Method for pressure oxidation leaching and enrichment of gold and silver from zinc sulphide concentrate

Abstract

This invention discloses a method for oxygen pressure leaching and gold and silver enrichment of zinc sulfide concentrate, belonging to the field of hydrometallurgical technology. The method includes the following steps: mixing zinc sulfide concentrate with lignin, adding water to form a slurry, and grinding to obtain a slurry; adding waste electrolyte to the slurry and introducing oxygen for high-temperature oxygen pressure leaching; after leaching, the slurry is flash-cooled and depressurized, followed by solid-liquid separation to obtain oxygen leaching solution and oxygen leaching residue; adjusting the oxygen leaching residue to obtain oxygen leaching residue slurry; flotating the oxygen leaching residue slurry to obtain sulfur concentrate and tailings slag; blending the tailings slag with coal, introducing oxygen-enriched air, and performing oxygen-enriched smelting to obtain smelting slag, gold and silver-rich dust, and flue gas. In this invention, a low liquid-to-solid ratio is used during oxygen pressure leaching, and the final acid concentration of the oxygen leaching solution is controlled at 1-2 g / L, thereby preventing the chlorine in the zinc concentrate from being dissolved by acid, ensuring that most of it remains in the oxygen leaching residue, thus laying the foundation for subsequent oxygen-enriched smelting and chlorine volatilization gold extraction processes.

Description

Technical Field

[0001] This invention belongs to the field of hydrometallurgical technology, specifically relating to a method for oxygen pressure leaching and gold and silver enrichment of zinc sulfide concentrate. Background Technology

[0002] Zinc sulfide concentrate is the core primary resource for zinc extraction. With the continuous growth of global demand for zinc resources and increasingly stringent environmental requirements, traditional pyrometallurgical zinc smelting processes, due to their high energy consumption, high environmental costs, and low utilization rate of valuable elements, are gradually being replaced by hydrometallurgical zinc smelting processes centered on oxygen pressure leaching. Oxygen pressure leaching directly leaches zinc sulfide concentrate under high temperature, high pressure, and oxygen-rich conditions, achieving highly efficient zinc leaching while simultaneously producing elemental sulfur. This avoids sulfur dioxide pollution generated by traditional roasting processes and offers significant advantages such as high zinc leaching rate, thorough sulfur recovery, low energy consumption per unit product, and environmental friendliness.

[0003] The main components of zinc sulfide concentrate are zinc, sulfur, iron and trace amounts of chlorine, and small amounts of metals such as lead, cadmium, copper and silver. Gold is generally found as an independent mineral in sphalerite, pyrite and galena, and sometimes as a native gold mineral in gangue minerals. Because the gold content is very low, most existing hydrometallurgical zinc smelting processes only enrich silver and rarely mention gold enrichment, which not only wastes resources but also affects the economic benefits of zinc smelting enterprises. Summary of the Invention

[0004] Therefore, the purpose of this invention is to provide a method for oxygen pressure leaching and gold and silver enrichment of zinc sulfide concentrate.

[0005] A method for oxygen pressure leaching and gold and silver enrichment of zinc sulfide concentrate includes the following steps:

[0006] S1: After mixing zinc sulfide concentrate with lignin, water is added to adjust the slurry and then the ore is ground to obtain a slurry;

[0007] S2: Add waste electrolyte to the slurry, control the liquid-to-solid ratio to 3-4 mL / g, and introduce oxygen for oxygen pressure leaching. Leach until the H2SO4 concentration in the slurry is 1-2 g / L, then stop leaching. After the slurry is cooled and depressurized by flash evaporation, solid and liquid are separated to obtain oxygen leaching solution and oxygen leaching residue.

[0008] S3: After adjusting the oxygen leaching residue into a slurry, an oxygen leaching residue slurry is obtained; the oxygen leaching residue slurry is subjected to flotation to obtain sulfur concentrate and tailings.

[0009] S4: After the tailings slag is blended with coal, oxygen-enriched air is introduced to carry out oxygen-enriched smelting, resulting in smelting slag, gold and silver rich dust and flue gas.

