JPS6348929B2

JPS6348929B2 -

Info

Publication number: JPS6348929B2
Application number: JP1483985A
Authority: JP
Inventors: Takeyoshi Shibazaki; Hiromi Mochida; Katsuji Tomaki
Original assignee: Mitsubishi Metal Corp
Current assignee: Mitsubishi Metal Corp
Priority date: 1985-01-29
Filing date: 1985-01-29
Publication date: 1988-10-03
Also published as: JPS61174341A

Description

[Detailed description of the invention]

（産業分野）本発明は銅電解スライムに石灰、珪砂等のフラ
ツクス及びコークス、石炭等の還元剤を添加して
加熱、溶解させ、鉛、アンチモン、ビスマス等の
大部分をスラグ化するとともに銀含有量の高い貴
鉛を生成せしめる銅電解スライムの製錬法に関す
る。（従来技術とその問題点）銅電解スライムには、金、銀等の貴金属のほか
に、銅、セレン、テルル、鉛、アンチモン、ビス
マス、砒素、等が含まれている。これらの中で、
銅は空気酸化を行ないつつ、硫酸浸出するか、あ
るいは酸化焙焼後、硫酸浸出を行なつて除去され
る。これを、電解工程で生成したままのスライム
と区別して脱銅スライムとよぶこともあるが、以
下、単に銅電解スライムをスライムとよぶ。銅電解スライムの処理工程は様々の異なる方式
が提案されているが、一般的方法としては次の如
きものである。 (1) 酸化焙焼又は硫酸化焙焼を行つてセレンを揮
発除去する工程、 (2) フラツクス及び還元剤を加えて加熱溶解し、
鉛、アンチモン、ビスマスを含むスラグと、
金、銀を含む貴鉛とを生成せしめる工程、 (3) 更に、貴鉛中の鉛、セレン、テルル、銅等を
酸化除去しDore′ metalと呼ばれる粗銀を生成
せしめる工程、の組合せよりなるものである。これらの工程中で第２工程ではスラグへの貴金
属の損失を最小にするためには、できるだけ多量
の還元剤を用いて完全に還元させることが望まし
いが、このようにすると、スライム中の鉛の大部
分が還元され、貴鉛の銀品位が著しく低下するの
で不適である。その結果、第３工程では多量の鉛を酸化してス
ラグ（密陀）として分離しなくてはならないから
反応時間が長くなり、しかも生成したスラグ中に
はなお高濃度の金、銀が含まれている。従つて再
処理を行なう必要があり、全体としてのプロセス
の効率的運用を阻害することになる。上記プロセスの効率的運用を可能ならしめる改
善策の一つとして、浮遊選鉱によるスライム中の
硫酸鉛の分離方法が提案されている（特開昭49−
123428号公報）。この方法はスライムの鉛含有量
が高い場合にはかなり有効な方法で、鉛含有量の
約1/2を、硫酸鉛を主成分とする尾鉱（tailing）
として、分離することができる。しかしながら、
スライムの鉛含有量がそれほど多くない場合はあ
まり有効ではない。また、上記スライム中の鉛をalkylene amine
たとえばdiethylene triamineによつて浸出する
方法（米国特許No.4283224号明細書）や酢酸塩に
よつて浸出する方法（特開昭57−149437号公報）
等が提案されている。これら湿式プロセスによる
処理では、鉛を炭酸塩その他の化合物として回収
できる利点がある反面、試薬、ロス、洗浄水の処
理等によるコストが大きいという問題点をかかえ
ている。更に、未焙焼スライムにソーダ灰等のフラツク
スを加えて非還元性雰囲気で加熱溶解し、鉛等を
スラグ化し、セレン、テルル、銀等を含む貴鉛を
生成する方法も提案されている（特公昭57−
23735号公報）。この方法は、従来のスライムを焙焼してセレン
を除去してから処理する方法に比し、貴鉛には多
量のセレンが含まれているため、これを酸化除去
するための、ソーダ灰及びチリ硝石等の試薬の使
用量が大で、かつ反応時間もまた長くなる。スラ
イムの溶解工程でソーダ系フラツクスを用いるの
で炉の耐火材の消耗も相当大きいという問題点が
ある。（発明の目的）本発明の目的は、上記の従来技術の問題点を解
決し、銅電解スライムより高品位の貴鉛を高収率
で得ることを可能ならしめる銅電解スライムの製
錬法を提供するにある。（発明の構成）すなわち、本発明によれば、基本的には、銅電
