Efficient recovery method of valuable elements in high gold and high silver zinc concentrate

Abstract

This invention relates to an efficient method for recovering valuable elements from high-gold, high-silver, and high-zinc concentrates, comprising the following steps: (1) gravity separation pretreatment; (2) oxidative roasting; (3) atmospheric pressure acid leaching: the roasted slag obtained in step (2) is ground, and the resulting roasted slag grinding slurry is subjected to atmospheric pressure sulfuric acid leaching, with 3 to 8 leaching stages, controlling the concentration of the leaching slurry at each stage to be 15% to 35%, the sulfuric acid concentration to be 98%, the amount of sulfuric acid to be 1.5 to 3 times the dry weight of the material before leaching, the leaching time to be 2 to 8 hours, and the stirring speed to be 200 to 800 rpm; (4) sulfuric acid leaching slurry separation; (5) water washing: obtaining wash water and water washing slag; (6) atmospheric pressure cyanide leaching: leaching gold and silver from the water washing slag grinding slurry to obtain a cyanide leaching slurry containing gold and silver; (7) cyanide leaching slurry separation. This method simplifies the process of recovering valuable metals from zinc concentrate, thereby avoiding the problems of high energy consumption in pyrometallurgical processes and long processes in hydrometallurgical processes, and the entire process generates almost no hazardous waste.

Description

Technical Field

[0001] This invention belongs to the field of polymetallic ore beneficiation technology, specifically relating to an efficient method for recovering valuable elements from high-gold, high-silver, and high-zinc concentrates. Background Technology

[0002] Existing zinc concentrate smelting processes are mainly divided into hydrometallurgical smelting and pyrometallurgical smelting. Pyrometallurgical smelting has disadvantages such as high energy consumption and strict operating conditions, and has been gradually phased out. Therefore, hydrometallurgical smelting is the main smelting process for zinc concentrate. Since polymetallic sulfide ores contain associated precious metals such as gold and silver, flotation processes will produce zinc concentrates with high gold and silver grades. However, conventional zinc concentrate smelting processes cannot directly recover this portion of gold and silver. These metals will be enriched in smelting slag or flue gas, and usually require subsequent pyrometallurgical treatment, hydrometallurgical treatment, and combined beneficiation treatment of the smelting slag for recovery. This not only prolongs the overall processing flow but also has disadvantages such as high energy consumption and small processing capacity.

[0003] Taking zinc concentrate hydrometallurgical smelting as an example, zinc concentrate hydrometallurgical smelting methods are divided into conventional leaching, hot acid leaching, high-pressure oxygen leaching, and atmospheric pressure oxygen leaching, the main purpose of which is to leach zinc. In hydrometallurgical smelting, iron, copper, and zinc in the leachate can be separated and recovered through iron precipitation, zinc powder replacement, and electrolysis. However, the leaching residue not only contains gold and silver, but also a considerable amount of other valuable metals such as iron, copper, and zinc. To recover these valuable metals, different methods such as pyrometallurgical treatment, hydrometallurgical treatment, pyrometallurgical-hydrometallurgical combined treatment, and beneficiation-metallurgical combined treatment are required to separate and recover the valuable metal elements, depending on the type of leaching residue. Among these methods, pyrometallurgical treatment consumes a lot of energy, hydrometallurgical treatment has a long process, and generates a lot of hazardous waste that is difficult to dispose of. Summary of the Invention

[0004] To address the aforementioned problems, this invention provides an efficient method for recovering valuable elements from high-gold and high-silver zinc concentrates. By pre-leaching zinc, copper, and iron from the zinc concentrate during acid leaching, the acid leaching residue no longer requires complex separation and recovery treatment. This simplifies the valuable metal recovery process from zinc concentrates, avoiding the problems of high energy consumption in pyrometallurgical processes and long hydrometallurgical processes. Moreover, the entire process generates almost no hazardous waste.

[0005] The specific technical solution is as follows: An efficient method for recovering valuable elements from high-gold, high-silver, and high-zinc concentrates includes the following steps: (1) Gravity separation pretreatment: The zinc concentrate is subjected to gravity separation pretreatment to obtain low gold zinc concentrate and high gold zinc concentrate.

