JP2020132957A - Silver recovery method - Google Patents
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本発明は、銀と共に鉛、スズ、ビスマス、アンチモン等を含む原料から、鉛およびスズ等を分離して銀を効率よく回収する方法に関する。 The present invention relates to a method for efficiently recovering silver by separating lead, tin, etc. from a raw material containing lead, tin, bismuth, antimony, etc. together with silver.
銅電解スライムを湿式塩化処理して得た残渣等には、銀および鉛が多量に含まれており、このような処理滓等から銀と鉛を効率よく分離して銀を回収することが求められる。一方、近年、金銀滓の処理量の増加に伴って、上記塩化処理滓に含まれるビスマス、スズ、アンチモン等の不純物が多く含まれるようになり、銀と鉛を分離すると共にこれらの不純物を経済的に効率よく分離する必要がある。 Residues obtained by wet chlorination of copper electrolytic slime contain a large amount of silver and lead, and it is required to efficiently separate silver and lead from such treatment slags to recover silver. Be done. On the other hand, in recent years, as the amount of gold and silver slag processed has increased, impurities such as bismuth, tin, and antimony contained in the chloride-treated slag have come to be contained in large amounts, and silver and lead are separated and these impurities are economically used. It is necessary to separate efficiently.
従来、銀鉛含有滓から鉛を分離して銀を回収する方法として、特許文献1(特許第4155176号公報)に記載されている銀回収方法が知られている。この方法は、銀鉛塩化物を含有する原料を希硫酸でスラリー化し、これに鉄粉を添加して塩化銀を還元し、メタリックの銀を析出させる工程、このスラリーにさらに硫酸を添加して塩化鉛を硫酸鉛に転化して沈澱化し、メタリックの銀と硫酸鉛を含む混合物を回収する工程、この混合物を乾式熔錬し、銀を含むメタルと硫酸鉛を含むスラグとを形成させ、メタルとスラグを分離する工程、回収したメタルをソーダスラグと共に酸化精製して粗銀を得る工程を有する銀の回収方法である。 Conventionally, as a method for separating lead from a silver lead-containing slag and recovering silver, the silver recovery method described in Patent Document 1 (Patent No. 4155176) is known. In this method, a raw material containing silver lead chloride is made into a slurry with dilute sulfuric acid, iron powder is added to the slurry to reduce silver chloride, and metallic silver is precipitated. Further, sulfuric acid is added to this slurry. The process of converting lead chloride to lead sulfate and precipitating it to recover a mixture containing metallic silver and lead sulfate, this mixture is dry-melted to form metal containing silver and slag containing lead sulfate, and metal. This is a silver recovery method having a step of separating sulfuric acid and sulfuric acid and a step of oxidizing and refining the recovered metal together with soda slag to obtain crude silver.
しかし、この処理方法は、原料に銀および鉛と共にスズ、ビスマス、アンチモン等が含まれると、乾式溶錬の際にスラグの粘度が増加して操業上のハンドリングが難しくなる問題がある。また、スズ、ビスマス、アンチモン等は硫酸に浸出し難いので、これらが十分に分離されずに粗銀中に残留する。また、この粗銀をソーダスラグと共に酸化溶錬して不純物をスラグ化して除去する場合、スズ、ビスマス、アンチモン等を十分に除去するのは難しい。 However, this treatment method has a problem that if the raw materials contain tin, bismuth, antimony, etc. together with silver and lead, the viscosity of the slag increases during dry smelting, which makes operational handling difficult. Further, since tin, bismuth, antimony and the like are difficult to leached into sulfuric acid, they are not sufficiently separated and remain in crude silver. Further, when this crude silver is oxidatively smelted together with soda slag to slag and remove impurities, it is difficult to sufficiently remove tin, bismuth, antimony and the like.
一方、特許文献2(特許第3792056号公報)には、鉛、スズ、ビスマスを含む鉛滓からこれらの金属を分離する方法が記載されている。この処理方法は、鉛、錫、ビスマスを含む鉛滓を炭酸化して炭酸化滓にし、該炭酸化滓を硝酸溶解して鉛を含む溶解液と錫とビスマスとを含む溶解残渣にし、鉛を含む溶解液から精鉛滓を回収し、錫とビスマスとを含む溶解残渣を塩酸溶解して錫を含む溶解残渣とビスマスを含む溶解液にし、これを固液分離して錫とビスマスを分離する方法である。 On the other hand, Patent Document 2 (Patent No. 3792056) describes a method for separating these metals from lead slag containing lead, tin and bismuth. In this treatment method, a lead slag containing lead, tin, and bismuth is carbonated to form a carbonated slag, and the carbonated slag is dissolved with nitrate to form a solution containing lead and a dissolution residue containing tin and bismuth, and lead is added. The refined lead slag is recovered from the containing solution, and the solution residue containing tin and bismuth is dissolved in hydrochloric acid to obtain a solution containing the solution residue containing tin and bismuth, which is solid-liquid separated to separate tin and bismuth. The method.
