JPS6319576B2 - - Google Patents
Info
- Publication number
- JPS6319576B2 JPS6319576B2 JP8071080A JP8071080A JPS6319576B2 JP S6319576 B2 JPS6319576 B2 JP S6319576B2 JP 8071080 A JP8071080 A JP 8071080A JP 8071080 A JP8071080 A JP 8071080A JP S6319576 B2 JPS6319576 B2 JP S6319576B2
- Authority
- JP
- Japan
- Prior art keywords
- tin
- lead
- stannate
- melt
- refining
- Prior art date
- Legal status (The legal status is an assumption and is not a legal conclusion. Google has not performed a legal analysis and makes no representation as to the accuracy of the status listed.)
- Expired
Links
- ATJFFYVFTNAWJD-UHFFFAOYSA-N Tin Chemical compound [Sn] ATJFFYVFTNAWJD-UHFFFAOYSA-N 0.000 claims description 32
- 238000007670 refining Methods 0.000 claims description 15
- 239000002893 slag Substances 0.000 claims description 12
- 238000000034 method Methods 0.000 claims description 9
- 239000000155 melt Substances 0.000 claims description 6
- NINIDFKCEFEMDL-UHFFFAOYSA-N Sulfur Chemical compound [S] NINIDFKCEFEMDL-UHFFFAOYSA-N 0.000 claims description 5
- HNQGTZYKXIXXST-UHFFFAOYSA-N calcium;dioxido(oxo)tin Chemical compound [Ca+2].[O-][Sn]([O-])=O HNQGTZYKXIXXST-UHFFFAOYSA-N 0.000 claims description 5
- TVQLLNFANZSCGY-UHFFFAOYSA-N disodium;dioxido(oxo)tin Chemical compound [Na+].[Na+].[O-][Sn]([O-])=O TVQLLNFANZSCGY-UHFFFAOYSA-N 0.000 claims description 4
- 238000002386 leaching Methods 0.000 claims description 4
- 229940079864 sodium stannate Drugs 0.000 claims description 4
- 239000011593 sulfur Substances 0.000 claims description 4
- 229910052717 sulfur Inorganic materials 0.000 claims description 4
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Substances O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 claims description 4
- 238000003723 Smelting Methods 0.000 claims description 3
- 239000011575 calcium Substances 0.000 claims description 3
- 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 claims description 2
- -1 soda compound Chemical class 0.000 claims description 2
- 238000011084 recovery Methods 0.000 claims 1
- 238000000926 separation method Methods 0.000 claims 1
- HEMHJVSKTPXQMS-UHFFFAOYSA-M Sodium hydroxide Chemical compound [OH-].[Na+] HEMHJVSKTPXQMS-UHFFFAOYSA-M 0.000 description 9
- 239000011734 sodium Substances 0.000 description 4
- VWDWKYIASSYTQR-UHFFFAOYSA-N sodium nitrate Chemical compound [Na+].[O-][N+]([O-])=O VWDWKYIASSYTQR-UHFFFAOYSA-N 0.000 description 4
- 238000002844 melting Methods 0.000 description 3
- 230000008018 melting Effects 0.000 description 3
- 235000011121 sodium hydroxide Nutrition 0.000 description 3
- AXCZMVOFGPJBDE-UHFFFAOYSA-L calcium dihydroxide Chemical compound [OH-].[OH-].[Ca+2] AXCZMVOFGPJBDE-UHFFFAOYSA-L 0.000 description 2
- 239000000920 calcium hydroxide Substances 0.000 description 2
- 235000011116 calcium hydroxide Nutrition 0.000 description 2
- 229910001861 calcium hydroxide Inorganic materials 0.000 description 2
- 239000000571 coke Substances 0.000 description 2
- 239000007788 liquid Substances 0.000 description 2