解スライムを脱銅しあるいは脱銅後、更に脱セレ
ン焙焼を行ない、これにフラツクス及び還元剤を
加え、次いで加熱、溶解せしめてなる銅電解スラ
イムの製錬法において、該銅電解スライムに、シ
リカ質フラツクスとライム質フラツクスよりなる
フラツクスと該銅電解スライムの１〜４％範囲の
炭素質物質よりなる還元剤とを加えて加熱して該
銅電解スライムを溶解させ、該銅電解スライム中
のベースメタルの実質的な大部分をスラグ中に移
行させるとともに高鉛品位のスラグと高銀品位の
貴鉛とを生成せしめることを特徴とする銅電解ス
ライムの製錬法、が得られる。本発明ではさらに、上記の基本的構成に加え
て、生成した該スラグを固化かつ微粉砕し、これ
を浮遊選鉱して該スラグ中の金、銀、セレン、テ
ルルを濃縮した精鉱として回収し、この精鉱を上
記溶解工程に繰り返すことによつて、高銀品位の
貴鉛を高収率で得ることを可能ならしめるもので
ある。次に、本発明をさらに詳述する。銅電解スライムには通常１〜３％のテルルと10
〜50％のセレンが含れている。セレンの含有率が
低い場合には脱セレンのための焙焼をしないで直
接加熱、溶解させることもできるが、セレンの含
有率が高い場合にはまず該スライムの酸化焙焼ま
たは硫酸化焙焼によつて含有セレンを除去した方
が後の工程の操作が容易となる。この焙焼はいか
なる方法によつて行なつてもよいが、特公昭49−
30328号公報に記載された焙焼方法を例にとつて
述べると、次の通りである。この焙焼炉としてはいわゆるシヨートロータリ
フアーネスが用いられている。このフアーネスは
予め炉温を昇温し、銅電解スライムを装入しつつ
温度を800〜900℃に保持し、かつ炉を回転させな
がら焙焼を行なう。スライムの１チヤージー分の
装入が終つてから、なお5Hr程度焙焼を行ない、
含有セレンの80〜90％を揮発させる。揮発したセ
レンは該セレンを含む排ガスをカ性ソーダ水溶液
でスクラビングして回収する。この段階で炉内の焙焼スライムは約900℃に保
持されているので、煙灰、密陀等の工程内繰返し
物、石灰石及び珪砂等よりなるフラツクス及び少
量のコークスを追加装入して1100℃〜1250℃に加
熱、溶解させる。溶解を促進するために、炉の回
転は引続き行なう。これら装入物が完全に溶解
し、還元反応が完了すると、炉の回転を停止し、
スラグと貴鉛をセツトリングによつて分離し、炉
を傾転してスラグはスラグレードルに抜出して固
化させ、貴鉛は貴鉛レードルに抜出して次の分銀
工程に送る。上記スライムを加熱、溶解させる際、ソーダ質
フラツクスを用いる方法と鉄−シリケート系スラ
グを生成させる方法があるが、前者ではスラグの
侵食性が大で、炉材の寿命が短いという欠点があ
り、後者では鉄を加えるため、スラグの生成量が
多いという欠点があつた。本発明の基本的工程、すなわち銅電解スライム
の溶解並びに高鉛品位のスラグと高銀品位の貴鉛
との生成を含む工程では、鉛のスラグ化を助長す
るため、シリケートベースのフラツクスを添加す
る。その際、スラグ生成量を少なくするため、ス
クラツプ鉄は添加しない。すなわち、従来法では
スラグ中の鉛は25〜30％の範囲であつたが、本発
明方法ではスラグ中の鉛を35〜45％まで高めるた
め、還元剤の添加量は最小限とする。即ち焙焼ス
ライムに対し１〜４重量％の範囲のコークスまた
は石炭添加量が必要である。還元剤の添加量が１
重量％未満では添加の効果が著しく低下し、４重
量％を越えると、効果の向上は最早見られず、コ
スト的に不利となる。鉛以外の主要なスラグ成分
はシリカ（SiO₂）ライム（CaO）であるが、ス
ラグ組成として、SiO₂8〜15％、CaO1〜３％の
範囲である。スライム中には若干量のSiOが含ま
れているため、フラツクス添加量は上記スラグを
生成するに足る量とすればよい。 Pb約20％、SiO約2.5％のスライムではコーク
スの添加量３％、珪砂１％、石灰石5.4％で上記
目標組成のスラグが得られる。この時、得られる
貴鉛中の銀は50〜60％である。スラグ中のPb％
と貴鉛中のAg％は第１図に例示するように、正
の相関があり、高鉛品位のスラグを生成すれば、
貴鉛中の銀品位は高められる。但し、それと同時
にスラグへの銀の損失も増加する。第２図には銀
のスラグ／貴鉛間の分配比と貴鉛の銀品位との関
係を概念的に示す。貴鉛は次の分銀工程で処理し、粗銀とする。貴
鉛の組成は上述の銀50〜60％以外の成分としては
鉛４〜10％、セレン10〜15％、テルル10〜20％、
銅５〜６％である。銀品位40％程度の貴鉛では鉛
が約20〜25％であるのに対し高銀貴鉛中の鉛は著
しく低く溶解工程における鉛のスラグ化がそれだ