[0010] Preferably, in step S1, the mass of the lignin is 3-5‰ of the mass of the zinc sulfide concentrate; the concentration of the slurry is 65-75wt%; and the proportion of mineral particles smaller than 40μm after grinding is ≥90%.

[0011] Preferably, in step S2, the concentration of sulfuric acid in the waste electrolyte is 170~180g / L, the purity of oxygen is greater than 99v%, the oxygen pressure leaching temperature is 145~155℃, the oxygen pressure leaching pressure is 1.2~1.3Mpa, and the oxygen pressure leaching time is 2~3h.

[0012] Preferably, in step S2, the oxygen leaching solution is sent to the subsequent production of electrolytic zinc.

[0013] Preferably, in step S3, the flotation includes a roughing stage, a cleaning stage, and a scavenging stage; the cleaned concentrate is sulfur concentrate, and the scavenging tailings are tailings slag.

[0014] Further optimization involves the rough concentrate from the roughing process entering the cleaning process, the tailings from the cleaning process returning to the roughing process, the tailings from the roughing process entering the scavenging process, and the scavenging foam returning to the roughing process.

[0015] For further optimization, the coarse selection time is 10-12 minutes, the fine selection time is 6-8 minutes, and the sweeping selection time is 22-24 minutes.

[0016] Preferably, in step S4, the amount of coal added is 10-12% of the tailings slag mass, the oxygen content in the oxygen-enriched air is 80-90%; the temperature of oxygen-enriched smelting is 1250-1300℃, and the smelting time is 2.5-3h.

[0017] More preferably, the oxygen-enriched smelting process is carried out by continuous feeding and intermittent slag discharge.

[0018] Preferably, the smelting slag is sold externally; the flue gas is used for waste heat utilization, and dust is collected to produce acid; the dust obtained from dust collection is gold and silver rich dust.

[0019] Compared with the prior art, one or more of the above technical solutions can achieve at least one of the following beneficial effects:

[0020] (1) In the method of the present invention, a lower liquid-solid ratio is used in the oxygen pressure leaching process, and the final acid concentration of the oxygen leaching solution is controlled at 1~2g / L (calculated as H2SO4), thereby avoiding the chlorine in the zinc concentrate from being dissolved by acid, so that most of it remains in the oxygen leaching residue, thus laying the foundation for the subsequent oxygen-enriched smelting chlorine volatilization gold extraction process.

[0021] (2) The oxygen pressure leaching method of the present invention, by precisely controlling the reaction conditions, enables the chlorine element in the raw material to be efficiently fixed in the oxygen leaching residue; it can not only effectively reduce the adverse effects of chlorine element on the oxygen pressure leaching process, but also significantly reduce the chlorine content in the leaching solution, thereby avoiding interference with the subsequent electrolytic zinc process; at the same time, it is also conducive to improving the enrichment and recovery efficiency of gold in the subsequent process.

[0022] (3) In the method of the present invention, lignin is mixed with zinc concentrate and then ground so that lignin is uniformly and fully adsorbed and dispersed on the surface of zinc concentrate particles, thereby effectively inhibiting the coating of elemental sulfur on the ZnS surface during the subsequent oxygen pressure leaching process, so as to improve the zinc leaching rate.

[0023] (4) The sulfur content in the tailings obtained by the method of the present invention is 15-20 wt%, which can be used as fuel in the oxygen-enriched smelting process, reducing the amount of coal used and thus reducing the coal consumption in the oxygen-enriched smelting process.

[0024] (5) The present invention adopts a wet pyrometallurgical process that combines oxygen pressure leaching, flotation to recover sulfur and oxygen-enriched smelting of tailings slag to process zinc sulfide concentrate. It can achieve a total zinc recovery rate of over 98%, a lead-zinc content (Pb+Zn) of <5wt% in the smelting slag, a gold grade of 3~5g / t in the produced dust, a gold recovery rate of over 80%, and a silver recovery rate of over 90%. It achieves efficient enrichment of gold and silver while effectively volatilizing and recovering lead and zinc metals, significantly improving economic benefits. Attached Figure Description

[0025] Figure 1 The process flow diagram of this invention. Detailed Implementation

[0026] To facilitate understanding of the present invention, the present invention will be described more fully and in detail below with reference to the accompanying drawings and preferred embodiments, but the scope of protection of the present invention is not limited to the following specific embodiments.

