JP5160163B2 - Tin recovery method - Google Patents
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- JP5160163B2 JP5160163B2 JP2007201656A JP2007201656A JP5160163B2 JP 5160163 B2 JP5160163 B2 JP 5160163B2 JP 2007201656 A JP2007201656 A JP 2007201656A JP 2007201656 A JP2007201656 A JP 2007201656A JP 5160163 B2 JP5160163 B2 JP 5160163B2
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- ATJFFYVFTNAWJD-UHFFFAOYSA-N Tin Chemical compound [Sn] ATJFFYVFTNAWJD-UHFFFAOYSA-N 0.000 title claims description 101
- 238000000034 method Methods 0.000 title claims description 38
- 238000011084 recovery Methods 0.000 title claims description 5
- HEMHJVSKTPXQMS-UHFFFAOYSA-M Sodium hydroxide Chemical compound [OH-].[Na+] HEMHJVSKTPXQMS-UHFFFAOYSA-M 0.000 claims description 153
- 238000002386 leaching Methods 0.000 claims description 80
- 235000011121 sodium hydroxide Nutrition 0.000 claims description 51
- 239000000243 solution Substances 0.000 claims description 35
- 239000000843 powder Substances 0.000 claims description 29
- 239000008187 granular material Substances 0.000 claims description 18
- 239000000463 material Substances 0.000 claims description 13
- 238000005363 electrowinning Methods 0.000 claims description 12
- 239000002245 particle Substances 0.000 claims description 11
- 239000000956 alloy Substances 0.000 claims description 10
- 229910045601 alloy Inorganic materials 0.000 claims description 10
- 230000001590 oxidative effect Effects 0.000 claims description 10
- QVGXLLKOCUKJST-UHFFFAOYSA-N atomic oxygen Chemical compound [O] QVGXLLKOCUKJST-UHFFFAOYSA-N 0.000 claims description 7
- 229910052760 oxygen Inorganic materials 0.000 claims description 7
- 239000001301 oxygen Substances 0.000 claims description 7
- 238000007664 blowing Methods 0.000 claims description 5
- 239000008151 electrolyte solution Substances 0.000 claims description 3
- 230000003647 oxidation Effects 0.000 claims description 2
- 238000007254 oxidation reaction Methods 0.000 claims description 2
- 239000003792 electrolyte Substances 0.000 claims 1
- 238000010298 pulverizing process Methods 0.000 claims 1
- 229910052718 tin Inorganic materials 0.000 description 83
- 229910052745 lead Inorganic materials 0.000 description 17
- 239000007788 liquid Substances 0.000 description 14
- 229910052751 metal Inorganic materials 0.000 description 14
- 239000002184 metal Substances 0.000 description 14
- 229910052787 antimony Inorganic materials 0.000 description 13
- 239000007864 aqueous solution Substances 0.000 description 12
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Substances O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 description 8
- 229940071182 stannate Drugs 0.000 description 6
- 239000002253 acid Substances 0.000 description 5
- 238000006243 chemical reaction Methods 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
- -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
- 238000005868 electrolysis reaction Methods 0.000 description 4
- 239000000155 melt Substances 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
- 125000005402 stannate group Chemical group 0.000 description 3
- 238000006467 substitution reaction Methods 0.000 description 3
- 239000003795 chemical substances by application Substances 0.000 description 2
- 230000007423 decrease Effects 0.000 description 2
- 238000001914 filtration Methods 0.000 description 2
- 239000002994 raw material Substances 0.000 description 2
- 239000004071 soot Substances 0.000 description 2
- 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
- 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
- AUYOHNUMSAGWQZ-UHFFFAOYSA-L dihydroxy(oxo)tin Chemical compound O[Sn](O)=O AUYOHNUMSAGWQZ-UHFFFAOYSA-L 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
- 238000010828 elution 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
- 239000000203 mixture Substances 0.000 description 1
- 239000007800 oxidant agent Substances 0.000 description 1
- 239000002244 precipitate 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
- 239000011734 sodium Substances 0.000 description 1
- 229910000679 solder Inorganic materials 0.000 description 1
- 238000003756 stirring Methods 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
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
したがって、本発明は、このような従来の問題点に鑑み、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 provides a method for recovering tin that can recover tin inexpensively and efficiently from a tin-containing material containing lead in addition to tin. Objective.
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ことにより、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, by using this leaching solution as an electrolytic solution and recovering tin by electrowinning, it was found that tin can be recovered inexpensively and efficiently from a tin-containing material containing lead in addition to tin. It came to complete.
