A method for synergistically extracting gold from complex gold concentrate by fire smelting-acid leaching-cyanidation

Abstract

The application belongs to the technical field of non-ferrous metal smelting and extraction, and particularly relates to a complex gold concentrate fire smelting-acid leaching-cyanidation synergistic gold extraction method, which comprises the following steps: (1) producing copper matte through complex gold concentrate fire smelting; (2) copper matte sulfuric acid roasting; (3) sulfuric acid roasting slag acid leaching; (4) returning the filter cake of the replacement and pressure filtration of the iron-containing leaching solution after copper removal to fire smelting; (5) cyanidation of the magnetic separation concentrate and acid leaching slag to extract gold; and (6) cyanide liquid gold extraction and mother liquor regeneration. The "smelting-acid leaching-cyanidation" process of the application can recover copper in the slag, form a closed loop with the smelting and gold extraction of low-grade gold concentrate, solve the pollution problem of acid leaching liquid recycling, and realize efficient recovery of copper and gold and resource recycling.

Description

Technical Field

[0001] This invention belongs to the field of non-ferrous metal smelting and extraction technology, specifically relating to a method for synergistic gold extraction from complex gold concentrates through pyrometallurgical smelting, acid leaching, and cyanidation. Background Technology

[0002] For complex gold concentrates (containing 10-60 g / t of gold, 20-200 g / t of silver, 2.5-3.5% of copper, 5-10% of arsenic, and 1-5% of carbon), the recovery rates of gold, silver, and copper are all low when using the traditional two-stage roasting cyanidation process, resulting in poor economic benefits.

[0003] The following drawbacks exist when using traditional pyrometallurgical processes: Low resource utilization: In direct pyrometallurgical smelting, a copper grade of less than 5% will result in a low matte yield (<15%), with a large amount of copper and gold remaining in the slag, and the overall recovery rate is less than 60%. High energy consumption costs: Low-grade concentrate requires additional fuel to maintain the smelting temperature, and the energy consumption per ton of ore is more than 30% higher than that of high-grade concentrate. Gold recovery pathway is limited: In traditional processes, gold is only recovered from the anode mud after copper matte blowing, and the gold resources in smelting by-products (such as slag and roasting slag) cannot be utilized. Significant environmental pressure: The smelting of low-grade ore produces high levels of SO2 emissions, and the residual copper and gold elements in the slag are difficult to treat harmlessly.

[0004] Therefore, there is an urgent need to develop a co-extraction process for low-grade gold concentrate to achieve efficient recovery of copper and gold and resource recycling. Summary of the Invention

[0005] This invention addresses the problems of low resource utilization, high energy consumption, and insufficient gold recovery rate in the pyrometallurgical smelting of complex gold concentrates mentioned above. It provides a synergistic gold extraction method for pyrometallurgical smelting, acid leaching, and cyanidation of complex gold concentrates. This method achieves efficient recovery of copper and gold and recycling of materials through a synergistic process of pyrometallurgical smelting to produce copper matte, sulfonation roasting of copper matte, acid leaching of sulfonation roasting residue, replacement of iron-containing leachate after copper removal, filter cake recycling (returning to pyrometallurgical smelting), cyanidation extraction of magnetic concentrate and acid leaching residue, and gold extraction from cyanide solution and mother liquor regeneration.

