CN111606308A - Method for efficiently separating and recycling tellurium from copper anode slime copper separation slag - Google Patents
Method for efficiently separating and recycling tellurium from copper anode slime copper separation slag Download PDFInfo
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Abstract
The invention belongs to the field of wet metallurgy in nonferrous metallurgy, and particularly relates to a method for efficiently separating and recycling tellurium from copper anode slime copper-separating slag, which can effectively realize efficient separation of tellurium from copper anode slime copper-separating slag. According to the method, copper anode mud copper separation slag is subjected to a hydrochloric acid oxidation system to realize a tellurium efficient leaching process, the tellurium leaching rate is more than 90%, the gold leaching rate is more than 99%, gold in a solution is preferentially reduced and precipitated in a mode of uniformly and slowly adding a weak reducing agent, the gold precipitation rate is more than 99%, tellurium is not precipitated basically, platinum and palladium in the solution are reduced and precipitated in a mode of uniformly and slowly adding the weak reducing agent, the concentrations of gold, platinum and palladium ions in a solution after platinum and palladium reduction can be reduced to be less than 0.001g/L, and the solution after platinum and palladium reduction is subjected to deep reduction and tellurium precipitation by adding the reducing agent, so that crude tellurium powder with the quality of more than 95% and the tellurium recovery rate of more than. The links are closely related, and the high-efficiency separation and recovery of gold and tellurium in the copper separation slag are realized under the combined action. The invention has the advantages of stable technical indexes, low labor intensity, low production cost and the like.
Description
Technical Field
The invention belongs to the field of wet metallurgy in nonferrous metallurgy, and particularly relates to a method for efficiently separating and recycling tellurium from copper anode slime copper-separating slag, which can effectively realize efficient separation of tellurium from copper anode slime copper-separating slag.
Background
Tellurium is an important rare metal, has wide and important application in the fields of semiconductor devices, refrigeration elements, photoelectric elements, special alloys and the like due to the unique physical and chemical properties, and has an outstanding position in the industries of national defense, aerospace and energy and chemical industry.
Tellurium mainly exists in the form of heavy metal associated ores and is produced as a byproduct in the heavy metal extraction process, the main associated ores are copper ores, and the common copper ore smelting process is to sequentially carry out pyrometallurgical smelting, pyrometallurgical blowing, pyrometallurgical refining and electrolytic refining on copper concentrate to produce electrolytic copper products. And tellurium is continuously enriched in the process and finally enters into the byproduct anode mud in the electrolytic refining stage. Therefore, the extraction of tellurium mainly takes copper anode slime as a raw material, about 80% of tellurium is extracted from the copper anode slime, and the copper anode slime generally contains 0.5% -15.0% of tellurium, 1.0% -10.0% of tellurium, 1.0% -8.0% of Sb1.01% -1.5% of tellurium and 5.0% -20.0% of Ags.
The copper anode slime contains a large amount of noble metals, and is a main noble metal extraction raw material. The treatment process of the copper anode slime generally comprises the steps of firstly carrying out pretreatment to remove base metals, and then adopting a pyrogenic process or a wet process to recover precious metals such as gold and silver. Wherein tellurium is mainly separated and recovered in a pretreatment stage.
The pretreatment process of the copper anode slime mainly removes elements such as Cu, Se, Te, Sb and the like, is mature at present, and generally comprises sulfating roasting selenium steaming to volatilize Se to enter flue gas, converting the copper anode slime into sulfate, and then performing water leaching or acid leaching copper removal to separate copper from a solution.
The slag after copper anode slime is used for separating copper is called copper separating slag, and is mainly a precious metal enrichment after the copper anode slime is used for separating Se and Cu. And carrying out alkaline leaching on the copper separation slag to remove tellurium, so that the tellurium enters the solution in the form of tellurite in the alkaline leaching process to be separated from the noble metal. And (3) removing impurities and purifying tellurium entering the solution, dissolving in alkali, and preparing refined tellurium through electrodeposition.
The tellurium separating slag is enriched with noble metals, and then the processes of gold separation by chlorination, sodium sulfite and silver separation are carried out to extract noble metals such as gold, silver and the like.
Tellurium is an amphoteric compound and can be dissolved in acid and alkali theoretically, but in production practice, the leaching effect of tellurium in sulfuric acid and alkali liquor is not ideal, the leaching rate of tellurium in the process of separating tellurium by pretreating copper anode slime is low, and the separation and recovery are difficult. The existing reports show that the comprehensive recovery rate of tellurium in copper anode slime of the current copper smelting enterprises is less than 60 percent, and even less than 40 percent in some enterprises.