- 239000007800 oxidant agent Substances 0.000 description 2
- 239000002244 precipitate Substances 0.000 description 2
- 235000010344 sodium nitrate Nutrition 0.000 description 2
- 239000004317 sodium nitrate Substances 0.000 description 2
- 239000003513 alkali Substances 0.000 description 1
- KGBXLFKZBHKPEV-UHFFFAOYSA-N boric acid Chemical compound OB(O)O KGBXLFKZBHKPEV-UHFFFAOYSA-N 0.000 description 1
- 239000004327 boric acid Substances 0.000 description 1
- 239000003638 chemical reducing agent Substances 0.000 description 1
- 239000012535 impurity Substances 0.000 description 1
- 238000002156 mixing Methods 0.000 description 1
- 230000001376 precipitating effect Effects 0.000 description 1
- 238000000746 purification Methods 0.000 description 1
- 239000011819 refractory material Substances 0.000 description 1
- 239000002904 solvent Substances 0.000 description 1
Description
本発明は鉛精錬において鉛中に随伴する錫を効
果的に分離回収する方法に関する。
鉛の電解精製において、錫は鉛と挙動を共にし
てカソードに析出するので、この電解精製では錫
を分離除去し難い。このため、得られたカソード
を溶融精製することが行なわれている。この溶融
精製には、例えばハリス法に従い、苛性ソーダや
硝酸ソーダを添加し、錫を含む精製滓を得る方法
が採用されている。
この鉛カソードの精製法を含め、錫含有鉛の溶
融体にソーダ化合物を添加して精製滓を得る場合
に、この精製滓中には鉛も多量に含まれてくるの
で、通常はこれを鉛還元炉に繰返して鉛分を粗鉛
として回収するのが一般であつた。しかし、この
場合には錫分も粗鉛中に入るので、再びカソード
に錫が濃縮され、カソードの精製に負荷が加わる
ことは避けられない。また、かかる精製滓を還元
炉に投入すると、炉内耐火物が精製滓中のアルカ
リにより浸蝕されるという問題も付随する。
本発明は、この鉛精錬工程で発生する精製滓か
ら錫を効率よく分離回収する方法を提供するもの
で、錫含有鉛の溶融体にソーダ化合物例えば苛性
ソーダ、硝酸ソーダを添加して得られる鉛精製滓
に硫黄を添加して水で浸出処理し、これによつて
錫を溶出させ、液中の錫をCa2+で錫酸カルシウ
ムとして沈殿させることによつて錫の分離を図る
ものであり、この沈殿物を還元溶融することによ
つて粗錫の回収を図るものである。
錫含有鉛の溶融体に苛性ソーダ、硝酸ソーダを
添加して溶融処理すると、錫は水溶性の錫酸ソー
ダ、Na2〔Sn(OH)6〕とすることができる。その
さい、錫の1部は錫酸鉛、Pb〔Sn(OH)6〕を形成
するが、これは若干の硫黄の添加によつて、下記
反応式により、水溶性の錫酸ソーダとすることが
できる。
6NaOH+4S→2Na2S+Na2S2O3+3H2O
Pb〔Sn(OH)6〕+Na2S→PbS+Na2〔Sn(OH)6〕
したがつて鉛精製滓に硫黄を添加して水で浸出
処理すれば、精製滓中の錫のほとんどは浸出する
ことができる。
浸出液中の錫の1部は、第1錫となつているの
で、H2O2等の酸化剤で第2錫としたあと、消石
灰等のCa2+により、水に不溶性の錫酸カルシウ
ムとして沈殿させれば、浸出液中の錫を効率よく
分離することができる。
この殿物を液から分離採取したあとは、これに
コークス等の還元剤を配合して還元溶融すること
によつて粗錫が回収できる。
以上のように、本発明によると従来鉛電解精製
の不純物にすぎなかつた錫を選択的に回収して粗
錫を得ることができる。また、鉛電解精製系で繰
返され濃縮される傾向にあつた錫を系外へ抜出す
ることができるので、溶融精製法で使用される溶
剤や酸化剤を減少させることもできる。
実施例 1
鉛電解精製カソードを溶融し、これにNaOH
とNaNO3を添加して得た精製滓5Kgに硫黄粉末
150g添加し、水10で浸出した。液温は反応熱
により、約50℃迄上昇した。8時間浸出し、その
結果を表1に示した。表1に見られるとおり錫浸
出率は86.0%であつた。
The present invention relates to a method for effectively separating and recovering tin accompanying lead in lead smelting. In electrolytic refining of lead, tin behaves in the same way as lead and is deposited on the cathode, so it is difficult to separate and remove tin in this electrolytic refining. For this reason, the obtained cathode is melt-purified. For this melt refining, a method is adopted, for example, according to the Harris method, in which caustic soda or sodium nitrate is added to obtain a refining slag containing tin. When obtaining refining slag by adding a soda compound to a melt of tin-containing lead, including this lead cathode refining method, the refining slag contains a large amount of lead, so it is usually It was common practice to repeatedly return the lead to a reduction furnace to recover