[0006] The specific technical solution is as follows: This invention provides a method for the synergistic gold extraction of complex gold concentrates through pyrometallurgical smelting, acid leaching, and cyanidation, comprising the following steps: (1) Pyrometallurgical smelting of complex gold concentrates to produce matte: Gold concentrate containing 10-60 g / t of gold, 20-200 g / t of silver, 2.5-3.5 wt% of copper, 0.3-1.0 wt% of lead, 5-10 wt% of arsenic, and 1-5 wt% of carbon is ground, mixed with flux and fuel, and fed into an oxygen-enriched side-blown smelting furnace. Oxygen-enriched air is introduced for smelting to produce matte containing 10-15 wt% of copper and 30-300 g / t of gold (relative to the gold concentrate). At the same time, slag containing ≤0.2 wt% of copper and ≤0.5 g / t of gold is discharged and is recorded as “initial slag”. (2) Sulphation roasting of copper matte: The matte produced in step (1) is crushed and sent to a fluidized bed roasting furnace. The air volume and furnace bottom pressure are controlled to roast the matte so that the cuprous sulfide Cu2S in the matte is converted into soluble copper sulfate CuSO4, and the ferrous sulfide FeS is converted into iron oxide Fe2O3. The product after roasting is called "sulfated roasting residue". (3) Acid leaching of sulfation roasted residue: Dilute sulfuric acid was added to the sulfation roasting residue in step (2), the liquid-solid ratio was controlled, and acid leaching and stirring were carried out under heating conditions. The copper acid leaching solution 1 containing 50~100g / L of copper was obtained by filtration through a blower filter press. Part of it was returned to step (1) for ore blending, and part of it was sent to the extraction electrodeposition section to produce cathode copper. At the same time, filter cake from the filter press was obtained. The filter cake was acid-leached and stirred with dilute sulfuric acid under heating conditions. After filtration, a copper acid leaching solution 2 containing 10~20 g / L of copper and 50~100 g / L of iron was obtained. The copper acid leaching solution 2 was subjected to sodium sulfide precipitation to produce a copper precipitate containing 50~70 wt% copper, which was returned to step (1) for ore blending. The iron-containing leaching solution after copper removal entered the next step. The acid leaching residue contained 100~2000 g / t of gold and ≤1 wt% of copper. (4) Replacement-pressing filter cake of iron-containing leachate after copper removal is returned to pyrometallurgical smelting: The iron-containing leaching solution obtained after copper removal in step (3) is added with reducing iron powder to carry out a displacement reaction. After displacement, it is filtered by pressure to obtain a pure ferrous sulfate solution. Ferrous sulfate is produced by evaporation, concentration and crystallization. After displacement, the filter cake is filtered by pressure and returned to the pyrometallurgical furnace in step (1) to replace part of the flux and realize the recycling of copper elements. The initial slag obtained in step (1) is enriched by magnetic separation to obtain magnetic concentrate. The magnetic tailings are used as building material raw materials. (5) Gold extraction by cyanidation of magnetic concentrate and acid leaching residue: The magnetic concentrate obtained in step (4) and the acid leaching residue obtained in step (3) are ground together. The liquid-solid ratio is controlled, sodium hydroxide is added to adjust the pH, ammonium bicarbonate and sodium cyanide are added, and the mixture is stirred and leached under heating conditions. The mixture is filtered to obtain a cyanide leaching solution containing 30~180mg / L of gold and 60~540mg / L of silver, which is referred to as "cyanide solution". A cyanide residue containing ≤5g / t of gold and 5~20wt% of lead is obtained, which is used as raw material for pyrometallurgical lead smelting to comprehensively recover gold and lead. (6) Gold extraction from cyanide solution and mother liquor regeneration: Zinc powder is added to the cyanide solution in step (5) for displacement, and precious gold and silver materials containing 30-50 wt% gold and 30-40 wt% silver are obtained by filter press; gold ingots and silver ingots are produced by wet separation and pyrometallurgical smelting, and the displacement solution is returned to step (5) for recycling.

[0007] The "smelting-acid leaching-cyanidation" process of this invention can not only recover copper from slag, but also form a closed loop with the smelting and gold extraction of low-grade gold concentrate. It also solves the pollution problem of recycling acid leaching solution and realizes efficient recovery and resource recycling of copper and gold.

[0008] Furthermore, in step (1), the flux is quartz, and the amount used is 15-20% of the mass of the gold concentrate; the fuel is pulverized coal, and the amount used is 8-12% of the mass of the gold concentrate.

[0009] Furthermore, in step (1), the gold concentrate is ground to a content of ≥80wt% at -200 mesh, and the volume concentration of the oxygen-enriched air is 30~40%.