しかし、この処理方法は鉛、スズ、ビスマスを分離することができるが、原料に銀が含まれていると、銀は溶解されずに錫の溶解滓に残るので、このままでは銀を分離回収することができない。 However, although lead, tin, and bismuth can be separated by this treatment method, if silver is contained in the raw material, silver is not dissolved and remains in the tin slag, so silver is separated and recovered as it is. Can't.
本発明は、従来の上記処理方法の問題を解決したものであり、銀と共に鉛、スズ、ビスマス、アンチモン等を含む原料から、鉛およびスズ等を分離して銀を効率よく回収する方法を提供する。 The present invention solves the problem of the above-mentioned conventional treatment method, and provides a method for efficiently recovering silver by separating lead, tin, etc. from a raw material containing lead, tin, bismuth, antimony, etc. together with silver. To do.
本発明は以下の銀回収方法に関する。
〔1〕銀、鉛、ビスマス、アンチモン、およびスズの塩化物を含む原料から銀を回収する方法において、該原料を炭酸化して該原料に含まれる塩化鉛を炭酸鉛にする炭酸化工程、炭酸化した原料に硝酸を加えて炭酸鉛を選択的に浸出して固液分離する硝酸浸出工程、硝酸浸出滓に強塩酸を加えてビスマスおよびアンチモンを浸出して固液分離する塩酸浸出工程、塩酸浸出滓に硫酸を加え、さらに鉄粉を加えて塩酸浸出滓の塩化銀を還元して粗銀にする鉄還元工程、鉄還元滓にソーダシリケートスラグを加えて酸化熔融し、スズおよび不純物をスラグに吸収させて粗銀を分離回収する酸化精製工程を有することを特徴とする銀回収方法。
〔2〕酸化精製工程において、Na/Siモル比が1.5〜4のソーダシリケートスラグを用い、鉄還元滓量に対して0.5〜1.0倍質量当量のソーダシリケートスラグを加えて、1100〜1300℃で酸化熔融する上記[1]に記載する銀回収方法。
The present invention relates to the following silver recovery method.
[1] In a method for recovering silver from a raw material containing chlorides of silver, lead, bismuth, antimony, and tin, a carbonization step of carbonizing the raw material to convert lead chloride contained in the raw material into lead carbonate, carbonic acid. Nitric acid leaching step of adding nitric acid to the converted raw material to selectively leaching lead carbonate to separate solid and liquid, hydrochloric acid leaching step of adding strong hydrochloric acid to the nitrate slag and leaching bismuth and antimony to separate solid and liquid, hydrochloric acid Nitric acid is added to the leachate, and iron powder is added to reduce the silver chloride in the hydrochloric acid leachate to make crude silver. The iron reduction slag is oxidatively melted by adding soda silicate slag to slag tin and impurities. A silver recovery method comprising an oxidation purification step of separating and recovering crude silver by absorbing it in nitric acid.
[2] In the oxidative purification step, soda silicate slag having a Na / Si molar ratio of 1.5 to 4 is used, and soda silicate slag having a mass equivalent of 0.5 to 1.0 times the amount of iron reducing slag is added. , The silver recovery method according to the above [1], which is oxidatively melted at 1100 to 1300 ° C.