[0027] Unless otherwise defined, all technical terms used herein have the same meaning as commonly understood by those skilled in the art. The technical terms used herein are for the purpose of describing particular embodiments only and are not intended to limit the scope of the invention.

[0028] As mentioned above, a method for oxygen pressure leaching and gold and silver enrichment of zinc sulfide concentrate includes the following steps:

[0029] S1: After mixing zinc sulfide concentrate with lignin, water is added to adjust the slurry and then the ore is ground to obtain a slurry;

[0030] S2: Add waste electrolyte to the slurry, control the liquid-to-solid ratio to 3-4 mL / g (including but not limited to: 3 mL / g, 3.1 mL / g, 3.2 mL / g, 3.3 mL / g, 3.4 mL / g, 3.5 mL / g, 3.6 mL / g, 3.7 mL / g, 3.8 mL / g, 3.9 mL / g, 4 mL / g, etc.), and introduce oxygen for high-temperature oxygen pressure leaching. Leach until the H2SO4 concentration in the slurry is 1-2 g / L (including but not limited to: 1 g / L, 1.1 g / L, 1.2 g / L, 1.3 g / L, 1.4 g / L, 1.5 g / L, 1.6 g / L, 1.7 g / L, 1.8 g / L, 1.9 g / L, 2 g / L, etc.), then stop leaching. After flash evaporation to cool and depressurize, the slurry is separated into solid and liquid components to obtain oxygen leaching solution and oxygen leaching residue.

[0031] S3: After adjusting the oxygen leaching residue into a slurry, an oxygen leaching residue slurry is obtained; the oxygen leaching residue slurry is subjected to flotation to obtain sulfur concentrate and tailings.

[0032] S4: After the tailings slag is blended with coal, oxygen-enriched air is introduced to carry out oxygen-enriched smelting, resulting in smelting slag, gold and silver rich dust and flue gas.

[0033] In the method of this invention, by precisely controlling the reaction conditions, specifically by employing a low liquid-to-solid ratio and controlling the final acid concentration of the oxygen leaching solution at 1-2 g / L (calculated as H2SO4), the chlorine element in the raw material is efficiently retained in the oxygen leaching residue. This not only effectively reduces the adverse effects of chlorine on the oxygen pressure leaching process and significantly reduces the chlorine content in the leaching solution, thereby avoiding interference with the subsequent electrolytic zinc process, but also lays a good foundation for the subsequent oxygen-enriched smelting chlorine volatilization gold extraction process, which is beneficial to improving the enrichment and recovery efficiency of gold.

[0034] In step S4, under high-temperature redox conditions, gold and silver volatilize due to their high vapor pressure at high temperatures. Specifically, chlorine in the raw materials is utilized; the core reaction is chlorination volatilization to extract gold, where gold reacts with chlorine at high temperatures to produce gaseous gold chloride. The chemical reaction formulas involved are as follows:

[0035] 2Au (solid) + 3Cl2 (gaseous) = 2AuCl3 (gaseous)

[0036] Au₂O₃ + 3CO = 2Au + 3CO₂

[0037] Ag₂O + CO = 2Ag + CO₂

[0038] Ag₂S + O₂ = 2Ag + SO₂.

[0039] In some embodiments, in step S1, the mass of the lignin is 3-5‰ of the mass of the zinc sulfide concentrate, including but not limited to: 3‰, 3.5‰, 4‰, 4.5‰, 5‰, etc.; the concentration of the slurry is 65-75wt%, including but not limited to: 65wt%, 66wt%, 67wt%, 68wt%, 69wt%, 70wt%, 71wt%, 72wt%, 73wt%, 74wt%, 75wt%, etc.; and the proportion of mineral particles smaller than 40μm after grinding is ≥90%.