[0010] Furthermore, in step (1), the melting temperature is 1200~1300℃ and the melting time is 4~6h.

[0011] Furthermore, in step (2), the copper matte is crushed to a -200 mesh with a content of ≥80wt%.

[0012] Furthermore, in step (2), the air volume is controlled at 12000~15000 m³ / h. 3 Firing at 8000~10000Pa at a furnace bottom pressure of 550~630℃ for 2~3 hours.

[0013] Further, in step (3), 1-2% dilute sulfuric acid is added to the sulfation roasted residue from step (2), the liquid-solid ratio is controlled to be 1-2, and acid leaching and stirring are carried out at 85-95℃ for 2-4 hours.

[0014] Furthermore, in step (3), the filter cake is treated with dilute sulfuric acid with a mass fraction of 15-20%, and the mixture is acid-soaked and stirred at 85-95°C for 4-6 hours. The amount of dilute sulfuric acid used is 2-4 times the mass of the filter cake.

[0015] Furthermore, in step (4), 3~5 kg / m³ of reduced iron powder is added. 3 The displacement reaction takes 2-4 hours.

[0016] Further, in step (5), the ore is ground to a -400 mesh content ≥95wt%, the liquid-solid ratio is controlled at 3~5, the pH is adjusted to 9~10 with sodium hydroxide, the amount of ammonium bicarbonate added is 3~5kg / t, the mass concentration of sodium cyanide is 1~2%, and the leaching is carried out by stirring at 35~45℃ for 36~48h.

[0017] Furthermore, in step (6), the amount of zinc powder used is 1.2 to 1.5 times the mass of gold and silver in the cyanide solution, and the replacement time is 2 to 3 hours.

[0018] Compared with the prior art, the beneficial effects of the present invention are as follows: (1) The gold extraction method of the present invention has a high resource recovery rate: through the closed-loop process of "smelting-acid leaching-cyanidation", the comprehensive copper recovery rate is ≥98% and the comprehensive gold recovery rate is ≥98%, which is 15~20% higher than the traditional pyrometallurgical process; (2) The gold extraction method of the present invention has low energy consumption and cost: after copper removal, the iron-containing leachate is replaced and the filter cake is returned to the smelting to replace the copper-containing material, reducing the energy consumption per ton of gold by 25-30%; (3) The gold extraction method of the present invention has significant environmental benefits: the cyanide mother liquor is regenerated and recycled, and there is no wastewater discharge; the primary furnace slag (magnetic separation concentrate) and acid leaching slag are mixed for cyanidation, realizing the resource utilization of smelting by-products; (4) The gold extraction method of the present invention has strong process adaptability: it can process complex gold concentrates of various grades, has high tolerance to raw material fluctuations, and has strong adaptability. Detailed Implementation

[0019] The principles and features of the present invention are described below with reference to examples. The examples are only used to explain the present invention and are not intended to limit the scope of the present invention.

