Dressing and smelting combined method for separating copper and silver minerals in alkaline system

Abstract

The invention discloses a beneficiation and metallurgy combined method for separating copper and silver minerals under an alkaline system. The beneficiation and metallurgy combined method comprises the following steps: grinding ores, carrying out copper and silver flotation on raw ores under alkaline conditions, carrying out alkaline roasting on flotation concentrate, carrying out ammonia leaching on roasted slag to extract copper and silver, carrying out copper and silver flotation on ammonia leaching slag, and carrying out replacement and silver extraction in ammonia leaching liquid. The method is particularly suitable for the silver-copper ore with tetrahedrite as the main component, production of metal products can be directly achieved on the mine site, and the method has the advantages of being simple in dressing and smelting process and remarkable in economic benefit.

Description

Technical Field

[0001] This invention belongs to the field of mineral processing technology, specifically relating to a combined mineral processing method for separating copper and silver minerals in an alkaline system. Background Technology

[0002] Secondary copper deposits with high silver content in nature are mainly represented by tetrahedrite, in which characteristic silver atoms are embedded in the copper mineral lattice, forming silver-copper symbiotic minerals. In sulfide mineral flotation collector systems, these minerals exhibit similar flotation behaviors between silver and copper minerals, making effective separation by physical methods difficult. Due to the natural physicochemical properties and floatability of copper and silver minerals, the obtained copper-silver concentrates are difficult to utilize at high value. Whether sold as copper or silver concentrate, the value of the other metal is usually discounted due to the difficulty in separation.

[0003] Currently, conventional methods for processing these copper-silver minerals typically employ pyrometallurgical processes, recovering crude copper and extracting silver from the anode slime obtained through electrolytic refining. However, while this process is highly adaptable, it suffers from high investment costs, high energy consumption, and complex precious metal recovery processes, particularly limiting the potential for further enhancing the economic value of silver. Therefore, developing a short-process, low-energy-consumption copper-silver mineral processing technology is a key requirement for the efficient utilization of this resource. This presents not only significant technical challenges but also substantial economic and research value. Summary of the Invention

[0004] The purpose of this invention is to overcome the defects or shortcomings of existing copper pyrometallurgical processes, such as long process length, high energy consumption, and significant environmental pollution, and to provide a combined beneficiation and smelting method for separating copper and silver minerals under an alkaline system, thereby achieving the goal of producing copper and silver metals locally in the concentrator.

[0005] To achieve the above objectives, the technical solution adopted by the present invention is as follows:

[0006] A combined beneficiation and metallurgical method for separating copper-silver minerals in an alkaline system includes the following steps:

[0007] S1. The copper-silver mineral-bearing ore is crushed and then fed into a ball mill for grinding. An alkaline modifier is added during the grinding process. The slurry after grinding is subjected to flotation in a closed-circuit process to obtain a flotation concentrate containing copper-silver minerals.

[0008] S2. The flotation concentrate obtained in S1 is transported to a deep cone thickener for concentration, and then sent to a plate and frame filter press for filtration to obtain filter cake.

[0009] S3. A basic modifier is uniformly added to the filter cake obtained in S2, and the mixture is sent to a rotary kiln for basic roasting to obtain roasted residue.

[0010] S4. After cooling the roasted residue obtained in S3, add water and ammonia water for ammonia leaching. Then send the ammonia-leached slurry to a plate and frame filter press for filtration to separate the ammonia leaching residue and ammonia leaching liquid.

[0011] S5. The ammonia leaching residue obtained in S4 is subjected to flotation to obtain flotation concentrate. The obtained flotation concentrate is returned to the concentration and filtration operation in S2, and then enters the roasting operation in S3.

[0012] S6. Add copper powder to the ammonia leaching solution obtained in S4 to carry out a silver replacement reaction. After filtration and precipitation, sponge silver and copper-rich solution are obtained.

[0013] Furthermore, in step S1, the portion of the slurry with a particle size less than 0.043 mm accounts for 70% to 75% of the total material, and the alkaline modifier sodium carbonate is used at a dosage of 2000 to 3000 g / t.

