JP2006083457A - Treatment method for zinc leach residue and the like - Google Patents

Treatment method for zinc leach residue and the like Download PDF

Info

Publication number
JP2006083457A
JP2006083457A JP2004271891A JP2004271891A JP2006083457A JP 2006083457 A JP2006083457 A JP 2006083457A JP 2004271891 A JP2004271891 A JP 2004271891A JP 2004271891 A JP2004271891 A JP 2004271891A JP 2006083457 A JP2006083457 A JP 2006083457A
Authority
JP
Japan
Prior art keywords
leaching
sulfuric acid
zinc
raw material
mixed
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.)
Pending
Application number
JP2004271891A
Other languages
Japanese (ja)
Inventor
Masami Ouchi
正美 大内
Ryoichi Taguchi
良一 田口
Akira Narumi
明 鳴海
Current Assignee (The listed assignees may be inaccurate. Google has not performed a legal analysis and makes no representation or warranty as to the accuracy of the list.)
Dowa Holdings Co Ltd
Original Assignee
Dowa Mining Co Ltd
Priority date (The priority date 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 date listed.)
Filing date
Publication date
Application filed by Dowa Mining Co Ltd filed Critical Dowa Mining Co Ltd
Priority to JP2004271891A priority Critical patent/JP2006083457A/en
Publication of JP2006083457A publication Critical patent/JP2006083457A/en
Pending legal-status Critical Current

Links

Images

Classifications

    • 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

Landscapes

  • Manufacture And Refinement Of Metals (AREA)

Abstract

<P>PROBLEM TO BE SOLVED: To provide a treatment method by which Sn and rare metals (particularly In) can be leached and recovered at high recovery rate by using, as a raw material to be leached, a leach residue etc. generated by the reduction and pressure leaching, by the use of SO<SB>2</SB>, of a zinc leach residue from the zinc hydrometallurgical process. <P>SOLUTION: In the treatment method for the zinc leach residue and the like, leaching treatment is carried out under the condition of a temperature of ≥60°C by using, as the raw material to be leached, the leach residue etc. generated by the reduction and pressure leaching, by the use of SO<SB>2</SB>, of the zinc leach residue from the zinc hydrometallurgical process and also using, as an original leaching solution, any of the followings: a mixed solution of sulfuric acid and hydrochloric acid; a mixed solution prepared by mixing a chloride salt with sulfuric acid; and a mixed solution prepared by mixing a chloride salt with sulfuric acid and hydrochloric acid. <P>COPYRIGHT: (C)2006,JPO&NCIPI

Description

本発明は，湿式亜鉛製錬において発生した亜鉛浸出残渣などといった少なくともＩｎまたはＳｎを含む被処理物からＩｎやＳｎを分離する方法に関する。   The present invention relates to a method for separating In and Sn from an object to be treated containing at least In or Sn such as zinc leaching residue generated in wet zinc smelting.

湿式亜鉛製錬においては，ＺｎＳ主体の亜鉛精鉱を焙焼して酸に可溶なＺｎＯの形態に変え，これを硫酸で浸出し，得られたＺｎリッチの浸出尾液は電解工程へ供される。一方，硫酸浸出で生じた亜鉛浸出残渣は，ＳＯ２を用いた還元加圧浸出法を採用することにより，硫酸浸出だけでは回収できなかったフェライト系のＺｎやＦｅ，レアメタルなどの有価金属を浸出させて回収している（特許文献１参照）。また，こうして有価金属を浸出回収した後の浸出残渣は，更に銀や鉛の原料として利用される（特許文献２参照）。 In hydro-zinc smelting, zinc concentrate mainly composed of ZnS is roasted into an acid-soluble form of ZnO, which is leached with sulfuric acid, and the resulting Zn-rich leaching tail solution is supplied to the electrolysis process. Is done. On the other hand, zinc leaching residue produced by sulfuric acid leaching uses a reduced pressure leaching method using SO 2 to leach valuable metals such as ferrite-based Zn, Fe, and rare metals that could not be recovered by sulfuric acid leaching alone. And collected (see Patent Document 1). Further, the leaching residue after leaching and recovering valuable metals is further used as a raw material for silver and lead (see Patent Document 2).