〔具体的な説明〕
本発明は、銀、鉛、ビスマス、アンチモン、およびスズの塩化物を含む原料から銀を回収する方法において、該原料を炭酸化して該原料に含まれる塩化鉛を炭酸鉛にする炭酸化工程、炭酸化した原料に硝酸を加えて炭酸鉛を選択的に浸出して固液分離する硝酸浸出工程、硝酸浸出滓に強塩酸を加えてビスマスおよびアンチモンを浸出して固液分離する塩酸浸出工程、塩酸浸出滓に硫酸を加え、さらに鉄粉を加えて塩酸浸出滓の塩化銀を還元して粗銀にする鉄還元工程、鉄還元滓にソーダシリケートスラグを加えて酸化熔融し、スズおよび不純物をスラグに吸収させて粗銀を分離回収する酸化精製工程を有することを特徴とする銀回収方法である。
本発明の処理方法の概略を図1に示す。
[Specific explanation]
The present invention is a method for recovering silver from a raw material containing chlorides of silver, lead, bismuth, antimony, and tin, in which the raw material is carbonated to convert the lead chloride contained in the raw material into lead carbonate. Nitric acid leaching step in which nitric acid is added to the carbonated raw material to selectively leaching lead carbonate and solid-liquid separation, hydrochloric acid leaching step in which strong hydrochloric acid is added to the nitrate slag to leach bismuth and antimon and solid-liquid separation An iron reduction step in which sulfuric acid is added to the hydrochloric acid slag and iron powder is further added to reduce silver chloride in the hydrochloric acid leaching slag to make crude silver. It is a silver recovery method characterized by having an oxidation purification step of absorbing crude silver into slag and separating and recovering crude silver.
The outline of the processing method of this invention is shown in FIG.
〔炭酸化工程〕
銀、鉛、ビスマス、アンチモン、およびスズの塩化物を含む原料を炭酸化して該原料に含まれる塩化鉛を炭酸鉛にする。銀、鉛、ビスマス、アンチモン、およびスズを含む原料として、例えば、銅電解スライムの塩化浸出滓などが用いられる。該原料に加える炭酸源は炭酸ナトリウムを用いると良い。原料1g当たり0.2〜0.5gの炭酸ナトリウムが好ましい。該原料を水スラリーにし、このスラリーを撹拌しながら炭酸ナトリウムを添加して原料に含まれている塩化鉛を炭酸鉛にする。銀、ビスマス、アンチモン、およびスズの塩化物は炭酸塩になり難いので、鉛が選択的に炭酸化される。
[Carbonation process]
A raw material containing chlorides of silver, lead, bismuth, antimony, and tin is carbonated to convert lead chloride contained in the raw material into lead carbonate. As a raw material containing silver, lead, bismuth, antimony, and tin, for example, chloride leaching slag of copper electrolytic slime is used. Sodium carbonate may be used as the carbonic acid source to be added to the raw material. 0.2 to 0.5 g of sodium carbonate per 1 g of the raw material is preferable. The raw material is made into a water slurry, and sodium carbonate is added while stirring the slurry to convert lead chloride contained in the raw material into lead carbonate. Chlorides of silver, bismuth, antimony, and tin are less likely to carbonate, so lead is selectively carbonated.
炭酸化のスラリーのpHが9〜10で安定することを確認し、固液分離して炭酸化滓を回収する。分離して液分はアルカリを含むので、次工程の硝酸浸出に影響しないように固液分離して液分を除くと良い。 After confirming that the pH of the carbonated slurry is stable at 9 to 10, solid-liquid separation is performed to recover the carbonated slag. Since the liquid is separated and contains alkali, it is advisable to separate the liquid by solid-liquid separation so as not to affect the leaching of nitric acid in the next step.
〔硝酸浸出工程〕
固液分離した炭酸化滓に水を加えてスラリーにし、これに硝酸を加えて炭酸鉛を浸出する。硝酸の濃度は市販品を用いれば良い。該水スラリーに硝酸を加えると反応中は発泡するので、発泡が止んでスラリーのpHが2以下になれば反応が終了する。該スラリーに含まれている銀、ビスマス、アンチモン、およびスズの塩化物は硝酸浸出せずに固形分に残り、炭酸鉛が選択的に浸出される。この鉛を含む液分を固液分離して硝酸浸出滓を回収する。
[Nitric acid leaching process]
Water is added to the solid-liquid separated carbonated slag to form a slurry, and nitric acid is added to this to leach lead carbonate. A commercially available product may be used for the concentration of nitric acid. When nitric acid is added to the water slurry, it foams during the reaction. Therefore, when the foaming stops and the pH of the slurry becomes 2 or less, the reaction is completed. Chlorides of silver, bismuth, antimony, and tin contained in the slurry remain in the solid content without leaching nitric acid, and lead carbonate is selectively leached. This lead-containing liquid is solid-liquid separated to recover the nitric acid leaching slag.
