JP5981821B2 - Tin recovery method - Google Patents
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- JP5981821B2 JP5981821B2 JP2012211634A JP2012211634A JP5981821B2 JP 5981821 B2 JP5981821 B2 JP 5981821B2 JP 2012211634 A JP2012211634 A JP 2012211634A JP 2012211634 A JP2012211634 A JP 2012211634A JP 5981821 B2 JP5981821 B2 JP 5981821B2
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- ATJFFYVFTNAWJD-UHFFFAOYSA-N Tin Chemical compound [Sn] ATJFFYVFTNAWJD-UHFFFAOYSA-N 0.000 title claims description 113
- 238000000034 method Methods 0.000 title claims description 46
- 238000011084 recovery Methods 0.000 title description 3
- HEMHJVSKTPXQMS-UHFFFAOYSA-M Sodium hydroxide Chemical compound [OH-].[Na+] HEMHJVSKTPXQMS-UHFFFAOYSA-M 0.000 claims description 189
- 238000002386 leaching Methods 0.000 claims description 80
- 235000011121 sodium hydroxide Nutrition 0.000 claims description 63
- 239000000243 solution Substances 0.000 claims description 49
- 239000000843 powder Substances 0.000 claims description 28
- 239000008187 granular material Substances 0.000 claims description 22
- 239000007864 aqueous solution Substances 0.000 claims description 20
- QVGXLLKOCUKJST-UHFFFAOYSA-N atomic oxygen Chemical compound [O] QVGXLLKOCUKJST-UHFFFAOYSA-N 0.000 claims description 17
- 229910052760 oxygen Inorganic materials 0.000 claims description 17
- 239000001301 oxygen Substances 0.000 claims description 17
- 239000000463 material Substances 0.000 claims description 15
- 230000001590 oxidative effect Effects 0.000 claims description 14
- 238000007664 blowing Methods 0.000 claims description 12
- 238000005363 electrowinning Methods 0.000 claims description 12
- 229910045601 alloy Inorganic materials 0.000 claims description 9
- 239000000956 alloy Substances 0.000 claims description 9
- 239000008151 electrolyte solution Substances 0.000 claims description 6
- 239000002245 particle Substances 0.000 claims description 5
- 238000010298 pulverizing process Methods 0.000 claims description 2
- 239000003792 electrolyte Substances 0.000 claims 1
- 229910052718 tin Inorganic materials 0.000 description 93
- 230000003647 oxidation Effects 0.000 description 34
- 238000007254 oxidation reaction Methods 0.000 description 34
- 229910052745 lead Inorganic materials 0.000 description 13
- 229910052787 antimony Inorganic materials 0.000 description 12
- 230000007423 decrease Effects 0.000 description 12
- 238000006243 chemical reaction Methods 0.000 description 11
- 229910052751 metal Inorganic materials 0.000 description 9
- 239000002184 metal Substances 0.000 description 9
- 229940071182 stannate Drugs 0.000 description 6
- 239000002253 acid Substances 0.000 description 5
- TVQLLNFANZSCGY-UHFFFAOYSA-N disodium;dioxido(oxo)tin Chemical compound [Na+].[Na+].[O-][Sn]([O-])=O TVQLLNFANZSCGY-UHFFFAOYSA-N 0.000 description 5
- 238000010828 elution Methods 0.000 description 5
- -1 soda compound Chemical class 0.000 description 5
- 229940079864 sodium stannate Drugs 0.000 description 5
- MHAJPDPJQMAIIY-UHFFFAOYSA-N Hydrogen peroxide Chemical compound OO MHAJPDPJQMAIIY-UHFFFAOYSA-N 0.000 description 4
- 239000000155 melt Substances 0.000 description 4
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Substances O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 description 4
- WMFOQBRAJBCJND-UHFFFAOYSA-M Lithium hydroxide Chemical compound [Li+].[OH-] WMFOQBRAJBCJND-UHFFFAOYSA-M 0.000 description 3
- KWYUFKZDYYNOTN-UHFFFAOYSA-M Potassium hydroxide Chemical compound [OH-].[K+] KWYUFKZDYYNOTN-UHFFFAOYSA-M 0.000 description 3
- 239000003513 alkali Substances 0.000 description 3
- 239000012670 alkaline solution Substances 0.000 description 3
- 239000000498 cooling water Substances 0.000 description 3
- 238000005868 electrolysis reaction Methods 0.000 description 3
- 239000007788 liquid Substances 0.000 description 3
- 125000005402 stannate group Chemical group 0.000 description 3
- 229910052785 arsenic Inorganic materials 0.000 description 2
- 239000003795 chemical substances by application Substances 0.000 description 2
- 229910052802 copper Inorganic materials 0.000 description 2
- 239000010949 copper Substances 0.000 description 2
- 230000003247 decreasing effect Effects 0.000 description 2
- 239000002244 precipitate Substances 0.000 description 2
- 239000002994 raw material Substances 0.000 description 2
- 239000004071 soot Substances 0.000 description 2