け高められていることを示す。分銀工程で酸化す
べき鉛の量が少ないため、分銀工程も反応時間が
短縮される。スラグは、銀の品位が高められた結果、生成率
は従来のPb25％のスラグを生成する場合に比し、
65〜70％にすぎない。その反面、銀等の含有率は
若干高く、Ag0.5〜３％、Se0.5〜1.5％、Te1〜
４％程度である。なお、従来法に比し、スラグ生
成量が少ないので、これらの含有量はそれほど増
加しているわけではないので、そのまま鉛製錬工
程に送ることもできる。次に、本発明方法の基本的工程につづく工程で
は、即ち基本的工程で生成した高鉛品位のスラグ
と高銀品位の貴鉛を有用化する工程では、スラグ
を固化、微粉砕し、浮遊選鉱により、金、銀、セ
レン、テルルを精鉱として濃縮回収し、それらを
さらに前記溶解工程に繰返し、スライムと同時処
理することにより、上記成分の収率を高めること
ができる。粉砕スラグの粒度は細かい方が選鉱成
績が良いが、−100ミクロン程度でも充分である。
スラリー濃度100−200ｇ／、PH８−10で、起泡
剤としてはパイン油、捕収剤としてはたとえば、
日本香料薬品株式会社製のDTP−８等を用いて
処理した１例ではAu0.07％、Ag2.2％、Se0.7％、
Te1.8％、Pb44％のスラグの場合ではAu＋Ag10
〜15％の精鉱がスラグの約15％得られ、テイリン
グはAu＜0.001％、Ag0.3〜0.35％、Se0.13％、
Te0.55％程度まで低下した。収率はAu＞99％、
Ag約90％、Se85％、Te75％程度である。上記パイン油及びDTP−８（商品名）は次のご
ときものである。パイプ油の主成分はターピネオ
ールでその構造式はである。一方、DTP−８の主成分はジチオフオ
スフエートでその構造式はである。テイリングへの鉛の移行率は約90％であ
り、このテイリングは鉛製錬に送り、鉛及び残留
する少量の金、銀をさらに回収することができ
る。精鉱はスライムの溶解工程に繰り返す。次に、本発明を実施例によつてより具体的に説
明するが、本発明の範囲は以下の実施例によつて
限定されるものではない。実施例１第１表に示すスライムA3.2t、スライムB2.2tを
900℃で酸化焙焼してセレンを除去し（推定除去
率85％）、次いで第１表に示す工程内繰返物及び
コークス160Kg、石灰石240ｇ、珪砂45Kgを追加装
入し、炉を回転させつつ、1150〜1200℃に昇温し
て溶解させた。約10時間で溶解し、反応も終了し
たので、約１時間静置し、スラク、貴鉛をそれぞ
れ抜き出した。これら産出物品位は第１表に示
す。産出物量は推定値を第２表に併記した。該ス
ラグ中のPbは42％、Sbは約４％、Biは約1.5％で
あつた。一方、該貴鉛は常法に従つて分銀工程で
処理したが、処理時間は従来のAg40％の貴鉛の
場合の約60％に短縮された。また、密陀、分銀工
程煙灰等の生成量は従来法に比し、密陀＋ソーダ
スラグは従来法の約60％、また煙灰は従来法の約
30％に減少 (Industrial Field) The present invention involves adding fluxes such as lime and silica sand, and reducing agents such as coke and coal to copper electrolytic slime, heating and melting the slime, and converting most of lead, antimony, bismuth, etc. into slag, as well as containing silver. This invention relates to a method for smelting copper electrolytic slime that produces a high amount of noble lead. (Prior art and its problems) Copper electrolytic slime contains copper, selenium, tellurium, lead, antimony, bismuth, arsenic, etc. in addition to precious metals such as gold and silver. Among these,