[0040] In some embodiments, in step S2, the concentration of sulfuric acid in the waste electrolyte is 170~180 g / L, including but not limited to: 170 g / L, 171 g / L, 172 g / L, 173 g / L, 174 g / L, 175 g / L, 176 g / L, 177 g / L, 178 g / L, 179 g / L, 180 g / L, etc.; the purity of oxygen is greater than 99 v%, and the oxygen pressure leaching temperature is 145~15 The temperature range is 5℃, including but not limited to: 145℃, 146℃, 147℃, 148℃, 149℃, 150℃, 151℃, 152℃, 153℃, 154℃, 155℃, etc.; the oxygen pressure leaching pressure is 1.2~1.3Mpa, including but not limited to: 1.2Mpa, 1.25Mpa, 1.3Mpa, etc.; the oxygen pressure leaching time is 2~3h, including but not limited to: 2h, 2.5h, 3h, etc.

[0041] In some embodiments, in step S2, the oxygen leaching solution is sent to the subsequent production of electrolytic zinc.

[0042] In some embodiments, in step S3, the flotation includes a roughing stage, a cleaning stage, and a scavenging stage; the cleaned concentrate is sulfur concentrate, and the scavenging tailings are tailings slag.

[0043] In some embodiments, the rough concentrate after roughing enters the cleaning process, the tailings after cleaning are returned to the roughing process, the tailings after roughing enter the scavenging process, and the scavenging foam is returned to the roughing process.

[0044] In some embodiments, the coarse selection time is 10-12 minutes, including but not limited to 10 minutes, 11 minutes, 12 minutes, etc.; the fine selection time is 6-8 minutes, including but not limited to 6 minutes, 11 minutes, 12 minutes, etc.; and the sweeping selection time is 22-24 minutes, including but not limited to 22 minutes, 23 minutes, 24 minutes, etc.

[0045] In some embodiments, in step S4, the amount of coal added is 10-12% of the tailings slag mass, including but not limited to: 10%, 11%, 12%, etc.; the oxygen content in the oxygen-enriched air is 80-90%; the oxygen-enriched smelting temperature is 1250-1300℃, including but not limited to: 1250℃, 1260℃, 1270℃, 1280℃, 1290℃, 1300℃, etc.; the smelting time is 2.5-3h, including but not limited to: 2.5h, 2.6h, 2.7h, 2.8h, 2.9h, 3h, etc.

[0046] In some embodiments, the oxygen-enriched smelting process is carried out by continuous feeding and intermittent slag discharge.

[0047] In some embodiments, the smelting slag is sold externally; the flue gas is used for waste heat recovery and dust collection to produce acid; the dust obtained from dust collection is gold and silver rich dust.

[0048] The process flow diagram of this invention is shown below. Figure 1 The specific steps can be found in the embodiments.

[0049] In this embodiment of the invention, the zinc sulfide concentrate composition is as follows: Zn: 48~53wt%, Pb: 1~2wt%, Fe: 6~11wt%, S: 28~32wt%, Ag: 70~80g / t, Au trace amount, Cl: 0.03~0.05wt%.

[0050] Example 1

[0051] S1: Add lignin (the amount of lignin added is 4‰ of the mass of zinc concentrate) to zinc concentrate (containing Zn: 51wt%, Pb: 1.6wt%, Fe: 7wt%, S: 31wt%, Cl: 0.03%, Ag: 80g / t, trace amount of Au), add water to adjust the slurry and grind it, control the solids concentration of the slurry to 75wt%, and the proportion of particles smaller than 40μm to more than 90% to obtain the slurry.

[0052] S2: The slurry is added to a high-pressure reactor, along with waste electrolyte (with an H2SO4 concentration of 170 g / L) and oxygen (99 v%) for oxygen pressure leaching. During oxygen pressure leaching, the liquid-to-solid ratio is controlled at 4 mL / g, the leaching temperature at 155℃, the pressure at 1.3 MPa, and the time at 2.5 h. The final acid concentration of the slurry is 2 g / L (based on H2SO4). After the reaction is complete, the slurry is flash-evaporated to cool and depressurize, and the solid and liquid are separated to obtain oxygen leaching solution and oxygen leaching residue. The oxygen leaching solution is sent to the subsequent production of electrolytic zinc.