[0014] Furthermore, the flotation process described in step S1 specifically consists of one roughing, two sweeping, and three cleaning stages;

[0015] Sodium sulfite is added as a modifier during the roughing stage;

[0016] The roughing stage and the two scavenging stages are combined with ethyl xanthate, isopropyl dithiophosphate and ethanol as a composite collector.

[0017] The synergistic effect of a composite collector consisting of ethyl xanthate, isopropyl dithiophosphate, and ethanol can improve the recovery and selectivity of target sulfide minerals. Ethyl xanthate has a strong collecting ability for sulfide minerals, while isopropyl dithiophosphate enhances the selectivity for target minerals, and ethanol improves reagent dispersibility and uniformity of action. Using only one of these components, or lacking the dispersing component, will lead to an imbalance between collection and selectivity, uneven reagent action, and a decrease in flotation efficiency and recovery.

[0018] Furthermore, the mass ratio of ethyl xanthate formate, isopropyl dithiophosphate, and ethanol in the composite collector is 5:4:1. The amount of the composite collector used in the roughing stage is 120-150 g / t, the amount of the composite collector used in the first scavenging stage is 30-40 g / t, and the amount of the composite collector used in the second scavenging stage is 10-20 g / t.

[0019] Furthermore, in step S2, the flotation concentrate is transported to a deep cone thickener and concentrated to a moisture content of 50% to 60%, and the filter cake has a moisture content of less than 15%.

[0020] Further, the alkaline adjuster in step S3 is a sodium hydroxide solution with a mass fraction of 1% to 5%, and the calcination conditions are a temperature of 550 to 650°C and a time of 1 to 2 hours.

[0021] Furthermore, the ammonia leaching in step S4 is carried out in a closed, room-temperature stirred leaching process, with the concentration of the ammonia water being 5–10 mol / L, the liquid-to-solid ratio being 6:1, and the leaching time being 1–2 h.

[0022] Furthermore, the flotation process described in step S5 specifically consists of one roughing, one sweeping, and three cleaning stages;

[0023] Sodium sulfite is added as a modifier during the roughing stage;

[0024] Dicyclohexylaminodithiophosphate and sodium isopentyl xanthate are added as a composite collector during the roughing stage.

[0025] The synergistic effect of the combined collectors dicyclohexylaminodithiophosphate and sodium isopentyl xanthate can improve the recovery and selectivity of target minerals (especially copper minerals). Dicyclohexylaminodithiophosphate exhibits high selectivity for target minerals, while sodium isopentyl xanthate enhances the collection ability of minerals. Using only one of them results in a single collecting performance, which can lead to insufficient recovery when selectivity is high or poor selectivity when collection is strong, thereby reducing flotation efficiency.

[0026] Furthermore, the amount of sodium sulfite used in the flotation stage is 800-1000 g / t, the mass ratio of dicyclohexylaminodithiophosphate and sodium isopentyl xanthate in the composite collector is 1:1, and the amount of the composite collector used in the roughing stage is 200-250 g / t.

[0027] Furthermore, in step S6, the amount of copper powder added is 2-2.5 g / L, and the replacement time is 1-2 h.

[0028] Compared with the prior art, the beneficial effects of the present invention are as follows:

[0029] Compared with traditional pyrometallurgical processes, this invention is suitable for copper mines with high silver content, especially mines in remote areas without surrounding smelters, in order to improve the economic benefits of the mines.

[0030] This invention addresses the complex mineralogical characteristics of copper and silver minerals by employing a highly efficient separation and comprehensive utilization method combining beneficiation and metallurgy. Through in-depth research on the physicochemical properties and floatability characteristics of the minerals, efficient separation and high-value utilization of copper and silver in flotation concentrates are achieved. Compared with existing technologies, this method has lower investment costs, lower energy consumption, and saves significant concentrate transportation costs. Detailed Implementation

[0031] The present invention is further illustrated below with reference to specific embodiments. These embodiments are for illustrative purposes only and are not intended to limit the scope of the invention. Experimental methods in the following embodiments that do not specify specific conditions are generally performed under conventional conditions in the art or as recommended by the manufacturer; the raw materials and reagents used, unless otherwise specified, are all commercially available from the conventional market. Any non-substantial changes and substitutions made by those skilled in the art based on the present invention are within the scope of protection claimed by the present invention.

