cu enriched Cu contained in sieve top A and Fe / Al shell rich in Fe and Al. And (v) Cu concentration treatment performed for the purpose of separating and recovering.
In the (v) Cu concentration treatment, the sieved product A is roasted in the range of 400 ° C. to 600 ° C., preferably in the range of 500 ° C. to 600 ° C., in an air atmosphere or an oxygen atmosphere.
After cooling, the oxidized roast is pulverized. For the pulverization, a pulverization apparatus using a known pulverization means such as impact, friction, shear or the like alone or in combination can be appropriately used. For example, the mixture is pulverized with a ball mill or the like, and then the pulverized product is sieved. The opening of the sieve is preferably 2.00 mm or less and 0.60 mm or more, particularly preferably 1.18 mm or less and 0.85 mm or more. If the sieve mesh is larger than 2.00 mm, the mixing ratio of impurities (Al, etc.) increases.
More specifically, for example, sieving is performed using a sieve having a sieve opening of 1.00 mm, and a Cu-rich Cu concentrate having a size of 1.00 mm or less is recovered as a sieve. On the other hand, Al foil and outer cans larger than 1.00 mm are separated and collected as Fe · Al shells.
In the invention of the development mode (II) of the present invention, only the hard and brittle copper oxide component is powdered by subjecting the sieved product by the grinding treatment A to “roasting treatment”, “grinding treatment”, and “sieving”. Thus, the Cu concentrate can be recovered as a sieving component.
なお、本発明の展開態様(II)の発明は、以下の背景からなされたものである。
リチウムイオン電池等の廃電池から有価金属を回収する方法は従来から提案されている。
その方法は、湿式法、乾式法、乾式法と湿式法の組み合わせ等が提案されている。これらの既存技術の主目的はレアメタル等の有価金属の回収である。そのため、これらの処理工程には、電池構成材料であるべースメタルを除去する工程が導入されている。
例えば、乾式法で回収した表面積が所定の範囲の粒状（ショット状）の有価金属を酸溶解し、電解法を用いてＣｕを除去する方法が提案されている。
しかしながら、電解法を用いたＣｕの除去方法は、高度な設備がないところでは実施困難である。また、当該技術は、コバルト等のレアメタルの回収を前提とした処理方法であり、今後、ベースメタルが電池構成材料の主流となった際には、採算性の問題から、当該技術を用いてＣｕを回収することは困難となることが予想される。
The invention of the development mode (II) of the present invention has been made from the following background.
Methods for recovering valuable metals from waste batteries such as lithium ion batteries have been proposed.
As the method, a wet method, a dry method, a combination of a dry method and a wet method, and the like have been proposed. The main purpose of these existing technologies is to collect valuable metals such as rare metals. Therefore, a process for removing base metal, which is a battery constituent material, is introduced into these treatment processes.
For example, there has been proposed a method in which a granular (shot-like) valuable metal having a surface area recovered by a dry method is acid-dissolved and Cu is removed using an electrolytic method.
However, the Cu removal method using the electrolytic method is difficult to implement in the absence of sophisticated equipment. In addition, this technology is a treatment method based on the premise of recovery of rare metals such as cobalt. When the base metal becomes the mainstream of battery constituent materials in the future, due to the problem of profitability, Cu It is expected that it will be difficult to recover.
上記の本展開態様(II)の発明は、上記課題を解決するためになされたものであり、リチウムイオン電池等の廃電池から簡単な処理でＣｕを回収する方法を提供することを目的とする。
そこで、本発明者らは、既に説明したように、（ｉｖ）篩分けＡにおいて篩下として廃電池中の正極活物質を1次濃縮物として分離する一方、篩上として回収された電池材料を酸化焙焼し、粉砕することで、Ｃｕが粉状となることを見出し、本展開態様(II)の発明を完成するに至った。
本展開態様(II)の発明によれば、高度な設備を要することなく、低コストでＣｕ濃縮物を回収することができる。
The invention of the present development mode (II) has been made in order to solve the above problems, and an object thereof is to provide a method for recovering Cu from a waste battery such as a lithium ion battery by a simple treatment. .