- RYGMFSIKBFXOCR-UHFFFAOYSA-N Copper Chemical compound [Cu] RYGMFSIKBFXOCR-UHFFFAOYSA-N 0.000 description 1
- 238000003723 Smelting Methods 0.000 description 1
- NINIDFKCEFEMDL-UHFFFAOYSA-N Sulfur Chemical compound [S] NINIDFKCEFEMDL-UHFFFAOYSA-N 0.000 description 1
- 229910001854 alkali hydroxide Inorganic materials 0.000 description 1
- 150000008044 alkali metal hydroxides Chemical class 0.000 description 1
- WATWJIUSRGPENY-UHFFFAOYSA-N antimony atom Chemical compound [Sb] WATWJIUSRGPENY-UHFFFAOYSA-N 0.000 description 1
- 239000011260 aqueous acid Substances 0.000 description 1
- RQNWIZPPADIBDY-UHFFFAOYSA-N arsenic atom Chemical compound [As] RQNWIZPPADIBDY-UHFFFAOYSA-N 0.000 description 1
- 239000006227 byproduct Substances 0.000 description 1
- 239000011575 calcium Substances 0.000 description 1
- HNQGTZYKXIXXST-UHFFFAOYSA-N calcium;dioxido(oxo)tin Chemical compound [Ca+2].[O-][Sn]([O-])=O HNQGTZYKXIXXST-UHFFFAOYSA-N 0.000 description 1
- 238000005119 centrifugation Methods 0.000 description 1
- 238000007796 conventional method Methods 0.000 description 1
- 239000013078 crystal Substances 0.000 description 1
- 238000002425 crystallisation Methods 0.000 description 1
- 230000008025 crystallization Effects 0.000 description 1
- IOUCSUBTZWXKTA-UHFFFAOYSA-N dipotassium;dioxido(oxo)tin Chemical compound [K+].[K+].[O-][Sn]([O-])=O IOUCSUBTZWXKTA-UHFFFAOYSA-N 0.000 description 1
- 238000004070 electrodeposition Methods 0.000 description 1
- 238000001704 evaporation Methods 0.000 description 1
- 230000008020 evaporation Effects 0.000 description 1
- 238000000227 grinding Methods 0.000 description 1
- QOSATHPSBFQAML-UHFFFAOYSA-N hydrogen peroxide;hydrate Chemical compound O.OO QOSATHPSBFQAML-UHFFFAOYSA-N 0.000 description 1
- MLOKPANHZRKTMG-UHFFFAOYSA-N lead(2+);oxygen(2-);tin(4+) Chemical compound [O-2].[O-2].[O-2].[Sn+4].[Pb+2] MLOKPANHZRKTMG-UHFFFAOYSA-N 0.000 description 1
- 229910052744 lithium Inorganic materials 0.000 description 1
- 229960001078 lithium Drugs 0.000 description 1
- 239000007800 oxidant agent Substances 0.000 description 1
- 238000000746 purification Methods 0.000 description 1
- 238000007670 refining Methods 0.000 description 1
- 239000012266 salt solution Substances 0.000 description 1
- 238000005070 sampling Methods 0.000 description 1
- 239000011734 sodium Substances 0.000 description 1
- 229910000679 solder Inorganic materials 0.000 description 1
- 239000000126 substance Substances 0.000 description 1
- 229910052717 sulfur Inorganic materials 0.000 description 1
- 239000011593 sulfur Substances 0.000 description 1
- CVNKFOIOZXAFBO-UHFFFAOYSA-J tin(4+);tetrahydroxide Chemical compound [OH-].[OH-].[OH-].[OH-].[Sn+4] CVNKFOIOZXAFBO-UHFFFAOYSA-J 0.000 description 1
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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)
- Electrolytic Production Of Metals (AREA)
Description
本発明は、錫の回収方法に関し、特に、錫の他に鉛などを含む錫含有物から錫を回収する方法に関する。 The present invention relates to a method for recovering tin, and more particularly to a method for recovering tin from a tin-containing material containing lead in addition to tin.
従来、錫の他に鉛などを含む錫含有物から錫を回収する方法として、錫と鉛を含む合金を３００〜５００℃程度の温度で溶融し、この溶融体に苛性ソーダを添加し、溶融体中の錫を錫酸ナトリウムとして苛性ソーダ中に抽出して鉛と分離した後、錫を抽出した苛性ソーダを水で溶解して、錫酸ナトリウムを含むアルカリ性溶液とし、電解などにより錫を回収する方法が知られている。 Conventionally, as a method of recovering tin from a tin-containing material containing lead in addition to tin, an alloy containing tin and lead is melted at a temperature of about 300 to 500 ° C., caustic soda is added to the melt, and the melt There is a method in which tin is extracted into sodium stannate as sodium stannate and separated from lead, and then the caustic soda extracted from tin is dissolved in water to form an alkaline solution containing sodium stannate, and tin is recovered by electrolysis or the like. Are known.
また、鉛精錬において鉛中の錫を分離回収する方法として、錫を含む鉛を溶融し、この溶融体をソーダ化合物と反応させて溶融体中の錫を錫酸ソーダとし、この錫酸ソーダと副生する錫酸鉛とを含む滓を溶融鉛から分離し、この鉛精製滓に硫黄を添加して水で浸出処理した後にＣａ２＋で錫を錫酸カルシウムとして沈澱させる方法も提案されている（例えば、特許文献１参照）。 Further, as a method for separating and recovering tin in lead in lead refining, lead containing tin is melted, and this melt is reacted with a soda compound, so that tin in the melt is converted to sodium stannate. A method has also been proposed in which soot containing by-product lead stannate is separated from molten lead, sulfur is added to the lead refined soot and leached with water, and then tin is precipitated as calcium stannate with Ca 2+ . (For example, refer to Patent Document 1).