Copper is removed by leaching with sulfuric acid while performing air oxidation, or by leaching with sulfuric acid after oxidative roasting. This is sometimes referred to as decoppered slime to distinguish it from the slime produced in the electrolytic process, but hereinafter copper electrolytic slime will be simply referred to as slime. Although various different methods have been proposed for the treatment process of copper electrolytic slime, the following is a general method. (1) Performing oxidation roasting or sulfation roasting to volatilize and remove selenium, (2) Adding flux and reducing agent and heating and dissolving,
Slag containing lead, antimony, and bismuth,
(3) A step of oxidizing and removing lead, selenium, tellurium, copper, etc. from the precious lead to produce crude silver called Dore' metal. It is something. In order to minimize the loss of precious metals to the slag in the second step of these steps, it is desirable to use as much reducing agent as possible to completely reduce the amount of lead in the slime. Most of the precious lead is reduced and the quality of silver in precious lead is significantly reduced, so it is unsuitable. As a result, in the third step, a large amount of lead has to be oxidized and separated as slag, which lengthens the reaction time, and the resulting slag still contains high concentrations of gold and silver. ing. Therefore, reprocessing is required, which impedes the efficient operation of the process as a whole. As one of the improvement measures to enable efficient operation of the above process, a method for separating lead sulfate from slime by flotation has been proposed (Japanese Patent Application Laid-Open No. 1983-1999-
Publication No. 123428). This method is quite effective when the lead content of the slime is high. Approximately 1/2 of the lead content is extracted from tailings containing lead sulfate as the main component.
It can be separated as however,
It is not very effective if the lead content of the slime is not very high. In addition, the lead in the slime mentioned above is replaced with alkylene amine.