特開昭５６−１６４０１８号公報JP 56-164018 A 特開平７−６２４５４号公報Japanese Patent Laid-Open No. 7-62454

しかしながら，亜鉛浸出残渣をＳＯ２で還元加圧浸出して有価金属を浸出回収した場合，残った浸出残渣中にはレアメタルがまだ相当量残されており，中でも付加価値の高いＩｎのロスが大きい。例えば亜鉛浸出残渣をＳＯ２で加圧浸出した場合，ＺｎやＦｅの浸出率が９５％程度であるのに対して，Ｉｎの浸出率は８８〜９０％程度と低い。 However, when the zinc leaching residue is reduced and leached with SO 2 and valuable metals are leached and recovered, a considerable amount of rare metal is still left in the remaining leaching residue, and the loss of In, which has a high added value, is particularly large. . For example, when zinc leaching residue is pressure leached with SO 2 , the leaching rate of In is as low as 88 to 90% while the leaching rate of Zn or Fe is about 95%.

また優良な亜鉛鉱石の枯渇化により，今後は不純物としてＳｎを含有した高Ｓｎ品位の亜鉛精鉱の処理が必要となることが予想される。亜鉛浸出残渣をＳＯ２で還元加圧浸出した場合，Ｓｎ浸出率は１０％程度以下に過ぎない。Ｚｎその他の有価金属を回収した後の浸出残渣は，次に銀や鉛の精錬工程に送られるが，特にＳｎは，鉛電解において不純物として製品に混入し易いため，浸出残渣中のＳｎ品位の更なる低減が望まれる。 In addition, due to the depletion of excellent zinc ore, it is expected that it will be necessary to treat high-grade zinc concentrate containing Sn as an impurity in the future. When the zinc leaching residue is reduced pressure leached with SO 2 , the Sn leaching rate is only about 10% or less. The leaching residue after recovering Zn and other valuable metals is then sent to the refining process for silver and lead. In particular, Sn is easily mixed into the product as an impurity in lead electrolysis. Further reduction is desired.

本発明の目的は，湿式亜鉛製錬において発生した亜鉛浸出残渣などといった少なくともＩｎまたはＳｎを含む被処理物からＳｎとレアメタル（特にＩｎ）を高い浸出率で浸出回収できる処理方法を提供することにある。   An object of the present invention is to provide a treatment method capable of leaching and recovering Sn and rare metals (particularly In) at a high leaching rate from an object to be treated containing at least In or Sn such as a zinc leaching residue generated in wet zinc smelting. is there.

従来は塩酸や塩化物塩を含む浸出元液は亜鉛製錬工程へＣｌ−が混入すると，設備の腐食やメタル亜鉛の生産性及び品質の低下の原因となるため，使用されていなかった。本発明では，Ｃｌ−を含んだ液を亜鉛製錬本工程と切り離した別系統処理することにより解決した。即ち本発明によれば，湿式亜鉛製錬において発生した亜鉛浸出残渣をＳＯ２を用いて還元加圧浸出することにより発生した浸出残渣などといった少なくともＩｎまたはＳｎを含む被処理物を浸出原料とし，硫酸と塩酸との混合溶液，硫酸に塩化物塩を混合した混合溶液，または硫酸と塩酸に塩化物塩を混合した混合溶液の何れかを浸出元液として，温度６０℃以上の条件で浸出処理することを特徴とする，亜鉛浸出残渣等の処理方法が提供される。 Conventional leaching source solution containing hydrochloric acid or chloride salt is Cl to zinc smelting process - when is mixed, it will cause corrosion of metal zinc productivity and decrease in quality of facilities, have not been used. In the present invention, the solution containing Cl − was solved by a separate system treatment separated from the zinc smelting process. That is, according to the present invention, the leaching raw material is an object to be treated containing at least In or Sn, such as a leaching residue generated by reducing pressure leaching of zinc leaching residue generated in wet zinc smelting using SO 2 . Leaching treatment is performed at a temperature of 60 ° C or higher using either a mixed solution of sulfuric acid and hydrochloric acid, a mixed solution of sulfuric acid mixed with chloride salt, or a mixed solution of sulfuric acid and hydrochloric acid mixed with chloride salt as the brewing source solution. A method for treating zinc leaching residue and the like is provided.

前記浸出元液における硫酸濃度は，１００ｇ／リットル以上であることが好ましい。また，前記浸出元液において混合された塩化物塩は，例えばＮａＣｌ，ＫＣｌまたはＣａＣｌの１種以上である。また，前記浸出元液におけるＣｌ−量は，前記浸出原料に含まれるＳｎの当量の１倍以上であり，かつ１０等量以下であることが好ましい。 The sulfuric acid concentration in the leaching source solution is preferably 100 g / liter or more. Further, the chloride salt mixed in the leaching source liquid is, for example, one or more of NaCl, KCl, or CaCl. In addition, the Cl − amount in the leaching source solution is preferably at least 1 time the equivalent of Sn contained in the leaching raw material and at most 10 equivalents.