[0032] Example 1

[0033] The raw ore was selected from a copper-silver mine in Xilingol League, Inner Mongolia, China. The copper grade of the raw ore was 0.53%, and the silver grade was 124 g / t.

[0034] The raw ore is crushed and ground, with the grinding fineness controlled to be less than 0.043 mm, accounting for 70%. During the grinding process, 2000 g / t of sodium carbonate is added as an alkaline modifier for flotation.

[0035] The flotation process employs a closed-circuit process with middlings sequentially returned after one roughing, two scavenging, and three cleaning stages. First, 120 g / t of a composite collector (ethyl xanthate:isopropyl dithiophosphate:ethanol mass ratio = 5:4:1) is added for the first roughing stage, yielding roughing concentrate and roughing tailings. Then, 40 g / t of the composite collector (ethyl xanthate:isopropyl dithiophosphate:ethanol mass ratio = 5:4:1) is added to the roughing tailings for the first scavenging stage, yielding scavenging concentrate 1 and scavenging tailings 1. Finally, 20 g / t of the composite collector (ethyl xanthate:isopropyl dithiophosphate:ethanol mass ratio = 5:4:1) is added to the scavenging tailings for the second scavenging stage, yielding scavenging concentrate 2 and scavenging tailings 2. The rougher concentrate is subjected to a first blank cleaning process to obtain concentrate 1 and tailings 1. The concentrate 1 is subjected to a second blank cleaning process to obtain concentrate 2 and tailings 2. The concentrate 2 is subjected to a third blank cleaning process to obtain concentrate 3 and tailings 3, and finally flotation concentrate is obtained.

[0036] The flotation concentrate was concentrated to approximately 60% moisture content using a deep cone thickener, and then filtered through a plate and frame filter press until the moisture content was less than 15%. The concentrate filter cake was uniformly mixed with a 1% (w / w) sodium hydroxide solution and calcined at 550°C for 2 hours. The resulting calcined residue was cooled and then subjected to ammonia leaching under the following conditions: ammonia concentration of 10 mol / L, liquid-to-solid ratio of 6:1, and a reaction time of 2 hours. The ammonia-leached slurry was filtered through a plate and frame filter press to separate the ammonia leaching residue and the ammonia leaching liquor. 2 g / L copper powder was added to the ammonia leaching liquor to replace silver for 2 hours, producing a copper-rich solution.

[0037] The ammonia leaching residue enters the flotation process. First, 800 g / t of sodium sulfite is added as a modifier, followed by 200 g / t of a composite collector (dicyclohexylaminodithiophosphate: sodium isopentyl xanthate = 1:1). After one roughing stage, rougher concentrate and rougher tailings are obtained. The rougher tailings undergo one scavenging stage, yielding scavenged concentrate and scavenged tailings. The rougher concentrate undergoes a first cleaning stage, yielding Cleaner 1 concentrate and Cleaner 1 tailings. Cleaner 1 concentrate undergoes a second cleaning stage, yielding Cleaner 2 concentrate and Cleaner 2 tailings. Cleaner 2 concentrate undergoes a third cleaning stage, yielding Cleaner 3 concentrate and Cleaner 3 tailings. Finally, the ammonia leaching residue flotation concentrate is returned to the concentration and filtration process of the original ore flotation concentrate for further processing.

[0038] In this embodiment, the copper grade in the raw ore flotation concentrate was 10.68%, with a copper recovery rate of 90.37%, and the silver grade was 3248 g / t, with a silver recovery rate of 87.59%. The copper grade in the ammonia leaching residue flotation concentrate was 15.36%, with a copper recovery rate of 97.28%, and the silver grade was 4367 g / t, with a silver recovery rate of 98.62%. After silver replacement, a sponge silver solution with a silver content of 72.69% and a silver recovery rate of 97.39% was obtained, along with a copper-rich solution with a copper content of 16.33 g / L and a copper leaching rate of 98.11%.

[0039] Example 2

[0040] The raw ore was selected from a copper-silver mine in Chifeng, Inner Mongolia, China. The copper grade in the raw ore was 0.87%, and the silver grade was 92.61 g / t.

