JP2008088470A - Method for recovering indium from indium-containing material - Google Patents
Method for recovering indium from indium-containing material Download PDFInfo
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- 229910052738 indium Inorganic materials 0.000 title claims abstract description 90
- APFVFJFRJDLVQX-UHFFFAOYSA-N indium atom Chemical compound [In] APFVFJFRJDLVQX-UHFFFAOYSA-N 0.000 title claims abstract description 86
- 238000000034 method Methods 0.000 title claims abstract description 53
- 239000000463 material Substances 0.000 title claims abstract description 18
- QAOWNCQODCNURD-UHFFFAOYSA-N Sulfuric acid Chemical compound OS(O)(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-N 0.000 claims abstract description 76
- 238000002386 leaching Methods 0.000 claims abstract description 66
- 239000002184 metal Substances 0.000 claims abstract description 22
- 229910052751 metal Inorganic materials 0.000 claims abstract description 22
- 230000033116 oxidation-reduction process Effects 0.000 claims description 9
- 238000005868 electrolysis reaction Methods 0.000 claims description 4
- 238000005987 sulfurization reaction Methods 0.000 claims description 4
- 230000002829 reductive effect Effects 0.000 claims description 2
- 239000012535 impurity Substances 0.000 abstract description 16
- 150000002739 metals Chemical class 0.000 abstract description 4
- 238000001914 filtration Methods 0.000 description 41
- 239000000706 filtrate Substances 0.000 description 39
- 238000006243 chemical reaction Methods 0.000 description 38
- 239000000243 solution Substances 0.000 description 20
- 239000002002 slurry Substances 0.000 description 17
- 229910052725 zinc Inorganic materials 0.000 description 13
- 239000011701 zinc Substances 0.000 description 13
- VEXZGXHMUGYJMC-UHFFFAOYSA-N Hydrochloric acid Chemical compound Cl VEXZGXHMUGYJMC-UHFFFAOYSA-N 0.000 description 12
- HEMHJVSKTPXQMS-UHFFFAOYSA-M Sodium hydroxide Chemical compound [OH-].[Na+] HEMHJVSKTPXQMS-UHFFFAOYSA-M 0.000 description 9
- HCHKCACWOHOZIP-UHFFFAOYSA-N Zinc Chemical compound [Zn] HCHKCACWOHOZIP-UHFFFAOYSA-N 0.000 description 9
- 239000010949 copper Substances 0.000 description 9
- 229910052602 gypsum Inorganic materials 0.000 description 9
- 239000010440 gypsum Substances 0.000 description 9
- 238000003723 Smelting Methods 0.000 description 8
- 238000004458 analytical method Methods 0.000 description 8
- 239000003795 chemical substances by application Substances 0.000 description 8
- 229910052802 copper Inorganic materials 0.000 description 8
- 238000011084 recovery Methods 0.000 description 8
- 229910052785 arsenic Inorganic materials 0.000 description 7
- 239000007788 liquid Substances 0.000 description 7
- 229910052782 aluminium Inorganic materials 0.000 description 6
- 150000004763 sulfides Chemical class 0.000 description 6
- RYGMFSIKBFXOCR-UHFFFAOYSA-N Copper Chemical group [Cu] RYGMFSIKBFXOCR-UHFFFAOYSA-N 0.000 description 5
- 229910052793 cadmium Inorganic materials 0.000 description 5
- XLYOFNOQVPJJNP-UHFFFAOYSA-M hydroxide Chemical compound [OH-] XLYOFNOQVPJJNP-UHFFFAOYSA-M 0.000 description 5
- 239000000047 product Substances 0.000 description 5
- 238000000746 purification Methods 0.000 description 5
- 238000000926 separation method Methods 0.000 description 5
- UCKMPCXJQFINFW-UHFFFAOYSA-N Sulphide Chemical compound [S-2] UCKMPCXJQFINFW-UHFFFAOYSA-N 0.000 description 4
- 239000002253 acid Substances 0.000 description 4
- 239000007789 gas Substances 0.000 description 4
- 229910052742 iron Inorganic materials 0.000 description 4
- 230000035484 reaction time Effects 0.000 description 4
- 229910021607 Silver chloride Inorganic materials 0.000 description 3
- NINIDFKCEFEMDL-UHFFFAOYSA-N Sulfur Chemical compound [S] NINIDFKCEFEMDL-UHFFFAOYSA-N 0.000 description 3
- 239000007864 aqueous solution Substances 0.000 description 3
- 239000002585 base Substances 0.000 description 3
- 230000000052 comparative effect Effects 0.000 description 3
- 230000000694 effects Effects 0.000 description 3
- 229910052733 gallium Inorganic materials 0.000 description 3
- 230000001376 precipitating effect Effects 0.000 description 3
- 238000001556 precipitation Methods 0.000 description 3
- HKZLPVFGJNLROG-UHFFFAOYSA-M silver monochloride Chemical compound [Cl-].[Ag+] HKZLPVFGJNLROG-UHFFFAOYSA-M 0.000 description 3
- 235000011121 sodium hydroxide Nutrition 0.000 description 3
- 238000003756 stirring Methods 0.000 description 3
- 239000000126 substance Substances 0.000 description 3
- 238000006467 substitution reaction Methods 0.000 description 3
- MHAJPDPJQMAIIY-UHFFFAOYSA-N Hydrogen peroxide Chemical compound OO MHAJPDPJQMAIIY-UHFFFAOYSA-N 0.000 description 2
- MKOYQDCOZXHZSO-UHFFFAOYSA-N [Cu].[Cu].[Cu].[As] Chemical compound [Cu].[Cu].[Cu].[As] MKOYQDCOZXHZSO-UHFFFAOYSA-N 0.000 description 2
- 229910021645 metal ion Inorganic materials 0.000 description 2
- 239000007800 oxidant agent Substances 0.000 description 2
- 230000001590 oxidative effect Effects 0.000 description 2
- 238000010979 pH adjustment Methods 0.000 description 2
- 239000000843 powder Substances 0.000 description 2
- 239000002244 precipitate Substances 0.000 description 2
- 239000002994 raw material Substances 0.000 description 2
- 239000007787 solid Substances 0.000 description 2
- 239000007858 starting material Substances 0.000 description 2
- GKCNVZWZCYIBPR-UHFFFAOYSA-N sulfanylideneindium Chemical compound [In]=S GKCNVZWZCYIBPR-UHFFFAOYSA-N 0.000 description 2
- 238000005486 sulfidation Methods 0.000 description 2
- 229910052717 sulfur Inorganic materials 0.000 description 2
- 239000011593 sulfur Substances 0.000 description 2