The new research of recovering tellurium from anode mud mainly focuses on the pretreatment process of copper anode mud, and mainly comprises processes of mixed acid leaching for removing tellurium, pressurized high-alkaline removing tellurium, oxidation leaching, circulating leaching for removing tellurium, extraction and the like. Although the processes can improve the leaching rate of tellurium in the base metal separation process of the copper anode slime to a certain extent, the processes have the problems of long process flow, complex operation, high equipment requirement and high production cost, and have difficulty in industrial popularization.
Disclosure of Invention
In order to overcome the defect of low recovery rate of directly extracting tellurium from copper anode slime copper separating slag, the invention provides a hydrometallurgical method for realizing high recovery rate of tellurium from copper anode slime copper separating slag, short process flow and simple operation.
In order to achieve the aim, the invention adopts the technical scheme that: adding an oxidant into hydrochloric acid solution with a certain concentration to leach copper-separating residues of copper anode mud, so that precious metals such as gold, platinum, palladium and the like and tellurium enter the solution, uniformly and slowly adding a reducing agent into the leachate, preferentially precipitating the gold and platinum palladium, further reducing the gold and platinum palladium, reducing the tellurium into tellurium powder, precipitating the tellurium powder, realizing efficient separation of the tellurium and the gold, and enabling the leached residues to enter a silver extraction process.
The essence of the invention is that a hydrochloric acid oxidation system for chlorination leaching is adopted to realize the high-efficiency leaching process of tellurium, gold and platinum palladium are sequentially reduced and precipitated in a step-by-step reduction mode of uniformly and slowly adding a weak reducing agent, and tellurium is subsequently and deeply reduced and precipitated, so that the high-efficiency separation process of gold and tellurium is realized. The links are closely related, and the aims of efficiently separating and recycling tellurium from copper anode slime by copper slag are achieved under the combined action.
The specific technological process and technological parameters are as follows:
1 copper anode mud copper-separating slag chlorination leaching
Adding water and hydrochloric acid into a reaction kettle, adjusting the concentration of the hydrochloric acid solution to be 2-5mol/L, and mixing copper anode mud copper-separating slag according to a liquid-solid ratio (the ratio of the liquid volume L to the dry weight Kg of solid) of 1-5: 1, heating to 65-85 ℃, keeping the stirring speed at 200-. And (3) after the reaction is finished, rapidly cooling to room temperature, stopping stirring, and performing filter pressing by using a plate-and-frame filter press to obtain the gold separation slag and the gold separation liquid.
The main reactions in the chlorination leaching process of copper anode slime copper-separating slag are as follows:
NaClO3+6HCl＝3Cl↑+NaCl+3H2O (1)
Cu2Te+2NaClO3+6HCl＝2CuCl2+H6TeO6+Cl2↑+2NaCl (2)
5Te+6NaClO3+6HCl+12H2O＝5H6TeO6+3Cl2↑+6NaCl (3)
5TeO2+NaClO3+2HCl+14H2O＝5H6TeO6+Cl2↑+2NaCl (4)
5Ag2Te+8NaClO3+18HCl+6H2O＝10AgCl+5H6TeO6+4Cl2↑+8NaCl (5)
2Au+3Cl2+2HCl＝2HAuCl4 (6)
Pd+Cl2+2HCl＝H2PdCl4 (7)
Pt+Cl2+2HCl＝H2PtCl4 (8)
(2) reduction of gold
Transferring the gold separating solution to a new reaction kettle, adjusting the temperature to be between 10 and 50 ℃, controlling the stirring speed to be 100-300r/min, uniformly and slowly adding a reducing agent, wherein the optional reducing agent comprises sodium sulfite, sodium bisulfite, sulfur dioxide and oxalic acid, the addition of the reducing agent is preferably in a sodium sulfite solution atomization spraying mode, and the addition amount is 1.4 to 1.6 times of the amount of the gold substances in the solution. And (4) stopping stirring after the reaction is stabilized for 0.1-2h, and then filtering to obtain gold reduction filter residue and gold reduction solution. And washing the gold reduction filter residue with hydrochloric acid with the concentration of 2-6mol/L and hot water in sequence to obtain crude gold powder.