Therefore, as described above, the present inventors separated the positive electrode active material in the waste battery as a primary concentrate as a sieve under (iv) sieving A, while the battery material recovered as on the sieve was separated. By oxidative roasting and pulverization, it was found that Cu became powdery, and the invention of the present development mode (II) was completed.
According to the invention of the present development mode (II), Cu concentrate can be recovered at low cost without requiring sophisticated equipment.
以下、実施例及び比較例により本発明を更に具体的に説明するが、本発明は、以下の実施例に限定されるものではない。
実施例１
まず、実施例について説明する。実施例においては、図１に示すように、３次濃縮工程に先行して、1次濃縮工程及び２次濃縮工程を行い、３次濃縮工程後は、４次濃縮工程を行った。
EXAMPLES Hereinafter, although an Example and a comparative example demonstrate this invention further more concretely, this invention is not limited to a following example.
Example 1
First, examples will be described. In the Examples, as shown in FIG. 1, the primary concentration step and the secondary concentration step were performed prior to the tertiary concentration step, and the fourth concentration step was performed after the tertiary concentration step.
１次濃縮物の回収
コバルトの回収を主目的として、主にＰＣ等で使用されていた円筒形の廃リチウムイオン電池を約１０ｋｇ使用し、１次濃縮工程で、図１に示すように、開孔処理、予備焙焼処理、粉砕処理Ａ、篩分けＡを施し、１次濃縮物を回収した。
まず、開孔処理は、針状刃先を有する冶具を用いて行い、予備焙焼処理は、開放タイプの回転式電気加熱炉にて、４００℃で５時間加熱して行った。予備焙焼後の電池の減量率は２５．４重量％であった。
次いで、予備焙焼した電池全量を、バッチ式のシュレッダーで粉砕した。
引き続き、粉砕品を振動篩（目開き：１.００ｍｍ）にて篩分け（篩分けＡ）し、篩下品としてコバルト等の有価金属を多く含む１次濃縮物を回収した。一方、篩上物Ａには、Ｆｅ、Ａｌ、Ｃｕが多く含まれていた（１次濃縮物：篩上物Ａ＝６１重量％：３９重量％）。
Recovery of primary concentrate Mainly for the recovery of cobalt, approximately 10 kg of cylindrical waste lithium ion batteries used mainly in PCs, etc. were used, and the primary concentration process was performed as shown in FIG. Hole treatment, preliminary roasting treatment, pulverization treatment A, and sieving A were performed, and the primary concentrate was recovered.
First, the hole-opening process was performed using a jig having a needle blade, and the preliminary roasting process was performed by heating at 400 ° C. for 5 hours in an open type rotary electric heating furnace. The weight loss rate of the battery after preliminary roasting was 25.4% by weight.
Next, the entire amount of the pre-baked battery was pulverized with a batch type shredder.
Subsequently, the pulverized product was sieved with a vibrating sieve (aperture: 1.00 mm) (sieving A), and a primary concentrate containing a large amount of valuable metals such as cobalt was recovered as a sieved product. On the other hand, the sieve top A contained a large amount of Fe, Al, and Cu (primary concentrate: sieve top A = 61 wt%: 39 wt%).
２次濃縮物
上記１次濃縮物のうち１１８．２ｇを分取し、その1次濃縮物に水５９１ｍｌを加え撹拌した。さらに、その溶液にヒーターを挿入・加温し、８０℃に保ちつつ、７１％硫酸１１７ｍｌを添加して、ｐＨを１．０に調整し、酸に溶解しないＣｕ及びＣをろ過によりろ過残渣として回収した。
上記除銅処理におけるろ液６２５ｍｌを２次濃縮物として３次濃縮工程へ供した。
Secondary Concentrate 118.2 g of the primary concentrate was fractionated, and 591 ml of water was added to the primary concentrate and stirred. Furthermore, a heater is inserted into the solution and heated, and while maintaining the temperature at 80 ° C., 117 ml of 71% sulfuric acid is added to adjust the pH to 1.0, and Cu and C that do not dissolve in the acid are filtered as filtration residues. It was collected.