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、30〜80g/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. It is characterized in that tin is recovered by electrowinning. In this tin recovery method, it is preferable to oxidize by blowing oxygen into an aqueous caustic soda solution. 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 30 to 80 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ことができる、錫の回収方法を提供することができる。 ADVANTAGE OF THE INVENTION According to this invention, the tin collection | recovery method which can collect | recover tin efficiently from the tin containing material containing lead etc. besides tin can be provided.
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。なお、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く、浸出終了時の遊離NaOH濃度は、0.1g/L(pH13)〜150g/Lであるのが好ましく、4〜80g/Lであるのがさらに好ましく、30〜80g/Lであるのが最も好ましい。 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. The initial appropriate concentration of free NaOH (initial concentration) is the pulp concentration (g / L) after adding the tin-containing powder or granule to the aqueous caustic soda solution, the tin-containing powder or granule Therefore, it is better to define the free NaOH concentration at the end of leaching, and the free NaOH concentration at the end of leaching is preferably 0.1 g / L (pH 13) to 150 g / L. More preferably, it is 80 g / L, and most preferably 30-80 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.
なお、この浸出後に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 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。
Na2[Sn(OH)4]→Sn+2NaOH+H2O+0.5O2
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.
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、表1に示すように、Sn品位37.55%、Pb品位58.48%、Sb(アンチモン)品位1.40%の合金塊をアトマイズした後、開き目150μmの篩で分級した金属粉を得た。
[Example 1]
First, as shown in Table 1, an alloy lump having an Sn grade of 37.55%, a Pb grade of 58.48%, and an Sb (antimony) grade of 1.40% was atomized and then classified with a sieve having an opening of 150 μm. Got.
次に、この金属粉(Sn品位37.55%(Sn21.0g)、Pb品位58.48%(Pb32.7g)、Sb品位1.40%(Sb0.8g)の合金粉)56gをNaOH濃度100g/Lの苛性ソーダ水溶液700mLに添加してパルプ濃度80g/Lとし、この水溶液の温度を80℃とし、水溶液中に酸素を吹き込みながらタービン羽根で撹拌してSnの酸化浸出を行った。なお、この酸化浸出の際に水分が蒸発して液量が減少するため、減少分の水をその都度追加した。この浸出時間に対する液中のSn、Pb、Sbの濃度を図2に示す。 Next, 56 g of this metal powder (alloy powder of Sn quality 37.55% (Sn 21.0 g), Pb quality 58.48% (Pb 32.7 g), Sb quality 1.40% (Sb 0.8 g)) was added with NaOH concentration. It was added to 700 mL of a 100 g / L aqueous solution of caustic soda to obtain a pulp concentration of 80 g / L. The temperature of this aqueous solution was set to 80 ° C., and oxygen leaching was performed by stirring with turbine blades while blowing oxygen into the aqueous solution. In this oxidation leaching, water is evaporated and the amount of liquid is reduced, so the reduced amount of water was added each time. The concentrations of Sn, Pb, and Sb in the liquid with respect to the leaching time are shown in FIG.
図2に示すように、浸出時間とともに液中のSn濃度は高くなったが、Pb濃度は殆ど変わらなかった。なお、浸出時間とともに液中のSb濃度も高くなったが、Sb濃度はSn濃度の100分の1程度であり、非常に低かった。 As shown in FIG. 2, the Sn concentration in the liquid increased with the leaching time, but the Pb concentration hardly changed. Although the Sb concentration in the liquid increased with the leaching time, the Sb concentration was about 1 / 100th of the Sn concentration and was very low.
この浸出を90分間行った後、ろ過して浸出液を得た。表2に示すように、得られた浸出液中に含まれるSn、Pb、Sb、遊離NaOHの濃度(量)は、それぞれ29.32g/L(20.5g)、0.72g/L(0.5g)、0.29g/L(0.2g)、79g/L(55g)であり、Sn濃度が高く且つPb濃度が低い浸出液が得られた。なお、浸出残渣の量は37.5gであり、浸出残渣中に含まれるSn、Pb、Sbの品位(量)は、それぞれ1.35%(0.51g)、85.98%(32.24g)、1.54%(0.58g)であった。また、Sn、Pb、Sbの浸出率は、それぞれ97.6%、1.5%、26.2%であった。 This leaching was performed for 90 minutes, followed by filtration to obtain a leaching solution. As shown in Table 2, the concentrations (amounts) of Sn, Pb, Sb, and free NaOH contained in the obtained leachate were 29.32 g / L (20.5 g) and 0.72 g / L (0. 5%), 0.29 g / L (0.2 g), 79 g / L (55 g), and a leachate having a high Sn concentration and a low Pb concentration was obtained. The amount of leaching residue is 37.5 g, and the grades (amounts) of Sn, Pb and Sb contained in the leaching residue are 1.35% (0.51 g) and 85.98% (32.24 g), respectively. ), 1.54% (0.58 g). Moreover, the leaching rates of Sn, Pb, and Sb were 97.6%, 1.5%, and 26.2%, respectively.