The main reactions taking place during the gold reduction are:
2HAuCl4+3Na2SO3+3H2O＝2Au+3Na2SO4+8HCl (9)
(3) reduction of platinum and palladium
Transferring the solution after gold reduction to a new reaction kettle, adjusting the temperature to 65-95 ℃, controlling the stirring speed to be 100-300r/min, uniformly and slowly adding a reducing agent, wherein the optional reducing agent comprises sodium sulfite, sodium bisulfite, sulfur dioxide and oxalic acid, the addition of the reducing agent is preferably in a sodium sulfite solution atomization spraying mode, and the addition amount is 1.0-1.5 times of the amount of the platinum-palladium total substances in the solution. And (4) stopping stirring after the stable reaction is carried out for 0.1-2h, and then filtering to obtain platinum-palladium reduction filter residue and platinum-palladium reduction solution. And washing the platinum-palladium reduction filter residue to obtain platinum-palladium concentrate.
The main reactions taking place during the reduction of platinum and palladium are:
H2PdCl4+Na2SO3+H2O＝Pd+Na2SO4+4HCl (10)
H2PtCl4+Na2SO3+H2O＝Pt+Na2SO4+4HCl (11)
3 tellurium reduction
The solution after platinum and palladium reduction is put into a reaction kettle, the temperature is adjusted to be 60-95 ℃, hydrochloric acid is added to lead the solution H to be+The concentration is 2.60-4.0mol/L, the stirring speed is controlled to be 100-300r/min, the reducing agent is uniformly and slowly added, the optional reducing agent comprises sodium sulfite, sulfur dioxide, oxalic acid, sodium bisulfite and the like, preferably sulfur dioxide, the adding amount is 1.0-2.0 times of the total tellurium content in the copper anode slime copper-separating slag, and the mixture is naturally cooled and continuously stirred for reaction for 1 hour after the ventilation is finished. Filtering after the reaction is finished to obtain coarse tellurium powder and a solution after tellurium separation.
The main reactions in the tellurium reduction process are as follows:
H6TeO6+3SO2＝Te+3H2SO4(12)
the hydrochloric acid, the sodium sulfite, the sodium bisulfite, the sulfur dioxide and the oxalic acid are all industrial reagents, wherein the mass percentage concentration of the sodium sulfite, the sodium bisulfite and the oxalic acid is not lower than 98.5 percent, the mass percentage concentration of the hydrochloric acid is not lower than 36.0 percent, and the volume percentage concentration of the sulfur dioxide gas is not lower than 98.5 percent.
The method is suitable for treating the copper anode slime copper separating slag, and the main component ranges are as follows by weight percent (%): te1.0-20.0, Sb1.0-10.0, Pb1.0-30.0, As1.0-10.0, Ag1.0-20.0 and Au0.01-2.0.
Compared with the traditional tellurium recovery process, the method has the following advantages:
1. the process of separating tellurium from the copper anode slime through pretreatment is omitted, and the process of extracting the noble metal from the copper anode slime is shortened. The process realizes high grade of crude gold powder, less tellurium content and comprehensive recovery rate of tellurium of more than 90 percent on the basis of flow simplification.
2. The hydrochloric acid system is adopted for chlorination leaching, the noble metal and the tellurium are leached together, the operation condition is simple, the realization is easy, and the gold leaching rate can reach 99.00 percent and the tellurium leaching rate can reach more than 90.00 percent.
3. The gold separating liquid is reduced by adopting weak reducing agents such as sodium sulfite solution with lower concentration and the like, and the uniform and slow adding mode solves the problem of gold and tellurium coprecipitation caused by overhigh concentration of local reducing agents in the traditional process, the gold and tellurium separation effect is good, and the gold recovery rate is more than 99.5 percent. The gold content of the crude gold powder is more than 98.0 percent, tellurium does not enter the crude gold powder basically, and the concentration of liquid gold ions can be reduced to be less than 0.02g/L after gold reduction. Then adopting weak reducing agents such as sodium sulfite solution with lower concentration and the like, reducing platinum and palladium in a uniform and slow adding mode, and reducing the concentration of noble metal ions such as gold, platinum, palladium and the like in the solution to be below 0.001 g/L. And then, continuously adding a reducing agent to reduce the tellurium, wherein the recovery rate of the tellurium is more than 90 percent, the quality of the coarse tellurium powder is more than 95 percent, and the concentration of the liquid tellurium after the tellurium reduction is reduced to be less than 0.05 g/L.
4. The invention adopts a step-by-step reduction mode of chlorination leaching and uniform and slow addition of weak reducing agent to realize the advantages of adjustable and controllable process, good gold and tellurium separation effect, high tellurium recovery rate, short process flow and less reagent consumption.