[0020] Example 1 A method for synergistic gold extraction from complex gold concentrates via pyrometallurgical smelting, acid leaching, and cyanidation includes the following steps: (1) Pyrometallurgical smelting of complex gold concentrates to produce matte: Gold concentrate containing 10 g / t of gold, 20 g / t of silver, 2.5 wt% of copper, 0.3 wt% of lead, 5 wt% of arsenic, and 1 wt% of carbon is ground to a -200 mesh with a content of 80 wt%. It is then mixed with quartz (flux, amounted to 15% of the gold concentrate mass) and pulverized coal (fuel, amounted to 8% of the gold concentrate mass) and fed into an oxygen-enriched side-blown smelting furnace. Oxygen-enriched air with a volume concentration of 30% is introduced and smelted at 1200℃ for 4 hours. The product is matte containing 10 wt% copper and 30 g / t of gold (relative to the gold concentrate). At the same time, slag containing 0.2 wt% copper and 0.5 g / t of gold is discharged and recorded as "initial slag". (2) Sulphation roasting of copper matte: The copper matte produced in step (1) is crushed and ground to -200 mesh with a content of 80wt%, and then fed into a fluidized bed roaster with a controlled air volume of 12000 m³ / h. 3 At a furnace bottom pressure of 8000 Pa and a roasting temperature of 550℃ for 2 hours, cuprous sulfide (Cu2S) in matte is converted into soluble copper sulfate (CuSO4), while ferrous sulfide (FeS) is converted into iron oxide (Fe2O3). The product after roasting is recorded as "sulfation roasting residue". (3) Acid leaching of sulfation roasted residue: Add 1% dilute sulfuric acid to the sulfation roasting residue in step (2), control the liquid-solid ratio to 1, and acid leaching and stirring at 85°C for 2 hours. After filtration by a blower filter press, copper acid leaching solution 1 containing 50g / L of copper is obtained. Part of it is returned to step (1) for ore blending, and part of it is sent to the extraction and electrowinning section to produce cathode copper. At the same time, filter cake from the filter press is obtained. The filter cake was acid-leached and stirred at 85℃ for 4 hours with 15% dilute sulfuric acid (the amount of dilute sulfuric acid was twice the mass of the filter cake). After filtration, a copper acid leaching solution 2 containing 10 g / L copper and 50 g / L iron was obtained. The copper acid leaching solution 2 was subjected to sodium sulfide precipitation to produce a copper precipitate containing 50 wt% copper, which was returned to step (1) for ore blending. The iron-containing leaching solution after copper removal entered the next step. The acid leaching residue contained 100 g / t gold and 1 wt% copper. (4) Replacement-pressing filter cake of iron-containing leachate after copper removal is returned to pyrometallurgical smelting: Add 3 kg / m³ of reduced iron powder to the copper-free iron-containing leachate obtained in step (3). 3 The displacement reaction lasted for 2 hours. After displacement, the solution was filtered to obtain a pure ferrous sulfate solution. The solution was then evaporated, concentrated, and crystallized to produce ferrous sulfate. After displacement, the filter cake was filtered and returned to the pyrometallurgical furnace in step (1) to replace part of the quartz stone and realize the recycling of copper. The initial slag obtained in step (1) was enriched by magnetic separation to obtain magnetic concentrate. The magnetic tailings were used as building material raw materials. (5) Gold extraction by cyanidation of magnetic concentrate and acid leaching residue: The magnetic concentrate obtained in step (4) and the acid leaching residue obtained in step (3) are ground to -400 mesh with a content of 95wt%. The liquid-solid ratio is controlled at 3. Sodium hydroxide is added to adjust the pH to 9. The amount of ammonium bicarbonate added is 3kg / t. The mass concentration of sodium cyanide is 1%. The mixture is stirred and leached at 35℃ for 36h. After filtration, a cyanide leaching solution containing 30mg / L of gold and 60mg / L of silver is obtained, which is referred to as "cyanide solution". A cyanide residue containing 5g / t of gold and 5wt% of lead is obtained. This residue is used as raw material for pyrometallurgical lead smelting to comprehensively recover gold and lead. (6) Gold extraction from cyanide solution and mother liquor regeneration: Add zinc powder to the cyanide solution in step (5). The amount of zinc powder is 1.2 times the mass of gold and silver in the cyanide solution. After 2 hours of replacement, the precious gold and silver material containing 30wt% gold and 30wt% silver is obtained by filter press. Gold ingots and silver ingots are produced by wet separation and pyrometallurgical smelting. The replacement liquid is returned to step (5) for recycling.