[0006] (2) Oxidative roasting: The high gold and zinc concentrate is subjected to oxidative roasting to obtain roasting residue and roasting flue gas; (3) Atmospheric pressure acid leaching: The roasted slag obtained in step (2) is ground and the resulting roasted slag grinding slurry is subjected to atmospheric pressure sulfuric acid leaching to react with sulfuric acid to obtain sulfuric acid leaching slurry; the number of leaching stages in the atmospheric pressure sulfuric acid leaching operation is 3 to 8, the concentration of leaching slurry in each stage is controlled at 15% to 35%, the sulfuric acid concentration is 98%, the amount of sulfuric acid used is 1.5 to 3 times the dry weight of the material before leaching, the leaching time is 2h to 8h, and the stirring speed is 200rpm to 800rpm.

[0007] (4) Separation of sulfuric acid leaching slurry: The sulfuric acid leaching slurry from step (3) is subjected to solid-liquid separation treatment to obtain acid leaching solution and acid leaching residue; (5) Water washing: The acid leaching residue obtained in step (4) is washed with water to obtain acid leaching residue water washing slurry. Then, the acid leaching residue water washing slurry is subjected to solid-liquid separation treatment to obtain washing water and water washing residue. (6) Atmospheric pressure cyanide leaching: The water-washed slag obtained in step (5) is ground to obtain water-washed slag grinding slurry. The water-washed slag grinding slurry is subjected to atmospheric pressure cyanide leaching to leach gold and silver from the water-washed slag grinding slurry to obtain cyanide leaching slurry containing gold and silver. (7) Cyanide leaching slurry separation: The cyanide leaching slurry obtained in step (6) is subjected to solid-liquid separation treatment to obtain cyanide leaching solution and cyanide residue.

[0008] In the above steps, in step (1), since the gold in the high-gold, high-silver-zinc concentrate raw material is mainly exposed and semi-exposed native gold, it can be separated by gravity separation based on the difference in specific gravity between gold and different minerals. After gravity separation pretreatment, low-gold-zinc concentrate and high-gold-zinc concentrate with further enriched gold grade can be obtained. The low-gold-zinc concentrate can be sold, while the high-gold-zinc concentrate can be used as raw material in step (2), thereby improving the gold recovery efficiency and reducing reagent consumption. At the same time, gravity separation pretreatment can enrich and improve the gold grade, correspondingly reducing the proportion of other elements, providing conditions for subsequent reactions.

[0009] Step (2) oxidative roasting converts copper, iron, zinc, and lead in zinc concentrate into oxide roasting slag. Combined with the grinding process in step (3), these oxides can be converted into sulfates in atmospheric pressure acid leaching. In step (3), atmospheric pressure acid leaching allows various oxides in the slurry to react fully with sulfuric acid, and zinc, iron, and copper are basically leached into the acid leaching solution. No additional temperature is required during acid leaching; the process relies solely on its own heat release. Moreover, due to the large amount of sulfuric acid used, the pH value will be extremely low during leaching, so no special requirements are needed for the acid leaching temperature and pH value.

[0010] Thus, the separated acid leaching solution can be used in existing technologies such as iron precipitation, zinc powder replacement, and electrolytic cell separation to recover copper, iron, and zinc. Compared to the hydrometallurgical smelting of zinc concentrate in existing technologies, the vast majority of copper and iron will not enter the acid leaching residue in this invention. The subsequent treatment of the acid leaching residue does not need to consider the influence of copper and iron, thereby simplifying the recovery of the acid leaching residue. Since zinc, copper, and iron are essentially removed from the acid leaching residue, not only is the grade of gold and silver further enriched, but the influence of zinc, copper, and iron elements consuming cyanide ions during cyanide leaching is also eliminated. Therefore, the recovery efficiency of gold and silver from subsequent cyanide leaching will be higher.

[0011] Furthermore, preferably, the gravity separation pretreatment adopts a shaking table and a spiral chute, and the separation process consists of one roughing and three to six scavenging processes. The roughing concentrate and the concentrate from each scavenging process are combined as a high gold and zinc concentrate, and the final scavenging tailings are used as a low gold and zinc concentrate.

[0012] Furthermore, preferably, the roasting flue gas in step (1) is prepared into sulfuric acid for use in the sulfuric acid leaching operation in step (2).

[0013] Furthermore, preferably, the calcination temperature in step (1) is 600℃~900℃ and the calcination time is 2h~5h.

[0014] This oxidative roasting step is carried out only in an air atmosphere, without the need for additional oxygen. It can essentially remove sulfur from the zinc concentrate, providing conditions for subsequent acid leaching.

[0015] Furthermore, preferably, the slurry concentration for grinding in steps (2) and (5) is 30% to 50%, and the grinding time is 5 min to 15 min.

