JP2020132957A - Silver recovery method - Google Patents

Abstract

To provide a method of efficiently recovering silver by separating lead and tin, from a raw material including silver and also lead, tin, bismuth, antimony and the like.SOLUTION: There is provided the silver recovery method for recovering silver by separating impurities such as lead and tin, from a raw material including chlorides of silver, lead, bismuth, antimony, and tin. The silver recovery method includes: a carbonation step of carbonating the raw material to convert lead chloride included in the raw material to lead carbonate; a nitric acid exudation step of adding nitric acid to a carbonated raw material to exude selectively lead carbonate, and performing solid-liquid separation; a hydrochloric acid exudation step of adding hydrochloric acid to a nitric acid exudation sullage to exude bismuth, and antimony, and performing solid-liquid separation; a silver reduction step of adding sulfuric acid to a hydrochloric exudation sullage, and further adding an iron powder to reduce the silver chloride of the hydrochloric exudation sullage to be crude silver; and an oxidation purification step of adding a soda-silicate slug to an iron reduction sullage, oxidation-melting, and making the slug absorb tin and impurities to separate and recover crude silver.SELECTED DRAWING: Figure 1

Description

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.

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.

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.

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.

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.

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.

Japanese Patent No. 4155176 Japanese Patent No. 3792056

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.

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.

[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.

[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.

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.

[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.

[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.

[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.

[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.

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.

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.

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.

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.

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.

Schematic process diagram of the processing method of the present invention.

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.

[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.

[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.

[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.

[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.

[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 a method of 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, a carbonated raw material. Nikkei leaching step of adding nitric acid to selectively leaching lead carbonate to separate solid and liquid, hydrochloric acid leaching step of adding strong hydrochloric acid to nitric acid slag and leaching bismuth and antimon to solid and liquid separation, Iron reduction step of adding sulfuric acid and further adding iron powder to reduce silver chloride in the hydrochloric acid leachate to make crude silver, adding soda silicate slag to the iron reducing slag and oxidatively melting it to absorb tin and impurities into the slag. A silver recovery method comprising an oxidation purification step of separating and recovering crude silver. 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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