5. The invention has the advantages of stable technical indexes of the process, low labor intensity, low production cost and the like.
Drawings
FIG. 1 is a flow chart of the method for efficiently separating and recovering tellurium from copper anode slime copper separation slag.
Detailed Description
The technical solution of the present invention is further explained with reference to the drawings and the embodiments.
As shown in fig. 1, the method for efficiently separating and recovering tellurium from copper anode slime copper separation slag of the invention specifically comprises the following steps:
s1) copper anode slime copper separating slag chlorination leaching: adding copper separation slag of copper anode mud into a hydrochloric acid solution, and adding an oxidant for chlorination leaching to obtain gold separation slag and a gold separation solution;
s2) gold reduction: adjusting the temperature of the gold separation liquid obtained in the step S1), slowly adding a reducing agent, filtering to obtain gold reduction filter residue and gold reduction liquid, and washing the gold reduction filter residue to obtain crude gold powder;
s3) platinum palladium reduction: uniformly and slowly adding a reducing agent into the solution obtained after the gold reduction in the step S2), and filtering to obtain platinum-palladium reduction filter residue and platinum-palladium reduction solution; washing the platinum-palladium reduction filter residue to obtain platinum-palladium concentrate;
s4) tellurium reduction: adding H in hydrochloric acid regulating solution into the solution obtained after reduction of the platinum and the palladium obtained in S3)+Slowly adding a reducing agent into the solution with the concentration, and filtering the solution to obtain coarse tellurium powder and a solution after tellurium separation; the quality of the obtained coarse tellurium powder is more than 95 percent, and the tellurium recovery rate is more than 90 percent.
The S1) comprises the following specific steps:
s1.1) separating copper slag from copper anode mud according to a liquid-solid ratio of 1-5: 1, adding the mixture into a hydrochloric acid solution with the concentration of 2-5 mol/L;
s1.2) heating to 65-85 ℃, continuously stirring at the rotation speed of 200-;
s1.3) cooling to room temperature quickly after the reaction is finished, stopping stirring, and performing filter pressing by using a plate-and-frame filter press to obtain gold separation slag and a gold separation liquid.
The S2) comprises the following specific steps:
s2.1) adjusting the temperature of the gold separating liquid obtained in S1.3) to be between 10 and 50 ℃, stirring at the rotating speed of 100-300r/min,
s2.2) uniformly and slowly adding a reducing agent, wherein the adding amount is 1.4-1.6 times of the amount of the gold substances in the solution. And (4) stopping stirring after the reaction is stabilized for 0.1-2h, and then filtering to obtain gold reduction filter residue and gold reduction solution.
The S3) comprises the following specific steps:
s3.1) adjusting the temperature of the gold reduced by the S2.2) to be between 65 and 95 ℃, and continuously stirring at the rotation speed of 100-300 r/min;
s3.2) uniformly and slowly adding a reducing agent, wherein the adding amount is 1.0-1.5 times of the amount of the platinum and palladium total substances in the solution after gold reduction. The stirring is stopped after the stable reaction is carried out for 0.1 to 2 hours;
and S3.3) filtering to obtain platinum-palladium reduction filter residues and platinum-palladium reduction solution, and washing the platinum-palladium reduction filter residues to obtain platinum-palladium concentrate.
The S4) comprises the following specific steps:
s4.1) adjusting the temperature of the solution obtained in S3.3) after reduction of the platinum and palladium to 60-95 ℃, and adding hydrochloric acid to ensure that the solution H+The concentration is 2.60-4.0 mol/L;
s4.2) continuously stirring at the rotating speed of 100-300r/min, and simultaneously uniformly and slowly adding a reducing agent, wherein the ratio of the reducing agent to the total tellurium substances in the copper-separating slag of the copper anode slime is 1.0-2.0: 1, adding a reducing agent, naturally cooling after the adding of the reducing agent is finished, continuously stirring and reacting for 0.8-1.2h, and filtering after the reaction is finished to obtain crude tellurium powder and a solution after tellurium separation.
The oxidant in S1.2) is one or a combination of more of sodium chlorate, oxygen, chlorine and hydrogen peroxide; the addition amount is as follows: the amount of the gold and tellurium in the raw material is 1.0 to 1.3 times of the theoretical amount of the oxidant consumed by oxidizing all the tellurium.