次に、得られた浸出液にSn粒(99.9%)を添加し、80℃で液面の空気を含まない程度に1時間撹拌し、置換反応によりPbとSbを沈澱させた後、ろ過して700mLの液(置換後浸出液)を得た。表3に示すように、得られた置換後浸出液中に含まれるSn、Pb、Sbの濃度(量)は、それぞれ31.05g/L(21.7g)、<0.001g/L、0.05g/L(0.0g)であり、Pbを殆ど含まない液であった。 Next, Sn particles (99.9%) are added to the obtained leachate, and the mixture is stirred at 80 ° C. for 1 hour so as not to contain air on the liquid surface, and Pb and Sb are precipitated by a substitution reaction, followed by filtration. As a result, 700 mL of liquid (leaching liquid after substitution) was obtained. As shown in Table 3, the concentrations (amounts) of Sn, Pb, and Sb contained in the obtained post-substitution leachate were 31.05 g / L (21.7 g), <0.001 g / L, and. The liquid was 05 g / L (0.0 g) and contained almost no Pb.
次に、アノードおよびカソードとしてSUS304を使用し、得られた液を80℃で電流密度100A/m3で電解採取を行い、表4に示すように、Sn品位、Pb品位、Sb品位がそれぞれ<99.9%、<10ppm、125ppmのSnメタル16.9gを得た。 Next, SUS304 was used as an anode and a cathode, and the obtained liquid was subjected to electrowinning at 80 ° C. and a current density of 100 A / m 3. As shown in Table 4, the Sn quality, Pb quality, and Sb quality were < 19.9 g of 99.9%, <10 ppm, 125 ppm Sn metal was obtained.
また、表5に示すように、電解採取後の液(電解后液)700mL中のSn、Pb、Sb、遊離NaOHは、それぞれ8.7g/L(4.9g)、0g/L(0.0g)、0.015g/L(0.0g)、91g/L(64g)であった。なお、電解採取の際に水分が蒸発して液量が減少するため、減少分の水をその都度追加した。 Moreover, as shown in Table 5, Sn, Pb, Sb, and free NaOH in 700 mL of the solution after electrolytic collection (post-electrolysis solution) were 8.7 g / L (4.9 g) and 0 g / L (0. 0 g), 0.015 g / L (0.0 g), and 91 g / L (64 g). It should be noted that water was evaporated during electrowinning and the amount of liquid was reduced, so the reduced amount of water was added each time.
なお、Snを電解採取した後の電解后液は、NaOHの消耗分を加えて、金属粉の浸出に使用することができる。 In addition, the post-electrolysis solution after electrolytically collecting Sn can be used for leaching of metal powder by adding consumption of NaOH.
[実施例2〜5]
表6に示すように、実施例2では粒径150μm未満の金属粉(Sn品位37.55%、Pb品位58.48%、Sb品位1.40%)、実施例3では粒径150〜400μmの金属粉(Sn品位37.94%、Pb品位58.23%、Sb品位1.38%)、実施例4では粒径400μm〜1mmの金属粉(Sn品位37.94%、Pb品位58.23%、Sb品位1.38%)、実施例5では粒径1.7〜3mmの金属粉(Sn品位49.12%、Pb品位49.17%、Sb品位1.32%)を用意し、これらの金属粉について、NaOH濃度50g/Lの苛性ソーダ水溶液を使用し、その水溶液の温度を60℃にした以外は、実施例1と同様の酸化浸出を行った。浸出時間に対するSnおよびPbの浸出率を、それぞれ図3および図4に示す。なお、実施例2〜5では、浸出終了時の苛性ソーダ水溶液中のNaOH濃度が、それぞれ28.6g/L、29.3g/L、33.0g/L、35.9g/Lであった。
[Examples 2 to 5]
As shown in Table 6, in Example 2, metal powder having a particle size of less than 150 μm (Sn grade 37.55%, Pb grade 58.48%, Sb grade 1.40%), and in Example 3,
図3および図4に示すように、実施例2〜4では、短時間で非常に高いSnの浸出率を得ることができるとともに、Pbの浸出率を抑えることができた。また、実施例5では、浸出時間が400分間程度であれば50%程度のSnの浸出率を得ることができるとともに、Pbの浸出率を抑えることができた。 As shown in FIGS. 3 and 4, in Examples 2 to 4, it was possible to obtain a very high Sn leaching rate in a short time and to suppress the Pb leaching rate. Further, in Example 5, when the leaching time was about 400 minutes, an Sn leaching rate of about 50% could be obtained, and the Pb leaching rate could be suppressed.
