1. A method for extracting silver from a high-iron high-tin high-indium flotation silver concentrate is characterized by comprising the following steps:

step 1: pretreatment of

Taking high-iron high-tin high-indium flotation silver concentrate, zinc electrolysis waste liquid and hydrogen peroxide, wherein the liquid-solid ratio of the zinc electrolysis waste liquid to the high-iron high-tin high-indium flotation silver concentrate is (6-10) L:1kg, the volume ratio of the zinc electrolysis waste liquid to the hydrogen peroxide is (25-40) to 1, reacting the three to obtain pretreatment liquid and pretreatment slag, and returning the pretreatment liquid to a leaching working section of a zinc system;

step 2: thiourea leaching

Adding the circulating agent obtained in the step 3 into the pretreatment slag obtained in the step 1, wherein the liquid-solid ratio of the circulating agent to the pretreatment slag is (5-7) L:1kg, the pH value is adjusted to 1.0-2.5, and a leaching reaction is carried out to obtain a thiourea leaching solution and thiourea leaching slag, wherein the thiourea leaching slag returns to the rotary kiln;

and step 3: replacement of

And (3) replacing the thiourea leaching solution obtained in the step (2) with zinc powder to obtain silver cotton, copper cotton and a replaced solution, adding thiourea into the replaced solution, wherein the solid-to-liquid ratio of the thiourea to the replaced solution is (29-33) g:1L to obtain a circulating agent, returning the circulating agent to the step (2), and selling the silver cotton and the copper cotton.

2. The method for extracting silver from the high-iron high-tin high-indium flotation silver concentrate according to claim 1, wherein in the step 1, the high-iron high-tin high-indium flotation silver concentrate comprises the following components in percentage by mass: 28-31% of Zn, 2-3% of Cu, 0.29-0.34% of Ag, 19-22% of Fe, 0.2-0.3% of As, 0.04-0.06% of In, 0.5-0.7% of Pb, 24-26% of S, 0.15-0.25% of Sn, 23-25% of O, and the sum of Si and Ca is less than 2.5%.

3. The method for extracting silver from the flotation silver concentrate with high iron, high tin and high indium as claimed in claim 1, wherein in the step 1, the initial acid of the zinc electrolysis waste liquid is 150g/L-200g/L, the manganese ion concentration is 3.5g/L-6g/L, the fluorine ion concentration is 20mg/L-100mg/L, and the chlorine ion concentration is 100mg/L-500 mg/L.

4. The method for extracting silver from the flotation silver concentrate with high iron, high tin and high indium as claimed in claim 1, wherein in the step 1, the hydrogen peroxide is added in 3 times at 30min intervals, and the mass fraction of the hydrogen peroxide is 20% -30%.

5. The method for extracting silver from the flotation silver concentrate with high iron, high tin and high indium as claimed in claim 1, wherein the reaction temperature in step 1 is 75-85 ℃ and the reaction time is 90-150 min.

6. The method for extracting silver from the high-iron high-tin high-indium flotation silver concentrate according to claim 1, wherein in the step 2, concentrated sulfuric acid is adopted for adjusting the pH value.

7. The method for extracting silver from the flotation silver concentrate with high iron, high tin and high indium according to any one of claims 1 to 6, wherein the temperature of the leaching reaction in the step 2 is 35 ℃ to 45 ℃ and the time is 3.5h to 5 h.

Background

In the process of zinc hydrometallurgy, a large amount of acid leaching residues are generated in the low acid leaching process, and basically all silver is enriched in the acid leaching residues. At present, the zinc smelting industry mostly adopts a combined process of selection and smelting to recover silver in zinc leaching residues and can also recover copper in the zinc leaching residues, thereby increasing the economic benefit of enterprises, reducing the discharge amount of residues and protecting the environment. The flotation process of the zinc leaching residues has the advantages of low cost, low investment and less pollution, and zinc smelting enterprises can perform resource utilization and improve enterprise competitiveness.