The reducing agent is one or a combination of sodium sulfite, sodium bisulfite, sulfur dioxide, ammonia water and oxalic acid.
The addition mode of the reducing agent is as follows: the solid reducing agent is prepared into solution and then atomized and sprayed, and the liquid reducing agent is diluted and atomized and sprayed, wherein the concentration of the reducing agent in the prepared solution is not higher than 1.0mol/L, the speed of spraying the reducing agent solution in each reaction solution in a physical and chemical mode is not higher than 20L/h, and the adding speed of the gas reducing agent in each reaction solution is not higher than 30L/min.
The sodium sulfite, the sodium bisulfite, the sulfur dioxide and the oxalic acid are all industrial reagents, wherein the mass percentage concentration of the sodium sulfite, the sodium bisulfite and the oxalic acid is not lower than 98.5 percent, and the volume percentage concentration of the sulfur dioxide gas is not lower than 98.5 percent.
The copper anode slime copper separating slag comprises the following components in percentage by weight: 1.0-20.0% of Te1.0-10.0% of Sb1.0-10.0%, 1.0-30.0% of Pb0%, 1.0-10.0% of As1.0%, 1.0-20.0% of Ag0.01-2.0% of Au0.
Example 1:
the copper anode mud comprises the following main components in percentage by weight: te4.56%, Pb12.3%, Se1.2%, Au0.90%, Ag11.69%; hydrochloric acid, sodium chlorate, sodium sulfite and sulfur dioxide are all industrial reagents, the mass percentage concentration of the hydrochloric acid, the sodium chlorate and the sodium sulfite is 31.0 percent, 99.9 percent and 99.9 percent, and the volume fraction of the sulfur dioxide is 99.9 percent.
Adding 4m into a reaction kettle3Adding hydrochloric acid solution of 3mol/L, adding 1220.0kg of dry weight of copper anode mud copper separation slag, starting stirring for 300r/min, uniformly heating to 80 ℃, slowly and uniformly adding 80kg of sodium chlorate, reacting for 2.5h, heating to 95 ℃ after the reaction is finished, keeping for 0.5h, rapidly cooling to room temperature by using circulating water after the reaction is finished, stopping stirring, carrying out pressure filtration on the solution, washing filter residue by using hot water of 85 ℃ to obtain gold separation slag, drying, weighing 900.0kg, uniformly sampling the gold separation slag, and detecting that the gold leaching rate is 99.66% and the Te leaching rate is 94.34%.
Transferring all the gold separation solution to a new reaction kettle, controlling the stirring speed to be 300r/min at the temperature of 25 ℃, uniformly adding 107L of 100g/L sodium sulfite solution by adopting an atomization spraying mode, wherein the adding process is 1.5h, then stably reacting for 0.5h, stopping stirring, filtering, washing filter residues by using 6mol/L hydrochloric acid and 90 ℃ hot water in sequence, then drying and weighing 11.03kg, containing 99.03% of gold and having a gold precipitation rate of 99.82%, and detecting the solution after gold reduction to show that the filtrate contains Au0.005g/L, Pt0.008g/L and Pd0.047g/L.
And (3) loading the gold reduced solution into a new reaction kettle, controlling the stirring speed to be 300r/min, heating to 75 ℃, uniformly adding 2.5L of 100g/L sodium sulfite solution in a spraying mode, reacting for 0.5h, stopping stirring, filtering, and washing filter residues to obtain platinum-palladium concentrate. The filtrate is the solution after reduction of platinum and palladium, and analysis and detection show that the solution contains Au of less than 0.001g/L, Pt of less than 0.001g/L, Pd of less than 0.001 g/L.
Loading the solution after platinum and palladium reduction into a reaction kettle, heating to 75 ℃, adjusting the stirring speed to 300r/min, and supplementing hydrochloric acid to adjust the solution H+The concentration is 2.70mol/L, and then SO is introduced at the speed of 150L/min2And (3) stopping heating after the gas is introduced for 3h, naturally cooling to room temperature, continuing to react for 1h, stopping stirring, filtering, washing filter residues with water to obtain 53.33kg of reduced tellurium powder, wherein the tellurium content is 95.17 percent, the tellurium recovery rate is 91.23 percent, and the tellurium content of the filtrate is 0.040 g/L.
Example 2:
the copper anode mud comprises the following main components in percentage by weight: te4.56%, Pb12.3%, Se1.2%, Au0.90%, Ag11.69%; hydrochloric acid, sodium chlorate, sodium sulfite and sulfur dioxide are all industrial reagents, the mass percentage concentration of the hydrochloric acid, the sodium chlorate and the sodium bisulfite is 31.0 percent, 99.9 percent and the volume fraction of the sulfur dioxide is 99.9 percent.