また、水酸化カリウム、水酸化ナトリウム及び水酸化リチウムのいずれかの水酸化アルカリ水溶液中に金属錫や錫を含む合金などの原料錫を投入し、水酸化アルカリ水溶液を撹拌または循環により原料錫の表面上に常時流動させながら所定の反応温度に維持し、且つ反応液中に反応促進剤として過酸化水素を滴下しながら反応を行って不溶解分を含む錫酸塩水溶液を得た後、不溶解分を濾別し、錫酸塩水溶液から減圧濃縮、蒸発、晶析又は遠心分離等により、錫酸カリウム、錫酸ナトリウム及び錫酸リチウムのいずれかの錫酸アルカリ化合物の結晶を得る方法が提案されている（例えば、特許文献２参照）。 Also, raw material tin such as an alloy containing metal tin or tin is put into an alkali hydroxide aqueous solution of any one of potassium hydroxide, sodium hydroxide and lithium hydroxide, and the aqueous tin hydroxide aqueous solution is stirred or circulated. The reaction is carried out while dripping hydrogen peroxide as a reaction accelerator in the reaction solution while maintaining a predetermined reaction temperature while constantly flowing on the surface to obtain a stannate aqueous solution containing an insoluble matter. There is a method in which the dissolved matter is filtered off and a crystal of an alkali stannate compound of any one of potassium stannate, sodium stannate and lithium stannate is obtained by concentration under reduced pressure, evaporation, crystallization or centrifugation from an aqueous stannate solution. It has been proposed (see, for example, Patent Document 2).
さらに、錫を主成分とする半田などから金属錫を回収する方法として、錫含有物に酸を添加して錫含有物を溶解した酸溶液とした後、この酸溶液にアルカリ剤を添加してｐＨ１２以上に調整することにより、錫含有物中の錫を溶解したアルカリ溶液とし、このアルカリ溶液を電解して錫を得る方法が提案されている（例えば、特許文献３参照）。 Furthermore, as a method of recovering metallic tin from solder containing tin as a main component, an acid is added to the tin-containing material to obtain an acid solution in which the tin-containing material is dissolved, and then an alkali agent is added to the acid solution. There has been proposed a method in which tin is obtained by adjusting the pH to 12 or more to obtain an alkaline solution in which tin in a tin-containing material is dissolved, and electrolyzing the alkaline solution (see, for example, Patent Document 3).
しかし、上述した錫と鉛を含む合金の溶融体に苛性ソーダを添加する従来の方法では、錫の量に対して１０倍等量程度の苛性ソーダを使用するため、苛性ソーダの使用量が非常に多くなる。また、錫を抽出した苛性ソーダを水で溶解した際に得られる苛性ソーダ水溶液を３５０℃以上で煮詰めて、水分を完全に蒸発させ、苛性ソーダを再利用することも可能であるが、多量の水分を蒸発させなければならないので、エネルギーコストが多大になる。同様に、特許文献１の方法の場合も、ソーダ化合物の使用量が非常に多くなり、ソーダ化合物を再利用しようとすると、エネルギーコストが多大になる。 However, in the conventional method of adding caustic soda to the above-described alloy melt containing tin and lead, caustic soda is used in an amount about 10 times as much as the amount of tin. . In addition, it is possible to boil the caustic soda solution obtained by dissolving tin extracted caustic soda with water at 350 ° C or higher, completely evaporate the water, and reuse the caustic soda. The energy cost becomes large because it has to be made. Similarly, in the case of the method of Patent Document 1, the amount of soda compound used is very large, and the energy cost becomes large when trying to reuse the soda compound.
また、特許文献２の方法は、錫酸アルカリ化合物を得る方法であり、錫を回収する方法ではないが、この方法によって得られた錫酸塩水溶液を利用して錫を回収しても、錫酸塩水溶液を得るまでに長時間を要する。さらに、特許文献２の方法では、過酸化水素水を使用する必要があり、特許文献３の方法では、錫含有物に酸を添加して得られた酸溶液にアルカリ剤を添加する必要があるので、さらに薬品コストを下げることができる方法が望まれている。
Further, the method of
このような従来の課題を解決するため、錫の他に鉛などを含む錫含有物から安価且つ効率的に錫を回収する方法として、錫と鉛を含む錫含有物の粉末または粒状物を、苛性ソーダ水溶液中で酸化しながら浸出して錫を含む浸出液を得た後、この浸出液を電解液として使用して電解採取により錫を回収する方法が提案されている（例えば、特許文献４参照）。 In order to solve such a conventional problem, as a method for recovering tin from a tin-containing material containing lead in addition to tin at a low cost, a tin-containing powder or granular material containing tin and lead, A method has been proposed in which a leach solution containing tin is obtained by leaching while oxidizing in an aqueous caustic soda solution, and then tin is recovered by electrowinning using this leach solution as an electrolyte solution (see, for example, Patent Document 4).