前記浸出原料を浸出処理した浸出尾液の一部または全部を，前記浸出元液として再利用することもできる。また，前記浸出原料を浸出処理した浸出尾液を，中和剤によりｐＨ ４〜９に調整した後，沈澱物を回収しても良い。   A part or all of the leaching tail liquid obtained by leaching the leaching raw material can be reused as the leaching source liquid. The leaching tail liquid obtained by leaching the leaching raw material may be adjusted to pH 4-9 with a neutralizing agent, and the precipitate may be recovered.

本発明によれば，湿式亜鉛製錬において発生した亜鉛浸出残渣をＳＯ２を用いて還元加圧浸出することにより発生した浸出残渣などといった少なくともＩｎまたはＳｎを含む被処理物を浸出原料として，ＳｎとＩｎ等のレアメタルを高い浸出率で回収することが可能となる。これにより，Ｉｎロスが減少する。また，浸出残渣中のＳｎ品位を低減することにより，後の銀や鉛の精錬工程における不純物の低減を図ることができるようになる。 According to the present invention, Sn to be treated as a leaching raw material is a material containing at least In or Sn, such as a leaching residue generated by reducing pressure leaching of zinc leaching residue generated in wet zinc smelting using SO 2. And rare metals such as In can be recovered at a high leaching rate. As a result, the In loss is reduced. Further, by reducing the Sn quality in the leaching residue, it becomes possible to reduce impurities in the subsequent silver and lead refining process.

図１に示すように，湿式亜鉛製錬においては，亜鉛精鉱を焙焼して得たＺｎＯを硫酸で浸出し，得られたＺｎリッチの浸出尾液は電解工程へ供される。一方，硫酸浸出で生じた亜鉛浸出残渣は，更にＳＯ２を用いて還元加圧浸出され，浸出尾液は有価金属の回収工程へ供される。また有価金属を浸出回収した後の浸出残渣は，本発明における浸出原料となる。かような浸出原料は，鉛を主成分とし，Ａｇなども含有するが，その他にＩｎやＳｎもまだ少なからず含んでおり，そのまま鉛や銀の原料として利用したのでは，Ｉｎのロスを生じてしまう。またＳｎ品位が高いと，銀や鉛の精錬工程において不純物として混入し易いため生産性を落すことになる。 As shown in FIG. 1, in wet zinc smelting, ZnO obtained by roasting zinc concentrate is leached with sulfuric acid, and the obtained Zn-rich leaching tail solution is subjected to an electrolysis process. On the other hand, the zinc leaching residue generated by sulfuric acid leaching is further subjected to reduction pressure leaching using SO 2 , and the leaching tail liquid is supplied to a valuable metal recovery step. Further, the leaching residue after leaching and recovering valuable metals becomes the leaching raw material in the present invention. Such a leaching material contains lead as a main component and also contains Ag and the like, but also contains a lot of In and Sn, and if used as it is as a raw material for lead and silver, loss of In occurs. End up. Moreover, when Sn grade is high, since it is easy to mix as an impurity in the refining process of silver or lead, productivity will fall.

そこで本発明では，このようにＩｎやＳｎをまだ含んでいる浸出原料を，硫酸と塩酸との混合溶液，硫酸に塩化物塩を混合した混合溶液，または硫酸と塩酸に塩化物塩を混合した混合溶液の何れかを浸出元液として，更に浸出処理する。硫酸に塩化物塩を混合する場合，及び硫酸と塩酸に塩化物塩を混合する場合，混合する塩化物塩として，例えばＮａＣｌ，ＫＣｌまたはＣａＣｌの１種以上を用いることができる。   Therefore, in the present invention, the leaching raw material that still contains In and Sn is mixed with sulfuric acid and hydrochloric acid, mixed solution of sulfuric acid with chloride salt, or mixed with sulfuric acid and hydrochloric acid with chloride salt. Further leaching is performed using any of the mixed solutions as a brewing source liquid. When a chloride salt is mixed with sulfuric acid, and when a chloride salt is mixed with sulfuric acid and hydrochloric acid, one or more of NaCl, KCl or CaCl can be used as the mixed chloride salt.