〔塩酸浸出工程〕
硝酸浸出滓に塩酸を加えてビスマスおよびアンチモンを浸出する。塩酸は5〜7mol/Lの濃塩酸が好ましい。約65〜85℃に加熱し、4〜24時間撹拌しながら反応させると良い。上記硝酸浸出滓に含まれる塩化ビスマスおよび塩化アンチモンは塩酸によって浸出され、該滓に含まれる塩化銀および塩化スズは浸出されずに残留する。ビスマスおよびアンチモンを含む液分を固液分離して塩酸浸出滓を回収する。この塩酸浸出滓には塩化銀および塩化スズが含まれる。
[Hydrochloric acid leaching process]
Hydrochloric acid is added to the nitric acid leaching residue to leach bismuth and antimony. The hydrochloric acid is preferably concentrated hydrochloric acid of 5 to 7 mol / L. It is good to heat to about 65-85 ° C. and react with stirring for 4 to 24 hours. Bismuth oxychloride and antimony chloride contained in the nitric acid leaching slag are leached by hydrochloric acid, and silver chloride and tin chloride contained in the slag remain without being leached out. The liquid containing bismuth and antimony is separated into solid and liquid to recover the hydrochloric acid leaching residue. This hydrochloric acid leaching residue contains silver chloride and tin chloride.
〔鉄還元工程〕
上記塩酸浸出滓には塩化銀および塩化スズが含まれているので、この塩酸浸出滓に硫酸を添加してpH1以下の酸性スラリーにし、該スラリーに鉄粉を加えて塩化銀を還元して鉄還元滓にする。pH1以下の酸性下にして塩化銀の還元を進める。鉄粉の量は水ラリー中の銀含有量の1〜2倍質量当量が好ましい。塩化銀の還元によってスラリー温度が上昇するが、必要に応じて60℃〜90℃に加熱して鉄の還元を進めると良い。反応後、固液分離して鉄還元滓を回収する。
[Iron reduction process]
Since the hydrochloric acid leaching slag contains silver chloride and tin chloride, sulfuric acid is added to the hydrochloric acid leaching slag to make an acidic slurry having a pH of 1 or less, and iron powder is added to the slurry to reduce silver chloride to iron. Make it a reducing slag. The reduction of silver chloride is promoted under acidic conditions of pH 1 or less. The amount of iron powder is preferably 1 to 2 times the mass equivalent of the silver content in the water rally. Although the slurry temperature rises due to the reduction of silver chloride, it is advisable to heat the slurry to 60 ° C. to 90 ° C. as necessary to promote the reduction of iron. After the reaction, solid-liquid separation is performed to recover the iron reducing slag.
〔酸化精製工程〕
上記鉄還元滓にソーダシリケートスラグを加えて酸化熔融し、スズおよび残留するアンチモンやビスマスなどの不純物をスラグに吸収させて粗銀と分離し、該粗銀を回収する。ソーダシリケートスラグは炭酸ソーダと硝酸ソーダおよび珪砂の混合物が好ましい。このソーダシリケートスラグの酸化力を利用して鉄還元滓中の鉛、スズ、ビスマス、アンチモン等の不純物を酸化してスラグ化し、これをソーダシリケートスラグに吸収させる。鉄還元滓に含まれる銀は酸化されずにメタルの粗銀として残るので上記不純物と分離することができる。この粗銀を回収する。
[Oxidation purification process]
Soda silicate slag is added to the iron reduction slag and oxidatively melted, and tin and residual impurities such as antimony and bismuth are absorbed by the slag to separate it from crude silver, and the crude silver is recovered. The soda silicate slag is preferably a mixture of sodium carbonate, sodium nitrate and silica sand. The oxidizing power of this soda silicate slag is used to oxidize impurities such as lead, tin, bismuth, and antimony in the iron reducing slag to form slag, which is absorbed by the soda silicate slag. Since the silver contained in the iron reduction slag remains as crude silver of the metal without being oxidized, it can be separated from the above impurities. Collect this crude silver.
ソーダシリケートスラグの炭酸ソーダと硝酸ソーダおよび珪砂の混合比は、Na/Siモル比が1.5〜4の範囲が好ましい。該Na/Siモル比が1.5未満ではスラグの粘性が高くなり、スラグと粗銀が分離し難くなる。一方、Na/Siモル比が4を超えると、相対的にシリカ量が少なく、鉛やスズ等の不純物を十分にスラグ化することができない。 The mixing ratio of sodium carbonate, sodium nitrate and silica sand in the soda silicate slag is preferably in the range of 1.5 to 4 in the Na / Si molar ratio. If the Na / Si molar ratio is less than 1.5, the viscosity of the slag becomes high, and it becomes difficult to separate the slag and the crude silver. On the other hand, when the Na / Si molar ratio exceeds 4, the amount of silica is relatively small, and impurities such as lead and tin cannot be sufficiently slagged.