the lead as crude lead. However, in this case, since tin also enters the crude lead, it is inevitable that tin will be concentrated in the cathode again and a load will be added to the purification of the cathode. Further, when such refining slag is put into a reduction furnace, there is also the problem that the refractories in the furnace are corroded by the alkali in the refining slag. The present invention provides a method for efficiently separating and recovering tin from the refining slag generated in the lead refining process. The method involves adding sulfur to the slag and leaching it with water, thereby eluting the tin, and precipitating the tin in the solution as calcium stannate with Ca 2+ to separate the tin. The purpose is to recover crude tin by reducing and melting this precipitate. When caustic soda and sodium nitrate are added to a melt of tin-containing lead and melted, tin can be converted into water-soluble sodium stannate, Na 2 [Sn(OH) 6 ]. At that time, a part of the tin forms lead stannate, Pb [Sn(OH) 6 ], which can be converted into water-soluble sodium stannate by the following reaction formula by adding a small amount of sulfur. Can be done. 6NaOH+4S→2Na 2 S+Na 2 S 2 O 3 +3H 2 O Pb [Sn(OH) 6 ]+Na 2 S→PbS+Na 2 [Sn(OH) 6 ] Therefore, sulfur was added to the lead refinery slag and leached with water. Most of the tin in the refinery slag can then be leached out. A portion of the tin in the leachate is in the form of stannous, so it is converted to stannous using an oxidizing agent such as H 2 O 2 and then converted to water-insoluble calcium stannate using Ca 2+ such as slaked lime. If precipitated, tin in the leachate can be efficiently separated. After this precipitate is separated and collected from the liquid, crude tin can be recovered by blending it with a reducing agent such as coke and reducing and melting it. As described above, according to the present invention, tin, which was only an impurity in conventional lead electrolytic refining, can be selectively recovered to obtain crude tin. Furthermore, since tin, which tends to be repeatedly concentrated in the lead electrolytic refining system, can be extracted from the system, it is also possible to reduce the amount of solvent and oxidizing agent used in the melt refining method. Example 1 Melt a lead electrolytically purified cathode and add NaOH to it.
Sulfur powder was added to 5 kg of refined slag obtained by adding NaNO 3 to
Added 150g and leached with 10ml of water. The liquid temperature rose to approximately 50°C due to the heat of reaction. Leaching was carried out for 8 hours, and the results are shown in Table 1. As shown in Table 1, the tin leaching rate was 86.0%.
【表】
実施例 2
実施例1で得られた浸出液に、消石灰700gを
添加して3時間撹拌を行い、錫を錫酸カルシウム
として沈殿させた。浸出液からの脱錫率は表2に
見られるとおり93.7%であつた。[Table] Example 2 700 g of slaked lime was added to the leachate obtained in Example 1 and stirred for 3 hours to precipitate tin as calcium stannate. As shown in Table 2, the tin removal rate from the leachate was 93.7%.