For example, a method of leaching with diethylene triamine (US Patent No. 4283224) and a method of leaching with acetate (Japanese Patent Application Laid-Open No. 149437/1983)
etc. have been proposed. Treatments using these wet processes have the advantage of recovering lead as carbonates and other compounds, but have the problem of high costs due to reagents, losses, treatment of washing water, etc. Furthermore, a method has been proposed in which flux such as soda ash is added to unroasted slime and heated and melted in a non-reducing atmosphere to turn lead, etc. into slag, and produce noble lead containing selenium, tellurium, silver, etc. ( Special Public Service 1977-
Publication No. 23735). Compared to the conventional method of roasting slime to remove selenium, this method uses soda ash and The amount of reagents such as saltpeter used is large, and the reaction time is also long. Since soda-based flux is used in the slime melting process, there is a problem in that the refractory material in the furnace is consumed considerably. (Object of the Invention) The object of the present invention is to provide a method for smelting copper electrolytic slime that solves the above-mentioned problems of the prior art and makes it possible to obtain noble lead of higher quality than copper electrolytic slime at a higher yield. It is on offer. (Structure of the Invention) That is, according to the present invention, basically, the copper electrolytic slime is decoppered or after decoppered, further roasted to remove selenium, flux and a reducing agent are added thereto, and then heated, In a method for smelting copper electrolytic slime by dissolving it, a flux consisting of a siliceous flux and a lime flux and a reducing agent consisting of a carbonaceous substance in a range of 1 to 4% of the copper electrolytic slime are added to the copper electrolytic slime. and heating to melt the copper electrolytic slime, transfer a substantial majority of the base metal in the copper electrolytic slime into slag, and produce high lead grade slag and high silver grade noble lead. A method for smelting copper electrolytic slime is obtained. In addition to the above-mentioned basic configuration, the present invention further includes solidifying and pulverizing the generated slag, flotating it, and recovering the gold, silver, selenium, and tellurium in the slag as a concentrated concentrate. By repeatedly subjecting this concentrate to the above-described melting process, it is possible to obtain noble lead with a high silver grade at a high yield. Next, the present invention will be explained in further detail. Copper electrolytic slime usually contains 1-3% tellurium and 10%
Contains ~50% selenium. If the selenium content is low, it is possible to directly heat and dissolve without roasting to remove selenium, but if the selenium content is high, the slime must first be oxidized or sulfated. If the selenium contained is removed by this method, the subsequent steps will be easier. This roasting may be done by any method, but
Taking the roasting method described in Publication No. 30328 as an example, it is as follows. A so-called short rotary furnace is used as this roasting furnace. In this furnace, the furnace temperature is raised in advance, the copper electrolytic slime is charged, the temperature is maintained at 800 to 900°C, and roasting is performed while the furnace is rotated. After charging one roast of slime, roast it for about 5 hours.
Volatizes 80-90% of the selenium contained. The volatilized selenium is recovered by scrubbing the selenium-containing exhaust gas with an aqueous caustic soda solution. At this stage, the roasted slime in the furnace is maintained at approximately 900°C, so we additionally charge the process-repeated materials such as smoke ash, mitzvah, etc., flux made of limestone and silica sand, and a small amount of coke to raise the temperature to 1100°C. Heat to ~1250℃ to dissolve. The furnace continues to rotate to promote melting. When these charges are completely melted and the reduction reaction is completed, the rotation of the furnace is stopped.