特許文献４の方法では、ＳｎとＰｂ含むＳｎ含有物の粉末または粒状物を、苛性ソーダ水溶液中で酸化しながら浸出してＳｎを含む浸出液を得る際に、アルカリ領域においてＳｎとＰｂの酸化還元電位の相違からＳｎが優先して浸出されるが、過剰に酸化浸出が行われるとＰｂの溶出量が増加したり、さらに過剰に酸化が進むとＳｎがＰｂと共沈してしまうため、Ｓｎの浸出が終了した段階で速やかに酸化浸出を終了する必要がある。また、必要以上に酸化浸出を続けると、時間や酸素のロスにもなるため、酸化浸出の終点を見極めることは非常に重要である。
In the method of
したがって、本発明は、このような従来の問題点に鑑み、錫と鉛を含む錫含有物の粉末または粒状物を、苛性ソーダ水溶液中で酸化しながら浸出して錫を含む浸出液を得た後、この浸出液を電解液として使用して電解採取により錫を回収する方法において、錫の浸出が終了した段階で速やかに酸化浸出を終了することができる、錫の回収方法を提供することを目的とする。 Therefore, in view of such a conventional problem, the present invention, after leaching a tin-containing powder or granule containing tin and lead in an aqueous caustic soda solution to obtain a leachate containing tin, An object of the present invention is to provide a method for recovering tin capable of quickly terminating oxidation leaching at the stage where the leaching of tin is completed in a method for recovering tin by electrowinning using this leaching solution as an electrolytic solution. .
本発明者らは、上記課題を解決するために鋭意研究した結果、錫と鉛を含む錫含有物の粉末または粒状物を、苛性ソーダ水溶液中で酸化しながら浸出して錫を含む浸出液を得た後、この浸出液を電解液として使用して電解採取により錫を回収する方法において、錫を含む浸出液を得る際に、苛性ソーダ水溶液の温度の上昇が停滞または温度が低下し始めたときに、酸化を停止することにより、錫の浸出が終了した段階で速やかに酸化浸出を終了することができる、錫の回収方法を提供することができることを見出し、本発明を完成するに至った。 As a result of diligent research to solve the above-mentioned problems, the present inventors obtained a leaching solution containing tin by leaching a tin-containing powder or granule containing tin and lead in an aqueous caustic soda solution while oxidizing. Later, in the method of recovering tin by electrowinning using this leaching solution as an electrolytic solution, when obtaining a leaching solution containing tin, when the rise in temperature of the caustic soda aqueous solution stagnates or the temperature begins to decrease, oxidation is performed. By stopping, it was found that it is possible to provide a method for recovering tin that can quickly complete oxidation leaching at the stage where the leaching of tin is completed, and the present invention has been completed.
すなわち、本発明による錫の回収方法は、錫と鉛を含む錫含有物の粉末または粒状物を、苛性ソーダ水溶液中で酸化しながら浸出して錫を含む浸出液を得た後、この浸出液を電解液として使用して電解採取により錫を回収する方法において、錫を含む浸出液を得る際に、苛性ソーダ水溶液の温度の上昇が停滞または温度が低下し始めたときに、酸化を停止することを特徴とする。この錫の回収方法において、苛性ソーダ水溶液中に酸素を吹き込むことによって酸化を行うのが好ましく、酸素の吹き込みを停止することによって酸化を停止するのが好ましい。また、錫と鉛を含む錫含有物の粉末または粒状物が、錫と鉛を含む合金塊から得られた粉末または粒状物であるのが好ましく、この錫と鉛を含む合金塊をアトマイズまたは粉砕することによって粉末または粒状物を得るのが好ましい。また、錫と鉛を含む錫含有物の粉末または粒状物の粒径が３ｍｍ以下であるのが好ましく、１ｍｍ以下であるのがさらに好ましい。また、浸出が終了した際の苛性ソーダ水溶液中のＮａＯＨ濃度が０．１〜１５０ｇ／Ｌであるのが好ましく、４〜８０ｇ／Ｌであるのがさらに好ましく、４〜３５ｇ／Ｌであるのが最も好ましい。また、浸出の際の苛性ソーダ水溶液の温度が５０〜１００℃であるのが好ましく、電解採取の際の電解液の温度が５０〜１００℃であるのが好ましい。さらに、電解採取前に浸出液に錫を添加して浸出液中の鉛を除去するのが好ましい。 That is, in the method for recovering tin according to the present invention, a tin-containing powder or granular material containing tin and lead is leached while oxidizing in an aqueous caustic soda solution to obtain a leachate containing tin. In the method of recovering tin by electrowinning, the oxidation is stopped when the temperature rise of the aqueous caustic soda solution stagnates or begins to decrease when obtaining a leachate containing tin. . In this tin recovery method, the oxidation is preferably performed by blowing oxygen into an aqueous caustic soda solution, and the oxidation is preferably stopped by stopping the blowing of oxygen. The tin-containing powder or granule containing tin and lead is preferably a powder or granule obtained from an alloy mass containing tin and lead, and the alloy mass containing tin and lead is atomized or pulverized. It is preferable to obtain a powder or a granular material. Further, the particle size of the tin-containing powder or granular material containing tin and lead is preferably 3 mm or less, and more preferably 1 mm or less. Further, the NaOH concentration in the aqueous caustic soda when leaching is completed is preferably 0.1 to 150 g / L, more preferably 4 to 80 g / L, and most preferably 4 to 35 g / L. preferable. Moreover, it is preferable that the temperature of the caustic soda aqueous solution in the case of leaching is 50-100 degreeC, and it is preferable that the temperature of the electrolyte solution in the case of electrowinning is 50-100 degreeC. Furthermore, it is preferable to add tin to the leachate to remove lead in the leachate before electrowinning.