In order to respond to national policies, reasonably utilize resources and protect the environment, and increase enterprise benefits, it is necessary to find more reasonable and more economic process schemes for smelting silver sponge and copper sponge for some complicated silver concentrate types, such as high-iron, high-tin and high-indium flotation silver concentrates, so as to bring more remarkable economic benefits.

At present, the recovery mode of the zinc leaching residue flotation silver concentrate is mainly divided into two types, namely an indirect method and a direct method. The indirect method mainly aims at enterprises of lead and copper smelting, and the flotation silver concentrate is added into the lead and copper concentrate in a burdening mode, and then silver is recovered from anode mud. The direct method mainly comprises a cyanidation method, a thiourea method, a chlorination method and a roasting-leaching-replacement method.

The method has the following disadvantages: the indirect method can only sell the flotation silver concentrate for enterprises without lead smelting or copper smelting, and the cost for recovering the zinc in the flotation silver concentrate is very high; the cyanidation method is more suitable for flotation of silver concentrate, has high gold content and general leaching rate of silver, and simultaneously, cyanide is extremely toxic, so the use is limited; the thiourea method can greatly increase the thiourea content when the flotation silver concentrate contains other impurities which react with thiourea. Thiourea is expensive compared to other agents, and at the same time, thiourea oxidizes readily, the higher the temperature the faster, resulting in high overall costs. The chlorination method mainly aims at flotation of silver concentrate containing lead and copper. The roasting-leaching-replacing method has long process flow, more filtering times, poorer separation of solid and liquid and low efficiency.

In view of the above, there is a need to provide a method for extracting silver from a high-iron, high-tin and high-indium flotation silver concentrate, so as to solve the deficiencies of the prior art.

Disclosure of Invention

The invention aims to make up the defects of the prior art and provides a method for extracting silver from high-iron, high-tin and high-indium flotation silver concentrate. The method can efficiently extract the silver sponge and the copper sponge from the high-iron, high-tin and high-indium flotation silver concentrate, has obvious economic benefit, simple extraction method, easy operation, low cost and wide market prospect, and is suitable for large-scale popularization and application.

The technical scheme for solving the technical problems is as follows: a method for extracting silver from a high-iron high-tin high-indium flotation silver concentrate comprises the following steps:

step 1: pretreatment of

Taking high-iron high-tin high-indium flotation silver concentrate, zinc electrolysis waste liquid and hydrogen peroxide, wherein the liquid-solid ratio of the zinc electrolysis waste liquid to the high-iron high-tin high-indium flotation silver concentrate is (6-10) L:1kg, the volume ratio of the zinc electrolysis waste liquid to the hydrogen peroxide is (25-40) to 1, reacting the three to obtain pretreatment liquid and pretreatment slag, and returning the pretreatment liquid to a leaching working section of a zinc system;

step 2: thiourea leaching

Adding the circulating agent obtained in the step 3 into the pretreatment slag obtained in the step 1, wherein the liquid-solid ratio of the circulating agent to the pretreatment slag is (5-7) L:1kg, the pH value is adjusted to 1.0-2.5, and a leaching reaction is carried out to obtain a thiourea leaching solution and thiourea leaching slag, wherein the thiourea leaching slag returns to the rotary kiln;

and step 3: replacement of

And (3) replacing the thiourea leaching solution obtained in the step (2) with zinc powder to obtain silver cotton, copper cotton and a replaced solution, adding thiourea into the replaced solution, wherein the solid-to-liquid ratio of the thiourea to the replaced solution is (29-33) g:1L to obtain a circulating agent, returning the circulating agent to the step (2), and selling the silver cotton and the copper cotton.

The principle of the invention is as follows:

in step 1 of the invention, the high-iron, high-tin and high-indium flotation silver concentrate is a complex silver concentrate and contains ZnS, ZnO and ZnSO₄、ZnFe₂O₄、Zn₂SiO₄、Fe₂O₃FeS, AgS, Ag, and the like. The silver phase analysis results are shown in Table 1, and the zinc phase analysis results are shown in Table 2.