ここで，浸出原料を浸出処理する浸出元液にあっては，硫酸濃度が１００ｇ／リットル以上であることが好ましい。浸出元液における硫酸濃度が１００ｇ／リットル以下の場合，Ｓｎが殆ど浸出されないからである。   Here, in the brewing source liquid for leaching the leaching raw material, the sulfuric acid concentration is preferably 100 g / liter or more. This is because when the sulfuric acid concentration in the leaching source solution is 100 g / liter or less, Sn is hardly leached.

また浸出元液には，塩酸由来または塩化物塩由来のＣｌ−が存在することになるが，浸出元液におけるＣｌ−量は，Ｓｎを溶解させるのに最低必要な量（１倍当量）以上で，かつ１０倍等量以下，好ましくは５倍等量以下が望ましい。Ｃｌ−を大量に添加しすぎる銀や鉛精錬用の銀や鉛まで浸出させることになってしまう。 In the leaching source solution, hydrochloric acid-derived or chloride salt-derived Cl − is present, but the amount of Cl − in the leaching source solution is more than the minimum necessary amount (1 equivalent) for dissolving Sn. And 10 times or less, preferably 5 times or less. Cl - becomes possible to leach up to large amounts of silver and lead for silver and lead smelting too much added.

また，かような浸出元液によって前記浸出原料を浸出処理するにあたり，温度は６０℃以上とする。６０℃以下では，Ｉｎの浸出が殆ど進まなくなってしまう。   Further, when the leaching raw material is leached with such a leaching source liquid, the temperature is set to 60 ° C. or higher. Below 60 ° C, In leaching hardly progresses.

こうして浸出処理することにより，前述の浸出原料中に存在していたＳｎの大部分を浸出させ，Ｉｎの殆ど全部を浸出させることが可能となる。そして，このように浸出原料を浸出処理したことによってＺｎ，Ｆｅなどと共にＳｎ，Ｉｎ等を含んだ浸出尾液を，中和剤によりｐＨ ４〜９に調整した後，沈澱物を回収することにより，Ｉｎなどの有価金属を少ないロスで回収することができるようになる。   By performing the leaching process in this manner, it is possible to leach most of Sn existing in the aforementioned leaching raw material and to leach almost all of In. By leaching the leaching raw material in this way, the leaching tail liquid containing Sn, In, etc., together with Zn, Fe, etc. is adjusted to pH 4-9 with a neutralizing agent, and then the precipitate is recovered. , In and other valuable metals can be recovered with little loss.

また，このように浸出原料を浸出処理した後の浸出尾液は，高い濃度の硫酸およびＣｌ−イオンを含んでおり，その一部または全部を浸出原料を浸出処理するための浸出元液として再利用することもできる。その場合，浸出原料を浸出処理した後の浸出尾液の一部を抜き出して浸出原料を浸出処理するための浸出元液として再利用し，残りは中和剤によりｐＨ ４〜９に調整した後，沈澱物を回収することにより，Ｉｎなどの有価金属を回収しても良い。なお中和後の液は，有価金属および重金属が殆ど入っていていないため，廃水処理施設へ送られ処理される。 In addition, the leaching tail liquid after leaching treatment of the leaching raw material contains high concentrations of sulfuric acid and Cl − ions, and a part or all of the leaching tail liquid is reused as a leaching source liquid for leaching the leaching raw material. It can also be used. In that case, a part of the leaching tail liquid after leaching the leaching raw material is extracted and reused as a brewing source liquid for leaching the leaching raw material, and the remainder is adjusted to pH 4-9 with a neutralizing agent. , Valuable metals such as In may be recovered by recovering the precipitate. The neutralized solution contains almost no valuable metals or heavy metals and is sent to a wastewater treatment facility for processing.

一方，浸出原料を浸出処理した後の浸出残渣は，更に銀や鉛の原料として利用される。本発明によれば，このようにＩｎなどの有価金属の回収ロスを低減できることに加え，浸出残渣中のＳｎ品位を低減することにより，後の銀や鉛の精錬工程における不純物処理の負荷軽減を図ることができるようになる。   On the other hand, the leaching residue after leaching treatment of the leaching raw material is further used as a raw material for silver or lead. According to the present invention, in addition to reducing the recovery loss of valuable metals such as In as described above, by reducing the Sn quality in the leaching residue, it is possible to reduce the load of impurity treatment in the subsequent silver and lead refining process. It becomes possible to plan.