[0016] Furthermore, preferably, the water washing process in step (4) is carried out at room temperature with a slurry concentration of 30% to 50% and a time of 15 min to 30 min.

[0017] Furthermore, preferably, in step (5), the number of leaching stages of atmospheric pressure cyanide leaching treatment is 5 to 10, the slurry concentration of each leaching stage is controlled at 10% to 30%, the sodium cyanide consumption is 10 kg / t to 50 kg / t, the leaching time is 12 h to 48 h, the stirring speed is 200 rpm to 800 rpm, air is continuously blown in during the leaching process, no additional heating is required, and lime is added to control the pH of the leaching process to be no lower than 11.

[0018] Furthermore, preferably, the zinc concentrate in step (1) has a gold grade greater than 10 g / t and a silver grade greater than 150 g / t.

[0019] The beneficial effects of this invention are as follows: Based on the initial oxidation roasting, the method involves two intermediate grinding processes, and finally, the zinc, copper, and iron in the zinc concentrate are leached together through an acid leaching process. The gold and silver content in the acid leaching residue will be further enriched, and the acid leaching residue contains very little zinc, copper, and iron. This allows the acid leaching residue to be directly cyanided to leach gold and silver, without the need for the complex process of separating and recovering various valuable metals from the acid leaching residue as is required in traditional processes.

[0020] Furthermore, since the acid leaching residue contains very little zinc, copper, and iron, the impact of zinc, copper, and iron ions on the subsequent cyanide leaching process can be reduced. This allows for more efficient recovery of gold and silver from the cyanide leaching. The separation and recovery processes for the acid leaching solution, cyanide leaching solution, and cyanide residue are mature. Valuable metals in the acid leaching solution can be further separated and recovered using a hydrometallurgical process. The cyanide residue, after harmless cyanide-breaking treatment, can be sold as a lead-containing secondary resource. Therefore, the method of this invention for recovering valuable metals not only optimizes the complex process of acid leaching residue treatment still required in traditional recycling, but also generates very little hazardous waste. Attached Figure Description

[0021] Figure 1 This is a flowchart of the present invention; Figure 2 This invention presents the main reaction equations and corresponding thermodynamic calculation data diagrams involved in the intermediate-temperature oxidative roasting of high-gold-zinc concentrate. Detailed Implementation

[0022] To make the technical problems and solutions solved by the present invention clearer, the present invention will be further described in detail below with reference to the accompanying drawings and embodiments. It should be understood that the specific embodiments described herein are merely for explaining the present invention and are not intended to limit the present invention.

[0023] High-gold, high-silver, and high-zinc concentrate from a certain region in Yunnan Province, produced through flotation, was used as raw material. The grades of various valuable elements are shown in Table 1. It can be seen that in addition to the high gold and silver content, the zinc concentrate also has slightly higher grades of copper, lead, and total iron. Meanwhile, the phase analysis data of gold in the raw material are shown in Table 2, indicating that most of the gold exists in the form of exposed and semi-exposed native gold. Using this raw material, the process flow described in this application was implemented. Figure 1 .

[0024] Table 1: Grade Detection Data of Valuable Elements in High-Gold, High-Silver, and High-Zinc Concentrate Raw Materials

[0025] Table 2: Phase analysis data of gold in high-gold, high-silver, and high-zinc concentrate raw materials Example 1

[0026] Five kilograms of high-gold, high-silver, and high-zinc concentrate were taken and subjected to a one-roughing, four-scavenging gravity separation pretreatment using a shaking table. The gravity separation pretreatment employed a shaking table and a spiral sluice. The roughing concentrate and the concentrates from each scavenging stage were combined to form the high-gold, high-zinc concentrate. The final scavenging tailings were used as the low-gold, high-zinc concentrate. The resulting low-gold, high-zinc concentrate had a gold grade below 5 g / t and a yield of 85.36%. The low-gold, high-zinc concentrate could be sold directly. The high-gold, high-zinc concentrate was used as raw material for oxidative roasting at 700°C for 3 hours to obtain roasting slag and roasting flue gas. The roasting flue gas was recovered and sent to a tail gas purification system to produce sulfuric acid. Figure 2 It can be seen that the reaction ΔG of sulfide minerals in high gold and zinc concentrate to oxides is less than 0 in the range of 100℃ to 1000℃, indicating that oxidative roasting can proceed spontaneously in this temperature range. Furthermore, by comparing the reaction ΔG of different minerals, it can be seen that pyrite is more likely to be converted into FeO.