TABLE 1 silver phase analysis results

TABLE 2 Zinc phase analysis results

The high-iron high-tin high-indium flotation silver concentrate is pretreated by zinc electrolysis waste liquid and hydrogen peroxide to generate pretreatment liquid and pretreatment slag. The zinc electrolysis waste liquid is a special solution, contains oxidative high-valence manganese besides high acid, and also contains fluorine chloride ions and the like, can play a comprehensive effect when used for pretreatment, and can further play a role in strengthening the comprehensive effect when matched with a small amount of hydrogen peroxide with strong oxidation without adding any impurities.

The chemical reaction involved in step 1 is as follows:

2ZnS+H₂O₂+2H₂SO₄＝2ZnSO₄+2H₂O+2S；

2FeS+3H₂O₂+4H₂SO₄＝4Fe₂(SO₄)₃+6H₂O+6S；

ZnS+Fe₂(SO₄)₃＝ZnSO₄+2S；

ZnFe₂O₄+4H₂SO₄＝ZnSO₄+Fe₂(SO₄)₃+4H₂O；

ZnO+H₂SO₄＝ZnSO₄+H₂O。

compared with the low-acid (initial acid is less than 100g/L) pretreatment method prepared by sulfuric acid in the prior art, the pretreatment of the step 1 of the invention is easier, more copper, tin, indium, iron and zinc in the high-iron, high-tin and high-indium flotation silver concentrate enter the pretreatment solution, and then the copper, tin, indium, iron and zinc enter the leaching working section of a zinc system for recovery; secondly, more silver in the high-iron high-tin high-indium flotation silver concentrate is enriched in the pretreatment slag; thirdly, the lattice structure of the original mineral is damaged to a greater extent; fourthly, the pretreatment slag is damaged to a greater extent due to the fact that the lattice structure of the pretreatment slag is damaged, and the pretreatment slag is easier to react with thiourea, so that the reaction temperature is lower, oxidation of the thiourea can be reduced, consumption of the thiourea is greatly reduced, and impurities in silver sponge and copper sponge obtained by replacement of zinc powder are greatly reduced; fifthly, the volume of the zinc system is not increased; sixthly, after the pretreatment reaction, the solid-liquid separation effect is better and the cost is lower.

In step 2 of the invention, the pretreatment slag is leached by thiourea to obtain thiourea leaching solution and thiourea leaching slag.

The chemical reaction involved in step 2 is as follows:

Ag+3CN₂H₄+Fe³⁺＝[Ag(CN₂H₄)₃]⁺+Fe²⁺；

Ag₂S+2H⁺+6CN₂H₄＝[Ag(CN₂H₄)₃]⁺+H₂S；

Ag₂S+2Fe³⁺+6CN₂H₄＝[Ag(CN₂H₄)₃]⁺+2Fe²⁺+S。

in the step 3 of the invention, the thiourea leachate is replaced by zinc powder to precipitate silver and copper, and the obtained silver sponge and copper sponge are sold.

The invention has the beneficial effects that the silver is extracted from the flotation silver concentrate with high iron, high tin and high indium:

1. the method adopts zinc electrolysis waste liquid and hydrogen peroxide to pre-treat the high-iron high-tin high-indium flotation silver concentrate, so that more copper, tin, indium, iron and zinc in the high-iron high-tin high-indium flotation silver concentrate enter the pretreatment liquid more easily and then return to a leaching working section of a zinc system for recovery; secondly, more silver in the high-iron high-tin high-indium flotation silver concentrate is enriched in the pretreatment slag; thirdly, the lattice structure of the original mineral is damaged to a greater extent; fourthly, the pretreatment slag is damaged to a greater extent due to the fact that the lattice structure of the pretreatment slag is damaged, and the pretreatment slag is easier to react with thiourea, so that the reaction temperature is lower, oxidation of the thiourea can be reduced, consumption of the thiourea is greatly reduced, and impurities in silver sponge and copper sponge obtained by replacement of zinc powder are greatly reduced; fifthly, the volume of the zinc system is not increased; sixthly, the solid-liquid separation effect is better, and the cost is lower.