625 ml of the filtrate in the above copper removal treatment was subjected to the tertiary concentration step as a secondary concentrate.
３次濃縮工程
次いで、３次濃縮工程において、水酸化処理として、２次濃縮物６２５ｍｌに、２０％ＮａＯＨを６３５ｍｌ添加しｐＨを１１に調整し、水酸化物（沈殿物：Ｃｏ（ＯＨ）2）を作製した。
その水酸化物をろ別し、６００℃、６時間の条件で酸化焙焼した。
酸化焙焼後の回収物の重量は１４６．３ｇであった。その回収物を約５００ｍｌのイオン交換水で３０分間、攪拌洗浄した（同様の操作を２回実施）。
洗浄・ろ過後、回収物を１００℃、１２時間の条件で乾燥処理を行い、３次濃縮物として５８．１ｇ回収した。表１にＩＣＰ（高周波誘導結合プラズマ）法で分析した結果を示す。なお、一連の操作を２回実施したので、２回目の結果も併記する。
ＩＣＰ（高周波誘導結合プラズマ）法による測定は、セイコーインスツル（株）製のＳＰＳ３５００（機器名）を用い、試料毎に積分時間２．０秒、積分回数３回にて行った。
Third concentration step Next, in the third concentration step, as a hydroxylation treatment, 635 ml of 20% NaOH was added to 625 ml of the secondary concentrate to adjust the pH to 11, and hydroxide (precipitate: Co (OH) 2 ) Was produced.
The hydroxide was filtered off and baked by oxidation under conditions of 600 ° C. and 6 hours.
The weight of the recovered product after oxidation roasting was 146.3 g. The collected product was stirred and washed with about 500 ml of ion-exchanged water for 30 minutes (the same operation was performed twice).
After washing and filtration, the recovered product was dried at 100 ° C. for 12 hours to recover 58.1 g as a tertiary concentrate. Table 1 shows the results of analysis by ICP (High Frequency Inductively Coupled Plasma) method. Since the series of operations was performed twice, the second result is also shown.
The measurement by the ICP (high frequency inductively coupled plasma) method was performed using an SPS3500 (equipment name) manufactured by Seiko Instruments Inc. with an integration time of 2.0 seconds and an integration count of 3 times for each sample.
４次濃縮工程
４次濃縮工程では、３次濃縮物５８．１ｇに、熔融助剤及び還元剤として高炉水砕スラグ２７．５ｇ、二酸化ケイ素（ＳｉＯ2）１．５ｇ、コークス８．４ｇを添加し混合したものをカーボンルツボに充填し、アルゴン雰囲気の高周波溶解炉にて、１５００℃、３０分間保持し、熔融処理を行った。熔融処理後、炉冷し、冷却後に金属とスラグを分離回収して、金属をＩＣＰ法により分析した。表２に金属の分析結果を示す。
Fourth concentration step In the fourth concentration step, 27.5 g of granulated blast furnace slag, 1.5 g of silicon dioxide (SiO 2 ), and 8.4 g of coke are added to 58.1 g of the third concentrate as a melting aid and a reducing agent. Then, the mixture was filled in a carbon crucible and held at 1500 ° C. for 30 minutes in a high-frequency melting furnace in an argon atmosphere to perform a melting treatment. After the melting treatment, the furnace was cooled, and after cooling, the metal and slag were separated and recovered, and the metal was analyzed by the ICP method. Table 2 shows the metal analysis results.