- 125000000101 thioether group Chemical group 0.000 description 2
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Substances O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 description 2
- 229910000925 Cd alloy Inorganic materials 0.000 description 1
- RWSOTUBLDIXVET-UHFFFAOYSA-N Dihydrogen sulfide Chemical compound S RWSOTUBLDIXVET-UHFFFAOYSA-N 0.000 description 1
- 229910000673 Indium arsenide Inorganic materials 0.000 description 1
- GRYLNZFGIOXLOG-UHFFFAOYSA-N Nitric acid Chemical compound O[N+]([O-])=O GRYLNZFGIOXLOG-UHFFFAOYSA-N 0.000 description 1
- 239000003929 acidic solution Substances 0.000 description 1
- 150000007513 acids Chemical class 0.000 description 1
- 239000003513 alkali Substances 0.000 description 1
- QVGXLLKOCUKJST-UHFFFAOYSA-N atomic oxygen Chemical compound [O] QVGXLLKOCUKJST-UHFFFAOYSA-N 0.000 description 1
- 230000015572 biosynthetic process Effects 0.000 description 1
- 239000006227 byproduct Substances 0.000 description 1
- BDOSMKKIYDKNTQ-UHFFFAOYSA-N cadmium atom Chemical group [Cd] BDOSMKKIYDKNTQ-UHFFFAOYSA-N 0.000 description 1
- 150000001875 compounds Chemical class 0.000 description 1
- 238000010586 diagram Methods 0.000 description 1
- 238000006073 displacement reaction Methods 0.000 description 1
- 238000005363 electrowinning Methods 0.000 description 1
- 229910000037 hydrogen sulfide Inorganic materials 0.000 description 1
- RPQDHPTXJYYUPQ-UHFFFAOYSA-N indium arsenide Chemical compound [In]#[As] RPQDHPTXJYYUPQ-UHFFFAOYSA-N 0.000 description 1
- 229910003437 indium oxide Inorganic materials 0.000 description 1
- PJXISJQVUVHSOJ-UHFFFAOYSA-N indium(iii) oxide Chemical compound [O-2].[O-2].[O-2].[In+3].[In+3] PJXISJQVUVHSOJ-UHFFFAOYSA-N 0.000 description 1
- 230000008595 infiltration Effects 0.000 description 1
- 238000001764 infiltration Methods 0.000 description 1
- 229910000765 intermetallic Inorganic materials 0.000 description 1
- 238000005342 ion exchange Methods 0.000 description 1
- XEEYBQQBJWHFJM-UHFFFAOYSA-N iron Substances [Fe] XEEYBQQBJWHFJM-UHFFFAOYSA-N 0.000 description 1
- 229910052745 lead Inorganic materials 0.000 description 1
- 229910052976 metal sulfide Inorganic materials 0.000 description 1
- 239000000203 mixture Substances 0.000 description 1
- 229910052759 nickel Inorganic materials 0.000 description 1
- 229910017604 nitric acid Inorganic materials 0.000 description 1
- 229910052760 oxygen Inorganic materials 0.000 description 1
- 239000001301 oxygen Substances 0.000 description 1
- 238000007670 refining Methods 0.000 description 1
- 239000004065 semiconductor Substances 0.000 description 1
- 238000000638 solvent extraction Methods 0.000 description 1
- 239000004071 soot Substances 0.000 description 1
- 230000019086 sulfide ion homeostasis Effects 0.000 description 1
- 239000000725 suspension Substances 0.000 description 1
- 239000010409 thin film Substances 0.000 description 1
- 230000037303 wrinkles Effects 0.000 description 1
- -1 zinc powder Chemical compound 0.000 description 1
Classifications
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- Y—GENERAL TAGGING OF NEW TECHNOLOGICAL DEVELOPMENTS; GENERAL TAGGING OF CROSS-SECTIONAL TECHNOLOGIES SPANNING OVER SEVERAL SECTIONS OF THE IPC; TECHNICAL SUBJECTS COVERED BY FORMER USPC CROSS-REFERENCE ART COLLECTIONS [XRACs] AND DIGESTS
- Y02—TECHNOLOGIES OR APPLICATIONS FOR MITIGATION OR ADAPTATION AGAINST CLIMATE CHANGE
- Y02P—CLIMATE CHANGE MITIGATION TECHNOLOGIES IN THE PRODUCTION OR PROCESSING OF GOODS
- Y02P10/00—Technologies related to metal processing
- Y02P10/20—Recycling
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- Manufacture And Refinement Of Metals (AREA)
Abstract
Description
本発明は、インジウム含有物からインジウムを回収する方法に関する。 The present invention relates to a method for recovering indium from an indium-containing material.
インジウムは、III-Ｖ族化合物半導体としてＩｎＰ、ＩｎＡｓ等の金属間化合物に、あるいは太陽電池用材料として錫をドープした酸化インジウム（ＩＴＯ）、透明導電性薄膜に利用されており、今後その需要は益々伸長するものと期待されている。 Indium is used as an III-V compound semiconductor for intermetallic compounds such as InP and InAs, or as a solar cell material for tin-doped indium oxide (ITO), and for transparent conductive thin films. It is expected to grow more and more.
元来、インジウムには主たる鉱石がなく、工業的には亜鉛製錬、鉛製錬の副産物、例えば、ばい煙中に濃縮されたインジウムを回収することにより生産されている。したがってインジウム回収の原料は、Ｚｎ、Ｆｅ、Ｃｕ、Ａｌ、Ｇａ、Ａｓ、Ｃｄ等の金属不純物を多く含んでおり、またこれら金属成分以外にも微量に含まれる成分の種類が多い。 Originally, indium has no main ore and is industrially produced by recovering byproducts of zinc smelting and lead smelting, for example, indium concentrated in soot. Accordingly, the raw material for indium recovery contains a large amount of metal impurities such as Zn, Fe, Cu, Al, Ga, As, and Cd, and there are many kinds of components contained in trace amounts other than these metal components.
したがって、これら金属不純物を除去し、高純度のインジウムを回収するには複雑な工程が必要となり、一般に上記インジウムの回収工程は、（Ａ）ｐＨ調整により水酸化物として沈殿させる方法、（Ｂ）硫化剤の添加により硫化物として沈殿させる方法、（Ｃ）金属Ａｌ、Ｚｎ、Ｃｄ、Ｚｎ−Ｃｄ合金等の添加により置換析出させる方法、（Ｄ）溶媒抽出によってインジウムを回収する方法、（Ｅ）イオン交換法によるインジウムの回収方法、等の化学精製と、電解製錬法との組み合わせにより行なわれている（例えば、特許文献１参照）。 Therefore, a complicated process is required to remove these metal impurities and recover high-purity indium. Generally, the indium recovery process includes (A) a method of precipitating as a hydroxide by pH adjustment, (B) A method of precipitating as a sulfide by adding a sulfiding agent, (C) a method of substituting and precipitating by adding metal Al, Zn, Cd, Zn—Cd alloy, etc., (D) a method of recovering indium by solvent extraction, (E) This is performed by a combination of chemical purification such as an indium recovery method by an ion exchange method and an electrolytic smelting method (see, for example, Patent Document 1).