[実施例6、7]
実施例2と同様の粒径150μm未満の金属粉(Sn品位37.55%、Pb品位58.48%、Sb品位1.40%)について、それぞれNaOH濃度100g/L(実施例6)および25g/L(実施例7)の苛性ソーダ水溶液を使用した以外は、実施例2と同様の酸化浸出を行った。実施例2、6および7において、浸出時間に対する浸出液中のSnおよびPbの濃度を、それぞれ図5および図6に示す。なお、実施例6および7では、浸出終了時の苛性ソーダ水溶液中のNaOH濃度が、それぞれ74.9g/L、7.6g/Lであった。
[Examples 6 and 7]
About the metal powder (Sn grade 37.55%, Pb grade 58.48%, Sb grade 1.40%) similar to Example 2 with a particle size of less than 150 μm, NaOH concentrations of 100 g / L (Example 6) and 25 g, respectively. The same oxidative leaching as in Example 2 was performed except that an aqueous caustic soda solution of / L (Example 7) was used. In Examples 2, 6 and 7, the concentrations of Sn and Pb in the leachate with respect to the leaching time are shown in FIGS. 5 and 6, respectively. In Examples 6 and 7, the NaOH concentrations in the aqueous caustic soda solution at the end of leaching were 74.9 g / L and 7.6 g / L, respectively.
図5および図6に示すように、実施例6では、90分間浸出を行った後に得られた浸出液中のSn、Pbおよび遊離NaOHの濃度は、それぞれ30.2g/L、13.9g/L、75g/Lであった。この実施例6では、Snの浸出率を実施例2よりも高いほぼ100%にすることができたが、Pbの浸出率が29.7%になり、Pbの浸出率が実施例2よりも高くなった。なお、60分間浸出を行った後に得られた浸出液中のSnおよびPbの濃度は、それぞれ28.85g/L(浸出率96%)、1.49g/L(浸出率3.2%)であり、60分後に浸出を止めれば、Pbの浸出が始まる前にSnの浸出をほぼ終了させることができる。 As shown in FIGS. 5 and 6, in Example 6, the concentrations of Sn, Pb and free NaOH in the leachate obtained after leaching for 90 minutes were 30.2 g / L and 13.9 g / L, respectively. 75 g / L. In Example 6, the Sn leaching rate could be almost 100% higher than that of Example 2, but the Pb leaching rate was 29.7%, and the Pb leaching rate was higher than that of Example 2. It became high. The concentrations of Sn and Pb in the leachate obtained after leaching for 60 minutes were 28.85 g / L (leaching rate 96%) and 1.49 g / L (leaching rate 3.2%), respectively. If the leaching is stopped after 60 minutes, the leaching of Sn can be almost completed before the leaching of Pb starts.
また、実施例7では、90分間浸出を行った後に得られた浸出液中のSn、Pbおよび遊離NaOHの濃度は、それぞれ24.22g/L、0.09g/L、7.6g/Lであった。この実施例7では、Snの浸出率が80%程度であったが、Pbの浸出率を0.2%に抑えることができた。なお、浸出液中にメタスズ酸と考えられる白濁の沈殿物が生じていた。 In Example 7, the concentrations of Sn, Pb and free NaOH in the leachate obtained after leaching for 90 minutes were 24.22 g / L, 0.09 g / L and 7.6 g / L, respectively. It was. In Example 7, the Sn leaching rate was about 80%, but the Pb leaching rate could be suppressed to 0.2%. In addition, the cloudy deposit considered to be metastannic acid was produced in the leachate.
また、80℃で酸化浸出を行った実施例1と、浸出温度60℃とした以外は実施例1と同様の酸化浸出を行った実施例6を比較すると、いずれも90分間浸出を行った後のSnの浸出率を高くすることができたが、実施例6では実施例1よりもPb浸出率も高くなった。しかし、実施例6では、上述したように、60分後に浸出を止めれば、Pbの浸出が始まる前にSnの浸出をほぼ終了させることができる。 Further, comparing Example 1 in which oxidative leaching was performed at 80 ° C. and Example 6 in which oxidative leaching was performed in the same manner as in Example 1 except that the leaching temperature was 60 ° C., both were leached for 90 minutes. Although the Sn leaching rate could be increased, the Pb leaching rate was higher in Example 6 than in Example 1. However, in Example 6, as described above, if leaching is stopped after 60 minutes, the leaching of Sn can be almost completed before the leaching of Pb starts.
Claims (12)
The method for recovering tin according to any one of claims 1 to 11, wherein tin is added to the leachate to remove lead in the leachate before the electrowinning.
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