[0021] Example 2 A method for synergistic gold extraction from complex gold concentrates via pyrometallurgical smelting, acid leaching, and cyanidation includes the following steps: (1) Pyrometallurgical smelting of complex gold concentrates to produce matte: Gold concentrate containing 35 g / t of gold, 110 g / t of silver, 3.0 wt% of copper, 0.65 wt% of lead, 7.5 wt% of arsenic, and 3 wt% of carbon is ground to -200 mesh with a content of 85 wt%. It is then mixed with quartz (flux, amounted to 17.5% of the gold concentrate mass) and pulverized coal (fuel, amounted to 10% of the gold concentrate mass) and fed into an oxygen-enriched side-blown smelting furnace. Oxygen-enriched air with a volume concentration of 35% is introduced and smelted at 1250℃ for 5 hours. The product is matte containing 12.5 wt% copper and 165 g / t of gold (relative to the gold concentrate). At the same time, slag containing 0.19 wt% copper and 0.4 g / t of gold is discharged and recorded as "initial slag". (2) Sulphation roasting of copper matte: The copper matte produced in step (1) is crushed and ground to -200 mesh with a content of 85wt%, and then fed into a fluidized bed roaster with a controlled air volume of 13500 m³ / h. 3 At a furnace bottom pressure of 9000 Pa and a roasting temperature of 595℃ for 2.5 hours, cuprous sulfide (Cu2S) in matte is converted into soluble copper sulfate (CuSO4), while ferrous sulfide (FeS) is converted into iron oxide (Fe2O3). The product after roasting is recorded as "sulfation roasting residue". (3) Acid leaching of sulfation roasted residue: Add 1.5% dilute sulfuric acid to the sulfation roasting residue in step (2), control the liquid-solid ratio to 1.5, and acid leaching and stirring at 90°C for 3 hours. After filtration by a blower filter press, copper acid leaching solution 1 containing 75g / L of copper is obtained. Part of it is returned to step (1) for ore blending, and part of it is sent to the extraction and electrowinning section to produce cathode copper. At the same time, filter cake from the filter press is obtained. The filter cake was acid-leached and stirred at 90℃ for 5 hours with 17.5% dilute sulfuric acid (the amount of dilute sulfuric acid was 3 times the mass of the filter cake). After filtration, a copper acid leaching solution 2 containing 15 g / L copper and 75 g / L iron was obtained. The copper acid leaching solution 2 was subjected to sodium sulfide precipitation to produce a copper precipitate containing 60 wt% copper, which was returned to step (1) for ore blending. The iron-containing leaching solution after copper removal entered the next step. The acid leaching residue contained 1050 g / t gold and 0.9 wt% copper. (4) Replacement-pressing filter cake of iron-containing leachate after copper removal is returned to pyrometallurgical smelting: Add 4 kg / m³ of reduced iron powder to the copper-free iron-containing leachate obtained in step (3). 3 The displacement reaction lasted for 3 hours. After displacement, the solution was filtered to obtain a pure ferrous sulfate solution. The solution was then evaporated, concentrated, and crystallized to produce ferrous sulfate. After displacement, the filter cake was filtered and returned to the pyrometallurgical furnace in step (1) to replace part of the quartz stone and realize the recycling of copper. The initial slag obtained in step (1) was enriched by magnetic separation to obtain magnetic concentrate. The magnetic tailings were used as building material raw materials. (5) Gold extraction by cyanidation of magnetic concentrate and acid leaching residue: The magnetic concentrate obtained in step (4) and the acid leaching residue obtained in step (3) are ground to -400 mesh with a content of 96wt%. The liquid-solid ratio is controlled at 4. Sodium hydroxide is added to adjust the pH to 9.5. The amount of ammonium bicarbonate added is 4kg / t. The mass concentration of sodium cyanide is 1.5%. The mixture is stirred and leached at 40℃ for 42h. After filtration, a cyanide leaching solution containing 105mg / L of gold and 300mg / L of silver is obtained, which is referred to as "cyanide solution". A cyanide residue containing 4.5g / t of gold and 12.5wt% of lead is obtained. This residue is used as raw material for pyrometallurgical lead smelting to comprehensively recover gold and lead. (6) Gold extraction from cyanide solution and mother liquor regeneration: Zinc powder is added to the cyanide solution in step (5). The amount of zinc powder is 1.4 times the mass of gold and silver in the cyanide solution. After displacement for 2.5 hours, precious gold and silver materials containing 40wt% gold and 35wt% silver are obtained by filter press. Gold ingots and silver ingots are produced by wet separation and pyrometallurgical smelting. The displacement solution is returned to step (5) for recycling.