比較例１
次に比較例について説明する。比較例においては、実施例と異なる点として、３次濃縮工程において低硫黄化処理としての酸化焙焼処理を行わなかった。
すなわち、比較例では、実施例と同様な工程により得られた１次濃縮物のうち３５．７ｇを分取し、水１７８ｍｌ、７１％硫酸３０ｍｌ、ｐＨ１．０で処理し、ろ液３４４ｍｌを２次濃縮物として回収した。
次いで、３次濃縮工程において、水酸化処理として、２次濃縮物３４４ｍｌに、１０％ＮａＯＨを３４４ｍｌ添加し、水酸化物（沈殿物：Ｃｏ（ＯＨ）2）を作製した。
該水酸化物をろ別し、ろ別後の回収物を約２００ｍｌのイオン交換水で３０分間、攪拌洗浄した（同様の操作を４回実施）。
前記攪拌洗浄後、ろ過により得られた回収物を１００℃、１２時間の条件で乾燥処理を行い、３次濃縮物として３６．０ｇ回収した。表３にＩＣＰ（高周波誘導結合プラズマ）法で分析した結果を示す。
ＩＣＰ（高周波誘導結合プラズマ）法による測定は、セイコーインスツル（株）製のＳＰＳ３５００（機器名）を用い、試料毎に積分時間２．０秒、積分回数３回にて行った。
Comparative Example 1
Next, a comparative example will be described. In the comparative example, as a different point from the example, the oxidation roasting treatment as the low sulfur treatment was not performed in the tertiary concentration step.
That is, in the comparative example, 35.7 g of the primary concentrate obtained by the same process as in the example was collected and treated with 178 ml of water, 30 ml of 71% sulfuric acid, pH 1.0, and 344 ml of the filtrate was added to 2 parts. It was recovered as the next concentrate.
Next, in the third concentration step, as a hydroxylation treatment, 344 ml of 10% NaOH was added to 344 ml of the second concentrate to produce a hydroxide (precipitate: Co (OH) 2 ).
The hydroxide was filtered off, and the recovered product after the filtration was stirred and washed with about 200 ml of ion-exchanged water for 30 minutes (the same operation was performed 4 times).
After stirring and washing, the recovered product obtained by filtration was dried at 100 ° C. for 12 hours to recover 36.0 g as a tertiary concentrate. Table 3 shows the results of analysis by ICP (High Frequency Inductively Coupled Plasma) method.
The measurement by the ICP (high frequency inductively coupled plasma) method was performed using an SPS3500 (equipment name) manufactured by Seiko Instruments Inc. with an integration time of 2.0 seconds and an integration count of 3 times for each sample.
次いで、回収された３次濃縮物成分について、実施例と同様にして４次濃縮工程の熔融処理を行い、金属とスラグを分離回収した。得られた金属をＩＣＰ法により分析した結果を表４に示す。 Next, the recovered third concentrate component was subjected to a melting treatment in the fourth concentration step in the same manner as in Example, and the metal and slag were separated and recovered. Table 4 shows the results of analyzing the obtained metal by the ICP method.
表１と表３の比較から解かるように、酸化焙焼処理を施さない比較例における３次濃縮物中のＳＯ4濃度の７．２７％を、実施例において酸化焙焼処理を水洗処理前に施すことによって、２．２６％に低減させることができることを確認できた。
また、表２と表４から、３次濃縮物中のＳＯ4濃度を低減させることで、４次濃縮工程の熔融処理において、表４に示す比較例の金属中に含まれるＳ濃度２．３５％が、表２に示すように実施例では０．１４％まで低減させることが可能となり、不純物の少ない金属の回収が可能となった。
As can be seen from the comparison between Table 1 and Table 3, 7.27% of the SO 4 concentration in the tertiary concentrate in the comparative example in which the oxidation roasting treatment is not performed, and the oxidation roasting treatment in the examples before the water washing treatment It was confirmed that it can be reduced to 2.26%.