2. The method can efficiently extract the silver sponge and the copper sponge from the high-iron, high-tin and high-indium flotation silver concentrate, has obvious economic benefit, simple extraction method, easy operation, low cost and wide market prospect, and is suitable for large-scale popularization and application.

On the basis of the technical scheme, the invention can be further improved as follows.

Further, in step 1, the flotation silver concentrate with high iron, high tin and high indium contains the following components by mass percent: 28-31% of Zn, 2-3% of Cu, 0.29-0.34% of Ag, 19-22% of Fe, 0.2-0.3% of As, 0.04-0.06% of In, 0.5-0.7% of Pb, 24-26% of S, 0.15-0.25% of Sn, 23-25% of O, and the sum of Si and Ca is less than 2.5%.

The adoption of the further beneficial effects is as follows: the flotation silver concentrate has high contents of iron, tin and indium, and is a complex silver concentrate.

Further, in the step 1, the starting acid of the zinc electrolysis waste liquid is 150g/L-200g/L, the manganese ion concentration is 3.5g/L-6g/L, the fluorine ion concentration is 20mg/L-100mg/L, and the chlorine ion concentration is 100mg/L-500 mg/L.

The adoption of the further beneficial effects is as follows: the zinc electrolysis waste liquid comes from a zinc system. The waste zinc electrolysis solution is a special solution, contains oxidizing high-valence manganese besides high acid, and contains fluorine chloride ions and the like, and can play a comprehensive effect when used for pretreatment.

Further, in the step 1, the hydrogen peroxide is added in 3 times at intervals of 30min, wherein the mass fraction of the hydrogen peroxide is 20-30%.

The adoption of the further beneficial effects is as follows: a small amount of hydrogen peroxide with strong oxidizing property is adopted to be matched with zinc electrolysis waste liquid, so that the pretreatment effect is further enhanced, impurities are not additionally brought into other strong oxidizing agents, and the price is relatively low.

Further, in the step 1, the reaction temperature is 75-85 ℃ and the reaction time is 90-150 min.

Further, in step 2, concentrated sulfuric acid is used for adjusting the pH value.

Further, in the step 2, the temperature of the leaching reaction is 35-45 ℃, and the time is 3.5-5 h.

Detailed Description

The principles and features of this invention are described below in conjunction with specific embodiments, which are set forth merely to illustrate the invention and are not intended to limit the scope of the invention.

Example 1

The method for extracting silver from the flotation silver concentrate with high iron, high tin and high indium comprises the following steps:

step 1: pretreatment of

The method comprises the following steps of taking high-iron high-tin high-indium flotation silver concentrate, zinc electrolysis waste liquid and hydrogen peroxide, wherein the content of the components of the high-iron high-tin high-indium flotation silver concentrate is shown in table 1. The initial acid of the zinc electrolysis waste liquid is 150g/L, the manganese ion concentration is 4g/L, the fluorine ion concentration is 20mg/L, and the chlorine ion concentration is 100 mg/L. The liquid-solid ratio of the zinc electrolysis waste liquid to the high-iron high-tin high-indium flotation silver concentrate is 6L:1kg, the volume ratio of the zinc electrolysis waste liquid to the hydrogen peroxide is 40:1, the mass fraction of the hydrogen peroxide is 25%, the zinc electrolysis waste liquid and the hydrogen peroxide are added in 3 times at 30min intervals, the three react for 120min at 80 ℃, and a pretreatment liquid and pretreatment slag are obtained, wherein the pretreatment liquid returns to a leaching working section of a zinc system.

Step 2: thiourea leaching

And (3) adding the recycling agent obtained in the step (3) into the pretreatment slag obtained in the step (1), wherein the liquid-solid ratio of the recycling agent to the pretreatment slag is 6L:1kg, adjusting the pH value to 2 by using concentrated sulfuric acid, carrying out leaching reaction at the temperature of 40 ℃ for 4h to obtain thiourea leaching liquid and thiourea leaching slag, and returning the thiourea leaching slag to the rotary kiln.