なお，以上では湿式亜鉛製錬においてＺｎＯを硫酸浸出して生じた亜鉛浸出残渣を更にＳＯ２で還元加圧浸出した後の浸出残渣を浸出原料とした例を説明したが，本発明では，亜鉛精鉱を酸素とＦｅ３＋を用いて直接浸出することにより発生した浸出残渣を浸出原料とすることもできる。また亜鉛精鉱を直接浸出する場合，高圧の酸素と１５０℃以上の温度で行う加圧浸出法を用いても良い。 In the above, an example in which the leaching residue obtained by further reducing and leaching the zinc leaching residue generated by sulfuric acid leaching of ZnO in wet zinc smelting with SO 2 was described as a leaching raw material. The leaching residue generated by directly leaching the concentrate using oxygen and Fe 3+ can be used as a leaching raw material. When zinc concentrate is leached directly, a pressure leaching method in which high pressure oxygen and a temperature of 150 ° C. or higher are used may be used.

亜鉛精鉱を焙焼したＺｎＯを硫酸浸出して得た亜鉛浸出残渣を更にＳＯ２で加圧浸出して有価金属を浸出回収した後の浸出残渣を浸出原料とした。表１に実施例の浸出原料とした浸出残渣の品位と対鉱石での浸出率を示す。 Zinc leaching residue obtained by leaching ZnO obtained by roasting zinc concentrate by sulfuric acid was further leached under pressure with SO 2 to leach and collect valuable metals, and the leaching residue was used as the leaching raw material. Table 1 shows the quality of the leaching residue used as the leaching raw material in the examples and the leaching rate against ore.

Figure 2006083457
Figure 2006083457

（実施例１）
先ず，この浸出原料（ｗｅｔ状態）３８６ｇに濃度５００ｇ／リットルの希硫酸１リットルを加え，更にＮａＣｌ ２９ｇを添加し，リパルプした。こうして得たスラリーを大気圧下で，温度９０℃，２ｈｒの加熱を行った。所定の時期にスラリーの一部を取り出してろ過，洗浄した残渣から浸出率を測定した結果を図２に示す。６０，１２０ｍｉｎを経過した時点におけるＳｎ及びＩｎの浸出率は各々８５％と９０％，鉱石から換算した総合浸出率は各々８６％，９９％であった。 Example 1
First, 1 liter of dilute sulfuric acid having a concentration of 500 g / liter was added to 386 g of this leaching raw material (wet state), and 29 g of NaCl was further added, followed by repulping. The slurry thus obtained was heated at 90 ° C. for 2 hours under atmospheric pressure. FIG. 2 shows the results of measuring the leaching rate from a residue obtained by taking out a portion of the slurry at a predetermined time, filtering and washing. The leaching rates of Sn and In after 60 and 120 min passed were 85% and 90%, respectively, and the total leaching rates converted from ore were 86% and 99%, respectively.

（実施例２）
実施例１と同じ浸出原料（ｗｅｔ状態）３８６ｇに濃度５００ｇ／リットルの希硫酸１リットルを加え，更にＮａＣｌ ５ｇを添加してリパルプした。こうして得たスラリーを大気圧下で，温度９０℃，２ｈｒの加熱を行った。所定の時期にスラリーの一部を取り出してろ過，洗浄した残渣から浸出率を測定した結果を図３に示す。６０，１２０ｍｉｎを経過した時点におけるＳｎ及びＩｎの浸出率は各々３５％（総合３％），６２％（総合９６％）であった。 (Example 2)
1 liter of dilute sulfuric acid having a concentration of 500 g / liter was added to 386 g of the same leaching raw material (wet state) as in Example 1, and 5 g of NaCl was further added for repulping. The slurry thus obtained was heated at 90 ° C. for 2 hours under atmospheric pressure. FIG. 3 shows the result of measuring the leaching rate from a residue obtained by removing a portion of the slurry at a predetermined time, filtering and washing. The leaching rates of Sn and In after 60 and 120 min were 35% (total 3%) and 62% (total 96%), respectively.

（比較例）
実施例１と同じ浸出原料（ｗｅｔ状態）３８６ｇに濃度５００ｇ／リットルの希硫酸１リットルを加えリパルプした。こうして得たスラリーを大気圧下で，温度９０℃，２ｈｒの加熱を行った。所定の時期にスラリーの一部を取り出してろ過，洗浄した残渣から浸出率を測定した結果を図４に示す。６０，１２０ｍｉｎを経過した時点におけるＳｎ及びＩｎの浸出率は各々１５％（総合２０％），３５％（総合９３％）であった。 (Comparative example)
1 liter of dilute sulfuric acid with a concentration of 500 g / liter was added to 386 g of the same leaching raw material (wet state) as in Example 1 and repulped. The slurry thus obtained was heated at 90 ° C. for 2 hours under atmospheric pressure. FIG. 4 shows the result of measuring the leaching rate from a residue obtained by removing a part of the slurry at a predetermined time, filtering and washing. The leaching rates of Sn and In after 60 and 120 minutes were 15% (20% overall) and 35% (93% overall), respectively.