Further, from Table 2 and Table 4, by reducing the SO 4 concentration in the tertiary concentrate, the S concentration 2.35 contained in the metal of the comparative example shown in Table 4 in the melting treatment in the fourth concentration step. % Can be reduced to 0.14% in the examples as shown in Table 2, and metal with less impurities can be recovered.
以下に本発明の展開態様（Ｉ）の発明及び展開態様(II)の発明について、それぞれ参考例により具体的に説明するが、本発明の展開態様（Ｉ）の発明及び展開態様(II)の発明は、以下の参考例に限定されるものではない。 The invention of the development aspect (I) and the invention of the development aspect (II) of the present invention will be specifically described below with reference examples. The invention of the development aspect (I) and the development aspect (II) of the present invention The invention is not limited to the following reference examples.
参考例１
実施例において分取した１次濃縮物１１８．２ｇ中のコバルト（Ｃｏ）及びＣの含有量は、それぞれ３８．９ｇ、３６．４ｇであった。このうち、Ｃは（viii）Ｃｕ・Ｃ分離処理にて回収することができる。また、前記１次濃縮物１１８．２ｇに相当する篩上物Ａの量は７６．５ｇであった。この篩上物Ａ７６．５ｇ中のＦｅ及びＡｌ含有量はそれぞれ４６．１ｇ、３．６ｇであった。なお、篩上物Ａに含まれるＣｕは、（ｖ）Ｃｕ濃縮処理により除去することができる。
１次濃縮物中のＣｏ３８．９ｇを有価金属Ｃｏとして回収する際に必要な還元剤の量は、
（１）還元剤Ｃの必要量は、式（II）の場合で８．０ｇ：
Ｃ＋１/２Ｏ2 → ＣＯ 式（Ｉ）
ＣｏＯ＋ＣＯ → Ｃｏ＋ＣＯ2 式（II）
ＣｏＯ+１/２Ｃ → Ｃｏ+１/２ＣＯ2 式（III）
（２）還元剤Ｆｅの必要量は式（IV）より３６．９ｇ：
ＣｏＯ＋Ｆｅ → Ｃｏ+ＦｅＯ 式（IＶ）
（３）還元剤Ａlの必要量は式（Ｖ）より１１．９ｇ:
ＣｏＯ+２／３Ａｌ → Ｃｏ+１／３Ａｌ2Ｏ3 式（Ｖ）
である。
上記の結果より、廃電池中のＦｅ、Ｃは、４次濃縮工程の熔融処理の還元剤として、化学量論的においても十分量存在することが確認でき、Ａｌにおいては、単独では必要還元剤量に満たないが、還元剤の一部として使用できることが確認できた。
Reference example 1
The contents of cobalt (Co) and C in 118.2 g of the primary concentrate fractionated in the examples were 38.9 g and 36.4 g, respectively. Among these, C can be recovered by (viii) Cu · C separation treatment. Moreover, the amount of the sieve top A corresponding to 118.2 g of the primary concentrate was 76.5 g. The Fe and Al contents in 76.5 g of the above sieved product A were 46.1 g and 3.6 g, respectively. In addition, Cu contained in the sieve top A can be removed by (v) Cu concentration treatment.
The amount of reducing agent required to recover 38.9 g of Co in the primary concentrate as valuable metal Co is:
(1) The required amount of reducing agent C is 8.0 g in the case of formula (II):
C + 1 / 2O 2 → CO Formula (I)
CoO + CO → Co + CO 2 formula (II)
CoO + 1 / 2C → Co + 1/2 CO 2 formula (III)
(2) The required amount of reducing agent Fe is 36.9 g from formula (IV):
CoO + Fe → Co + FeO Formula (IV)
(3) The required amount of reducing agent Al is 11.9 g from formula (V):
CoO + 2 / 3Al → Co + 1 / 3Al 2 O 3 formula (V)
It is.
From the above results, it can be confirmed that Fe and C in the waste battery are present in a sufficient amount stoichiometrically as a reducing agent for the melting treatment in the quaternary concentration step. Although it was less than the amount, it was confirmed that it could be used as a part of the reducing agent.