And step 3: replacement of

And (3) replacing the thiourea leaching solution obtained in the step (2) with zinc powder to obtain silver sponge, copper sponge and a replaced solution, adding thiourea into the replaced solution, wherein the solid-to-liquid ratio of the thiourea to the replaced solution is 29g:1L, namely the concentration of the thiourea in the replaced solution is 29g/L to obtain a circulating agent, and returning the circulating agent to the step (2) for selling the silver sponge and the copper sponge.

The components and leaching rates of the pretreated slag obtained in step 1 were measured, as shown in table 1.

TABLE 1

The leaching rate of the pretreatment is [1- (the content of elements in the silver concentrate-the content of elements in the pretreatment slag)/the content of elements in the silver concentrate ]. times.100%

And (3) detecting the pretreatment liquid and the pretreatment slag obtained in the step (1), wherein the leaching rate of silver in the pretreatment liquid is 0.645%, and can be almost ignored.

And (3) detecting the thiourea leaching solution and thiourea leaching slag obtained in the step (2), as shown in table 2.

TABLE 2

The leaching rate of thiourea is 1- (the content of elements in the pretreatment slag-the content of elements in the thiourea slag)/the content of elements in the pretreatment slag ] x 100 percent

And (3) detecting the thiourea leaching solution and the thiourea leaching residues obtained in the step (2), wherein the leaching rate of silver in the thiourea solution is 91.5%.

Example 2

The method for extracting silver from the flotation silver concentrate with high iron, high tin and high indium comprises the following steps:

step 1: pretreatment of

And (3) taking the high-iron high-tin high-indium flotation silver concentrate, zinc electrolysis waste liquid and hydrogen peroxide, wherein the component contents of the high-iron high-tin high-indium flotation silver concentrate are shown in table 3. The initial acid of the zinc electrolysis waste liquid is 175g/L, the manganese ion concentration is 5g/L, the fluorine ion concentration is 60mg/L, and the chlorine ion concentration is 300 mg/L. The liquid-solid ratio of the zinc electrolysis waste liquid to the high-iron high-tin high-indium flotation silver concentrate is 6L:1kg, the volume ratio of the zinc electrolysis waste liquid to the hydrogen peroxide is 40:1, the mass fraction of the hydrogen peroxide is 20%, the zinc electrolysis waste liquid and the hydrogen peroxide are added in 3 times at 30min intervals, the zinc electrolysis waste liquid, the hydrogen peroxide and the hydrogen peroxide react for 150min at 75 ℃, a pretreatment liquid and pretreatment slag are obtained, and the pretreatment liquid returns to a leaching working section of a zinc system.

Step 2: thiourea leaching

And (3) adding the recycling agent obtained in the step (3) into the pretreatment slag obtained in the step (1), wherein the liquid-solid ratio of the recycling agent to the pretreatment slag is 5L:1kg, adjusting the pH value to 1.0 by using concentrated sulfuric acid, carrying out leaching reaction for 5h at the temperature of 35 ℃ to obtain thiourea leaching liquid and thiourea leaching slag, and returning the thiourea leaching slag to the rotary kiln.

And step 3: replacement of

And (3) replacing the thiourea leaching solution obtained in the step (2) with zinc powder to obtain silver sponge, copper sponge and a replaced solution, adding thiourea into the replaced solution, wherein the solid-to-liquid ratio of the thiourea to the replaced solution is 31g:1L, namely the concentration of the thiourea in the replaced solution is 31g/L to obtain a circulating agent, and returning the circulating agent to the step (2) for selling the silver sponge and the copper sponge.

The components and leaching rates of the pretreated slag obtained in step 1 were measured, and are shown in Table 3.

TABLE 3

The leaching rate of the pretreatment is [1- (the content of elements in the silver concentrate-the content of elements in the pretreatment slag)/the content of elements in the silver concentrate ]. times.100%

And (3) detecting the pretreatment liquid and the pretreatment slag obtained in the step (1), wherein the leaching rate of silver in the pretreatment liquid is 0.564 percent and can be almost ignored.

And (3) detecting the thiourea leaching solution and thiourea leaching slag obtained in the step (2), as shown in table 4.