（実施例３）
実施例１と同じ浸出原料（ｗｅｔ状態）３８６ｇに所定濃度の希硫酸１リットルを加え，更にＮａＣｌ２９ｇ添加してリパルプした。こうして得たスラリーを大気圧下，温度４０，６０，９０℃の各温度でそれぞれ２ｈｒ浸出後にろ過，洗浄した残渣から浸出率を測定した。各浸出温度に対する浸出率は図５のようになった。浸出温度４０℃ではＳｎおよびＩｎの浸出率は１０％以下であった。 (Example 3)
1 liter of dilute sulfuric acid with a predetermined concentration was added to 386 g of the same leaching raw material (wet state) as in Example 1, and 29 g of NaCl was further added for repulping. The leaching rate was measured from the residue obtained by leaching the slurry thus obtained at a temperature of 40, 60, and 90 ° C. under atmospheric pressure for 2 hours, followed by filtration and washing. The leaching rate with respect to each leaching temperature is as shown in FIG. At a leaching temperature of 40 ° C., the leaching rate of Sn and In was 10% or less.

（実施例４）
実施例１と同じ浸出原料（ｗｅｔ状態）３８６ｇに濃度７０，１００，２００，５００ｇ／リットルの各希硫酸１リットルをそれぞれ加え，更にＮａＣｌ２９ｇ添加してリパルプした。こうして得た各スラリーを大気圧下，温度９０℃でそれぞれ２ｈｒ浸出後にろ過，洗浄した残渣から浸出率を測定した。各硫酸濃度に対する浸出率は図６のようになった。硫酸濃度１００ｇ／リットルでのＳｎ及びＩｎの浸出率は各々２０％（総合２５％），４２％（総合９４％）であった。浸出原料を浸出処理する浸出元液にあっては，硫酸濃度が１００ｇ／リットル以上であることが好ましい。浸出元液における硫酸濃度が１００ｇ／リットル未満の場合，Ｓｎが殆ど浸出されない。 Example 4
1 liter of each diluted sulfuric acid having a concentration of 70, 100, 200, and 500 g / liter was added to 386 g of the same leaching raw material (wet state) as in Example 1, and 29 g of NaCl was further added for repulping. Each slurry thus obtained was leached for 2 hours at 90 ° C. under atmospheric pressure and then filtered and washed, and the leaching rate was measured. The leaching rate for each sulfuric acid concentration was as shown in FIG. The leaching rates of Sn and In at a sulfuric acid concentration of 100 g / liter were 20% (total 25%) and 42% (total 94%), respectively. In the leaching source liquid for leaching the leaching raw material, the sulfuric acid concentration is preferably 100 g / liter or more. When the sulfuric acid concentration in the leaching source solution is less than 100 g / liter, Sn is hardly leached.

本発明は，湿式亜鉛製錬において発生する亜鉛浸出残渣の処理などに利用される。   The present invention is used for treatment of zinc leaching residue generated in wet zinc smelting.

発明の実施の形態にかかる処理方法を説明するためのフローチャートである。It is a flowchart for demonstrating the processing method concerning embodiment of invention. 実施例１における経過時間と浸出率の関係を示すグラフである。It is a graph which shows the relationship between the elapsed time in Example 1, and a leaching rate. 実施例２における経過時間と浸出率の関係を示すグラフである。It is a graph which shows the relationship between the elapsed time in Example 2, and a leaching rate. 比較例における経過時間と浸出率の関係を示すグラフである。It is a graph which shows the relationship between the elapsed time and the leaching rate in a comparative example. 実施例３における浸出温度と浸出率の関係を示すグラフである。It is a graph which shows the relationship between the leaching temperature in Example 3, and a leaching rate. 実施例４における硫酸濃度と浸出率の関係を示すグラフである。It is a graph which shows the relationship between the sulfuric acid concentration and the leaching rate in Example 4.