しかしながら前記回収工程のうち、（Ａ）の方法は、金属イオンの水酸化物生成ｐＨ領域の違いを利用したものであり、例えばＺｎ、ＡｌとＩｎの分離法としてはｐＨを１２以上にすることによってＺｎ、Ａｌを溶解し、Ｉｎを水酸化物として沈殿させて回収する方法がある。しかしこの方法では、生成したＩｎの水酸化物は濾過性が極めて悪いため濾過設備が大きくなり、操作も長時間となる。またこの方法ではＦｅ、Ｃｕ、Ａｓ、Ｃｄ等の不純物とＩｎとの分離は困難である。 However, among the recovery steps, the method (A) utilizes the difference in the metal ion hydroxide formation pH range. For example, as a method for separating Zn, Al and In, the pH is set to 12 or more. There is a method in which Zn and Al are dissolved, and In is precipitated and recovered as a hydroxide. However, in this method, the produced In hydroxide is very poor in filterability, so the filtration equipment becomes large and the operation takes a long time. In this method, it is difficult to separate In from impurities such as Fe, Cu, As, and Cd.
（Ｂ）の方法は、金属硫化物の溶解度積の違いを利用したものであるが、前述のような様々な金属不純物を含むため純度の低い硫化物が大量に発生する。これらの硫化物は一般に濾過性が悪く、また得られたＩｎの硫化物を浸出する場合、硫酸のみではＩｎを完全に浸出することができないため、この方法には、湿式亜鉛工程に応用し難いという欠点がある。 The method (B) uses the difference in solubility product of metal sulfides, but contains a large amount of low-purity sulfides because it contains various metal impurities as described above. These sulfides generally have poor filterability, and when leaching the obtained In sulfide, In cannot be completely leached with sulfuric acid alone, so this method is difficult to apply to the wet zinc process. There is a drawback.
（Ｃ）については、インジウムより貴な不純物を含む場合にはその金属とＩｎの分離は不可能である。またＩｎが置換析出する場合に生成するスポンジは塊状化するため好ましい粉状金属が得られない。 As for (C), in the case where impurities more precious than indium are contained, the separation of the metal and In is impossible. In addition, since the sponge produced when In is deposited by substitution is agglomerated, a preferable powder metal cannot be obtained.
（Ｄ）、（Ｅ）についてはＩｎと分離する不純物によっては前処理に負担がかかりまたランニングコストが高いという問題がある。 With respect to (D) and (E), depending on the impurities separated from In, there is a problem that a pretreatment is burdened and the running cost is high.
上記いずれの化学精製方法においても、不純物金属の分離が不十分であるため、これと組み合わせる電解製錬方法も簡便な電解採取法（水溶液中に目的金属を浸出させておき不溶性の陽極を用いて電気分解し、一挙に陰極に高純度の金属を得る）を採用できず、煩雑な電解精製法（粗金属を陽極に、高純度金属を陰極において電気分解して精製を行なう）を採用せざるを得なかった。 In any of the above chemical refining methods, the separation of impurity metals is insufficient, so the electrolytic smelting method combined therewith is also a simple electrowinning method (leaving the target metal in an aqueous solution and using an insoluble anode. Electrolysis can not be used to obtain a high-purity metal at the cathode at once, and a complicated electrolytic purification method (crude metal is used as an anode and high-purity metal is electrolyzed at the cathode for purification) Did not get.
したがって、上記いずれの方法もそれぞれ欠点を有しており、実際の回収には上記の方法を組み合わせたものが使用されており、高純度Ｉｎを回収するためには工程が複雑でかつ煩雑となり、また、雑多に含まれる金属成分を除去するための条件を設定する際に、雑多な金属が含まれるため、良好な条件の予測は不可能であり、経済的な方法はまだ提案されていなかった。 Accordingly, each of the above methods has its respective drawbacks, and a combination of the above methods is used for actual recovery, and the process is complicated and complicated for recovering high purity In, In addition, when setting conditions for removing metal components contained in miscellaneous conditions, miscellaneous metals are included, so it is impossible to predict good conditions, and an economical method has not yet been proposed. .
本発明は、従来における諸問題を解決し、以下の目的を達成することを課題とする。即ち、本発明は、様々な不純物を含むインジウム含有物から高純度インジウムを効率よく回収する方法を提供することを目的とする。 An object of the present invention is to solve various problems in the prior art and achieve the following objects. That is, an object of the present invention is to provide a method for efficiently recovering high-purity indium from an indium-containing material containing various impurities.
本発明者らは、上記の課題を解決すべく鋭意研究を続け、試行錯誤の結果本発明に到達することができた。 The inventors of the present invention have continued intensive studies to solve the above problems, and have reached the present invention as a result of trial and error.
前記課題を解決する手段としては、以下の通りである。即ち、
＜１＞ インジウムと前記インジウム以外の金属とを含むインジウム含有物から、インジウムを回収する方法において、前記インジウムを硫酸を含む溶液中に浸出させる浸出工程を含み、インジウム浸出液の硫酸濃度を３５〜８０ｇ／Ｌとした、インジウム含有物からインジウムを回収する方法である。
＜２＞ 浸出温度を７０〜８０℃とした、前記＜１＞に記載のインジウムを回収する方法である。
＜３＞ インジウム浸出液の酸化還元電位を５００ｍＶ以下とした、前記＜１＞から＜２＞のいずれかに記載のインジウムを回収する方法である。
＜４＞ インジウム浸出液から、インジウム以外の金属を硫化および還元浸出により分離し、電解を行う、前記＜１＞から＜３＞のいずれかに記載のインジウムを回収する方法である。
Means for solving the above problems are as follows. That is,
<1> In a method for recovering indium from an indium-containing material containing indium and a metal other than indium, the method includes a leaching step of leaching the indium into a solution containing sulfuric acid, and the sulfuric acid concentration of the indium leaching solution is 35 to 80 g. / L is a method of recovering indium from the indium-containing material.
<2> The method for recovering indium according to <1>, wherein the leaching temperature is 70 to 80 ° C.
<3> The method for recovering indium according to any one of <1> to <2>, wherein an oxidation-reduction potential of the indium leaching solution is 500 mV or less.
<4> A method for recovering indium according to any one of <1> to <3>, wherein a metal other than indium is separated from the indium leaching solution by sulfurization and reduction leaching and electrolysis is performed.
本発明の方法によれば、多種、多様の金属不純物を含むインジウム含有物から、電解精製を要しない簡略な工程で、効率よく、しかも純度が５Ｎ以上の高純度のインジウムを回収することができる。また、各工程の濾過における濾過性を向上することができ、もって濾過設備を小さくすると共に操作を短時間で行うことができる。 According to the method of the present invention, high-purity indium having a purity of 5N or more can be efficiently recovered from indium-containing materials containing various metal impurities in a simple process that does not require electrolytic purification. . Moreover, the filterability in filtration of each process can be improved, so that the filtration facility can be reduced and the operation can be performed in a short time.