[0053] S3: The oxygen leaching residue is added to the flotation cell for roughing, cleaning and scavenging flotation (the concentrate from the roughing process goes to the cleaning process, the tailings from the cleaning process are returned to the roughing process, the tailings from the roughing process go to the scavenging process, and the froth from the scavenging process is returned to the roughing process). The roughing time is 12 minutes, the cleaning time is 8 minutes, and the scavenging time is 24 minutes. The sulfur concentrate produced by the cleaning process is sent to the downstream process to produce sulfur, and the tailings produced by the scavenging process contain S: 18wt%, Ag: 286g / t, Au: 0.4g / t, and water content of 20wt%).

[0054] S4: Tailings slag is added to the smelting furnace at a rate of 2.8 t / h and coal at a rate of 0.28 t / h. Oxygen-enriched air (O2 content in the oxygen-enriched air is 90%) is introduced for smelting. The smelting temperature is controlled at 1250℃ for 2.5 h. The smelting slag is discharged at a rate of 2.3 t / h using a continuous feeding and intermittent slag discharge method (the smelting slag contains 1.7 wt% Pb and 2.5 wt% Zn). The flue gas generated during the smelting process is used for waste heat utilization, dust collection and acid production. The dust collected is gold and silver rich dust (the gold and silver rich dust contains 4.2 g / t Au and 3312 g / t Ag). The production rate of gold and silver rich dust is 0.22 t / h.

[0055] In summary, the overall zinc recovery rate in this embodiment is 98.9%, the gold recovery rate is 82.5%, and the silver recovery rate is 91%. The economic benefit generated by the recovered gold is 0.22t / h × 4.2g / t × 330 days × 24h × 1000 yuan / g × 0.55 ÷ 10000 = 4.025 million yuan / year. Specifically: for gold content of 3-4g / t, a discount of 0.50 is required based on the cost of recovery; for gold content of 4-5g / t, a discount of 0.55 is also required based on the cost of recovery.

[0056] Comparative Example 1

[0057] The process is basically the same as in Example 1, except that in step S2, the liquid-to-solid ratio is controlled at 5 mL / g, which corresponds to a final acid concentration of 3.1 g / L (calculated as H2SO4) after leaching.

[0058] In summary, the total recovery rate of zinc in this comparative example is 98.8%, the recovery rate of gold is 78%, and the recovery rate of silver is 90.8%. The economic benefits generated by the recovered gold each year are 0.22t / h × 3.97g / t × 330 days × 24h × 1000 yuan / g × 0.50 ÷ 10000 = 3.456 million yuan / year. Compared with Example 1, the economic benefits are reduced by 14.1%.

[0059] Comparing Example 1 and Comparative Example 1, the gold recovery rate in Example 1 is significantly improved compared to Comparative Example 1. This may be because the liquid-solid ratio was controlled and the concentration of the final acid was adjusted, which allowed more chlorine to remain in the leaching residue, thereby improving the gold recovery rate.

[0060] Comparative Example 2

[0061] The process is basically the same as in Example 1, except that in step S2, the liquid-to-solid ratio is controlled at 2 mL / g, which corresponds to a final acid concentration of 0.88 g / L (calculated as H2SO4) after leaching.

[0062] In summary, the total recovery rate of zinc in this embodiment is 98.6%, the recovery rate of gold is 78.5%, and the recovery rate of silver is 90.7%. The economic benefit generated by the recovered gold each year is 0.22t / h × 3.996g / t × 330 days × 24h × 1000 yuan / g × 0.50 ÷ 10000 = 3.481 million yuan / year. Compared with Example 1, the economic benefit is reduced by 13.5%.

[0063] Comparing Example 1 and Comparative Example 2, the gold recovery rate in Example 1 is significantly improved compared to Comparative Example 2. This may be because the final acid concentration is too low, resulting in a reduction in the gold content in the leaching residue, which in turn affects the gold recovery rate.