The slag and precious lead are separated by settling, and the furnace is tilted to take out the slag into a slag ladle where it solidifies, and the precious lead into a noble lead ladle to be sent to the next silver separation process. When heating and melting the slime, there are two methods: using soda flux and producing iron-silicate slag, but the former has the drawbacks of high erosiveness of the slag and short lifespan of the furnace material. The latter method had the disadvantage of producing a large amount of slag due to the addition of iron. In the basic process of the present invention, which involves melting the copper electrolytic slime and producing high lead grade slag and high silver grade precious lead, a silicate-based flux is added to promote lead slagging. . At this time, scrap iron is not added to reduce the amount of slag produced. That is, in the conventional method, the lead content in the slag was in the range of 25 to 30%, but in the method of the present invention, the amount of reducing agent added is minimized in order to increase the lead content in the slag to 35 to 45%. That is, it is necessary to add coke or coal in an amount of 1 to 4% by weight based on the roasted slime. The amount of reducing agent added is 1
If it is less than 4% by weight, the effect of the addition is significantly reduced, and if it exceeds 4% by weight, no improvement in the effect can be seen and it becomes disadvantageous in terms of cost. The main slag components other than lead are silica (SiO ₂ ) and lime (CaO), and the slag composition ranges from 8 to 15% SiO ₂ and 1 to 3% CaO. Since the slime contains a small amount of SiO, the amount of flux added may be sufficient to generate the above-mentioned slag. For slime containing approximately 20% Pb and approximately 2.5% SiO, slag having the above target composition can be obtained by adding 3% coke, 1% silica sand, and 5.4% limestone. At this time, the silver content in the noble lead obtained is 50-60%. Pb% in slag
As shown in Figure 1, there is a positive correlation between Ag% and Ag% in noble lead, and if high lead grade slag is produced,
The quality of silver in precious lead is increased. However, at the same time, the loss of silver to the slag also increases. FIG. 2 conceptually shows the relationship between the silver slag/noble lead distribution ratio and the silver quality of the noble lead. Precious lead is processed in the next silver separation process to become coarse silver. The composition of noble lead is 4-10% lead, 10-15% selenium, 10-20% tellurium, in addition to the 50-60% silver mentioned above.
Copper content is 5-6%. In noble lead with a silver grade of about 40%, the lead content is about 20-25%, whereas in high-silver noble lead, the lead content is significantly lower, indicating that the slag formation of lead during the melting process is increased accordingly. Since the amount of lead to be oxidized during the silver separation process is small, the reaction time of the silver separation process is also shortened. As a result of the improved quality of silver, the slag production rate is higher than that of conventional slag with 25% Pb.
Only 65-70%. On the other hand, the content of silver, etc. is slightly high, with Ag0.5~3%, Se0.5~1.5%, Te1~
It is about 4%. In addition, since the amount of slag produced is small compared to the conventional method, the content of these slags does not increase significantly, so it can be sent as is to the lead smelting process. Next, in the process following the basic process of the method of the present invention, that is, in the process of making the high lead grade slag and high silver grade precious lead produced in the basic process useful, the slag is solidified, pulverized, and suspended. The yield of the above components can be increased by concentrating and recovering gold, silver, selenium, and tellurium as concentrates through ore beneficiation, repeating the melting process, and treating them simultaneously with the slime. The finer the particle size of the crushed slag, the better the beneficiation results, but a particle size of -100 microns is sufficient.
The slurry concentration is 100-200g/, pH8-10, pine oil is used as the foaming agent, and e.g.
In one case treated using DTP-8 etc. manufactured by Nippon Flavor Yakuhin Co., Ltd., Au0.07%, Ag2.2%, Se0.7%,
In the case of slag with Te1.8% and Pb44%, Au + Ag10
~15% concentrate is obtained about 15% of the slag, with tailings containing Au<0.001%, Ag0.3-0.35%, Se0.13%,
Te decreased to about 0.55%. Yield is Au>99%,
The content is approximately 90% Ag, 85% Se, and 75% Te. The above pine oil and DTP-8 (trade name) are as follows. The main component of pipe oil is terpineol, whose structural formula is It is. On the other hand, the main component of DTP-8 is dithiophosphate, and its structural formula is It is. The migration rate of lead into the tailings is approximately 90%, and the tailings can be sent to lead smelting to further recover the lead and the remaining small amounts of gold and silver. The concentrate is repeated in the slime melting process. Next, the present invention will be explained in more detail with reference to examples, but the scope of the present invention is not limited by the following examples. Example 1 Slime A 3.2t and slime B 2.2t shown in Table 1 were
Selenium was removed by oxidation roasting at 900℃ (estimated removal rate 85%), and then the in-process repeats shown in Table 1, 160 kg of coke, 240 g of limestone, and 45 kg of silica sand were additionally charged, and the furnace was rotated. At the same time, the temperature was raised to 1150-1200°C to dissolve the mixture. The solution was dissolved in about 10 hours and the reaction was completed, so the solution was allowed to stand for about 1 hour and the slurry and noble lead were extracted. The quality of these products is shown in Table 1. Estimated values for the amount of production are also listed in Table 2. The slag contained 42% Pb, approximately 4% Sb, and approximately 1.5% Bi. On the other hand, the noble lead was treated in a silver separation process according to a conventional method, but the treatment time was shortened to about 60% of the conventional noble lead with 40% Ag. In addition, compared to the conventional method, the amount of smoke ash generated during the Mitsuda and silver separation process is approximately 60% of that of the conventional method, and the amount of smoke ash generated by the Mitsuda + soda slag is approximately 60% of that of the conventional method.