なお、特許文献1の方法は、乾式熔錬で回収したメタルを、シリカを含まないソーダスラグと共に酸化溶融して粗銀を回収しているが、シリカを含まないソーダスラグを用いると、スズ、ビスマス、アンチモン等の不純物は十分にスラグ化されず、粗銀中の不純物濃度が高くなる。 In the method of Patent Document 1, crude silver is recovered by oxidizing and melting the metal recovered by dry smelting together with soda slag containing no silica. However, when soda slag containing no silica is used, tin, bismuth, etc. Impurities such as antimony are not sufficiently slagged, and the concentration of impurities in crude silver increases.
ソーダシリケートスラグの添加量は鉄還元滓量に対して0.5〜1.0倍質量当量が適量である。ソーダシリケートスラグの添加量がこれより少ないと上記不純物が十分にスラグ化されない。一方、ソーダシリケートスラグの添加量がこれより多いとスラグ量が増大し、後処理の負担が増す。 The appropriate amount of soda silicate slag added is 0.5 to 1.0 times the mass equivalent of the amount of iron reduced slag. If the amount of soda silicate slag added is less than this, the above impurities are not sufficiently slagged. On the other hand, if the amount of soda silicate slag added is larger than this, the amount of slag increases and the burden of post-treatment increases.
酸化精製工程の熔錬温度は1100℃〜1300℃が好ましい。この酸化熔錬による鉄還元滓の精製によって、回収される粗銀中の不純物は概ね0.1質量%以下になり、電解精製に適する不純物レベルの粗銀を得ることができる。 The melting temperature in the oxidative purification step is preferably 1100 ° C to 1300 ° C. By refining the iron reducing slag by this oxidative melting, the impurities in the recovered crude silver become about 0.1% by mass or less, and crude silver at an impurity level suitable for electrolytic purification can be obtained.
本発明の処理方法によれば、原料に含まれているビスマスおよびアンチモンの約80%以上を系外に除去することができ、スズは40%を除去することができる。
この結果、不純物の負荷によって生じる酸化精製工程での銀及び金を含む貴金属の損失を大幅に低減することができる。
According to the treatment method of the present invention, about 80% or more of bismuth and antimony contained in the raw material can be removed from the system, and 40% of tin can be removed.
As a result, the loss of precious metals including silver and gold in the oxidative purification process caused by the loading of impurities can be significantly reduced.
以下、本発明の実施例および比較例を示す。
なお、原料から鉄還元滓への移行率は次式[1]に従って算出した。
移行率[%]=[滓中の金属質量]/[原料の金属質量]〕×100 ・・・[1]
粗銀の品位は鉛、ビスマスアンチモン、スズなどの不純物濃度を100質量%から差し引いた値である。
Hereinafter, examples and comparative examples of the present invention will be shown.
The transfer rate from the raw material to the iron reduction slag was calculated according to the following equation [1].
Transition rate [%] = [Metal mass in slag] / [Metal mass of raw material]] × 100 ・ ・ ・ [1]
The grade of crude silver is a value obtained by subtracting the concentration of impurities such as lead, bismuth antimony, and tin from 100% by mass.