Claims (7)

少なくともＩｎまたはＳｎを含む被処理物を浸出原料とし，硫酸と塩酸との混合溶液，硫酸に塩化物塩を混合した混合溶液，または硫酸と塩酸に塩化物塩を混合した混合溶液の何れかを浸出元液として，温度６０℃以上の条件で浸出処理することを特徴とする，Ｉｎ及び/またはＳｎの分離方法。   A material to be treated containing at least In or Sn is used as a leaching raw material, and either a mixed solution of sulfuric acid and hydrochloric acid, a mixed solution of sulfuric acid mixed with a chloride salt, or a mixed solution of sulfuric acid and hydrochloric acid mixed with a chloride salt is used. A method for separating In and / or Sn, wherein leaching is performed as a leaching source solution at a temperature of 60 ° C. or higher. 前記浸出元液における硫酸濃度が，１００ｇ／リットル以上であることを特徴とする，請求項１に記載のＩｎ及び/またはＳｎの分離方法。   The method for separating In and / or Sn according to claim 1, wherein the sulfuric acid concentration in the leaching source solution is 100 g / liter or more. 前記浸出元液において混合された塩化物塩が，ＮａＣｌ，ＫＣｌまたはＣａＣｌの１種以上であることを特徴とする，請求項１または２に記載のＩｎ及び/またはＳｎの分離方法。   The method for separating In and / or Sn according to claim 1 or 2, wherein the chloride salt mixed in the leaching source liquid is one or more of NaCl, KCl, or CaCl. 前記浸出元液におけるＣｌ−量が，前記浸出原料に含まれるＳｎの当量の１倍以上であり，かつ１０倍当量以下であることを特徴とする，請求項１，２または３に記載のＩｎ及び/またはＳｎの分離方法。 Cl in the leaching source liquid - amounts, the not less than one times the equivalent of Sn contained in the leaching material, and wherein the 10-fold is equivalent or less, In of claim 1, 2 or 3 And / or Sn separation method. 前記浸出原料を浸出処理した浸出尾液の一部または全部を，前記浸出元液として再利用することを特徴とする，請求項１，２，３または４に記載のＩｎ及び/またはＳｎの分離方法。   The separation of In and / or Sn according to claim 1, 2, 3 or 4, wherein a part or all of the leaching tail liquid obtained by leaching the leaching raw material is reused as the leaching source liquid. Method. 前記浸出原料を浸出処理した浸出尾液を，中和剤によりｐＨ ４〜９に調整した後，沈澱物を回収することを特徴とする，請求項１，２，３または４に記載のＩｎ及び/またはＳｎの分離方法。   The indium and effluent according to claim 1, 2, 3 or 4, wherein the leaching tail liquid obtained by leaching the leaching raw material is adjusted to pH 4-9 with a neutralizing agent, and then the precipitate is recovered. / Or Sn separation method. 湿式亜鉛製錬において発生した亜鉛浸出残渣をＳ０２を用いて還元加圧浸出することにより発生した浸出残渣を浸出原料とすることを特徴とする，請求項１，２，３，４，５または６に記載のＩｎ及び/またはＳｎの分離方法。 Characterized by leaching residue generated by output reduction pressure leaching with an S0 2 zinc leach residue generated in wet zinc smelting and leaching material, according to claim 1, 2, 3, 4 or 7. The method for separating In and / or Sn according to 6.
JP2004271891A 2004-09-17 2004-09-17 Treatment method for zinc leach residue and the like Pending JP2006083457A (en)

Priority Applications (1)

Application Number Priority Date Filing Date Title
JP2004271891A JP2006083457A (en) 2004-09-17 2004-09-17 Treatment method for zinc leach residue and the like

Applications Claiming Priority (1)

Application Number Priority Date Filing Date Title
JP2004271891A JP2006083457A (en) 2004-09-17 2004-09-17 Treatment method for zinc leach residue and the like

Publications (1)

Publication Number Publication Date
JP2006083457A true JP2006083457A (en) 2006-03-30

Family

ID=36162205

Family Applications (1)

Application Number Title Priority Date Filing Date
JP2004271891A Pending JP2006083457A (en) 2004-09-17 2004-09-17 Treatment method for zinc leach residue and the like

Country Status (1)

Country Link
JP (1) JP2006083457A (en)