reduced to 30%

【表】した。実施例２第１表に示す原料を焙焼し、繰返物とコークス
80Kg、石灰石240Kg、珪砂40Kgを追加装入して加
熱溶解した。産出物の組成を第２表に示す。【expressed. Example 2 The raw materials shown in Table 1 were roasted to produce repeated products and coke.
Additional charges of 80 kg, 240 kg of limestone, and 40 kg of silica sand were heated and melted. The composition of the output is shown in Table 2.

【表】貴鉛はさらに分銀工程で処理した。スラグは固
化し、浮遊選鉱処理した。粉砕粒度は−100ミク
ロン80％、スラリー濃度160ｇ／、PH８〜10、
起泡剤としてパイン油、捕集剤としてDTP−８
を用いて浮選した結果を第３表に示す。精鉱の生
成量は処理スラグ量の約15％であり、かつ該スラ
イム溶解工程に繰返されるテイリングへの鉛の移
行率は約90％であり、該精鉱は鉛製錬に送られ
る。金、銀は90％以上分離されるので、スライム
工程における収率は極めて高い。アンチモン、ビ
スマスのテイリングへの移行率は鉛とほぼ同様で
85〜90％であつた。[Table] Noble lead was further processed in a silver separation process. The slag was solidified and subjected to flotation treatment. Grinding particle size is -100 microns 80%, slurry concentration 160g/, PH8~10,
Pine oil as a foaming agent, DTP-8 as a collection agent
The results of flotation are shown in Table 3. The amount of concentrate produced is about 15% of the amount of processed slag, and the transfer rate of lead to the tailings repeated in the slime melting process is about 90%, and the concentrate is sent to lead smelting. Since over 90% of gold and silver are separated, the yield in the slime process is extremely high. The migration rate of antimony and bismuth to tailings is almost the same as that of lead.
It was 85-90%.