[0027] The obtained roasted slag was ground using a ball mill with a slurry concentration of 40% and a grinding time of 10 minutes. The slurry was then fed into a leaching unit for a three-stage atmospheric pressure sulfuric acid leaching process. The amount of sulfuric acid used in each stage was controlled to be twice the dry weight of the sample before leaching, with a sulfuric acid concentration of 98%. The slurry concentration was diluted to 33%, and the leaching time was 5 hours with a stirring speed of 500 rpm. The sulfuric acid leaching slurry was then fed into a filter press for solid-liquid separation to obtain acid leaching solution and acid leaching residue. The acid leaching solution can be used for the separation and recovery of iron, zinc, and copper through general processes such as iron precipitation, zinc powder replacement, and electrolysis.

[0028] The acid leaching residue is fed into a water washing device for ambient temperature stirring and washing. The water washing slurry concentration is 35% and the time is 20 minutes. After solid-liquid separation treatment, the obtained acid leaching residue water washing slurry is obtained as wash water and water washing residue. The wash water can be returned to the atmospheric pressure acid leaching for recycling.

[0029] The washed slag was ground using a ball mill to achieve a slurry concentration of 40% for 10 minutes. The resulting slurry was then fed into a leaching unit for a 7-stage atmospheric pressure cyanide leaching process. The sodium cyanide consumption per stage was controlled at 30 kg / t, the slurry concentration was diluted to 33%, the leaching time was 24 hours, the stirring rate was 500 rpm, and lime was added to maintain a pH above 11 during leaching. Air was continuously bubbled into the leaching process, eliminating the need for additional heating. The resulting cyanide leaching slurry, containing gold and silver, underwent solid-liquid separation to obtain cyanide leaching solution and cyanide slag. The cyanide leaching solution can be processed for gold and silver extraction to recover gold and silver, while the cyanide slag can be treated to obtain lead-containing secondary resources.

[0030] Table 3: Grade Detection Data of Valuable Elements in the Recovery Process of Example 1

[0031] Among them, the yield of high gold and zinc concentrate was 14.64%, the yield of roasted slag after roasting was 96.21%, the yield of acid leaching slag was 11.25%, and the yield of cyanide slag was 98.59%.

[0032] As can be seen from the table above, the gold grade in the high-gold-zinc concentrate after gravity separation pretreatment is significantly enriched; the silver grade in the roasting slag after oxidative roasting is reduced, the grades of other metal elements do not change much, while the sulfur grade decreases significantly, indicating that most of the sulfur is removed in the form of sulfur dioxide gas, while a small portion of sulfur is converted into sulfate form due to insufficient oxygen and other reasons; after three stages of atmospheric pressure acid leaching, zinc, copper, and iron are basically leached out, while gold, silver, and lead are basically not leached out; after seven stages of atmospheric pressure cyanide leaching, the gold leaching rate can exceed 99%, the silver leaching rate can exceed 98%, while lead is not leached out and remains in the cyanide slag. Example 2

[0033] Five kilograms of high-gold, high-silver, and high-zinc concentrate were taken and subjected to a one-roughing and five-scavenging gravity separation pretreatment using a shaking table. The gravity separation pretreatment employed a shaking table and a spiral sluice. The roughing concentrate and the concentrates from each scavenging stage were combined to form the high-gold, high-zinc concentrate, and the final scavenging tailings were used as the low-gold, high-zinc concentrate. The low-gold, high-zinc concentrate had a gold grade of less than 5 g / t and could be sold directly. The high-gold, high-zinc concentrate was used as raw material for oxidative roasting at 750°C for 4 hours to obtain roasting slag and roasting flue gas. The roasting flue gas was recovered and sent to the tail gas purification system to produce sulfuric acid.

[0034] The obtained roasted slag was ground using a ball mill with a slurry concentration of 33% and a grinding time of 8 minutes. The ground slurry was then fed into a leaching device for a four-stage atmospheric pressure sulfuric acid leaching operation. The amount of sulfuric acid used in each stage was controlled to be 1.5 times the dry weight of the sample before leaching, with a sulfuric acid concentration of 98%, a slurry concentration of 25%, a leaching time of 4 hours, and a stirring speed of 500 rpm. The sulfuric acid leaching slurry was then fed into a filter press for solid-liquid separation to obtain acid leaching solution and acid leaching residue. The acid leaching solution can be used for the separation and recovery of iron, zinc, and copper through general processes such as iron precipitation, zinc powder replacement, and electrolysis.