[0064] Comparative Example 3

[0065] This method is basically the same as Example 1, except that the lignin in step S1 is added in step S2. The specific steps of the method are as follows:

[0066] S1: Add water to zinc concentrate (containing Zn: 51wt%, Pb: 1.6wt%, Fe: 7wt%, S: 31wt%, Cl: 0.03wt%, Ag: 80g / t, trace amount of Au) to form a slurry and grind it. Control the solids concentration of the slurry to 75wt% and the particle size of less than 40μm to account for more than 90% to obtain the slurry.

[0067] S2: The slurry is added to a high-pressure reactor, along with waste electrolyte (with an H2SO4 concentration of 170 g / L), lignin (at a rate of 4‰ of the zinc concentrate mass), and oxygen (99 v%) for oxygen pressure leaching. During oxygen pressure leaching, the liquid-to-solid ratio is controlled at 4 mL / g, the leaching temperature at 155℃, the pressure at 1.3 MPa, and the time at 2.5 h. The final acid concentration of the slurry is 2 g / L (based on H2SO4). After the reaction is complete, the slurry is flash-evaporated to cool and depressurize, and the solid and liquid are separated to obtain oxygen leaching solution and oxygen leaching residue. The oxygen leaching solution is sent to subsequent electrolytic zinc production.

[0068] S3: Same as Example 1.

[0069] S4: Same as Example 1.

[0070] In summary, the total recovery rate of zinc in this comparative example is 97.2%, the recovery rate of gold is 81.8%, and the recovery rate of silver is 90.8%. The economic benefits generated by the recovered gold each year are 0.22t / h × 4.164g / t × 330 days × 24h × 1000 yuan / g × 0.55 ÷ 10000 = 3.99 million yuan / year. However, the recovery rate of the main metal zinc is 1.7% lower than that of Example 1, resulting in an economic loss of 13.464 million yuan / year.

[0071] Comparing Example 1 and Comparative Example 3, adding lignin during the grinding stage can significantly improve the zinc recovery rate. This may be because adding lignin during the grinding stage enhances the adsorption of lignin to the slurry particles, which can better prevent sulfur from adsorbing onto zinc sulfide, thereby further improving the zinc recovery rate.

[0072] Example 2

[0073] S1: Add lignin (the amount of lignin added is 3‰ of the mass of zinc concentrate) to zinc concentrate (containing Zn: 50wt%, Pb: 1.7wt%, Fe: 8wt%, S: 30wt%, Cl: 0.04wt%, Ag: 75g / t, trace amount of Au), add water to adjust the slurry and grind it, control the solids concentration of the slurry to 70wt%, and the proportion of particles smaller than 40μm to more than 90% to obtain the slurry.

[0074] S2: The slurry is added to a high-pressure reactor, along with waste electrolyte (with an H2SO4 concentration of 175 g / L) and oxygen (99 v%) for oxygen pressure leaching. During oxygen pressure leaching, the liquid-to-solid ratio is controlled at 3.5 mL / g, the leaching temperature at 150℃, the pressure at 1.2 MPa, and the time at 3 hours. The final acid concentration of the slurry is 1.0 g / L (based on H2SO4). After the reaction is complete, the slurry is flash-evaporated to cool and depressurize, and the solid and liquid are separated to obtain oxygen leaching solution and oxygen leaching residue. The oxygen leaching solution is sent to the subsequent production of electrolytic zinc.

[0075] S3: The oxygen leaching residue is added to the flotation cell for roughing, cleaning and scavenging flotation (the concentrate from the roughing flotation enters the cleaning flotation, the tailings from the cleaning flotation return to the roughing flotation, the tailings from the roughing flotation enter the scavenging flotation, and the froth from the scavenging flotation returns to the roughing flotation). The roughing flotation time is 11 min, the cleaning flotation time is 7 min, and the scavenging flotation time is 23 min. The sulfur concentrate produced by the cleaning flotation is sent to the downstream process to produce sulfur, and the tailings produced by the scavenging flotation (containing S: 17wt%, Ag: 278g / t, Au: 0.35g / t, and water content 19wt%).