[0041] After being crushed, the raw ore enters the grinding stage, with 75% of the ore having a fineness of less than 0.043 mm. During the grinding process, 3000 g / t of sodium carbonate is added as an alkaline modifier for flotation.

[0042] The flotation process employs a closed-circuit process with middlings sequentially returned after one roughing, two scavenging, and three cleaning stages. First, 150 g / t of a composite collector (ethyl xanthate:isopropyl dithiophosphate:ethanol mass ratio = 5:4:1) is added for the first roughing stage, yielding a roughing concentrate and roughing tailings. Then, 40 g / t of the composite collector (ethyl xanthate:isopropyl dithiophosphate:ethanol mass ratio = 5:4:1) is added to the roughing tailings for the first scavenging stage, yielding scavenging concentrate 1 and scavenging tailings 1. Finally, 20 g / t of the composite collector (ethyl xanthate:isopropyl dithiophosphate:ethanol mass ratio = 5:4:1) is added to the scavenging tailings for the second scavenging stage, yielding scavenging concentrate 2 and scavenging tailings 2. The rougher concentrate is subjected to a first blank cleaning process to obtain concentrate 1 and tailings 1. The concentrate 1 is subjected to a second blank cleaning process to obtain concentrate 2 and tailings 2. The concentrate 2 is subjected to a third blank cleaning process to obtain concentrate 3 and tailings 3, and finally flotation concentrate is obtained.

[0043] The flotation concentrate was concentrated to approximately 60% moisture content using a deep cone thickener, and then filtered through a plate and frame filter press until the moisture content was less than 15%. Subsequently, the concentrate filter cake was uniformly mixed with a 3% (w / w) sodium hydroxide solution and then fed into a roasting furnace for roasting at 600℃ for 2 hours. The roasted slag was cooled and then subjected to ammonia leaching. The ammonia concentration in the leaching solution was 10 mol / L, the liquid-to-solid ratio was 6:1, and the leaching time was 2 hours. The ammonia-leached slurry was filtered through a plate and frame filter press to obtain ammonia leaching residue and ammonia leaching solution. 2.5 g / L copper powder was added to the ammonia leaching solution for silver extraction via displacement, with a displacement time of 2 hours. The silver extraction tailings were then the copper-rich solution.

[0044] The ammonia leaching residue enters the flotation process. First, 1000 g / t of sodium sulfite is added as a modifier, followed by 240 g / t of a composite collector (dicyclohexylaminodithiophosphate: sodium isopentyl xanthate = 1:1). After one roughing stage, rougher concentrate and rougher tailings are obtained. The rougher tailings undergo one scavenging stage, yielding scavenged concentrate and scavenged tailings. The rougher concentrate undergoes a first cleaning stage, yielding Cleaner 1 concentrate and Cleaner 1 tailings. Cleaner 1 concentrate undergoes a second cleaning stage, yielding Cleaner 2 concentrate and Cleaner 2 tailings. Cleaner 2 concentrate undergoes a third cleaning stage, yielding Cleaner 3 concentrate and Cleaner 3 tailings. Finally, the ammonia leaching residue flotation concentrate is returned to the concentration and filtration process of the original ore flotation concentrate for further processing.

[0045] In this embodiment, the copper grade in the raw ore flotation concentrate is 13.43%, with a copper recovery rate of 92.37%, and the silver grade is 4368 g / t, with a silver recovery rate of 90.24%. The copper grade in the ammonia leaching residue flotation concentrate is 16.47%, with a copper recovery rate of 96.17%, and the silver grade is 5248 g / t, with a silver recovery rate of 97.49%. After silver replacement, a sponge silver solution with a silver content of 79.18% and a silver recovery rate of 98.41% is obtained, along with a copper-rich solution with a copper content of 19.51 g / L and a copper leaching rate of 98.74%.

[0046] Comparative Example 1

[0047] The raw ore was selected from a copper-silver mine in Chifeng, Inner Mongolia, China. The copper grade of the raw ore was 0.87%, and the silver grade was 92.61 g / t.

[0048] After being crushed, the raw ore enters the grinding stage, with 75% of the ore having a fineness of less than 0.043 mm. During the grinding process, 3000 g / t of sodium carbonate is added as an alkaline modifier for flotation.