[0035] The acid leaching residue is fed into a water washing device for ambient temperature stirring and washing. The water washing slurry concentration is 40% and the time is 25 minutes. After solid-liquid separation treatment, the obtained acid leaching residue water washing slurry is obtained as wash water and water washing residue. The wash water can be returned to the atmospheric pressure acid leaching for recycling.

[0036] The washed slag was ground using a ball mill to a pulp concentration of 33% for 8 minutes. The resulting pulp was then fed into a leaching unit for 8-stage atmospheric pressure cyanide leaching. The sodium cyanide consumption per stage was controlled at 20 kg / t, the pulp concentration at 25%, the leaching time at 24 hours, and the stirring speed at 500 rpm. Lime was added to maintain the pH at least 11 during the leaching process. Air was continuously bubbled in during leaching, eliminating the need for additional heating. The resulting cyanide leaching pulp, containing gold and silver, underwent solid-liquid separation to obtain cyanide leaching solution and cyanide slag. The cyanide leaching solution can be processed for gold and silver extraction to recover gold and silver. The cyanide slag, after harmless cyanide removal treatment, yields lead-containing secondary resources.

[0037] Table 4: Grade Detection Data of Valuable Elements in the Recovery Process of Example 2

[0038] Among them, the yield of high gold and zinc concentrate was 17.82%, the yield of roasted slag after roasting was 94.43%, the yield of acid leaching slag was 9.89%, and the yield of cyanide slag was 97.89%.

[0039] As can be seen from the table above, the gold grade in the high-zinc gold concentrate after gravity separation pretreatment is significantly enriched; the sulfur grade in the roasting slag after oxidative roasting is significantly reduced; after four stages of atmospheric pressure acid leaching, zinc, copper, and iron are basically leached out, while gold, silver, and lead are basically not leached out; after eight stages of atmospheric pressure cyanide leaching, the gold leaching rate can exceed 99%, the silver leaching rate can exceed 98%, while lead is not leached out and remains in the cyanide slag. Example 3

[0040] A pilot-scale continuous operation was conducted for 4 weeks on high-gold, high-silver, and high-zinc concentrate raw materials, with a processing capacity of 100 kg / h. A shaking table was used for one roughing and four scavenging stage as a gravity separation pretreatment. The resulting high-gold and high-zinc concentrate was then subjected to medium-temperature oxidative roasting at 750℃ for 5 hours in a tunnel kiln. The roasted residue was ground in a ball mill with a grinding concentration of 40% for 10 minutes. After diluting the grinding slurry to a concentration of 25%, four-stage sulfuric acid leaching at room temperature was performed. The amount of sulfuric acid used in each stage was controlled at 1.5 times the pre-leaching dry weight, with a sulfuric acid concentration of 98%, a leaching time of 5 hours, and a stirring speed of 600 rpm. After leaching, a filter press was used for solid-liquid separation.

[0041] The separated acid leaching residue was subjected to water washing and solid-liquid separation. The water washing concentration was 40%, and the time was 30 minutes. The water-washed residue was then regrinded for 10 minutes using a ball mill with a grinding concentration controlled at 40%. The regrinded slurry was diluted to a concentration of 25% and then subjected to 8-stage ambient temperature cyanide leaching. The sodium cyanide consumption per stage was controlled at 35 kg / t, the slurry concentration was 25%, the leaching time was 24 hours, the stirring speed was 500 rpm, and lime was added to control the pH of the leaching process to be no lower than 11. The high gold, high silver, and high zinc concentrate raw materials, the oxidized roasted samples, the acid leaching residue, and the cyanide residue were sampled and tested 3 times a day. The average was used as the daily data, and the data for every 7 days were averaged to obtain the weekly data, as shown in Tables 5 to 8.

[0042] Table 5. Grade detection data of relevant valuable elements during the first week of continuous pilot-scale operation.

[0043] Table 6. Grade data of relevant valuable elements during the second week of continuous pilot-scale operation.

[0044] Table 7. Grade data of relevant valuable elements during the third week of continuous pilot-scale operation.

[0045] Table 8. Grade detection data of relevant valuable elements during the 4th week of continuous pilot-scale operation.

[0046] After four weeks of continuous operation on a large scale, it was found that the invention remained effective in large-scale continuous production.