本発明によれば、錫と鉛を含む錫含有物の粉末または粒状物を、苛性ソーダ水溶液中で酸化しながら浸出して錫を含む浸出液を得た後、この浸出液を電解液として使用して電解採取により錫を回収する方法において、錫の浸出が終了した段階で速やかに酸化浸出を終了することができる。 According to the present invention, a tin-containing powder or granular material containing tin and lead is leached while oxidizing in an aqueous caustic soda solution to obtain a leaching solution containing tin, and then the leaching solution is used as an electrolytic solution for electrolysis. In the method of recovering tin by sampling, oxidation leaching can be completed promptly when the leaching of tin is completed.
以下、図１を参照して本発明による錫の回収方法の実施の形態について説明する。 Hereinafter, an embodiment of a method for recovering tin according to the present invention will be described with reference to FIG.
まず、錫（Ｓｎ）と鉛（Ｐｂ）を含む錫含有物が塊状の場合には、微細化して錫含有物の粉末または粒状物を得る。錫含有物が錫と鉛を含有する合金塊の場合には、アトマイズや粉砕などによって粉末にして金属粉を得る。この錫含有物の粉末または粒状物の粒径は３ｍｍ以下であるのが好ましく、１ｍｍ以下であるのがさらに好ましい。 First, when the tin-containing material containing tin (Sn) and lead (Pb) is in the form of a lump, it is refined to obtain a powder or granular material of the tin-containing material. In the case where the tin-containing material is an alloy lump containing tin and lead, metal powder is obtained by atomizing or grinding. The particle diameter of the tin-containing powder or granule is preferably 3 mm or less, and more preferably 1 mm or less.
次に、得られた錫含有物の粉末または粒状物を苛性ソーダ水溶液に添加して、この水溶液中に酸素を吹き込みながら撹拌して酸化浸出によりＳｎを選択的に浸出する。なお、この水溶液中に空気を吹き込んでＳｎの浸出を行うこともできるが、ＮａＯＨの炭酸化を防ぐために酸素を吹き込むのが好ましい。 Next, the obtained tin-containing powder or granule is added to an aqueous caustic soda solution and stirred while blowing oxygen into the aqueous solution, and Sn is selectively leached by oxidative leaching. Note that Sn can be leached out by blowing air into the aqueous solution, but oxygen is preferably blown in order to prevent carbonation of NaOH.
なお、錫含有物の粉末または粒状物の粒径が数ｍｍ以上であると、酸化浸出の際に水溶液中に酸素を吹き込んでもＳｎの浸出速度が遅く、酸化浸出に長時間を要し、また、Ｓｎの浸出率が５０％を超える前にＰｂの浸出が起こる。このような場合、十分な浸出速度を得るために、過酸化水素などの酸化剤を添加する必要がある。 If the particle size of the tin-containing powder or granule is several mm or more, the leaching rate of Sn is slow even if oxygen is blown into the aqueous solution during the oxidative leaching, and the oxidative leaching takes a long time. , Pb leaching occurs before the Sn leaching rate exceeds 50%. In such a case, it is necessary to add an oxidizing agent such as hydrogen peroxide in order to obtain a sufficient leaching rate.
この浸出に使用する苛性ソーダ水溶液中の遊離ＮａＯＨ濃度（初期濃度）は、１０〜２００ｇ／Ｌであるのが好ましく、５０〜１００ｇ／Ｌであるのがさらに好ましい。苛性ソーダ水溶液中の遊離ＮａＯＨ濃度が低いと、浸出途中でＳｎが酸化物として沈澱して浸出率が低下し、苛性ソーダ水溶液中の遊離ＮａＯＨ濃度が高いと、Ｐｂの溶解度が上がってＰｂの溶出量が増大することにより浄液コストが増大するとともに、Ｓｎの溶解度が低下し、電解採取時の電流効率も悪くなる。 The free NaOH concentration (initial concentration) in the aqueous caustic soda solution used for the leaching is preferably 10 to 200 g / L, and more preferably 50 to 100 g / L. When the free NaOH concentration in the aqueous caustic soda solution is low, Sn precipitates as an oxide during the leaching, and the leaching rate decreases. When the free NaOH concentration in the aqueous caustic soda solution is high, the solubility of Pb increases and the amount of Pb eluted increases. By increasing, the cost of liquid purification increases, the solubility of Sn decreases, and the current efficiency during electrowinning also deteriorates.
初期の適正な遊離ＮａＯＨの濃度（初期濃度）は、錫含有物の粉末または粒状物を苛性ソーダ水溶液に添加した後のパルプ濃度（ｇ／Ｌ）や、錫含有物の粉末または粒状物中のＳｎ品位によって異なるので、浸出終了時の遊離ＮａＯＨ濃度を規定する方がよい。この浸出終了時の遊離ＮａＯＨ濃度は、０．１ｇ／Ｌ（ｐＨ１３）〜１５０ｇ／Ｌであるのが好ましい。また、浸出中には、遊離ＮａＯＨ濃度が高くても、Ｐｂの溶出が抑えられているが、浸出終了後には、遊離ＮａＯＨ濃度に対応した溶解度までＰｂの溶出が進むので、浸出終了時のＰｂの溶解度の上昇を抑えるためには、浸出終了時の遊離ＮａＯＨ濃度は、４〜８０ｇ／Ｌであるのが好ましく、４〜３５ｇ／Ｌであるのがさらに好ましく、４〜２０ｇ／Ｌであるのが最も好ましい。 The initial appropriate concentration of free NaOH (initial concentration) depends on the pulp concentration (g / L) after the tin-containing powder or granule is added to the aqueous caustic soda solution, and Sn in the tin-containing powder or granule. Since it depends on the quality, it is better to define the free NaOH concentration at the end of leaching. The free NaOH concentration at the end of the leaching is preferably 0.1 g / L (pH 13) to 150 g / L. Further, during the leaching, the elution of Pb is suppressed even if the free NaOH concentration is high, but after the leaching is completed, the elution of Pb proceeds to the solubility corresponding to the free NaOH concentration. In order to suppress the increase in the solubility of the leaching, the free NaOH concentration at the end of the leaching is preferably 4 to 80 g / L, more preferably 4 to 35 g / L, and 4 to 20 g / L. Is most preferred.