[0049] The flotation process employs a closed-circuit process with middlings sequential return after one roughing, two scavenging, and three cleaning stages. First, 150 g / t of ethyl xanthate formate is added as a collector. After one roughing stage, rougher concentrate and rougher tailings are obtained. Then, 40 g / t of ethyl xanthate formate is added to the rougher tailings for the first scavenging stage, yielding scavenger concentrate 1 and scavenger tailings 1. Next, 20 g / t of ethyl xanthate formate is added to the scavenger tailings for the second scavenging stage, yielding scavenger concentrate 2 and scavenger tailings 2. A first blank cleaning stage is performed on the rougher concentrate, yielding cleaner concentrate 1 and cleaner tailings 1. A second blank cleaning stage is performed on the cleaner concentrate 1, yielding cleaner concentrate 2 and cleaner tailings 2. Finally, a third blank cleaning stage is performed on the cleaner concentrate 2, yielding cleaner concentrate 3 and cleaner tailings 3, ultimately obtaining the flotation concentrate.

[0050] The flotation concentrate was concentrated to approximately 60% moisture content using a deep cone thickener, and then filtered through a plate and frame filter press until the moisture content was less than 15%. Subsequently, the concentrate filter cake was uniformly mixed with a 3% (w / w) sodium hydroxide solution and then fed into a roasting furnace for roasting at 600℃ for 2 hours. The roasted slag was cooled and then subjected to ammonia leaching. The ammonia concentration in the leaching solution was 10 mol / L, the liquid-to-solid ratio was 6:1, and the leaching time was 2 hours. The ammonia-leached slurry was filtered through a plate and frame filter press to obtain ammonia leaching residue and ammonia leaching solution. 2.5 g / L copper powder was added to the ammonia leaching solution for silver extraction via displacement, with a displacement time of 2 hours. The silver extraction tailings were then the copper-rich solution.

[0051] The ammonia leaching residue enters the flotation process. First, 1000 g / t of sodium sulfite is added as a modifier, followed by 240 g / t of sodium isopentyl xanthate as a collector. After one roughing stage, rougher concentrate and rougher tailings are obtained. The rougher tailings undergo a scavenging stage, yielding scavenged concentrate and scavenged tailings. The rougher concentrate undergoes a first cleaning stage, yielding Cleaner 1 concentrate and Cleaner 1 tailings. Cleaner 1 concentrate undergoes a second cleaning stage, yielding Cleaner 2 concentrate and Cleaner 2 tailings. Cleaner 2 concentrate undergoes a third cleaning stage, yielding Cleaner 3 concentrate and Cleaner 3 tailings. Finally, the ammonia leaching residue flotation concentrate is returned to the concentration and filtration process of the original ore flotation concentrate for further processing.

[0052] In this comparative example, the copper grade in the raw ore flotation concentrate was 9.82%, with a copper recovery rate of 87.65%, and the silver grade was 3674 g / t, with a silver recovery rate of 85.27%. The copper grade in the ammonia leaching residue flotation concentrate was 14.28%, with a copper recovery rate of 91.46%, and the silver grade was 4774 g / t, with a silver recovery rate of 93.66%. After silver replacement, a sponge silver solution with a silver content of 78.57% and a silver recovery rate of 95.61% was obtained, along with a copper-rich solution with a copper content of 14.38 g / L and a copper leaching rate of 93.27%.

[0053] Comparative Example 2

[0054] The raw ore was selected from a copper-silver mine in Xilingol League, Inner Mongolia, China. The copper grade of the raw ore was 0.53%, and the silver grade was 124 g / t.

[0055] After being crushed, the raw ore enters the grinding stage, with 70% of the ore having a fineness of less than 0.043 mm. During the grinding process, 2000 g / t of sodium carbonate is added as an alkaline modifier for flotation.