Adding 4m into a reaction kettle3Adding hydrochloric acid solution of 3mol/L, adding 1220.0kg of copper anode slime by dry weight, starting stirring for 300r/min, uniformly heating to 80 ℃, slowly and uniformly adding 80kg of sodium chlorate, reacting for 2.5h, heating to 95 ℃ after the reaction is finished, keeping for 0.5h, quickly cooling to room temperature by using circulating water after the reaction is finished, stopping stirring, carrying out pressure filtration on the solution, washing filter residue by using hot water of 85 ℃ to obtain gold separating residue, drying, weighing 901.2kg, uniformly sampling the gold separating residue, and detecting that the gold leaching rate is 99.56% and the Te leaching rate is 93.85%.
Transferring all gold separation liquid to a new reaction kettle, controlling the stirring speed at 300r/min at the temperature of 25 ℃, uniformly adding 87L of 100g/L sodium bisulfite solution by adopting an atomization spraying mode, wherein the adding process takes 1.5h, then stably reacting for 0.5h, stopping stirring, filtering, washing filter residues by using 6mol/L hydrochloric acid and 90 ℃ hot water in sequence, then drying and weighing 11.02kg, 99.10 percent of gold, 99.90 percent of gold precipitation rate, and detecting the liquid after gold reduction to show that the filtrate contains Au0.003g/L, Pt0.008g/L and Pd0.049g/L.
And (3) putting the gold reduced solution into a new reaction kettle, controlling the stirring speed to be 300r/min, heating to 75 ℃, uniformly adding 2.7L of 100g/L sodium bisulfite solution in a spraying mode, reacting for 0.5h, stopping stirring, filtering, and washing filter residues with water to obtain platinum-palladium concentrate. The filtrate is the solution after reduction of platinum and palladium, and analysis and detection show that the solution contains Au of less than 0.001g/L, Pt of less than 0.001g/L, Pd of less than 0.001 g/L.
Loading the solution after platinum and palladium reduction into a reaction kettle, heating to 75 ℃, adjusting the stirring speed to 300r/min, and supplementing hydrochloric acid to adjust the solution H+The concentration is 2.70mol/L, and then SO is introduced at the speed of 150L/min2And (3) stopping heating after the gas is introduced for 3h, naturally cooling to room temperature, continuing to react for 1h, stopping stirring, filtering, washing filter residues with water to obtain 52.82kg of reduced tellurium powder, wherein the tellurium content is 97.04%, the tellurium recovery rate is 92.14%, and the tellurium content of the filtrate is 0.040 g/L.
Example 3:
the copper anode mud comprises the following main components in percentage by weight: te4.56%, Pb12.3%, Se1.2%, Au0.90%, Ag11.69%; hydrochloric acid, sodium chlorate, sodium sulfite and sulfur dioxide are all industrial reagents, the mass percentage concentration of the hydrochloric acid and the sodium chlorate is 31.0 percent, 99.9 percent and 99.9 percent, and the volume fraction of the sulfur dioxide is 99.9 percent.
Adding 4m into a reaction kettle3Adding hydrochloric acid solution of 3mol/L, adding 1220.0kg of dry weight of copper anode mud copper separation slag, starting stirring for 300r/min, uniformly heating to 80 ℃, slowly and uniformly adding 80kg of sodium chlorate, reacting for 2.5h, heating to 95 ℃ after the reaction is finished, keeping for 0.5h, rapidly cooling to room temperature by using circulating water after the reaction is finished, stopping stirring, carrying out pressure filtration on the solution, washing filter residue by using hot water of 85 ℃ to obtain gold separation slag, drying, weighing 899.8kg, uniformly sampling the gold separation slag, and detecting that the gold leaching rate is 99.72% and the Te leaching rate is 96.54%.
Transferring the gold separating liquid to a new reaction kettle at the temperatureAt 25 ℃, the stirring speed is controlled to be 300r/min, and SO is introduced at 30L/min2Gas is used for 1h, then stable reaction is carried out for 0.5h, stirring is stopped, filtering is carried out, filter residue is washed by 6mol/L hydrochloric acid and 90 ℃ hot water in sequence, then drying and weighing are carried out, 11.00kg of filter residue is obtained, gold is 99.15 percent, the gold precipitation rate is 99.61 percent, and liquid after gold reduction detection shows that the filter solution contains Au0.011g/L, Pt0.008g/L and Pd0.050g/L.