[0076] S4: Tailings slag is added to the smelting furnace at a rate of 2.7 t / h and coal at a rate of 0.3 t / h. Oxygen-enriched air (O2 content in the oxygen-enriched air is 85%) is introduced for smelting. The smelting temperature is controlled at 1280℃ for 2.5 hours. The smelting slag is discharged at a rate of 2.3 t / h using a continuous feeding and intermittent slag discharge method (the smelting slag contains 1.8 wt% Pb and 2.6 wt% Zn). The flue gas generated during the smelting process is used for waste heat utilization, dust collection and acid production. The dust collected is gold and silver rich dust (the gold and silver rich dust contains 3.7 g / t Au and 3245 g / t Ag). The production rate of gold and silver rich dust is 0.21 t / h.

[0077] Based on comprehensive calculations, the total recovery rate of zinc in this embodiment is 98.8%, the recovery rate of gold is 81.6%, and the recovery rate of silver is 90.8%. The economic benefits generated by the recovered gold each year are 0.21t / h × 3.7g / t × 330 days × 24h × 1000 yuan / g × 0.5 ÷ 10000 = 3.077 million yuan / year.

[0078] Example 3

[0079] S1: Add lignin (the amount of lignin added is 5‰ of the mass of zinc concentrate) to zinc concentrate (containing Zn: 49wt%, Pb: 1.8wt%, Fe: 9wt%, S: 29wt%, Cl: 0.05wt%, Ag: 70g / t, trace amount of Au), add water to adjust the slurry and grind it, control the solids concentration of the slurry to 65wt%, and the proportion of particles smaller than 40μm to more than 90% to obtain the slurry.

[0080] S2: The slurry is added to a high-pressure reactor, along with waste electrolyte (with an H2SO4 concentration of 180 g / L) and oxygen (99 v%) for oxygen pressure leaching. During oxygen pressure leaching, the liquid-to-solid ratio is controlled at 3 mL / g, the leaching temperature at 145℃, the pressure at 1.2 MPa, and the time at 2 hours. The final acid concentration of the slurry is 1.5 g / L (based on H2SO4). After the reaction is complete, the slurry is flash-evaporated to cool and depressurize, and the solid and liquid are separated to obtain oxygen leaching solution and oxygen leaching residue. The oxygen leaching solution is sent to the subsequent production of electrolytic zinc.

[0081] S3: The oxygen leaching residue is added to the flotation cell for roughing, cleaning and scavenging flotation (the roughing concentrate goes to the cleaning, the cleaning tailings are returned to the roughing, the roughing tailings go to the scavenging, and the scavenging froth goes to the roughing). The roughing time is 10 minutes, the cleaning time is 6 minutes, and the scavenging time is 22 minutes. The roughing concentrate goes to the cleaning, the cleaning tailings go to the scavenging, and the scavenging concentrate goes to the roughing. The sulfur concentrate produced by the cleaning is sent to the downstream process to produce sulfur, and the tailings produced by the scavenging (containing S: 16wt%, Ag: 269g / t, Au: 0.3g / t, and water content 18wt%).

[0082] S4: Tailings slag is added to the smelting furnace at a rate of 2.6 t / h and coal at a rate of 0.31 t / h. Oxygen-enriched air (O2 content in the oxygen-enriched air is 85%) is introduced for smelting. The smelting temperature is controlled at 1300℃ for 3 hours. The smelting slag is discharged at a rate of 2.2 t / h using a continuous feeding and intermittent slag discharge method (the smelting slag contains 1.8 wt% Pb and 2.7 wt% Zn). The flue gas generated during the smelting process is used for waste heat utilization, dust collection and acid production. The dust collected is gold and silver rich dust (the gold and silver rich dust contains 3.2 g / t Au and 3161 g / t Ag). The production rate of gold and silver rich dust is 0.21 t / h.

[0083] Based on comprehensive calculations, the total recovery rate of zinc in this embodiment is 98.7%, the recovery rate of gold is 81.3%, and the recovery rate of silver is 90.4%. The economic benefits generated by the recovered gold each year are 0.21t / h × 3.2g / t × 330 days × 24h × 1000 yuan / g × 0.5 ÷ 10000 = 2.661 million yuan / year.