Cited By (7)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
WO2006080565A1 (en) * 2005-01-31 2006-08-03 Dowa Metals & Mining Co., Ltd. Method for recovering indium
JP2009155717A (en) * 2007-12-28 2009-07-16 Dowa Eco-System Co Ltd Method for recovering indium
US8992759B1 (en) 2014-02-20 2015-03-31 Honeywell International Inc. Metal refining process using mixed electrolyte
CN105907978A (en) * 2016-04-20 2016-08-31 武汉长海高新技术有限公司 Method of recycling indium from indium-containing smoke dust
CN109628746A (en) * 2019-01-03 2019-04-16 江西铜业技术研究院有限公司 The extracting method of tin in a kind of silver separating residues
CN110016547A (en) * 2019-04-23 2019-07-16 王柯娜 A kind of method of comprehensive utilization using sodium jarosite
CN115896487A (en) * 2022-11-25 2023-04-04 沈阳有色金属研究院有限公司 Method for enriching and extracting indium from tin smelting smoke dust

Cited By (10)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
WO2006080565A1 (en) * 2005-01-31 2006-08-03 Dowa Metals & Mining Co., Ltd. Method for recovering indium
JP2009155717A (en) * 2007-12-28 2009-07-16 Dowa Eco-System Co Ltd Method for recovering indium
US8992759B1 (en) 2014-02-20 2015-03-31 Honeywell International Inc. Metal refining process using mixed electrolyte
CN105907978A (en) * 2016-04-20 2016-08-31 武汉长海高新技术有限公司 Method of recycling indium from indium-containing smoke dust
CN105907978B (en) * 2016-04-20 2018-06-29 武汉长海高新技术有限公司 A kind of method of recovery indium in flue dust containing indium
CN109628746A (en) * 2019-01-03 2019-04-16 江西铜业技术研究院有限公司 The extracting method of tin in a kind of silver separating residues
CN109628746B (en) * 2019-01-03 2021-02-26 江西铜业技术研究院有限公司 Method for extracting tin from silver separating slag
CN110016547A (en) * 2019-04-23 2019-07-16 王柯娜 A kind of method of comprehensive utilization using sodium jarosite
CN110016547B (en) * 2019-04-23 2020-11-13 王柯娜 Comprehensive utilization method of jarosite slag
CN115896487A (en) * 2022-11-25 2023-04-04 沈阳有色金属研究院有限公司 Method for enriching and extracting indium from tin smelting smoke dust

Similar Documents

Publication Publication Date Title
CN103966450A (en) All-wet process pretreatment method for copper anode mud
CA2943483C (en) Copper removal method for aqueous nickel chloride solution
JP4765062B2 (en) Wet treatment method for zinc leaching residue
JP2011021219A (en) Method for recovering copper from copper/iron-containing material
JP3052535B2 (en) Treatment of smelting intermediates
JP2010138490A (en) Method of recovering zinc
EP2902510A1 (en) A new method for leaching of electric arc furnace dust (EAFD) with sulphuric acid
JP2006083457A (en) Treatment method for zinc leach residue and the like
JP3411320B2 (en) Zinc smelting method
JP5296482B2 (en) Non-ferrous smelting arsenic removal method and non-ferrous smelting concentrate obtained by the method
JP5200588B2 (en) Method for producing high purity silver
JP6233478B2 (en) Purification method of bismuth
JP2008115429A (en) Method for recovering silver in hydrometallurgical copper refining process
JP5821775B2 (en) Processing method of copper smelting ash
JP6015824B2 (en) Processing method of copper smelting ash
JP2015113503A (en) Method of separating and collecting selenium and tellurium in transition metal-containing aqueous solution
JP2020105587A (en) Treatment method of acidic solution containing noble metal, selenium and tellurium
JP2010202457A (en) Method for removing chlorine in acidic liquid
JP5423592B2 (en) Method for producing low chlorine nickel sulfate / cobalt solution
JP7119551B2 (en) Method for producing aqueous solution of cobalt chloride
JP2011195935A (en) Method for separating and recovering platinum group element
JP2007154249A (en) Method for recovering silver in wet type copper refining process
JP3937273B2 (en) Method for treating slurry containing colloidal silica
JP4982662B2 (en) Treatment method of leachate in wet zinc smelting process
JP4399567B2 (en) Liquid purification method of wet zinc smelting

Legal Events

Date Code Title Description
A621 Written request for application examination

Free format text: JAPANESE INTERMEDIATE CODE: A621

Effective date: 20070806

A977 Report on retrieval

Free format text: JAPANESE INTERMEDIATE CODE: A971007

Effective date: 20091001

A131 Notification of reasons for refusal

Free format text: JAPANESE INTERMEDIATE CODE: A131

Effective date: 20091013

A02 Decision of refusal

Free format text: JAPANESE INTERMEDIATE CODE: A02

Effective date: 20100302