なお、初期の遊離ＮａＯＨ濃度と浸出終了時の遊離ＮａＯＨ濃度は、［初期の遊離ＮａＯＨ濃度（ｇ／Ｌ）］＝［錫含有物をＮａＯＨ水溶液に添加した後のパルプ濃度（ｇ／Ｌ）］×［錫含有物中のＳｎ品位（％）／１００］×［（実績により得られた想定の）浸出率（％）／１００］÷（Ｓｎの原子量１１８．７）×（ＮａＯＨの分子量４０）×（Ｓｎ１モルに対するＮａＯＨの消費モル数２）＋［浸出終了時の遊離ＮａＯＨ濃度（ｇ／Ｌ）］により調整することができる。 The initial free NaOH concentration and the free NaOH concentration at the end of leaching are [initial free NaOH concentration (g / L)] = [pulp concentration after adding tin-containing material to NaOH aqueous solution (g / L)] X [Sn grade in tin-containing material (%) / 100] x [leaching rate (%) / 100] (assumed by actual results) / (Sn atomic weight 118.7) x (NaOH molecular weight 40) It can be adjusted by × (Na2 mole consumption of NaOH relative to 1 mol of Sn) + [Free NaOH concentration at the end of leaching (g / L)].
また、この浸出の際の水溶液の温度は、５０〜１００℃であるのが好ましく、７０〜９０℃であるのがさらに好ましい。この温度が低いとＳｎの浸出終了前にＰｂの溶出が開始し、温度が高いとエネルギーコストが増大する。 Further, the temperature of the aqueous solution at the time of leaching is preferably 50 to 100 ° C, and more preferably 70 to 90 ° C. When this temperature is low, elution of Pb starts before the end of Sn leaching, and when the temperature is high, the energy cost increases.
この酸化浸出では、Ｓｎの浸出に伴う酸化熱により溶液の温度が上昇するが、Ｓｎの浸出が終了して溶液中のＳｎ濃度の上昇が止まると、反応熱が急激に減少して、溶液の温度の上昇が停滞または溶液の温度が下降に転ずる。そのため、この酸化浸出の溶液の温度を測定し、溶液の温度の上昇が停滞し始めたとき、または溶液の温度が低下し始めたときに、酸素の吹き込みを停止する。このときに酸素の吹き込みを停止すれば、過剰な酸化浸出によるＰｂの溶出を最小限に抑えることができ且つＳｎを十分に浸出することができるタイミングで反応を停止することができる。 In this oxidative leaching, the temperature of the solution rises due to the oxidation heat accompanying the leaching of Sn, but when the leaching of Sn is finished and the increase in the Sn concentration in the solution stops, the reaction heat decreases rapidly, The rise in temperature stagnates or the temperature of the solution starts to fall. Therefore, the temperature of the solution of this oxidative leaching is measured, and when the increase in the temperature of the solution starts to stagnate or when the temperature of the solution starts to decrease, the blowing of oxygen is stopped. If the blowing of oxygen is stopped at this time, the reaction can be stopped at a timing at which Pb elution due to excessive oxidation leaching can be minimized and Sn can be sufficiently leached.
なお、この酸化浸出後にＰｂメタルが残渣として残り、このＰｂメタルを鉛製錬原料として利用することができる。 In addition, after this oxidation leaching, Pb metal remains as a residue, and this Pb metal can be used as a lead smelting raw material.
次に、この浸出により得られたＳｎを含む浸出液に（粉末、ショット、板などの）Ｓｎを投入して、セメンテーションにより浸出液中のＰｂを除去する。 Next, Sn (powder, shot, plate, etc.) is introduced into the leachate containing Sn obtained by this leaching, and Pb in the leachate is removed by cementation.
次に、得られた液を用いて電解採取によりＳｎメタルを回収する。なお、この電解採取の際の液の温度は、５０〜１００℃であるのが好ましく、７０〜９０℃であるのがさらに好ましい。５０℃より低いと殆ど電着せず、また、７０℃より低いと電流効率が悪くなる。また、Ｓｎを電解採取した後の電解后液は、以下の反応によって苛性ソーダ水溶液を再生するため、Ｓｎの浸出に繰り返し使用することができる。 Next, Sn metal is collect | recovered by electrowinning using the obtained liquid. In addition, the temperature of the liquid at the time of this electrowinning is preferably 50 to 100 ° C, and more preferably 70 to 90 ° C. If it is lower than 50 ° C., almost no electrodeposition is performed, and if it is lower than 70 ° C., current efficiency is deteriorated. Further, the post-electrolysis solution after electrolytically collecting Sn can be used repeatedly for Sn leaching because it regenerates the aqueous caustic soda solution by the following reaction.