本発明ではインジウムを含有するものを広く出発原料として採用し得るが、ここでは湿式亜鉛製錬に際して副生する中和石膏に適用する場合について説明することにする。本発明の方法によるインジウム回収の工程を図１に示す。 In the present invention, a material containing indium can be widely used as a starting material, but here, a case where it is applied to neutralized gypsum by-produced during wet zinc smelting will be described. The process of indium recovery by the method of the present invention is shown in FIG.
（１）の工程では、中和石膏を硫酸で浸出すると、Ｉｎと共にＣｕ、Ａｓ、Ａｌ、Ｆｅ、Ｚｎ、Ｇａ等の酸に可溶な不純物金属イオンが浸出され、不溶性石膏とのスラリーを形成する。浸出に使用する酸としては、硫酸の他に塩酸、硝酸等を使用でき、硫酸に制限されるものではないが硫酸が最も安価である。Ｉｎ浸出液の硫酸濃度を３５〜８０ｇ／Ｌとし、酸化還元電位を５００ｍＶ以下とし、浸出温度を６０〜８０℃とする。 In the step (1), when neutralized gypsum is leached with sulfuric acid, impurity metal ions soluble in acids such as Cu, As, Al, Fe, Zn, and Ga are leached together with In to form a slurry with insoluble gypsum. To do. As the acid used for the leaching, hydrochloric acid, nitric acid or the like can be used in addition to sulfuric acid, and although it is not limited to sulfuric acid, sulfuric acid is the cheapest. The sulfuric acid concentration of the In leachate is 35 to 80 g / L, the oxidation-reduction potential is 500 mV or less, and the leaching temperature is 60 to 80 ° C.
（２）の工程では、（１）の工程で得られたＩｎ浸出スラリーに、硫化剤として例えばＨ２Ｓ、ＮａＳＨを酸化還元電位（以下Ｅｈと言う）が５０〜３２０ｍＶ（Ａｇ／ＡｇＣｌ電極使用）の範囲内に入るようにコントロールしながら添加し、Ｃｕ、Ａｓ等の不純物を硫化物として沈殿除去する。このとき硫酸濃度も２０〜４０ｇ／ＬにコントロールするためＩｎは沈殿しない。 In the step (2), the In leach slurry obtained in the step (1) has a redox potential (hereinafter referred to as Eh) of 50 to 320 mV (using an Ag / AgCl electrode), for example, H 2 S and NaSH as sulfiding agents. ) While being controlled so as to be within the range of), impurities such as Cu and As are precipitated and removed as sulfides. At this time, since the sulfuric acid concentration is also controlled to 20 to 40 g / L, In does not precipitate.
（１）および（２）の工程の処理により中和石膏中に含まれるＩｎの９０％以上が硫酸酸性溶液中に移行するので、例えばフィルタープレス等を用いて沈殿物（銅残渣）を固液分離する。この時浸出時の不溶性石膏が濾過助剤の働きをするため、一般には悪い硫化物の濾過性が著しく改善される。銅残渣は亜鉛製錬の本系統へ送られる。 Since 90% or more of In contained in the neutralized gypsum is transferred into the sulfuric acid acidic solution by the treatment in the steps (1) and (2), the precipitate (copper residue) is solid-liquid using, for example, a filter press. To separate. At this time, insoluble gypsum at the time of leaching functions as a filter aid, so that generally the filterability of bad sulfides is remarkably improved. Copper residue is sent to the main line of zinc smelting.
（３）の工程では、（２）の工程で得られたＩｎ含有水溶液に硫化剤例えばＨ２Ｓ、ＮａＳＨを硫酸と同時に添加し、Ｉｎを硫化物として沈殿させ、フィルタープレス等を用いて固液分離し、液中に残るＺｎ、Ｆｅ、Ａｌ、Ｇａ等の不純物を分離除去する。Ｉｎの沈殿への回収率は９５％以上である。濾液（硫化后液）は排水系統へ送られる。 In the step (3), a sulfurizing agent such as H 2 S or NaSH is added to the In-containing aqueous solution obtained in the step (2) simultaneously with sulfuric acid to precipitate In as a sulfide, which is solidified using a filter press or the like. Liquid separation is performed, and impurities such as Zn, Fe, Al, and Ga remaining in the liquid are separated and removed. The recovery rate of In into precipitation is 95% or more. The filtrate (after sulfidation) is sent to the drainage system.
（４）の工程では、（３）の工程で得られた硫化インジウムに、硫酸酸性下でＳＯ２ガスを吹き込みながらＩｎを浸出する。 In the step (4), In is leached while injecting SO 2 gas into the indium sulfide obtained in the step (3) under sulfuric acid acidity.
硫化物の酸浸出法には一般に、（ａ）硫化水素発生型、（ｂ）硫黄生成型、（ｃ）硫酸生成型の３通りの型があるが、硫化インジウムを浸出する場合、（ａ）の反応では溶解度積が小さいため、Ｉｎを完全に浸出することができず、（ｂ）、（ｃ）の反応では酸化剤として酸素を用いる場合、反応温度、圧力をそれぞれ１５０℃、１２ｋｇ／ｃｍ２のように高くする必要があるためオートクレーブ等の圧力容器を反応槽としなければならない。また、この方法でＩｎを完全に浸出することは可能であるが、酸化力が強力であるため含有している不純物も同様に完全に浸出されてしまう。 There are generally three types of acid leaching methods for sulfides: (a) hydrogen sulfide generation type, (b) sulfur generation type, and (c) sulfuric acid generation type. When indium sulfide is leached, (a) In this reaction, the solubility product is small, so that In cannot be completely leached. In the reactions (b) and (c), when oxygen is used as the oxidizing agent, the reaction temperature and pressure are 150 ° C. and 12 kg / cm, respectively. the pressure vessel such as an autoclave it is necessary to increase as 2 to do with the reaction vessel. In addition, although it is possible to completely leaze In by this method, since the oxidizing power is strong, the contained impurities are also completely leached.
本発明の方法では、酸化剤としてＳＯ２を用いることで（ａ）と（ｂ）との反応の組み合わせを行ない、酸化力を適度にコントロールしＩｎは浸出しつつ他の不純物の浸出を抑える、つまり選択的にＩｎを浸出する。この時の温度は常温でもよく、圧力も大気圧でよいため通常の反応槽を使用することができる。反応後Ｉｎの９０％以上が浸出液に移行するためフィルタープレス等を用いて固液分離する。ケーキ（硫黄残渣）は亜鉛製錬の本系統へ送られる。 In the method of the present invention performs a combination of reactions by using SO 2 (a) and (b) as an oxidizing agent, and appropriately controlling the oxidizing power In order to suppress the leaching of other impurities while leaching, That is, In is leached selectively. At this time, the temperature may be room temperature and the pressure may be atmospheric pressure, so that a normal reaction vessel can be used. Since 90% or more of In is transferred to the leachate after the reaction, it is solid-liquid separated using a filter press or the like. The cake (sulfur residue) is sent to the main line of zinc smelting.