[0022] Example 3 A method for synergistic gold extraction from complex gold concentrates via pyrometallurgical smelting, acid leaching, and cyanidation includes the following steps: (1) Pyrometallurgical smelting of complex gold concentrates to produce matte: Gold concentrate containing 60 g / t gold, 200 g / t silver, 3.5 wt% copper, 1.0 wt% lead, 10 wt% arsenic, and 5 wt% carbon is ground to -200 mesh with a content of 86 wt%. It is then mixed with quartz (flux, amounted to 20% of the gold concentrate mass) and pulverized coal (fuel, amounted to 12% of the gold concentrate mass) and fed into an oxygen-enriched side-blown smelting furnace. Oxygen-enriched air with a volume concentration of 40% is introduced and smelted at 1300℃ for 6 hours. The product is matte containing 15 wt% copper and 300 g / t gold (relative to the gold concentrate). At the same time, slag containing 0.18 wt% copper and 0.3 g / t gold is discharged and recorded as "initial slag". (2) Sulphation roasting of copper matte: The copper matte produced in step (1) is crushed and ground to -200 mesh with a content of 86wt%, and then fed into a fluidized bed roaster with a controlled air volume of 15000 m³ / h.3 At a furnace bottom pressure of 10000Pa and a roasting temperature of 630℃ for 3 hours, the cuprous sulfide (Cu2S) in the matte is converted into soluble copper sulfate (CuSO4), and the ferrous sulfide (FeS) is converted into iron oxide (Fe2O3). The product after roasting is recorded as "sulfation roasting residue". (3) Acid leaching of sulfation roasted residue: Add 2% dilute sulfuric acid to the sulfation roasting residue in step (2), control the liquid-solid ratio to 2, and acid leaching and stirring at 95°C for 4 hours. After filtration by a blower filter press, copper acid leaching solution 1 containing 100g / L of copper is obtained. Part of it is returned to step (1) for ore blending, and part of it is sent to the extraction and electrowinning section to produce cathode copper. At the same time, filter cake from the filter press is obtained. The filter cake was acid-leached and stirred at 95℃ for 6 hours with 20% dilute sulfuric acid (the amount of dilute sulfuric acid was 4 times the mass of the filter cake). After filtration, a copper acid leaching solution 2 containing 20 g / L copper and 100 g / L iron was obtained. The copper acid leaching solution 2 was subjected to sodium sulfide precipitation to produce a copper precipitate containing 70 wt% copper, which was returned to step (1) for ore blending. The iron-containing leaching solution after copper removal entered the next step. The acid leaching residue contained 2000 g / t gold and 0.8 wt% copper. (4) Replacement-pressing filter cake of iron-containing leachate after copper removal is returned to pyrometallurgical smelting: Add 5 kg / m³ of reduced iron powder to the copper-free iron-containing leachate obtained in step (3). 3 The displacement reaction lasted for 4 hours. After displacement, the solution was filtered to obtain a pure ferrous sulfate solution. The solution was then evaporated, concentrated, and crystallized to produce ferrous sulfate. After displacement, the filter cake was filtered and returned to the pyrometallurgical furnace in step (1) to replace part of the quartz stone and realize the recycling of copper. The initial slag obtained in step (1) was enriched by magnetic separation to obtain magnetic concentrate. The magnetic tailings were used as building material raw materials. (5) Gold extraction by cyanidation of magnetic concentrate and acid leaching residue: The magnetic concentrate obtained in step (4) and the acid leaching residue obtained in step (3) are ground to -400 mesh with a content of 97wt%. The liquid-solid ratio is controlled at 5. Sodium hydroxide is added to adjust the pH to 10. The amount of ammonium bicarbonate added is 5kg / t. The mass concentration of sodium cyanide is 2%. The mixture is stirred and leached at 45℃ for 48h. After filtration, a cyanide leaching solution containing 180mg / L of gold and 540mg / L of silver is obtained, which is referred to as "cyanide solution". A cyanide residue containing 4g / t of gold and 20wt% of lead is obtained. This residue is used as raw material for pyrometallurgical lead smelting to comprehensively recover gold and lead. (6) Gold extraction from cyanide solution and mother liquor regeneration: Add zinc powder to the cyanide solution in step (5). The amount of zinc powder is 1.5 times the mass of gold and silver in the cyanide solution. After 3 hours of replacement, the precious gold and silver material containing 50wt% gold and 40wt% silver is obtained by filter press. Gold ingots and silver ingots are produced by wet separation and pyrometallurgical smelting. The replacement liquid is returned to step (5) for recycling.