〔実施例1〕
銀、鉛、ビスマス、アンチモン、スズを含む塩化浸出滓を用いた。原料の成分を表1に示す。この原料600gに水を2L加えてスラリー化した。この水スラリーを攪拌しながら炭酸ナトリウムを150g添加した。pHが9〜10で安定することを確認した後、固液分離を行い、炭酸化滓620gを回収した。
この炭酸化滓に水を2L加えてスラリー化した。この水スラリーにpH2になるまで硝酸を添加した。反応中は発泡しているが、発泡が止まってpH2以下になったところで反応終了とした。反応物を固液分離し、硝酸浸出滓260gを回収した。
この硝酸浸出滓に、6mol/Lの塩酸を2L添加し、攪拌しながら75℃まで加温した。4時間後、反応を終了させ、固液分離を行い、塩酸浸出滓230gを回収した。
塩酸浸出滓に水を1L加え、さらにpHが1以下になるまで硫酸を添加して硫酸スラリーにした。これに鉄粉を40g添加してスラリー中の塩化銀を還元した。反応終了後、固液分離を行い、鉄還元滓を得た。この鉄還元滓193gにソーダシリケートスラグ(Na/Siモル比4)200g加え、1150℃で1時間加熱熔融させ、スラグを分離して粗銀を得た。スラグ中に混入する銀は2質量%未満であった。回収した粗銀の品位は99.5質量%、粗銀中の不純物濃度は何れも0.1質量%未満であった。これらの結果を表1に示す。
[Example 1]
Chloride leaching slag containing silver, lead, bismuth, antimony and tin was used. The components of the raw material are shown in Table 1. 2 L of water was added to 600 g of this raw material to form a slurry. 150 g of sodium carbonate was added while stirring this water slurry. After confirming that the pH was stable at 9 to 10, solid-liquid separation was carried out, and 620 g of carbonated slag was recovered.
2 L of water was added to the carbonated slag to form a slurry. Nitric acid was added to this water slurry until the pH reached 2. Although it foamed during the reaction, the reaction was terminated when the foaming stopped and the pH became 2 or less. The reaction product was separated into solid and liquid, and 260 g of nitric acid leaching slag was recovered.
To this nitric acid leaching residue, 2 L of 6 mol / L hydrochloric acid was added, and the mixture was heated to 75 ° C. with stirring. After 4 hours, the reaction was terminated, solid-liquid separation was performed, and 230 g of hydrochloric acid leaching slag was recovered.
1 L of water was added to the hydrochloric acid leaching slag, and sulfuric acid was further added until the pH became 1 or less to prepare a sulfuric acid slurry. 40 g of iron powder was added thereto to reduce silver chloride in the slurry. After completion of the reaction, solid-liquid separation was carried out to obtain an iron reducing slag. 200 g of soda silicate slag (Na / Si molar ratio 4) was added to 193 g of this iron reducing slag and melted by heating at 1150 ° C. for 1 hour, and the slag was separated to obtain crude silver. The amount of silver mixed in the slag was less than 2% by mass. The grade of the recovered crude silver was 99.5% by mass, and the concentration of impurities in the crude silver was less than 0.1% by mass. These results are shown in Table 1.
〔実施例2、3〕
実施例1の塩酸浸出時間を8時間(実施例2)、24時間(実施例3)およびソーダシリケートスラグ添加量を100g(実施例3)にした以外は実施例1と同様にして粗銀とスラグを回収した。いずれもスラグ中に混入する銀は2質量%未満であり、回収した粗銀の品位は99.2質量%、99.7質量%であり、粗銀中の不純物濃度は何れも0.1質量%未満であった。これらの結果を表1に示す。
[Examples 2 and 3]
With crude silver in the same manner as in Example 1, except that the hydrochloric acid leaching time of Example 1 was 8 hours (Example 2), 24 hours (Example 3), and the amount of soda silicate slag added was 100 g (Example 3). The slag was recovered. In each case, the amount of silver mixed in the slag was less than 2% by mass, the grade of the recovered crude silver was 99.2% by mass and 99.7% by mass, and the impurity concentration in the crude silver was 0.1% by mass. Was less than%. These results are shown in Table 1.
〔比較例1〕
実施例1と同様の原料を使用し、実施例1と同様に炭酸化処理および硝酸浸出処理を行った後に、硝酸浸出滓B260gについて実施例1と同様の鉄還元処理を行い、鉄還元滓を回収した。この鉄還元滓195gにソーダスラグ(炭酸ソーダ:硝酸ソーダ=2.5:1))200gを加え、1150℃で1時間加熱熔融させ、スラグを分離して粗銀を得た。スラグの粘度が高く粗銀とスラグの分離が難しく、スラグ中に混入する銀は5質量%以上であった。回収した粗銀の品位は95質量%、粗銀中の鉛、ビスマス、スズ、アンチモン濃度は0.1〜0.8質量%であった。これらの結果を表2に示す。
[Comparative Example 1]
Using the same raw materials as in Example 1, carbonation treatment and nitric acid leaching treatment were performed in the same manner as in Example 1, and then the same iron reduction treatment as in Example 1 was performed on 260 g of nitric acid leaching slag B to obtain iron reducing slag. Recovered. 200 g of soda slag (sodium carbonate: sodium nitrate = 2.5: 1) was added to 195 g of this iron reducing slag and melted by heating at 1150 ° C. for 1 hour, and the slag was separated to obtain crude silver. The viscosity of the slag was high and it was difficult to separate the crude silver and the slag, and the amount of silver mixed in the slag was 5% by mass or more. The grade of the recovered crude silver was 95% by mass, and the concentrations of lead, bismuth, tin and antimony in the crude silver were 0.1 to 0.8% by mass. These results are shown in Table 2.