以下、展開態様（II）の発明について参考例２により更に具体的に説明するが、本展開態様（II）の発明は、以下の参考例２に限定されるものではない。
参考例２
参考例２では実施例と同様な工程により得られた篩上物Ａ７７．９ｇに含まれるＦｅ、Ａｌ、Ｃｕ量を測定した（表５参照。）。
Hereinafter, the invention of the development mode (II) will be described more specifically with reference example 2. However, the invention of the development mode (II) is not limited to the following reference example 2.
Reference example 2
In Reference Example 2, the amounts of Fe, Al, and Cu contained in 77.9 g of sieved product A obtained by the same process as in Example were measured (see Table 5).
（ｖ）Ｃｕ濃縮処理では、篩上物Ａを、６００℃、５時間の条件で酸化焙焼(図示省略)した。
その後、２５φアルミナボールを２０個入れたポットに酸化焙焼後の篩上物Ａを入れ、３０分間、ボールミルによる粉砕処理を行った。粉砕処理後、再度、篩目開き１．００ｍｍの篩を用いて篩分け（図示省略）を行い、篩下のＣｕ濃縮物と、篩上物のＦｅ・Ａｌ殻に分離した。
Ｃｕ濃縮物、Ｆｅ・Ａｌ殻それぞれに含まれるＦｅ、Ａｌ、Ｃｕ量を表６に示す。
(V) In the Cu concentration treatment, the sieved product A was oxidized and roasted (not shown) at 600 ° C. for 5 hours.
Thereafter, the sieved product A after oxidative roasting was placed in a pot containing 20 25φ alumina balls, and pulverized by a ball mill for 30 minutes. After the pulverization treatment, sieving (not shown) was performed again using a sieve having a sieve opening of 1.00 mm, and separated into a Cu concentrate under the sieve and an Fe · Al shell of the sieve top.
Table 6 shows the amounts of Fe, Al, and Cu contained in the Cu concentrate and the Fe / Al shell, respectively.
成分分析は全てＩＣＰ法を用いた。
表５、６から解るように、篩上物ＡのＣｕの量１５．５ｇに対して、Ｃｕ濃縮処理を施すことで１２．１ｇを回収することができ、Ｃｕの回収率を７８．１％〔（１２．１÷１５．５）×１００）〕とすることが可能となった。
なお、Ｃｕ濃縮物中の「その他」は、外装缶（Ｆｅ、Ａｌ）に咬み混まれた態様で存在したＣｏが、（ｖ）Ｃｕ濃縮処理における酸化焙焼処理、粉砕処理により粉体となり、前記Ｃｕ濃縮物中に取り込まれたものと想定される。よって、前記Ｃｕ濃縮物に対し、（ｖi）酸溶解処理を施すことにより、純度の高いＣｕを残渣として回収することができる。また、ろ液に対し、（iｘ）水酸化処理以降の工程を施すことにより、Ｃｏを回収することができる。
前記Ｃｕ濃縮物は、処理が容易な乾式法によりＦｅ、Ａｌ等が少なく、かつＣｕ濃度が高い濃縮物とすることができる。また、前記Ｃｕ濃縮物は粉体であるので、その後の湿式処理時間が短く、産業利用可能性が高いものとなっている。
All component analyzes used the ICP method.
As can be seen from Tables 5 and 6, 12.1 g can be recovered by applying Cu concentration to 15.5 g of Cu in sieved product A, and the Cu recovery rate is 78.1%. [(12.1 ÷ 15.5) × 100)].
In addition, “others” in the Cu concentrate is Co present in the form bitten by the outer can (Fe, Al), and becomes (v) a powder by oxidation roasting treatment and pulverization treatment in the Cu concentration treatment, It is assumed that it was taken up in the Cu concentrate. Therefore, high purity Cu can be recovered as a residue by subjecting the Cu concentrate to (vi) acid dissolution treatment. Further, Co can be recovered by subjecting the filtrate to a process after (ix) hydroxylation.