（５）の工程では、（４）の工程で得られたＩｎ浸出液をアルカリ例えば苛性ソーダ等で中和し、好ましくはｐＨを１〜３．５の範囲に調整する。ｐＨが１より低いと後の工程で置換剤として加える亜鉛末の使用量が過剰に必要となり、ｐＨが３．５を超えるとＩｎが水酸化物を生成してしまうためである。ｐＨの調整後、インジウムよりイオン化傾向の大きい金属の粉末、例えば亜鉛末を添加してインジウムスポンジを置換析出させる。（４）の工程で浸出にＳＯ２を使用しているため（５）の工程に供するＩｎ浸出液中にはＳＯ２が溶存している。この濃度を０．０５〜０．３ｇ／Ｌにコントロールすることによりインジウムスポンジの塊状化を防止することができ、粉状のインジウムスポンジを得ることができる。置換后液は前記（３）の工程へ繰り返される。 In the step (5), the In leachate obtained in the step (4) is neutralized with an alkali such as caustic soda, and the pH is preferably adjusted in the range of 1 to 3.5. This is because if the pH is lower than 1, an excessive amount of zinc powder to be added as a substitution agent in the subsequent step is required, and if the pH exceeds 3.5, In generates hydroxide. After the pH adjustment, a metal powder having a higher ionization tendency than indium, such as zinc powder, is added to displace and deposit the indium sponge. (4) The of In leaching solution to be subjected to a process for using the SO 2 leaching in step (5) is SO 2 is dissolved. By controlling this concentration to 0.05 to 0.3 g / L, the indium sponge can be prevented from being agglomerated and a powdery indium sponge can be obtained. The post-replacement solution is repeated to the step (3).
（６）の工程では、（５）の工程で得られたインジウムスポンジを塩酸でｐＨを０．５〜１．５の範囲内、Ｅｈを−４００〜−５００ｍＶの範囲内にそれぞれコントロールして浸出する。この時Ｉｎの９０％以上が浸出液に移行するためフィルタープレス等を用いて固液分離する。浸出残分（スポンジ滓）にはＣｄ、Ｐｂ、Ｎｉ、Ａｓ等の微量金属が濃縮されて除去できる。スポンジ滓は前記（４）の工程へ繰り返される。 In step (6), the indium sponge obtained in step (5) was leached with hydrochloric acid within a pH range of 0.5 to 1.5 and Eh within a range of -400 to -500 mV. To do. At this time, since 90% or more of In is transferred to the leachate, solid-liquid separation is performed using a filter press or the like. Trace metals such as Cd, Pb, Ni, As, etc. can be concentrated and removed from the leaching residue (sponge sponge). The sponge wrinkle is repeated to the step (4).
（７）の工程では、（６）の工程で得られたＩｎ浸出液にまだＣｄ、Ａｓ等が残留している場合、硫化剤例えばＨ２Ｓガスを吹き込み、最終浄液を行ない、固液分離して濾液を電解元液とする。ケーキ（カドミ残渣）は前記（４）の工程へ繰り返される。 In the step (7), when Cd, As, etc. still remain in the In leachate obtained in the step (6), a sulfidizing agent, for example, H 2 S gas is blown, the final purification is performed, and solid-liquid separation is performed. Thus, the filtrate is used as an electrolytic base solution. The cake (cadomi residue) is repeated to the step (4).
（８）の工程では、（７）の工程で得られた電解元液から、アノードにＤＳＡ（寸法適格陽極）、カソードにＴｉ板を用いて電解採取を行ない、高純度の金属インジウムを得る。 In the step (8), electrolysis is performed from the electrolytic source solution obtained in the step (7) using a DSA (dimension qualified anode) for the anode and a Ti plate for the cathode to obtain high purity metallic indium.
（実施例１）
湿式亜鉛製錬工程で副生する中和石膏を出発原料としてインジウムの回収処理を行なった。
(Example 1)
The indium recovery process was performed using neutralized gypsum by-produced in the wet zinc smelting process as a starting material.
この中和石膏を５００ｍＬ採取して、容器内に投入した。この容器には、温度制御、撹拌機が取り付けられており、それぞれ制御可能となっている。該中和石膏の入った容器に、硫酸を５５ｍＬを２ｍＬ/分の速度で加え、反応温度６０℃、攪拌機の回転数１９６ｒｐｍにて混合し、浸出反応させ、浸出スラリーを得た。 500 mL of this neutralized gypsum was collected and put into a container. This container is provided with a temperature control and a stirrer, and can be controlled respectively. To the container containing the neutralized gypsum, 55 mL of sulfuric acid was added at a rate of 2 mL / min, and the mixture was mixed at a reaction temperature of 60 ° C. and a rotation speed of a stirrer of 196 rpm to cause a leaching reaction to obtain a leaching slurry.
浸出反応させた後、浸出スラリーの濾過試験を実施し、濾過速度及び濾過時間を測定した。濾過試験は、浸出スラリー（５００ｍＬ）を容量が１．５Ｌの容器にて、容器内を５ｋｇ/ｃｍ２の加圧しながら濾過することにより実施した。濾過速度は、濾液の量を、該濾液の量を濾過するまでの時間（濾過時間）で除算した値（濾液の量／濾過時間）である。濾過時間は、加圧初めから濾液が４６０ｍＬに達するまでの時間である。即ち、この濾過時間が短い程、濾過性が優れている。また、上記濾過により得た濾液のｐＨ、酸化還元電位（ｍＶ）、遊離硫酸濃度（ｇ／Ｌ）、及びインジウム浸出率（％）を測定した。インジウム浸出率（％）は、化学分析により得たインジウム含有量と該中和石膏中のインジウム含有量との比（％）で表す。 After the leaching reaction, the leaching slurry was subjected to a filtration test, and the filtration rate and the filtration time were measured. The filtration test was performed by filtering the leach slurry (500 mL) in a 1.5 L container while pressurizing the inside of the container at 5 kg / cm 2 . The filtration rate is a value (amount of filtrate / filtration time) obtained by dividing the amount of the filtrate by the time until the amount of the filtrate is filtered (filtration time). The filtration time is the time from the beginning of pressurization until the filtrate reaches 460 mL. That is, the shorter the filtration time, the better the filterability. The pH, redox potential (mV), free sulfuric acid concentration (g / L), and indium leaching rate (%) of the filtrate obtained by the filtration were measured. The indium leaching rate (%) is expressed as a ratio (%) between the indium content obtained by chemical analysis and the indium content in the neutralized gypsum.
その結果、濾過時間は７６秒であり、濾液のｐＨは０．６２であり、濾液の酸化還元電位は４７４ｍＶであり、濾液の遊離硫酸濃度は４１ｇ／Ｌであり、インジウム浸出率は９７．３％であった。この結果より、実施例１は、良好な濾過性を得ることができ、インジウムが十分に浸出されていることが分かった。 As a result, the filtration time was 76 seconds, the pH of the filtrate was 0.62, the redox potential of the filtrate was 474 mV, the free sulfuric acid concentration of the filtrate was 41 g / L, and the indium leaching rate was 97.3. %Met. From this result, it was found that Example 1 was able to obtain good filterability, and indium was sufficiently leached.