[0056] The flotation process employs a closed-circuit process with middlings sequential return after one roughing, two scavenging, and three cleaning stages. First, 120 g / t of isopropyl dithiophosphate is added as a collector. After one roughing stage, rougher concentrate and rougher tailings are obtained. Then, 40 g / t of isopropyl dithiophosphate is added to the rougher tailings for the first scavenging stage, yielding scavenger concentrate 1 and scavenger tailings 1. Next, 20 g / t of isopropyl dithiophosphate is added to the scavenger tailings for the second scavenging stage, yielding scavenger concentrate 2 and scavenger tailings 2. A first blank cleaning stage is performed on the rougher concentrate, yielding cleaner concentrate 1 and cleaner tailings 1. A second blank cleaning stage is performed on the cleaner concentrate 1, yielding cleaner concentrate 2 and cleaner tailings 2. Finally, a third blank cleaning stage is performed on the cleaner concentrate 2, yielding cleaner concentrate 3 and cleaner tailings 3, ultimately obtaining the flotation concentrate.

[0057] The flotation concentrate was concentrated to approximately 60% moisture content using a deep cone thickener, and then filtered through a plate and frame filter press until the moisture content was less than 15%. Subsequently, the concentrate filter cake was uniformly mixed with a 1% sodium hydroxide solution and then fed into a roasting furnace for roasting at 550℃ for 2 hours. The roasted slag was cooled and then subjected to ammonia leaching. The ammonia concentration in the leaching solution was 10 mol / L, the liquid-to-solid ratio was 6:1, and the leaching time was 2 hours. The ammonia-leached slurry was filtered through a plate and frame filter press to obtain ammonia leaching residue and ammonia leaching solution. 2 g / L copper powder was added to the ammonia leaching solution for silver extraction via displacement, with a displacement time of 2 hours. The silver extraction tailings were then the copper-rich solution.

[0058] The ammonia leaching residue enters the flotation process. First, 800 g / t of sodium sulfite is added as a modifier, followed by 200 g / t of dicyclohexanedithiophosphate as a collector. After one roughing stage, rougher concentrate and rougher tailings are obtained. The rougher tailings undergo one scavenging stage, yielding scavenger concentrate and scavenger tailings. The rougher concentrate undergoes a first cleaning stage, yielding Cleaner 1 concentrate and Cleaner 1 tailings. Cleaner 1 concentrate undergoes a second cleaning stage, yielding Cleaner 2 concentrate and Cleaner 2 tailings. Cleaner 2 concentrate undergoes a third cleaning stage, yielding Cleaner 3 concentrate and Cleaner 3 tailings. Finally, the ammonia leaching residue flotation concentrate is returned to the concentration and filtration process of the original ore flotation concentrate for further processing.

[0059] In this comparative example, the copper grade in the raw ore flotation concentrate was 8.46%, with a copper recovery rate of 85.72%, and the silver grade was 3077 g / t, with a silver recovery rate of 82.88%. The copper grade in the ammonia leaching residue flotation concentrate was 13.66%, with a copper recovery rate of 94.87%, and the silver grade was 4173 g / t, with a silver recovery rate of 95.33%. After silver replacement, a sponge silver solution with a silver content of 70.59% and a silver recovery rate of 97.11% was obtained, along with a copper-rich solution with a copper content of 15.17 g / L and a copper leaching rate of 97.24%.

[0060] Therefore, it can be seen that adding a single reagent in the flotation process will lead to a significant decrease in the recovery rate of the operation; and the decrease in the recovery rate of the operation will further lead to a decrease in the final recovery rate.

[0061] Analysis of the detection results of the raw ore flotation concentrate, ammonia leaching residue flotation concentrate, and silver-replaced products separated in Examples 1 and 2 and Comparative Examples 1 and 2 of this invention shows that the method of this invention significantly improves the grade and recovery rate of copper and silver in the flotation concentrate, realizing the high-value utilization of copper and silver. Using the method of this invention not only effectively improves the overall recovery efficiency of copper and silver but also significantly enhances the separation accuracy of silver and copper, providing reliable technical support for the efficient utilization of resources.