The Cu concentrate can be made into a concentrate with low Fe, Al, etc. and high Cu concentration by a dry process that is easy to process. Moreover, since the said Cu concentrate is a powder, the subsequent wet processing time is short and the industrial applicability is high.
本発明の廃電池からの有価金属の回収方法は、３次濃縮工程における低硫黄化処理により、４次濃縮工程で回収する有価金属への硫黄分の混入を低減させ、有価金属の純度価値を向上させるとともに、硫黄由来の低融点組成スラグによる不定形耐火材等の損傷を防ぐことができる方法として有効に利用できる。
また、本発明の展開態様(I)の発明は、熔融処理工程の還元剤を回収系から供給することによって、回収コストを低減できる有価物の回収方法として利用できる。
さらに、本発明の展開態様(II)の発明に基づけば、処理が容易な乾式法によりＣｕ濃縮物を得ることができる。また、酸溶解にてＣｕを回収する際、Ｃｕ濃縮物は篩上物Ａに比べ、細かく粉砕されているため、Ｃｕ以外の不純物の酸溶解処理時間が短く、Ｃｕを短時間で回収する方法として利用できる。
The method for recovering valuable metals from waste batteries according to the present invention reduces the contamination of valuable metals recovered in the quaternary concentration step by reducing sulfur in the tertiary concentration step, thereby reducing the purity value of the valuable metals. While improving, it can utilize effectively as a method which can prevent damage to the amorphous refractory material etc. by the low melting-point composition slag derived from sulfur.
Further, the invention of the development mode (I) of the present invention can be used as a valuable material recovery method capable of reducing the recovery cost by supplying the reducing agent of the melting treatment step from the recovery system.
Furthermore, based on the invention of the development mode (II) of the present invention, a Cu concentrate can be obtained by a dry process that is easy to process. Further, when Cu is recovered by acid dissolution, since the Cu concentrate is finely pulverized compared to the sieve top A, a method for recovering Cu in a short time since the acid dissolution treatment time of impurities other than Cu is short. Available as
Claims (6)
（１）予備焙焼処理、粉砕処理及び篩分け処理を経て、前記有価金属の１次濃縮物を得る１次濃縮工程、
（２）前記１次濃縮物を硫酸で溶解処理し、該溶解液を２次濃縮物として得る２次濃縮工程、
（３）前記２次濃縮物を水酸化処理後、低硫黄化処理し、前記有価金属の３次濃縮物を得る３次濃縮工程、及び
（４）前記３次濃縮物を熔融し、前記有価金属を回収する４次濃縮工程、
を含むことを特徴とする有価金属の回収方法。 A method for recovering the valuable metal from waste batteries or process waste containing at least one valuable metal of cobalt or nickel,
(1) A primary concentration step for obtaining a primary concentrate of the valuable metal through preliminary roasting treatment, pulverization treatment and sieving treatment,
(2) a secondary concentration step of dissolving the primary concentrate with sulfuric acid to obtain the solution as a secondary concentrate;
(3) A tertiary concentration step for obtaining a tertiary concentrate of the valuable metal by subjecting the secondary concentrate to a low sulfur treatment after hydroxylation, and (4) melting the tertiary concentrate, A fourth concentration step for recovering the metal;
A method for recovering valuable metals, comprising:
（１）前記硫酸溶解液を水酸化処理する工程、
（２）前記水酸化処理後の沈殿物を酸化焙焼処理する工程、
（３）水洗処理する工程、
を含むことを特徴とする有価金属の回収方法。 A method for recovering the valuable metal from a sulfuric acid solution of a valuable metal containing at least one of cobalt and nickel,
(1) A step of hydroxylating the sulfuric acid solution,
(2) a step of oxidizing and roasting the precipitate after the hydroxylation treatment,
(3) A step of washing with water,
A method for recovering valuable metals, comprising:
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