なお、攪拌機の回転数の浸出反応に対する影響は少なく、攪拌機の回転数を２倍（３９２ｒｐｍ）にして実施しても同様な結果であった。 The effect of the number of revolutions of the stirrer on the leaching reaction was small, and similar results were obtained even when the number of revolutions of the stirrer was doubled (392 rpm).
（実施例２）
反応温度を７０℃に変更した以外は、実施例１と同様に浸出反応させた後、この浸出スラリーの濾過試験を実施し、濾過時間を測定した。
(Example 2)
Except that the reaction temperature was changed to 70 ° C., the leaching reaction was carried out in the same manner as in Example 1, and then the filtration test of this leached slurry was performed to measure the filtration time.
その結果、濾過時間は７０秒であり、濾液のｐＨは０．６６であり、濾液の酸化還元電位は４７９ｍＶであり、濾液の遊離硫酸濃度は４３ｇ／Ｌであり、インジウム浸出率は９７．８％であった。この結果より、実施例２は、良好な濾過性を得ることができ、インジウムが十分に浸出されていることが分かった。 As a result, the filtration time was 70 seconds, the pH of the filtrate was 0.66, the redox potential of the filtrate was 479 mV, the free sulfuric acid concentration of the filtrate was 43 g / L, and the indium leaching rate was 97.8. %Met. From this result, it was found that Example 2 was able to obtain good filterability, and indium was sufficiently leached.
（実施例３）
反応温度を８０℃に変更した以外は、実施例１と同様に浸出反応させた後、この浸出スラリーの濾過試験を実施し、濾過時間を測定した。
(Example 3)
Except that the reaction temperature was changed to 80 ° C., the leaching reaction was carried out in the same manner as in Example 1, and then the filtration test of the leached slurry was performed to measure the filtration time.
その結果、濾過時間は３８秒であり、濾液のｐＨは０．８０であり、濾液の酸化還元電位は４７８ｍＶであり、濾液の遊離硫酸濃度は３８ｇ／Ｌであり、インジウム浸出率は９７．５％であった。この結果より、実施例３は、良好な濾過性を得ることができ、インジウムが十分に浸出されていることが分かった。 As a result, the filtration time was 38 seconds, the pH of the filtrate was 0.80, the redox potential of the filtrate was 478 mV, the free sulfuric acid concentration of the filtrate was 38 g / L, and the indium leaching rate was 97.5. %Met. From this result, it was found that Example 3 was able to obtain good filterability, and indium was sufficiently leached.
実施例１〜３の結果から、反応温度を６０〜８０℃、好ましくは７０〜８０℃にすると、良好な濾過性を得ることができ、インジウムの浸出が十分に得られることが分かった。また、反応温度を８０℃とすると、反応温度が６０〜７０℃のときと比べて、顕著に濾過性が向上されることが分かった。 From the results of Examples 1 to 3, it was found that when the reaction temperature is 60 to 80 ° C., preferably 70 to 80 ° C., good filterability can be obtained and infiltration of indium is sufficiently obtained. Moreover, when reaction temperature was 80 degreeC, it turned out that filterability is improved notably compared with the case where reaction temperature is 60-70 degreeC.
（比較例１）
硫酸添加量を４５ｍＬに変更した以外は、実施例１と同様に浸出反応させた後、この浸出スラリーの濾過試験を実施し、濾過時間を測定した。
(Comparative Example 1)
Except that the amount of sulfuric acid added was changed to 45 mL, the leaching reaction was carried out in the same manner as in Example 1, and then the filtration test of this leached slurry was performed to measure the filtration time.
その結果、濾過時間は１，０００秒以上（濾過性が低くて途中で中断）であり、濾液のｐＨは１．４２であり、濾液の酸化還元電位は４４７ｍＶであり、濾液の遊離硫酸濃度は不明であり、インジウム浸出率は不明であった。この結果より、反応温度が６０℃であっても、硫酸量が少ないと濾過性が低くなることが分かった。 As a result, the filtration time was 1,000 seconds or more (interruption was low due to low filterability), the pH of the filtrate was 1.42, the redox potential of the filtrate was 447 mV, and the free sulfuric acid concentration of the filtrate was Unknown and indium leaching rate was unknown. From this result, it was found that even when the reaction temperature was 60 ° C., the filterability was lowered when the amount of sulfuric acid was small.
（実施例４）
硫酸添加量を６５ｍＬに変更した以外は、実施例１と同様に浸出反応させた後、この浸出スラリーの濾過試験を実施し、濾過時間を測定した。
Example 4
Except that the amount of sulfuric acid added was changed to 65 mL, the leaching reaction was carried out in the same manner as in Example 1, and then the leaching slurry was subjected to a filtration test to measure the filtration time.
その結果、濾過時間は５２秒であり、濾液のｐＨは０．３１であり、濾液の酸化還元電位は４９０ｍＶであり、濾液の遊離硫酸濃度は３８ｇ／Ｌであり、インジウム浸出率は９８．５％であった。この結果より、実施例４は、良好な濾過性を得ることができ、インジウムが十分に浸出されていることが分かった。 As a result, the filtration time was 52 seconds, the pH of the filtrate was 0.31, the redox potential of the filtrate was 490 mV, the free sulfuric acid concentration of the filtrate was 38 g / L, and the indium leaching rate was 98.5. %Met. From this result, it was found that in Example 4, good filterability was obtained and indium was sufficiently leached.
実施例１，４及び比較例１の結果から、硫酸添加量を５５ｍＬ以上とすると、濾過性のが著しく改善されることが分かった。 From the results of Examples 1 and 4 and Comparative Example 1, it was found that the filterability was remarkably improved when the amount of sulfuric acid added was 55 mL or more.
（実施例５）
電位調整剤として過酸化水素を用いて、酸化還元電位の調整を行ったこと以外は、実施例１と同様に浸出反応させた後、この浸出スラリーの濾過試験を実施し、濾過時間を測定した。
(Example 5)
Except that the oxidation-reduction potential was adjusted using hydrogen peroxide as a potential adjusting agent, the leaching reaction was carried out in the same manner as in Example 1, and then the filtration test of this leached slurry was carried out to measure the filtration time. .
その結果、濾過時間は４９秒であり、濾液のｐＨは０．６６であり、濾液の酸化還元電位は７８２ｍＶであり、濾液の遊離硫酸濃度は４３ｇ／Ｌであり、インジウム浸出率は９６．７％であった。この結果より、実施例５は、良好な濾過性を得ることができ、インジウムが十分に浸出されていることが分かった。また、単純に濾液の酸化還元電位を上げても、浸出率の向上は図れないことが分かった。 As a result, the filtration time was 49 seconds, the pH of the filtrate was 0.66, the redox potential of the filtrate was 782 mV, the free sulfuric acid concentration of the filtrate was 43 g / L, and the indium leaching rate was 96.7. %Met. From this result, it was found that Example 5 was able to obtain good filterability, and indium was sufficiently leached. It was also found that the leach rate could not be improved by simply raising the redox potential of the filtrate.