Ｎａ２［Ｓｎ（ＯＨ）４］→Ｓｎ＋２ＮａＯＨ＋Ｈ２Ｏ＋０．５Ｏ２ Na 2 [Sn (OH) 4 ] → Sn + 2NaOH + H 2 O + 0.5O 2
以下、本発明による錫の回収方法の実施例について詳細に説明する。 Examples of the method for recovering tin according to the present invention will be described in detail below.
［実施例１］
まず、Ｓｎ品位４３．４４％、Ｐｂ品位４９．１２％、Ｓｂ（アンチモン）品位２．６８％、Ｃｕ（銅）品位１．０１％、Ａｓ（砒素）０．４６％の合金塊をエアアトマイズ法により微粉化した後、開き目１５０μｍの篩で分級して金属粉を得た。
[Example 1]
First, air atomize an alloy lump with Sn grade 43.44%, Pb grade 49.12%, Sb (antimony) grade 2.68%, Cu (copper) grade 1.01%, As (arsenic) 0.46% After pulverization by the method, classification was performed with a sieve having an opening of 150 μm to obtain metal powder.
次に、この金属粉１３．５ｋｇを遊離ＮａＯＨ濃度５５．４ｇ／Ｌの苛性ソーダ水溶液９０Ｌに添加し、水溶液中に１５Ｌ／分の流量で酸素を吹き込みながら２段タービン羽根により２８０ｒｐｍ（５２．５Ｈｚ）で撹拌してＳｎの酸化浸出を１４０分間行った。なお、このＳｎの酸化浸出の開始時の水溶液の温度は４３．６℃であった。 Next, 13.5 kg of this metal powder is added to 90 L of a caustic soda aqueous solution having a free NaOH concentration of 55.4 g / L, and oxygen is blown into the aqueous solution at a flow rate of 15 L / min, and 280 rpm (52.5 Hz) by a two-stage turbine blade. And oxidative leaching of Sn was performed for 140 minutes. The temperature of the aqueous solution at the start of the oxidation leaching of Sn was 43.6 ° C.
このＳｎの酸化浸出の開始後、２０分毎に水溶液の温度を測定するとともに、２０分毎に水溶液の一部を抽出してＳｎ、ＰｂおよびＳｂの濃度を測定した。これらの結果を図２に示す。なお、Ｓｎの酸化浸出の開始から１４０分経過後の遊離ＮａＯＨ濃度は、Ｓｎの酸化浸出により消費されて１６．８ｇ／Ｌまで低下していた。 After the start of the oxidation leaching of Sn, the temperature of the aqueous solution was measured every 20 minutes, and a part of the aqueous solution was extracted every 20 minutes to measure the concentrations of Sn, Pb and Sb. These results are shown in FIG. The free NaOH concentration after 140 minutes from the start of the oxidation leaching of Sn was consumed by the oxidation leaching of Sn and decreased to 16.8 g / L.
図２に示すように、Ｓｎの酸化浸出の開始から８０分経過後までは温度が上昇しているが、その後に温度が低下している。また、８０分経過後にはＳｎの濃度も殆ど上昇しておらず、ＰｂとＳｂの濃度も低かった。これらの結果からわかるように、温度の上昇が停滞または低下し始めた段階で酸素の吹き込みを止めれば、浸出液中のＰｂの濃度を低く抑えることができる。 As shown in FIG. 2, the temperature increases until 80 minutes after the start of the oxidation leaching of Sn, but the temperature decreases thereafter. Further, after 80 minutes, the Sn concentration hardly increased and the Pb and Sb concentrations were low. As can be seen from these results, the concentration of Pb in the leachate can be kept low by stopping the blowing of oxygen at the stage where the rise in temperature starts to stagnate or decrease.
［実施例２］
金属粉の重量を１４．４ｋｇ、遊離ＮａＯＨ濃度を５８．２ｇ／Ｌ、酸素の流量を２０Ｌ／分とした以外は、実施例１と同様の方法により、Ｓｎの酸化浸出を１６０分間行い、水溶液の温度とＳｎ、ＰｂおよびＳｂの濃度を測定した。なお、水溶液の温度が８０℃前後に保たれるように、８２℃以上になると反応槽に冷却水を流し、７８℃以下になると冷却水の供給を停止するようにした。これらの結果を図３に示す。なお、Ｓｎの酸化浸出の開始から１４０分経過後の遊離ＮａＯＨ濃度は、Ｓｎの酸化浸出により消費されて１３．８ｇ／Ｌまで低下していた。
[Example 2]
Except that the weight of the metal powder was 14.4 kg, the free NaOH concentration was 58.2 g / L, and the flow rate of oxygen was 20 L / min, oxidation leaching of Sn was performed for 160 minutes by the same method as in Example 1 to obtain an aqueous solution. And the concentrations of Sn, Pb and Sb were measured. In order to keep the temperature of the aqueous solution at around 80 ° C., the cooling water was allowed to flow into the reaction vessel when the temperature was 82 ° C. or higher, and the cooling water supply was stopped when the temperature was 78 ° C. or lower. These results are shown in FIG. The free NaOH concentration after 140 minutes from the start of the oxidation leaching of Sn was consumed by the oxidation leaching of Sn and was reduced to 13.8 g / L.