[0047] The present invention has been described in detail above through specific and preferred embodiments. However, those skilled in the art should understand that the present invention is not limited to the embodiments described above. Any modifications, equivalent substitutions, etc., made within the spirit and principles of the present invention should be included within the protection scope of the present invention.

Claims

1. A method for efficiently recovering valuable elements from high-gold, high-silver, and high-zinc concentrates, characterized in that, Includes the following steps: (1) Gravity separation pretreatment: The zinc concentrate is subjected to gravity separation pretreatment to obtain low gold and zinc concentrate and high gold and zinc concentrate; (2) Oxidative roasting: The high gold and zinc concentrate is subjected to oxidative roasting to obtain roasting residue and roasting flue gas; (3) Atmospheric pressure acid leaching: The roasted slag obtained in step (2) is ground, and the resulting roasted slag grinding slurry is subjected to atmospheric pressure sulfuric acid leaching to react with sulfuric acid to obtain sulfuric acid leaching slurry; the number of leaching stages in the atmospheric pressure sulfuric acid leaching operation is 3 to 8, the concentration of leaching slurry in each stage is controlled at 15% to 35%, the sulfuric acid concentration is 98%, the amount of sulfuric acid used is 1.5 to 3 times the dry weight of the material before leaching, the leaching time is 2 to 8 hours, and the stirring speed is 200 rpm to 800 rpm; (4) Separation of sulfuric acid leaching slurry: The sulfuric acid leaching slurry from step (3) is subjected to solid-liquid separation treatment to obtain acid leaching solution and acid leaching residue; (5) Water washing: The acid leaching residue obtained in step (4) is washed with water to obtain acid leaching residue water washing slurry. Then, the acid leaching residue water washing slurry is subjected to solid-liquid separation treatment to obtain washing water and water washing residue. (6) Atmospheric pressure cyanide leaching: The water-washed slag obtained in step (5) is ground to obtain water-washed slag grinding slurry. The water-washed slag grinding slurry is subjected to atmospheric pressure cyanide leaching to leach gold and silver from the water-washed slag grinding slurry to obtain cyanide leaching slurry containing gold and silver. (7) Cyanide leaching slurry separation: The cyanide leaching slurry obtained in step (6) is subjected to solid-liquid separation treatment to obtain cyanide leaching solution and cyanide residue.

2. The efficient recovery method for valuable elements in high-gold, high-silver, and high-zinc concentrates according to claim 1, characterized in that: The gravity separation pretreatment uses a shaking table and a spiral chute. The separation process consists of one roughing and three to six scavenging processes. The roughing concentrate and the concentrate from each scavenging process are combined to form a high gold and zinc concentrate, and the final scavenging tailings are used as a low gold and zinc concentrate.

3. The efficient recovery method for valuable elements in high-gold, high-silver, and high-zinc concentrates according to claim 1, characterized in that: The roasting flue gas in step (2) is used to prepare sulfuric acid for sulfuric acid leaching in step (3).

4. The efficient recovery method for valuable elements in high-gold, high-silver, and high-zinc concentrates according to claim 1, characterized in that: In step (2), the calcination temperature is 600℃~900℃ and the calcination time is 2h~5h.

5. The efficient recovery method for valuable elements in high-gold, high-silver, and high-zinc concentrates according to claim 1, characterized in that: In steps (3) and (6), the slurry concentration for grinding is 30% to 50%, and the grinding time is 5 min to 15 min.

6. The efficient recovery method for valuable elements in high-gold, high-silver, and high-zinc concentrates according to claim 1, characterized in that: The water washing process in step (5) is carried out at room temperature with a slurry concentration of 30% to 50% and a time of 15 to 30 minutes.

7. The efficient recovery method for valuable elements in high-gold, high-silver, and high-zinc concentrates according to claim 1, characterized in that: In step (6), the number of leaching stages in the atmospheric pressure cyanide leaching treatment is 5 to 10, the slurry concentration of each leaching stage is controlled at 10% to 30%, the sodium cyanide consumption is 10 kg / t to 50 kg / t, the leaching time is 12 h to 48 h, the stirring speed is 200 rpm to 800 rpm, air is continuously blown in during the leaching process, and lime is added to control the pH of the leaching process to be no less than 11.

8. The efficient recovery method for valuable elements in high-gold, high-silver, and high-zinc concentrates according to claim 1, characterized in that: In step (1), the zinc concentrate has a gold grade greater than 10 g / t and a silver grade greater than 150 g / t.

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