Loading the gold reduced solution into a new reaction kettle, controlling the stirring speed to 300r/min, heating to 75 ℃, and introducing SO at the speed of 6L/min2And (3) reacting for 0.5h for 10min, stopping stirring, filtering, and washing filter residues to obtain platinum-palladium concentrate. The filtrate is the solution after platinum and palladium reduction, and analysis and detection show that the solution contains Au<0.001g/L、Pt<0.001g/L、Pd<0.001g/L。
Loading the solution after platinum and palladium reduction into a reaction kettle, heating to 75 ℃, adjusting the stirring speed to 300r/min, and supplementing hydrochloric acid to adjust the solution H+The concentration is 2.70mol/L, and then SO is introduced at the speed of 150L/min2And (3) stopping heating after the gas is introduced for 3h, naturally cooling to room temperature, continuing to react for 1h, stopping stirring, filtering, washing filter residues with water to obtain 53.00kg of reduced tellurium powder, 96.84% of tellurium, the recovery rate of tellurium of 92.26%, and detecting the filtrate to detect the content of tellurium of 0.059 g/L.
The method for efficiently separating and recovering tellurium from copper anode slime copper separation slag provided by the embodiment of the application is described in detail above. The above description of the embodiments is only for the purpose of helping to understand the method of the present application and its core ideas; meanwhile, for a person skilled in the art, according to the idea of the present application, there may be variations in the specific embodiments and the application scope, and in summary, the content of the present specification should not be construed as a limitation to the present application.
As used in the specification and claims, certain terms are used to refer to particular components. As one skilled in the art will appreciate, manufacturers may refer to a component by different names. This specification and claims do not intend to distinguish between components that differ in name but not function. In the following description and in the claims, the terms "include" and "comprise" are used in an open-ended fashion, and thus should be interpreted to mean "include, but not limited to. "substantially" means within an acceptable error range, and a person skilled in the art can solve the technical problem within a certain error range to substantially achieve the technical effect. The description which follows is a preferred embodiment of the present application, but is made for the purpose of illustrating the general principles of the application and not for the purpose of limiting the scope of the application. The protection scope of the present application shall be subject to the definitions of the appended claims.
It is also noted that the terms "comprises," "comprising," or any other variation thereof, are intended to cover a non-exclusive inclusion, such that a good or system that comprises a list of elements does not include only those elements but may include other elements not expressly listed or inherent to such good or system. Without further limitation, an element defined by the phrase "comprising an … …" does not exclude the presence of other like elements in a commodity or system that includes the element.
It should be understood that the term "and/or" as used herein is merely one type of association that describes an associated object, meaning that three relationships may exist, e.g., a and/or B may mean: a exists alone, A and B exist simultaneously, and B exists alone. In addition, the character "/" herein generally indicates that the former and latter related objects are in an "or" relationship.
The foregoing description shows and describes several preferred embodiments of the present application, but as aforementioned, it is to be understood that the application is not limited to the forms disclosed herein, but is not to be construed as excluding other embodiments and is capable of use in various other combinations, modifications, and environments and is capable of changes within the scope of the application as described herein, commensurate with the above teachings, or the skill or knowledge of the relevant art. And that modifications and variations may be effected by those skilled in the art without departing from the spirit and scope of the application, which is to be protected by the claims appended hereto.
Claims (10)
1. The method for efficiently separating and recovering tellurium from copper anode slime copper separation slag is characterized by comprising the following steps:
s1) copper anode slime copper separating slag chlorination leaching: adding copper separation slag of copper anode mud into a hydrochloric acid solution, and adding an oxidant for chlorination leaching to obtain gold separation slag and a gold separation solution;
s2) gold reduction: adjusting the temperature of the gold separation liquid obtained in the step S1), slowly adding a reducing agent, filtering to obtain gold reduction filter residue and gold reduction liquid, and washing the gold reduction filter residue to obtain crude gold powder;
s3) platinum palladium reduction: uniformly and slowly adding a reducing agent into the solution obtained after the gold reduction in the step S2), and filtering to obtain platinum-palladium reduction filter residue and platinum-palladium reduction solution; washing the platinum-palladium reduction filter residue to obtain platinum-palladium concentrate;
s4) tellurium reduction: adding H in hydrochloric acid regulating solution into the solution obtained after reduction of the platinum and the palladium obtained in S3)+Slowly adding a reducing agent into the solution with the concentration, and filtering the solution to obtain coarse tellurium powder and a solution after tellurium separation; the quality of the obtained coarse tellurium powder is more than 95 percent, and the tellurium recovery rate is more than 90 percent.