（比較例２）
電位調整剤として砒化銅を用いて、酸化還元電位の調整を行ったこと以外は、実施例１と同様に浸出反応させた後、この浸出スラリーの濾過試験を実施し、濾過時間を測定した。
(Comparative Example 2)
A leaching reaction was carried out in the same manner as in Example 1 except that copper arsenide was used as the potential adjusting agent to adjust the oxidation-reduction potential, and then a filtration test of this leached slurry was performed to measure the filtration time.
その結果、濾過時間は１，０００秒以上（濾過性が低くて途中で中断）であり、濾液のｐＨは０．６０であり、濾液の酸化還元電位は１５４ｍＶであり、濾液の遊離硫酸濃度は不明であり、インジウム浸出率は不明であった。この結果より、酸化還元電位が２００ｍＶ以下では、インジウムの回収が不可能であり、酸化還元電位が４００ｍＶ以上であれば、十分に効果が得られることが分かった。 As a result, the filtration time was 1,000 seconds or longer (interruption was low due to low filterability), the pH of the filtrate was 0.60, the redox potential of the filtrate was 154 mV, and the free sulfuric acid concentration of the filtrate was Unknown and indium leaching rate was unknown. From this result, it was found that indium cannot be recovered when the oxidation-reduction potential is 200 mV or less, and that a sufficient effect can be obtained when the oxidation-reduction potential is 400 mV or more.
（実施例６）
反応温度を７０℃に変更し、硫酸添加量を６５ｍＬに変更した以外は、実施例１と同様に浸出反応させた後、この浸出スラリーの濾過試験を実施し、濾過時間を測定した。
(Example 6)
Except that the reaction temperature was changed to 70 ° C. and the addition amount of sulfuric acid was changed to 65 mL, the leaching reaction was carried out in the same manner as in Example 1, and then the filtration test of the leached slurry was performed to measure the filtration time.
その結果、濾過時間は１２秒であり、濾液のｐＨは０．２６であり、濾液の酸化還元電位は４５９ｍＶであり、濾液の遊離硫酸濃度は７７ｇ／Ｌであり、インジウム浸出率は９９．６％であった。この結果より、実施例６は、顕著に良好な濾過性であり、インジウム浸出率も顕著に向上されていることが分かった。 As a result, the filtration time was 12 seconds, the pH of the filtrate was 0.26, the redox potential of the filtrate was 459 mV, the free sulfuric acid concentration of the filtrate was 77 g / L, and the indium leaching rate was 99.6. %Met. From this result, it was found that Example 6 has remarkably good filterability and the indium leaching rate is remarkably improved.
（実施例７）
反応温度を８０℃に変更し、硫酸添加量を７０ｍＬに変更し、電位調整剤として砒化銅を用いて、酸化還元電位の調整を行い、凝集剤を１０ｍＬ添加した以外は実施例１と同様に浸出反応させた後、この浸出スラリーの濾過試験を実施し、濾過時間を測定した。
(Example 7)
The reaction temperature was changed to 80 ° C., the amount of sulfuric acid added was changed to 70 mL, the oxidation-reduction potential was adjusted using copper arsenide as the potential adjusting agent, and 10 mL of the flocculant was added, as in Example 1. After the leaching reaction, the leaching slurry was subjected to a filtration test, and the filtration time was measured.
その結果、濾過時間は６５秒であり、濾液のｐＨは０．３０であり、濾液の酸化還元電位は１５０ｍＶであり、濾液の遊離硫酸濃度は７７ｇ／Ｌであり、インジウム浸出率は９９.４％であった。
すなわち、反応温度、硫酸添加量、酸化還元電位、及び凝集剤の添加などは、単純にそれらの効果が加算されるものではないことが分かった。
That is, it was found that the reaction temperature, the addition amount of sulfuric acid, the oxidation-reduction potential, the addition of the flocculant, and the like do not simply add the effects.
（１）酸浸出
前記実施例６の条件で浸出を実施した。
この浸出により得られた浸出液を浸出液１とする。なお、この浸出では、従来と比べ５分の１程度の濾過時間で濾過が完了し、生産性向上が図られた。
原料および得られた浸出液１のＩｎ、Ｚｎ、Ｃｕ、Ａｓの含有率と分配率を表３に示す。
(1) Acid leaching Leaching was performed under the conditions of Example 6.
The leachate obtained by this leaching is designated as leachate 1. In this leaching, the filtration was completed in about one-fifth the filtration time compared with the conventional one, and the productivity was improved.
Table 3 shows the contents of In, Zn, Cu, and As and the distribution ratio of the raw materials and the obtained leachate 1.
（２）Ｃｕ等の除去
上記浸出工程で得られた浸出スラリーに、Ｅｈが３００ｍＶ（Ａｇ／ＡｇＣｌ電極使用）になるまでＮａＳＨを添加して硫化反応を行った。反応時間は２時間、反応温度は６０℃であった。反応終了後、得られたスラリーを濾過し、ケーキを銅残渣、濾液を脱銅液とした。それぞれの分析結果を表４に示す。
(2) Removal of Cu, etc. NaSH was added to the leaching slurry obtained in the above leaching step until Eh reached 300 mV (using an Ag / AgCl electrode), and a sulfurization reaction was performed. The reaction time was 2 hours and the reaction temperature was 60 ° C. After completion of the reaction, the obtained slurry was filtered, the cake was used as a copper residue, and the filtrate was used as a copper removal solution. Each analysis result is shown in Table 4.
（３）硫化沈殿
上記脱銅液（Ｉｎ含有水溶液）を撹拌機で撹拌しながら、硫酸でｐＨを０．８の一定レベルに保ち、Ｅｈが−２０ｍＶ（Ａｇ／ＡｇＣｌ電極使用）になるまでＮａＳＨを添加してＩｎを硫化物として沈殿させた。反応は６０℃の温度で５時間行った。反応終了後、得られたスラリーを濾過し、ケーキを硫化残渣、濾液を硫化后液とした。それぞれの分析結果と物質収支を表５に示す。
(3) Sulfidation precipitation While stirring the copper removal solution (In-containing aqueous solution) with a stirrer, the pH is kept at a constant level of 0.8 with sulfuric acid, and NaSH is used until Eh becomes -20 mV (using an Ag / AgCl electrode). Was added to precipitate In as sulfide. The reaction was carried out at a temperature of 60 ° C. for 5 hours. After completion of the reaction, the resulting slurry was filtered to obtain a cake as a sulfide residue, and the filtrate as a solution after sulfurization. Table 5 shows the analysis results and material balance.