[0023] In summary, this invention achieves efficient recovery of copper and gold and recycling of materials through a synergistic gold extraction method involving pyrometallurgical smelting to produce copper matte, sulfation roasting of copper matte, acid leaching of sulfation roasting residue, replacement-pressing filter cake recycling smelting of iron-containing leachate after copper removal, cyanidation gold extraction from magnetic concentrate and acid leaching residue, and gold extraction from cyanide solution and mother liquor regeneration.

[0024] The above description is only a preferred embodiment of the present invention and is not intended to limit the present invention. Any modifications, equivalent substitutions, improvements, 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 synergistic gold extraction from complex gold concentrates via pyrometallurgical smelting, acid leaching, and cyanidation, characterized in that, Includes the following steps: (1) Pyrometallurgical smelting of complex gold concentrates to produce matte: Gold concentrate containing 10-60 g / t of gold, 20-200 g / t of silver, 2.5-3.5 wt% of copper, 0.3-1.0 wt% of lead, 5-10 wt% of arsenic, and 1-5 wt% of carbon is ground, mixed with flux and fuel, and fed into an oxygen-enriched side-blown smelting furnace. Oxygen-enriched air is introduced for smelting to produce matte containing 10-15 wt% copper and 30-300 g / t of gold relative to the gold concentrate. At the same time, slag containing ≤0.2 wt% copper and ≤0.5 g / t of gold is discharged and is recorded as "initial slag". (2) Sulphation roasting of copper matte: The matte produced in step (1) is crushed and sent to a fluidized bed roasting furnace. The air volume and furnace bottom pressure are controlled to roast the matte so that the cuprous sulfide Cu2S in the matte is converted into soluble copper sulfate CuSO4, and the ferrous sulfide FeS is converted into iron oxide Fe2O3. The product after roasting is called "sulfation roasting residue". (3) Acid leaching of sulfation roasted residue: Dilute sulfuric acid was added to the sulfation roasting residue in step (2), the liquid-solid ratio was controlled, and acid leaching and stirring were carried out under heating conditions. After pressure filtration, copper acid leaching solution 1 containing 50~100g / L of copper was obtained. Part of it was returned to step (1) for ore blending, and part of it was sent to the extraction and electrowinning section to produce cathode copper. At the same time, filter cake was obtained. The filter cake was acid-leached and stirred with dilute sulfuric acid under heating conditions. After filtration, a copper acid leaching solution 2 containing 10~20 g / L of copper and 50~100 g / L of iron was obtained. The copper acid leaching solution 2 was subjected to sodium sulfide precipitation to produce a copper precipitate containing 50~70 wt% copper, which was returned to step (1) for ore blending. The iron-containing leaching solution after copper removal entered the next step. The acid leaching residue contained 100~2000 g / t of gold and ≤1 wt% of copper. (4) Replacement-pressing filter cake of iron-containing leachate after copper removal is returned to pyrometallurgical smelting: The iron-containing leaching solution obtained after copper removal in step (3) is added with reducing iron powder to carry out a displacement reaction. After displacement, it is filtered by pressure to obtain a pure ferrous sulfate solution. Ferrous sulfate is produced by evaporation, concentration and crystallization. After displacement, the filter cake is filtered by pressure and returned to the pyrometallurgical furnace in step (1) to replace part of the flux and realize the recycling of copper elements. The primary slag obtained in step (1) is enriched by magnetic separation to obtain magnetic concentrate. (5) Gold extraction by cyanidation of magnetic concentrate and acid leaching residue: The magnetic concentrate obtained in step (4) and the acid leaching residue obtained in step (3) are ground together. The liquid-solid ratio is controlled, sodium hydroxide is added to adjust the pH, ammonium bicarbonate and sodium cyanide are added, and the mixture is stirred and leached under heating conditions. The mixture is filtered to obtain a cyanide leaching solution containing 30~180mg / L of gold and 60~540mg / L of silver, which is referred to as "cyanide solution". A cyanide residue containing ≤5g / t of gold and 5~20wt% of lead is obtained. (6) Gold extraction from cyanide solution and mother liquor regeneration: Zinc powder was added to the cyanide solution in step (5) for displacement, and the solution was filtered through a filter press to obtain precious gold and silver material containing 30-50 wt% gold and 30-40 wt% silver.