TABLE 4

The leaching rate of thiourea is 1- (the content of elements in the pretreatment slag-the content of elements in the thiourea slag)/the content of elements in the pretreatment slag ] x 100 percent

And (3) detecting the thiourea leaching solution and the thiourea leaching residues obtained in the step (2), wherein the leaching rate of silver in the thiourea solution is 91.86%.

Example 3

The method for extracting silver from the flotation silver concentrate with high iron, high tin and high indium comprises the following steps:

step 1: pretreatment of

And (3) taking the high-iron high-tin high-indium flotation silver concentrate, zinc electrolysis waste liquid and hydrogen peroxide, wherein the component contents of the high-iron high-tin high-indium flotation silver concentrate are shown in table 5. The initial acid of the zinc electrolysis waste liquid is 200g/L, the manganese ion concentration is 6g/L, the fluorine ion concentration is 100mg/L, and the chlorine ion concentration is 500 mg/L. The liquid-solid ratio of the zinc electrolysis waste liquid to the high-iron high-tin high-indium flotation silver concentrate is 6L:1kg, the volume ratio of the zinc electrolysis waste liquid to the hydrogen peroxide is 40:1, the mass fraction of the hydrogen peroxide is 30%, the zinc electrolysis waste liquid and the hydrogen peroxide are added in 3 times at 30min intervals, the zinc electrolysis waste liquid, the hydrogen peroxide and the hydrogen peroxide react for 90min at 85 ℃, a pretreatment liquid and pretreatment slag are obtained, and the pretreatment liquid returns to a leaching working section of a zinc system.

Step 2: thiourea leaching

And (3) adding the recycling agent obtained in the step (3) into the pretreatment slag obtained in the step (1), wherein the liquid-solid ratio of the recycling agent to the pretreatment slag is 7L:1kg, adjusting the pH value to 2.5 by using concentrated sulfuric acid, carrying out leaching reaction at the temperature of 45 ℃ for 3.5h to obtain thiourea leaching liquid and thiourea leaching slag, and returning the thiourea leaching slag to the rotary kiln.

And step 3: replacement of

And (3) replacing the thiourea leaching solution obtained in the step (2) with zinc powder to obtain silver sponge, copper sponge and a replaced solution, adding thiourea into the replaced solution, wherein the solid-to-liquid ratio of the thiourea to the replaced solution is 33g:1L, namely the concentration of the thiourea in the replaced solution is 33g/L to obtain a circulating agent, and returning the circulating agent to the step (2) for selling the silver sponge and the copper sponge.

The components and leaching rates of the pretreated slag obtained in step 1 were measured, and are shown in Table 5.

TABLE 5

The leaching rate of the pretreatment is [1- (the content of elements in the silver concentrate-the content of elements in the pretreatment slag)/the content of elements in the silver concentrate ]. times.100%

And (3) detecting the pretreatment liquid and the pretreatment slag obtained in the step (1), wherein the leaching rate of silver in the pretreatment liquid is 0.927%, and can be almost ignored.

And (3) detecting the thiourea leaching solution and the thiourea leaching slag obtained in the step (2), as shown in table 6.

TABLE 6

The leaching rate of thiourea is 1- (the content of elements in the pretreatment slag-the content of elements in the thiourea slag)/the content of elements in the pretreatment slag ] x 100 percent

And (3) detecting the thiourea leaching solution and the thiourea leaching residues obtained in the step (2), wherein the leaching rate of silver in the thiourea solution is 92.16%.

Comparative example 1

Unlike example 1, comparative example 1 had no step 1 pretreatment and step 3 displacement, i.e. only the thiourea leach of step 2. The method comprises the following steps:

and (3) leaching the high-iron high-tin high-indium flotation silver concentrate by using thiourea, wherein the concentration of the thiourea is 29g/L, the pH value is adjusted to be 1.5 by using concentrated sulfuric acid, leaching reaction is carried out for 3.5h at the temperature of 35 ℃, so as to obtain thiourea leaching liquid and thiourea leaching slag, and the thiourea leaching slag is returned to the rotary kiln.