Claims

1. A combined beneficiation and metallurgical method for separating copper-silver minerals in an alkaline system, characterized in that, Includes the following steps: S1. The copper-silver mineral-bearing ore is crushed and then fed into a ball mill for grinding. An alkaline modifier is added during the grinding process. The slurry after grinding is subjected to flotation in a closed-circuit process to obtain a flotation concentrate containing copper-silver minerals. S2. The flotation concentrate obtained in S1 is transported to a deep cone thickener for concentration, and then sent to a plate and frame filter press for filtration to obtain filter cake. S3. A basic modifier is uniformly added to the filter cake obtained in S2, and the mixture is sent to a rotary kiln for basic roasting to obtain roasted residue. S4. After cooling the roasted residue obtained in S3, add water and ammonia water for ammonia leaching. Then send the ammonia-leached slurry to a plate and frame filter press for filtration to separate the ammonia leaching residue and ammonia leaching liquid. S5. The ammonia leaching residue obtained in S4 is subjected to flotation to obtain flotation concentrate. The obtained flotation concentrate is returned to the concentration and filtration operation in S2, and then enters the roasting operation in S3. S6. Add copper powder to the ammonia leaching solution obtained in S4 to carry out a silver replacement reaction. After filtration and precipitation, sponge silver and copper-rich solution are obtained.

2. The combined beneficiation and metallurgical method for separating copper-silver minerals in an alkaline system according to claim 1, characterized in that, In step S1, the portion of the slurry with a particle size less than 0.043 mm in the grinding process accounts for 70% to 75% of the total material. The alkaline modifier is sodium carbonate, and the amount of sodium carbonate used is 2000 to 3000 g / t.

3. The combined beneficiation and metallurgical method for separating copper-silver minerals in an alkaline system according to claim 1, characterized in that, The flotation process described in step S1 specifically consists of one roughing, two sweeping, and three cleaning stages; Sodium sulfite is added as a modifier during the roughing stage; The roughing stage and the two scavenging stages are combined with ethyl xanthate, isopropyl dithiophosphate and ethanol as a composite collector.

4. The combined beneficiation and metallurgical method for separating copper-silver minerals in an alkaline system according to claim 3, characterized in that, The mass ratio of ethyl xanthate formate, isopropyl dithiophosphate, and ethanol in the composite collector is 5:4:

1. The amount of the composite collector used in the roughing stage is 120-150 g / t, the amount of the composite collector used in the first scavenging stage is 30-40 g / t, and the amount of the composite collector used in the second scavenging stage is 10-20 g / t.

5. The combined beneficiation and metallurgical method for separating copper-silver minerals in an alkaline system according to claim 1, characterized in that, In step S2, the flotation concentrate is transported to a deep cone thickener and concentrated to a moisture content of 50% to 60%, and the filter cake has a moisture content of less than 15%.

6. The combined beneficiation and metallurgical method for separating copper-silver minerals in an alkaline system according to claim 1, characterized in that, The alkaline adjuster in step S3 is a sodium hydroxide solution with a mass fraction of 1% to 5%, and the calcination conditions are a temperature of 550 to 650°C and a time of 1 to 2 hours.

7. The combined beneficiation and metallurgical method for separating copper-silver minerals in an alkaline system according to claim 1, characterized in that, The ammonia leaching in step S4 is carried out in a closed, room-temperature stirred leaching process. The concentration of the ammonia solution is 5-10 mol / L, the liquid-to-solid ratio is 6:1, and the leaching time is 1-2 hours.

8. The combined beneficiation and metallurgical method for separating copper-silver minerals in an alkaline system according to claim 1, characterized in that, The flotation process described in step S5 specifically consists of one roughing, one sweeping, and three cleaning stages; Sodium sulfite is added as a modifier during the roughing stage; Dicyclohexylaminodithiophosphate and sodium isopentyl xanthate are added as a composite collector during the roughing stage.

9. The combined beneficiation and metallurgical method for separating copper-silver minerals in an alkaline system according to claim 8, characterized in that, In the flotation stage, the amount of sodium sulfite used is 800-1000 g / t, and the mass ratio of dicyclohexylaminodithiophosphate and sodium isopentyl xanthate in the composite collector is 1:

1. In the roughing stage, the amount of the composite collector used is 200-250 g / t.

10. The combined beneficiation and metallurgical method for separating copper-silver minerals in an alkaline system according to claim 1, characterized in that, The amount of copper powder added in step S6 is 2-2.5 g / L, and the replacement time is 1-2 h.