【表】
実施例 3
実施例2における錫酸カルシウムとして沈殿し
た残渣2.5Kgに、硼酸3Kgとコークス450gとを混
合し、還元溶融行なつて粗錫を得た。表3に見ら
れるとおり還元率96.7%で錫を採取することがで
きた。[Table] Example 3 2.5 kg of the residue precipitated as calcium stannate in Example 2 was mixed with 3 kg of boric acid and 450 g of coke, and subjected to reduction melting to obtain crude tin. As shown in Table 3, tin could be collected with a reduction rate of 96.7%.
Claims (1)
いし、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あとCa2+で錫を錫酸カル
シウムとして沈殿させることからなる鉛精錬にお
ける錫の分離回収法。1. When separating and recovering tin from lead in lead smelting, lead containing tin is melted, this melt is reacted with a soda compound, the tin in the melt is converted to sodium stannate, and this sodium stannate and In lead smelting, the slag containing raw lead stannate is separated from the molten lead, sulfur is added to this lead refining slag, leaching treatment is performed with water, and then tin is precipitated as calcium stannate using Ca 2+ . Tin separation and recovery method.
Priority Applications (1)
Application Number | Priority Date | Filing Date | Title |
---|---|---|---|
JP8071080A JPS575829A (en) | 1980-06-14 | 1980-06-14 | Separating and recovering method for tin in lead refining process |
Applications Claiming Priority (1)
Application Number | Priority Date | Filing Date | Title |
---|---|---|---|
JP8071080A JPS575829A (en) | 1980-06-14 | 1980-06-14 | Separating and recovering method for tin in lead refining process |
Publications (2)
Publication Number | Publication Date |
---|---|
JPS575829A JPS575829A (en) | 1982-01-12 |
JPS6319576B2 true JPS6319576B2 (en) | 1988-04-23 |
Family
ID=13725884
Family Applications (1)
Application Number | Title | Priority Date | Filing Date |
---|---|---|---|
JP8071080A Granted JPS575829A (en) | 1980-06-14 | 1980-06-14 | Separating and recovering method for tin in lead refining process |
Country Status (1)
Country | Link |
---|---|
JP (1) | JPS575829A (en) |
Families Citing this family (10)
Publication number | Priority date | Publication date | Assignee | Title |
---|---|---|---|---|
JPS59180892A (en) * | 1983-03-31 | 1984-10-15 | Toshiba Corp | Semiconductor memory |
GB8916673D0 (en) * | 1989-07-21 | 1989-09-06 | Alcan Int Ltd | Method of making alkali metal stannates |
JP4876221B2 (en) * | 2005-05-18 | 2012-02-15 | Dowaメタルマイン株式会社 | Metal recovery method |
US8211207B2 (en) * | 2006-12-05 | 2012-07-03 | Stannum Group LLC | Process for refining lead bullion |
JP5507310B2 (en) * | 2010-03-31 | 2014-05-28 | 三井金属鉱業株式会社 | Method for producing valuable metals |
US8105416B1 (en) | 2010-05-05 | 2012-01-31 | Stannum Group LLC | Method for reclaiming lead |
CN102776386B (en) * | 2012-07-20 | 2014-02-12 | 北京科技大学 | Method for recycling stannic oxide from tin-containing lead slag |
CN105366713B (en) * | 2015-12-10 | 2016-12-07 | 柳州百韧特先进材料有限公司 | A kind of method utilizing stannum waste residue to produce high-purity sodium stannate |
CN107142376A (en) * | 2017-04-01 | 2017-09-08 | 中南大学 | A kind of method that efficiently concentrating separates valuable metals from complex lead-containing precious metal material |
JP7243458B2 (en) * | 2019-05-31 | 2023-03-22 | 三菱マテリアル株式会社 | Tin recovery method |
-
1980
- 1980-06-14 JP JP8071080A patent/JPS575829A/en active Granted
Also Published As
Publication number | Publication date |
---|---|
JPS575829A (en) | 1982-01-12 |
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