2. The method as claimed in claim 1, wherein the specific steps of S1) are:
s1.1) separating copper slag from copper anode mud according to a liquid-solid ratio of 1-5: 1, adding the mixture into a hydrochloric acid solution with the concentration of 2-5 mol/L;
s1.2) heating to 65-85 ℃, continuously stirring at the rotation speed of 200-;
s1.3) cooling to room temperature quickly after the reaction is finished, stopping stirring, and performing filter pressing by using a plate-and-frame filter press to obtain gold separation slag and a gold separation liquid.
3. The method as claimed in claim 2, wherein the specific steps of S2) are:
s2.1) adjusting the temperature of the gold separating liquid obtained in S1.3) to be between 10 and 50 ℃, stirring at the rotating speed of 100-300r/min,
s2.2) uniformly and slowly adding a reducing agent, wherein the adding amount is 1.4-1.6 times of the amount of the gold substances in the solution; and (4) stopping stirring after the reaction is stabilized for 0.1-2h, and then filtering to obtain gold reduction filter residue and gold reduction solution.
4. The method as claimed in claim 3, wherein the specific steps of S3) are as follows:
s3.1) adjusting the temperature of the gold reduced by the S2.2) to be between 65 and 95 ℃, and continuously stirring at the rotation speed of 100-300 r/min;
s3.2) uniformly and slowly adding a reducing agent, wherein the adding amount is 1.0-1.5 times of the amount of the platinum-palladium total substances in the solution after gold reduction, and stopping stirring after stable reaction for 0.1-2 h;
and S3.3) filtering to obtain platinum-palladium reduction filter residues and platinum-palladium reduction solution, and washing the platinum-palladium reduction filter residues to obtain platinum-palladium concentrate.
5. The method as claimed in claim 4, wherein the specific steps of S4) are as follows:
s4.1) adjusting the temperature of the solution obtained in S3.3) after reduction of the platinum and palladium to 60-95 ℃, and adding hydrochloric acid to ensure that the solution H+The concentration is between 2.60 and 4.0 mol/L;
s4.2) continuously stirring at the rotating speed of 100-300r/min, and simultaneously uniformly and slowly adding a reducing agent, wherein the ratio of the reducing agent to the total tellurium substances in the copper-separating slag of the copper anode slime is 1.0-2.0: 1, adding a reducing agent, naturally cooling after the adding of the reducing agent is finished, continuously stirring and reacting for 0.8-1.2h, and filtering after the reaction is finished to obtain crude tellurium powder and a solution after tellurium separation.
6. The method as claimed in claim 2, wherein the oxidant in S1.2) is one or more of sodium chlorate, oxygen, chlorine and hydrogen peroxide; the addition amount is as follows: the amount of the gold and tellurium in the raw material is 1.0 to 1.3 times of the theoretical amount of the oxidant consumed by oxidizing all the tellurium.
7. The method according to any one of claims 3, 4 or 5, wherein the reducing agent is one or a combination of sodium sulfite, sodium bisulfite, sulfur dioxide, ammonia water, oxalic acid.
8. The method of claim 7, wherein the reducing agent is added by: the solid reducing agent is prepared into solution and then atomized and sprayed, and the liquid reducing agent is diluted and atomized and sprayed, wherein the concentration of the reducing agent in the prepared solution is not higher than 1.0mol/L, the speed of spraying the reducing agent solution in each reaction solution in a physical and chemical mode is not higher than 20L/h, and the adding speed of the gas reducing agent in each reaction solution is not higher than 30L/min.
9. The method of claim 7, wherein the sodium sulfite, sodium bisulfite, sulfur dioxide and oxalic acid are all technical grade reagents, wherein the mass percentage concentration of the sodium sulfite, the sodium bisulfite and the oxalic acid is not lower than 98.5%, and the volume percentage concentration of the sulfur dioxide gas is not lower than 98.5%.
10. The method according to claim 1, wherein the copper anode slime copper separating slag comprises the following components in percentage by weight: 1.0-20.0% of Te1.0-10.0% of Sb1.0-10.0%, 1.0-30.0% of Pb0%, 1.0-10.0% of As1.0%, 1.0-20.0% of Ag0.01-2.0% of Au0.
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