2. The method for synergistic gold extraction from complex gold concentrates via pyrometallurgical smelting, acid leaching, and cyanidation according to claim 1, characterized in that, In step (1), the flux is quartz, and the amount used is 15-20% of the mass of gold concentrate; the fuel is pulverized coal, and the amount used is 8-12% of the mass of gold concentrate.

3. The method for synergistic gold extraction from complex gold concentrates via pyrometallurgical smelting, acid leaching, and cyanidation according to claim 1, characterized in that, In step (1), the gold concentrate is ground to a content of ≥80wt% at -200 mesh, and the volume concentration of the oxygen-enriched air is 30~40%.

4. The method for synergistic gold extraction from complex gold concentrates via pyrometallurgical smelting, acid leaching, and cyanidation according to claim 1, characterized in that, In step (1), the melting temperature is 1200~1300℃ and the melting time is 4~6h.

5. The method for synergistic gold extraction from complex gold concentrates via pyrometallurgical smelting, acid leaching, and cyanidation according to claim 1, characterized in that, In step (2), the air volume is controlled at 12000~15000 m³ / h. 3 Firing at 8000~10000Pa at a furnace bottom pressure of 550~630℃ for 2~3 hours.

6. The method for synergistic gold extraction from complex gold concentrates via pyrometallurgical smelting, acid leaching, and cyanidation according to claim 1, characterized in that, In step (3), 1-2% dilute sulfuric acid is added to the sulfation roasting residue from step (2), the liquid-solid ratio is controlled to be 1-2, and acid leaching and stirring are carried out at 85-95℃ for 2-4 hours.

7. The method for synergistic gold extraction from complex gold concentrates via pyrometallurgical smelting, acid leaching, and cyanidation according to claim 1, characterized in that, In step (3), the filter cake is acid-soaked and stirred at 85-95℃ for 4-6 hours with a mass fraction of 15-20% dilute sulfuric acid. The amount of dilute sulfuric acid used is 2-4 times the mass of the filter cake.

8. The method for synergistic gold extraction from complex gold concentrates via pyrometallurgical smelting, acid leaching, and cyanidation according to claim 1, characterized in that, In step (4), add 3~5 kg / m³ of reduced iron powder. 3 The displacement reaction takes 2-4 hours.

9. The method for synergistic gold extraction from complex gold concentrates via pyrometallurgical smelting, acid leaching, and cyanidation according to claim 1, characterized in that, In step (5), the ore is ground to a -400 mesh content ≥95wt%, the liquid-solid ratio is controlled at 3~5, the pH is adjusted to 9~10 with sodium hydroxide, the amount of ammonium bicarbonate added is 3~5kg / t, the mass concentration of sodium cyanide is 1~2%, and the leaching is carried out by stirring at 35~45℃ for 36~48h.

10. The method for synergistic gold extraction from complex gold concentrates via pyrometallurgical smelting, acid leaching, and cyanidation according to claim 1, characterized in that, In step (6), the amount of zinc powder used is 1.2 to 1.5 times the mass of gold and silver in the cyanide solution, and the replacement time is 2 to 3 hours.