（４）ＳＯ２浸出
上記（１）〜（３）の工程を繰り返して得られた硫化残渣を集めて４１７．７ｇとし、これに水を加えて固体濃度１１９ｇ／Ｌのパルプとし、撹拌機で撹拌しながら硫酸を加えて硫酸濃度を５１ｇ／Ｌとし、溶存ＳＯ２濃度が８ｇ／ＬになるようにＳＯ２ガスを吹き込んだ。反応は、８０℃の温度で２時間行った。反応終了後、得られたスラリーを濾過し、ケーキを硫黄残渣、濾液をＳＯ２浸出液とした。それぞれの分析結果と物質収支を表６に示す。
(4) SO 2 leaching The sulfide residue obtained by repeating the steps (1) to (3) above was collected to 417.7 g, and water was added thereto to obtain a pulp having a solid concentration of 119 g / L. Sulfuric acid was added while stirring to make the sulfuric acid concentration 51 g / L, and SO 2 gas was blown so that the dissolved SO 2 concentration became 8 g / L. The reaction was carried out at a temperature of 80 ° C. for 2 hours. After completion of the reaction, the resulting slurry was filtered, the cake was sulfur residue, and the filtrate was SO 2 leachate. Table 6 shows the results of each analysis and the material balance.
（５）置換析出
上記ＳＯ２浸出液に空気を吹き込んで溶存ＳＯ２濃度が０．２ｇ／Ｌになるまで脱気し、ｐＨが２．５になるまでＮａＯＨを加えて中和したものを置換元液とした。得られた置換元液３，０００ｍＬに、Ｉｎに対して１．８当量の亜鉛末を添加し、Ｉｎスポンジを置換析出させた。反応温度は６０℃、反応時間は１時間であった。各産物の分析結果と物質収支を表７に示す。
(5) Displacement precipitation The air was blown into the SO 2 leaching solution to deaerate until the dissolved SO 2 concentration reached 0.2 g / L, and neutralized by adding NaOH until the pH reached 2.5. A liquid was used. To 3,000 mL of the obtained substitution source solution, 1.8 equivalents of zinc powder with respect to In were added, and In sponge was substituted and deposited. The reaction temperature was 60 ° C., and the reaction time was 1 hour. Table 7 shows the analysis results and material balance of each product.
（６）塩酸浸出工程
上記の諸工程を繰り返して集めたスポンジＩｎ２３８．１ｇに水を加えて固体濃度１４４ｇ／Ｌのパルプとし、撹拌機で撹拌しながら、ｐＨが１、Ｅｈが−４８０ｍＶ（Ａｇ／ＡｇＣｌ電極使用）となるように塩酸を添加してインジウムを浸出した。反応温度は６５℃、反応時間は３時間であった。各産物の分析結果と物質収支を表８に示す。
(6) Hydrochloric acid leaching step Water was added to 238.1 g of sponge In collected by repeating the above steps to obtain a pulp having a solid concentration of 144 g / L. While stirring with a stirrer, pH was 1 and Eh was -480 mV (Ag Indium was leached out by adding hydrochloric acid so that the following was obtained. The reaction temperature was 65 ° C., and the reaction time was 3 hours. Table 8 shows the analysis results and material balance of each product.
（７）Ｃｄ等除去工程
上記塩酸浸出工程で得られた塩酸浸出液（１，５００ｍＬ）にＮａＯＨを加えてｐＨ１．５まで中和した後、この液に１．５ＬのＨ２Ｓガスを吹き込んでＣｄ等の不純物を硫化物として沈殿させた。反応温度は４０℃、反応時間は０．５時間であった。反応後の懸濁液を濾過し、ケーキをカドミ残渣、濾液を脱Ｃｄ液とした。各産物の分析結果と物質収支を表９に示す。
(7) Cd etc. removing step After adding NaOH to the hydrochloric acid leaching solution (1,500 mL) obtained in the hydrochloric acid leaching step to neutralize to pH 1.5, 1.5 L of H 2 S gas was blown into this solution. Impurities such as Cd were precipitated as sulfides. The reaction temperature was 40 ° C. and the reaction time was 0.5 hour. The suspension after the reaction was filtered, the cake was a cadmium residue, and the filtrate was a de-Cd solution. Table 9 shows the analysis results and material balance of each product.
（８）電解採取工程
上記（７）の工程で得られた脱Ｃｄ液を電解元液とし、温度４０℃、電流密度１５０Ａ／ｍ２で４８時間電解採取を行った。アノードにはＤＳＡを、カソードにはＴｉ板を使用した。電解元液および得られたインジウムと電解尾液の分析結果と物質収支を表１０に示す。
(8) Electrolytic collection step The de-Cd solution obtained in the step (7) was used as an electrolytic base solution, and electrolytic collection was performed at a temperature of 40 ° C. and a current density of 150 A / m 2 for 48 hours. DSA was used for the anode and a Ti plate was used for the cathode. Table 10 shows the analysis results and material balance of the electrolytic base solution and the obtained indium and electrolytic tail solution.
Claims (4)
前記インジウムを硫酸を含む溶液中に浸出させる浸出工程を含み、インジウム浸出液の硫酸濃度を３５〜８０ｇ／Ｌとした、インジウム含有物からインジウムを回収する方法。 In a method for recovering indium from an indium-containing material containing indium and a metal other than indium,
A method for recovering indium from an indium-containing material, comprising a leaching step of leaching the indium into a solution containing sulfuric acid, wherein the sulfuric acid concentration of the indium leaching solution is 35 to 80 g / L.
The method for recovering indium according to any one of claims 1 to 3, wherein a metal other than indium is separated from the indium leaching solution by sulfurization and reductive leaching, and electrolysis is performed.
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| US20100124344A1 (en) * | 2008-11-17 | 2010-05-20 | Atsuhito Mizutani | Piezoelectric body module and manufacturing method therefor |
| JP2012052215A (en) * | 2010-08-31 | 2012-03-15 | Jx Nippon Mining & Metals Corp | Method for collecting indium |
| JP2016156065A (en) * | 2015-02-25 | 2016-09-01 | Ｊｘ金属株式会社 | Method for recovering indium and gallium |
| CN110106352A (en) * | 2019-05-22 | 2019-08-09 | 中国恩菲工程技术有限公司 | Sponge indium producing equipment |
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| Publication number | Priority date | Publication date | Assignee | Title |
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| US20100124344A1 (en) * | 2008-11-17 | 2010-05-20 | Atsuhito Mizutani | Piezoelectric body module and manufacturing method therefor |
| US8311247B2 (en) * | 2008-11-17 | 2012-11-13 | Panasonic Corporation | Piezoelectric body module and manufacturing method therefor |
| JP2012052215A (en) * | 2010-08-31 | 2012-03-15 | Jx Nippon Mining & Metals Corp | Method for collecting indium |
| JP2016156065A (en) * | 2015-02-25 | 2016-09-01 | Ｊｘ金属株式会社 | Method for recovering indium and gallium |
| CN110106352A (en) * | 2019-05-22 | 2019-08-09 | 中国恩菲工程技术有限公司 | Sponge indium producing equipment |
| CN110106352B (en) * | 2019-05-22 | 2024-01-05 | 中国恩菲工程技术有限公司 | Sponge indium preparation equipment |
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