The leaching rate of silver is 8-15% which is seriously lower than that of silver in the example 1 by detecting the thiourea leaching solution and the thiourea leaching residue, so the method of the comparative example 1 is not advisable.

Comparative example 2

Unlike example 1, in step 1 of comparative example 2, the zinc electrolysis waste liquid was not used, and concentrated sulfuric acid having a concentration of 100g/L was used, and the rest was the same. When the components and the leaching rate of the pretreated slag obtained in step 1 were measured, as shown in Table 7, the time was increased by about 45min under the same filtration conditions, and the solid-liquid separation effect was poor.

TABLE 7

The leaching rate of the pretreatment is [1- (the content of elements in the silver concentrate-the content of elements in the pretreatment slag)/the content of elements in the silver concentrate ]. times.100%

And (3) detecting the pretreatment liquid and the pretreatment slag obtained in the step (1), wherein the leaching rate of silver in the pretreatment liquid is 0.39%, and can be almost ignored.

And (3) detecting the thiourea leaching solution and the thiourea leaching slag obtained in the step (2), as shown in table 8.

TABLE 8

The leaching rate of thiourea is 1- (the content of elements in the pretreatment slag-the content of elements in the thiourea slag)/the content of elements in the pretreatment slag ] x 100 percent

When the thiourea leaching solution and the thiourea leaching residue obtained in the step 2 were tested, the leaching rate of silver in the thiourea solution was 52.62%, which was also significantly lower than that of silver in example 1, so the method of comparative example 2 was not preferable.

Comparative example 3

Unlike example 1, in step 1 of comparative example 3, the zinc electrolysis waste liquid was not used, and concentrated sulfuric acid having a concentration of 150g/L was used, and the rest was the same. The components and leaching rates of the pretreated slag obtained in step 1 were measured and are shown in Table 9.

TABLE 9

The leaching rate of the pretreatment is [1- (the content of elements in the silver concentrate-the content of elements in the pretreatment slag)/the content of elements in the silver concentrate ]. times.100%

And (3) detecting the pretreatment liquid and the pretreatment slag obtained in the step (1), wherein the leaching rate of silver in the pretreatment liquid is 0.564 percent and can be almost ignored.

The thiourea leaching solution and thiourea leaching residue obtained in step 2 were tested, as shown in table 10.

Watch 10

The leaching rate of thiourea is 1- (the content of elements in the pretreatment slag-the content of elements in the thiourea slag)/the content of elements in the pretreatment slag ] x 100 percent

And (3) detecting the thiourea leaching solution obtained in the step (2) and the thiourea leaching residue, wherein the leaching rate of silver in the thiourea liquid is 84.35%, which is slightly lower than that of silver in the example 1. .

Compared with the comparative examples 1 to 3, the invention adopts the zinc electrolysis waste liquid and the hydrogen peroxide to pre-treat the high-iron high-tin high-indium flotation silver concentrate, so that more copper, tin, indium, iron and zinc in the high-iron high-tin high-indium flotation silver concentrate enter the pre-treatment liquid more easily and then return to the leaching working section of the zinc system for recovery; secondly, more silver in the high-iron high-tin high-indium flotation silver concentrate is enriched in the pretreatment slag; thirdly, the lattice structure of the original mineral is damaged to a greater extent; fourthly, the pretreatment slag is damaged to a greater extent due to the fact that the lattice structure of the pretreatment slag is damaged, and the pretreatment slag is easier to react with thiourea, so that the reaction temperature is lower, oxidation of the thiourea can be reduced, consumption of the thiourea is greatly reduced, and impurities in silver sponge and copper sponge obtained by replacement of zinc powder are greatly reduced; fifthly, the volume of the zinc system is not increased; sixthly, the solid-liquid separation effect is better, and the cost is lower.

The above description is only for the purpose of illustrating the preferred embodiments of the present invention and is not to be construed as limiting the invention, and any modifications, equivalents, improvements and the like that fall within the spirit and principle of the present invention are intended to be included therein.