【表】（発明の効果）本発明は上記構成をとることによつて、次の効
果を奏することができる。 (1) シリケートベースのスラグを生成することに
より、スライム中の鉛、アンチモン、ビスムス
の大部分をスラグ化することができ、そのため
貴鉛中の銀品位が向上し、後処理工程の性能が
著しく向上する。また、密陀等の分解工程中間
物の生成量は著しく減少するので、スライム溶
解工程の負荷も低下する。 (2) ライム質フラツクスを併用することにより、
比較的低い温度で操業できる。ソーダ系フラツ
クスを用いる場合に問題となるスラグによる煉
瓦の侵食はシリカ質およびライム質フラツクス
を併用しているため、大幅に緩和され、炉の寿
命が長くなり、稼動率が高められる。 (3) スラグの鉛品位を35〜45％まで高めることが
できるため、従来の該スラグの鉛品位25％程度
の場合に比べて、スラグの生成量は65〜70％に
減少する。このため、スラグの再処理コストが
低減される。 (4) スラグへの金、銀、セレン、テルルの損失は
従来法に比してやや増加する傾向が認められる
が、スラグを固化、微粉砕し、浮遊選鉱処理を
行なうことにより、これら成分の75％〜99％を
精鉱として濃縮回収することができる。テイリ
ングへの鉛の移行率は約90％であり、上記浮選
精鉱をスライム溶解工程で混合処理しても、ス
ラグ生成量の増加はわずかである。[Table] (Effects of the Invention) By adopting the above configuration, the present invention can achieve the following effects. (1) By producing silicate-based slag, most of the lead, antimony, and bismuth in the slime can be converted into slag, which improves the silver quality in noble lead and significantly improves the performance of the post-treatment process. improves. Furthermore, since the amount of intermediates produced in the decomposition process, such as sludge, is significantly reduced, the load on the slime dissolution process is also reduced. (2) By using lime flux in combination,
Can operate at relatively low temperatures. The corrosion of bricks by slag, which is a problem when using soda-based flux, is greatly alleviated because siliceous and lime-based fluxes are used together, extending the life of the furnace and increasing its operating rate. (3) Since the lead quality of the slag can be increased to 35-45%, the amount of slag produced is reduced to 65-70% compared to the conventional case where the lead content of the slag is about 25%. Therefore, the cost of reprocessing the slag is reduced. (4) Although the loss of gold, silver, selenium, and tellurium in slag tends to increase slightly compared to the conventional method, by solidifying and pulverizing the slag and performing flotation treatment, 75% of these components can be removed. % to 99% can be concentrated and recovered as concentrate. The rate of lead transfer to the tailings is approximately 90%, and even if the flotation concentrate is mixed in the slime melting process, the amount of slag produced is only slightly increased.

[Brief explanation of the drawing]

第１図はスラグの鉛品位と貴鉛の銀品位の関係
を示す図、第２図は銀のスラグと貴鉛への分配比
を示す図である。 FIG. 1 is a diagram showing the relationship between the lead quality of slag and the silver quality of noble lead, and FIG. 2 is a diagram showing the distribution ratio of silver to slag and noble lead.

Claims

[Scope of Claims] 1. A method for smelting copper electrolytic slime by decoppering copper electrolytic slime or after decoppering, further roasting to remove selenium, adding flux and a reducing agent, and then heating and melting the slime. In the copper electrolytic slime, a flux consisting of a siliceous flux and a lime flux and 1 to 4 of the copper electrolytic slime are added.
% of the carbonaceous material and heated to dissolve the copper electrolytic slime, transferring a substantial majority of the base metal in the copper electrolytic slime into the slag, and transferring a high lead grade to the slag. A method for smelting copper electrolytic slime, which is characterized by producing slag and noble lead of high silver grade. 2. In a method of smelting copper electrolytic slime, which involves decoppering the copper electrolytic slime or further roasting to remove selenium after decoppering, and adding flux and a reducing agent to dissolve the slime, a siliceous flux is added to the copper electrolytic slime. A flux consisting of lime flux and a reducing agent consisting of a carbonaceous substance in the range of 1 to 4% of the copper electrolytic slime are added and heated to dissolve the copper electrolytic slime, and the base metal in the copper electrolytic slime is dissolved. A substantial part of the lead is transferred into the slag, and a high lead grade slag and a high silver grade noble lead are produced, and then the slag is solidified and pulverized,
Concentrate and recover gold, silver, selenium, and tellurium as concentrates, and use the base metals as tailings.
A method for smelting copper electrolytic slime, characterized in that the concentrate is reprocessed in the copper electrolytic slime melting step. 3. The method according to claim 1 or 2, wherein the siliceous flux is silica sand or silica stone. 4. The method according to claim 1 or 2, wherein the lime flux is limestone or quicklime. 5. The method according to claim 1 or 2, wherein the base metal is at least one of lead, antimony, and bismuth. 6. The method according to claim 1 or 2, wherein the carbonaceous material is coke or lime.