〔比較例2〕
実施例1と同様の原料を使用し、実施例1と同様に炭酸化工程、硝酸浸出工程を行い、塩酸浸出に代えて硫酸浸出を行い、鉄還元工程を行って鉄還元滓を回収した。
この鉄還元滓195gに比較例1と同様のソーダスラグ200gを加え、1150℃で1時間加熱熔融させ、スラグを分離して粗銀を得た。スラグの粘性が高く粗銀とスラグの分離が難しく、スラグ中に混入する銀は4質量%以上であった。粗銀品位及び不純物濃度は比較例1と同様の値であった。
[Comparative Example 2]
Using the same raw materials as in Example 1, a carbonization step and a nitric acid leaching step were performed in the same manner as in Example 1, sulfuric acid leaching was performed instead of hydrochloric acid leaching, and an iron reduction step was performed to recover the iron reduction slag.
200 g of soda slag similar to that of Comparative Example 1 was added to 195 g of this iron reduction slag and melted by heating at 1150 ° C. for 1 hour, and the slag was separated to obtain crude silver. The viscosity of the slag was high and it was difficult to separate the crude silver from the slag, and the amount of silver mixed in the slag was 4% by mass or more. The crude silver grade and the impurity concentration were the same values as in Comparative Example 1.
〔比較例3〕
実施例1の炭酸化工程および硝酸化工程を省略し、原料600gに硫酸2Lを添加して、75℃に加熱し、8時間硫酸浸出を行い、硫酸浸出滓を回収した。この硫酸浸出滓255gに鉄粉40gを添加して鉄還元を行い、鉄還元滓を回収した。
この鉄還元滓180gに比較例1と同様のソーダスラグを加え、1150℃で1時間加熱熔融させ、スラグを分離して粗銀を得た。スラグの粘性が高く、粗銀とスラグの分離が困難となり、スラグ中に混入する銀は25質量%以上であり、粗銀中にも多くのスラグが混在していた。
[Comparative Example 3]
The carbonation step and the nitrate step of Example 1 were omitted, 2 L of sulfuric acid was added to 600 g of the raw material, the mixture was heated to 75 ° C., sulfuric acid was leached for 8 hours, and the sulfuric acid leaching slag was recovered. 40 g of iron powder was added to 255 g of the sulfuric acid leaching slag to reduce iron, and the iron reducing slag was recovered.
The same soda slag as in Comparative Example 1 was added to 180 g of this iron reduction slag and melted by heating at 1150 ° C. for 1 hour, and the slag was separated to obtain crude silver. The viscosity of the slag was high, making it difficult to separate the crude silver and the slag. The amount of silver mixed in the slag was 25% by mass or more, and a large amount of slag was also mixed in the crude silver.
Claims (2)
In the oxidative purification step, soda silicate slag having a Na / Si molar ratio of 1.5 to 4 is used, and soda silicate slag having a mass equivalent of 0.5 to 1.0 times the amount of iron reducing slag is added to 1100 to The silver recovery method according to claim 1, wherein oxidative melting is performed at 1300 ° C.
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JP2000109939A (en) * | 1998-10-05 | 2000-04-18 | Nippon Mining & Metals Co Ltd | Separation of lead, tin and bismuth from lead slag |
JP2001316736A (en) * | 2000-03-03 | 2001-11-16 | Nippon Mining & Metals Co Ltd | Method for recovering silver |
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JP2000109939A (en) * | 1998-10-05 | 2000-04-18 | Nippon Mining & Metals Co Ltd | Separation of lead, tin and bismuth from lead slag |
JP2001316736A (en) * | 2000-03-03 | 2001-11-16 | Nippon Mining & Metals Co Ltd | Method for recovering silver |
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CN116377240A (en) * | 2023-04-23 | 2023-07-04 | 四川轻化工大学 | Method for purifying Ag by washing precipitate obtained by OLED silver target hydrochloric acid |
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