図３に示すように、Ｓｎの酸化浸出の開始から１００分経過後には、冷却水の供給を停止しても、温度が上昇しなくなって下降に転じている。また、１００分経過後にはＳｎの濃度も上昇しておらず、Ｓｎの酸化浸出反応が終了していたと考えられ、また、ＰｂとＳｂの濃度も低かった。これらの結果からわかるように、温度の上昇が停滞または低下し始めた段階で酸素の吹き込みを止めれば、浸出液中のＰｂの濃度を低く抑えることができる。 As shown in FIG. 3, after 100 minutes from the start of the oxidation leaching of Sn, even if the supply of the cooling water is stopped, the temperature does not increase and starts to decrease. Further, after 100 minutes, the Sn concentration did not increase, it was considered that the Sn leaching reaction was completed, and the Pb and Sb concentrations were low. As can be seen from these results, the concentration of Pb in the leachate can be kept low by stopping the blowing of oxygen at the stage where the rise in temperature starts to stagnate or decrease.
［実施例３］
金属粉の重量を１１．７ｋｇ、遊離ＮａＯＨ濃度を７２．４ｇ／Ｌ、酸素の流量を２０Ｌ／分とした以外は、実施例１と同様の方法により、Ｓｎの酸化浸出を１００分間行い、水溶液の温度とＳｎ、ＰｂおよびＳｂの濃度を測定した。これらの結果を図４に示す。なお、Ｓｎの酸化浸出の開始から１４０分経過後の遊離ＮａＯＨ濃度は、Ｓｎの酸化浸出により消費されて３７．６ｇ／Ｌまで低下していた。
[Example 3]
Except that the weight of the metal powder was 11.7 kg, the free NaOH concentration was 72.4 g / L, and the flow rate of oxygen was 20 L / min, oxidation leaching of Sn was performed for 100 minutes in the same manner as in Example 1 to obtain an aqueous solution. And the concentrations of Sn, Pb and Sb were measured. These results are shown in FIG. It should be noted that the
図４に示すように、Ｓｎの酸化浸出の開始から７０分経過後までは温度が上昇しているが、その後に温度が低下している。また、７０分経過後にはＳｎの濃度も殆ど上昇しておらず、Ｓｎの酸化浸出反応が終了していたと考えられ、また、ＰｂとＳｂの濃度も低かった。これらの結果からわかるように、温度の上昇が停滞または低下し始めた段階で酸素の吹き込みを止めれば、浸出液中のＰｂの濃度を低く抑えることができる。但し、本実施例では、遊離ＮａＯＨ濃度が高過ぎたため、浸出液中のＰｂの濃度が若干上がってしまった。 As shown in FIG. 4, the temperature rises after 70 minutes from the start of the oxidation leaching of Sn, but thereafter the temperature falls. Further, after 70 minutes, the Sn concentration hardly increased, it was considered that the Sn leaching reaction was completed, and the Pb and Sb concentrations were low. As can be seen from these results, the concentration of Pb in the leachate can be kept low by stopping the blowing of oxygen at the stage where the rise in temperature starts to stagnate or decrease. However, in this example, since the free NaOH concentration was too high, the concentration of Pb in the leachate slightly increased.
［実施例４］
遊離ＮａＯＨ濃度を３４．０ｇ／Ｌとした以外は、実施例１と同様の方法により、Ｓｎの酸化浸出を１００分間行い、水溶液の温度とＳｎ、ＰｂおよびＳｂの濃度を測定した。これらの結果を図５に示す。なお、Ｓｎの酸化浸出の開始から１４０分経過後の遊離ＮａＯＨ濃度は、Ｓｎの酸化浸出により消費されて３．２ｇ／Ｌまで低下していた。
[Example 4]
Except for the free NaOH concentration of 34.0 g / L, Sn was leached for 100 minutes in the same manner as in Example 1, and the temperature of the aqueous solution and the concentrations of Sn, Pb, and Sb were measured. These results are shown in FIG. The free NaOH concentration after 140 minutes from the start of the oxidation leaching of Sn was consumed by the oxidation leaching of Sn and was reduced to 3.2 g / L.
図５に示すように、Ｓｎの酸化浸出の開始から４０分経過後までは温度が上昇しているが、その後に温度が低下している。また、４０分経過後にはＳｎ濃度も殆ど上昇しておらず、Ｓｎの酸化浸出反応が終了していたと考えられ、また、ＰｂとＳｂの濃度も低かった。これらの結果からわかるように、温度の上昇が停滞または低下し始めた段階で酸素の吹き込みを止めれば、浸出液中のＰｂの濃度を低く抑えることができる。但し、本実施例では、遊離ＮａＯＨ濃度が低過ぎたため、浸出液中のＳｎの濃度が低いままその濃度の上昇が停滞し、十分なＳｎを浸出することができなかった。また、酸化浸出を続けると、コロイド状の白濁した沈殿物が発生してろ過性が悪かった。 As shown in FIG. 5, the temperature rises after 40 minutes from the start of the oxidation leaching of Sn, but thereafter the temperature falls. Further, after 40 minutes, the Sn concentration hardly increased, it was considered that the Sn leaching reaction was completed, and the concentrations of Pb and Sb were low. As can be seen from these results, the concentration of Pb in the leachate can be kept low by stopping the blowing of oxygen at the stage where the rise in temperature starts to stagnate or decrease. However, in this example, since the concentration of free NaOH was too low, the increase in the concentration stagnated while the concentration of Sn in the leachate was low, and sufficient Sn could not be leached. Further, when the oxidative leaching was continued, colloidal white turbid precipitates were generated and the filterability was poor.
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| JP5188768B2 (en) * | 2007-09-20 | 2013-04-24 | Ｄｏｗａメタルマイン株式会社 | Tin recovery method |
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