[0084] The above embodiments are only for illustrating the technical concept and features of the present invention, and are intended to enable those skilled in the art to understand the content of the present invention and implement it accordingly. They should not be construed as limiting the scope of protection of the present invention. All equivalent changes or modifications made in accordance with the spirit and essence of the present invention should be covered within the scope of protection of the present invention.

Claims

1. A method for oxygen pressure leaching and gold and silver enrichment from zinc sulfide concentrate, characterized in that, Includes the following steps: S1: After mixing zinc sulfide concentrate with lignin, water is added to adjust the slurry and then the ore is ground to obtain a slurry; S2: Add waste electrolyte to the slurry, control the liquid-to-solid ratio to 3-4 mL / g, and introduce oxygen for oxygen pressure leaching. Leach until the H2SO4 concentration in the slurry is 1-2 g / L, then stop leaching. After the slurry is cooled and depressurized by flash evaporation, solid and liquid are separated to obtain oxygen leaching solution and oxygen leaching residue. S3: After adjusting the oxygen leaching residue into a slurry, an oxygen leaching residue slurry is obtained; the oxygen leaching residue slurry is subjected to flotation to obtain sulfur concentrate and tailings. S4: After the tailings slag is blended with coal, oxygen-enriched air is introduced to carry out oxygen-enriched smelting, resulting in smelting slag, gold and silver rich dust and flue gas.

2. The method for oxygen pressure leaching and gold and silver enrichment of zinc sulfide concentrate according to claim 1, characterized in that, In step S1, the mass of the lignin is 3-5‰ of the mass of the zinc sulfide concentrate; the concentration of the slurry is 65-75wt%; and the proportion of mineral particles smaller than 40μm after grinding is ≥90%.

3. The method for oxygen pressure leaching and gold and silver enrichment of zinc sulfide concentrate according to claim 1, characterized in that, In step S2, the concentration of sulfuric acid in the waste electrolyte is 170~180g / L, the purity of oxygen is not less than 99v%, the oxygen pressure leaching temperature is 145~155℃, the oxygen pressure leaching pressure is 1.2~1.3MPa, and the oxygen pressure leaching time is 2~3h.

4. The method for oxygen pressure leaching and gold and silver enrichment of zinc sulfide concentrate according to claim 1, characterized in that, In step S2, the oxygen leaching solution is sent to the subsequent production of electrolytic zinc.

5. The method for oxygen pressure leaching and gold and silver enrichment of zinc sulfide concentrate according to claim 1, characterized in that, In step S3, the flotation includes a roughing stage, a cleaning stage, and a scavenging stage; the cleaned concentrate is sulfur concentrate, and the scavenging tailings are tailings slag.

6. The method for oxygen pressure leaching and gold and silver enrichment of zinc sulfide concentrate according to claim 5, characterized in that, The rough concentrate after roughing enters the cleaning process, the tailings after cleaning are returned to the roughing process, the tailings after roughing enter the scavenging process, and the scavenging foam is returned to the roughing process.

7. The method for oxygen pressure leaching and gold and silver enrichment of zinc sulfide concentrate according to claim 5 or 6, characterized in that, The coarse selection takes 10-12 minutes, the fine selection takes 6-8 minutes, and the sweeping selection takes 22-24 minutes.

8. The method for oxygen pressure leaching and gold and silver enrichment of zinc sulfide concentrate according to claim 1, characterized in that, In step S4, the amount of coal added is 10-12% of the tailings slag mass, the oxygen content in the oxygen-enriched air is 80-90%; the temperature of oxygen-enriched smelting is 1250-1300℃, and the smelting time is 2.5-3h.

9. The method for oxygen pressure leaching and gold and silver enrichment of zinc sulfide concentrate according to claim 1, characterized in that, In step S4, the oxygen-enriched smelting process is carried out by continuous feeding and intermittent slag discharge.

10. The method for oxygen pressure leaching and gold and silver enrichment of zinc sulfide concentrate according to claim 1, characterized in that, In step S4, the smelting slag is sold externally; the flue gas is used for dust collection and acid production after waste heat utilization; the flue gas obtained from dust collection is gold and silver rich flue gas.

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