Method for secondary resource utilization of cyanidation dump leaching tailings in gold industry

Abstract

The invention discloses a method for secondary resource utilization of cyanided dump leaching tailings in the gold industry, which adopts the technical route of dump leaching tailings grinding, flotation, flotation gold concentrate roasting, roasted slag cyanidation leaching, leaching liquid carbon adsorption, carbon desorption, desorption liquid electrodeposition and gold ingot production, and the flotation tailings are backfilled and / or a building cement material is used. The gold concentrate cyanidation leaching residues are subjected to harmless treatment and then backfilled, effective secondary recycling of the cyanidation dump leaching tailings is achieved, metal and nonmetal in the tailings are fully recycled, economic benefits are increased, the occupied area of stockpiling of the tailings is reduced, and meanwhile the ecological environment is protected.

Description

Technical Field

[0001] The invention relates to the technical fields of ore dressing, metallurgy and environmental protection, and in particular to a method for secondary resource utilization of cyanide heap leaching tailings in the gold industry. Background Art

[0002] The conventional treatment process for gold mine heap leaching tailings is usually reclamation, planting strong plants and agricultural and forestry crops on the surface of abandoned dumps, decomposing the cyanide in the slag as a carbon and nitrogen source, and converting it into natural fertilizer for plant growth, which has a positive effect on protecting the environment and preventing pollution.

[0003] However, the metal and non-metal recovery in tailings is not complete. For example, a certain amount of gold, copper, lead, zinc, sulfur, etc. remain. Abandonment reclamation will cause a waste of resources. In addition, the amount of waste increases year by year, and most mining companies stockpile tailings. Therefore, the development and utilization of these secondary resources can not only reduce heavy metal emissions and protect the environment, but also generate economic benefits.

[0004] The most effective way to directly utilize tailings is to use tailings to backfill the goaf. Especially for those mining enterprises that are not convenient to set up tailings ponds and the local building materials market is relatively small, the comprehensive utilization method of backfilling the goaf after the tailings are utilized has great environmental and economic significance. Summary of the invention

[0005] In view of the deficiencies of the prior art, the present invention aims to provide a method for secondary resource utilization of cyanide heap leaching tailings in the gold industry.

[0006] In order to achieve the above object, the present invention adopts the following technical solution:

[0007] A method for secondary resource utilization of cyanide heap leaching tailings in the gold industry, comprising the following steps:

[0008] S1. Grinding the cyanide heap leaching tailings to obtain a ground sample;

[0009] S2, performing preliminary flotation on the sample obtained in step S1 to obtain a coarse gold concentrate and flotation tailings, and the flotation tailings are transferred to step S12 for treatment; the coarse gold concentrate is regrinded and closed-circuit flotation is performed using amyl xanthate as a collector to obtain a gold concentrate;

[0010] S3, roasting the gold concentrate obtained in step S2 in stages, absorbing the flue gas generated during the staged roasting with an alkaline solution to achieve desulfurization, dearsenicization and demercuration, and discharging the tail gas after it meets the standards;

[0011] S4, adding sodium cyanide to the roasted slag finally obtained by staged roasting in step S3 for cyanide leaching, and adjusting the pH value with alkali;

[0012] S5, performing solid-liquid separation on the leached slurry obtained after cyanide leaching in step S4 to obtain a leaching solution and a leaching residue, the leaching solution enters step S6 for treatment, and the leaching residue enters step S11 for treatment;

[0013] S6, the leaching solution obtained in step S5 is subjected to carbon adsorption to obtain gold-loaded carbon and adsorption liquid, the gold-loaded carbon is desorbed by the electrodeposition lean solution from step S7 to obtain desorption liquid and gold-removed carbon, the desorption liquid is transferred to step S7 for treatment, and the gold-removed carbon is transferred to step S8 for treatment; sodium cyanide and alkali are added to the adsorption liquid and then returned to step S4 to participate in cyanide leaching;

[0014] S7, the desorption liquid obtained in step S6 is continuously circulated and flows through the electrolytic cell, where an electrolytic reaction occurs, and the gold complex ions in the desorption liquid are converted into gold element and precipitated on the cathode, thereby obtaining gold mud, which is transferred to step S9 for treatment, and the electrolytic lean liquid enters step S6;

[0015] S8, the gold-free carbon obtained in step S6 is calcined and regenerated and then returned to step S6 for carbon adsorption;

[0016] S9: The gold mud obtained in step S7 is successively washed with nitric acid and washed with water for deacidification, and then dried, and then enters the process of step S10; the pickling liquid obtained by nitric acid washing is precipitated, and the precipitated slag obtained by the precipitation is smelted to recover silver, and Na2S is added to the precipitated liquid obtained by the precipitation to precipitate copper to obtain copper precipitated slag;

[0017] S10, adding flux to the dried gold mud obtained in step S9, and then entering the medium frequency furnace for smelting to obtain alloy gold;

[0018] S11, the pH value of the leached residue obtained in step S5 is adjusted with lime, the pH value is maintained above 8, and then sodium pyrosulfite and copper sulfate are added to break the cyanide, the broken cyanide residue is filtered and dried in the shade, and the CN of the filter residue is - After the content reaches the standard, the mining pit is backfilled, and the filtrate is returned to the cyanide breaking section for recycling, and the residual sodium cyanide and alkali in the filtrate are recycled;

[0019] S12, thickening the flotation tailings obtained in step S2 to obtain a thickened underflow and a thickened overflow, the thickened underflow is recycled and / or returned to the mining pit, and the thickened overflow is returned to the preliminary flotation and / or closed-circuit flotation of step S2 to adjust the flotation pulp concentration and utilize the return water.

[0020] Furthermore, in step S1, the grinding particle size is -75 μm, accounting for 60%-70%; in step S2, the coarse gold concentrate is re-ground to a fineness of -20 μm, accounting for 60%-80%.

[0021] Furthermore, in step S3, the staged roasting includes a first stage roasting and a second stage roasting; the first stage roasting temperature is 550-650°C, and the first stage roasting time is 0.5-3h; the second stage roasting temperature is 650-750°C, and the second stage roasting time is 0.5-3h.

[0022] Furthermore, in step S4, the alkali is sodium hydroxide, the pH value of cyanide leaching is 9.5-11, the mass concentration of sodium cyanide is 0.5‰-1‰, and the time is 24-48h.

[0023] Furthermore, in step S6, the desorbent is 4-8 g / L sodium hydroxide solution, the desorption temperature is 120-150°C, the desorption column pressure is 0.6-0.8 MPa, and the desorbent flow rate is 6-8 m 3 / h, the desorption time is 20-30h.

[0024] Furthermore, in step S7, the voltage of the electrowinning cell is 4V; the mass content of gold in the gold mud finally obtained is 40%-70%.

[0025] Furthermore, in step S8, the specific process of the gold-removed carbon undergoing pyrometallurgical regeneration is as follows: the gold-removed carbon is dried at 100-150° C. in a regeneration kiln, then carbonized at 650-750° C. for 1-2 hours, and then water quenched.

[0026] Furthermore, in step S10, the solvent is borax and potassium nitrate, the smelting temperature is 1250° C., the smelting time is 2 to 2.5 hours, and the gold mass content of the alloy gold is 75% to 95%.

[0027] Furthermore, in step S12, the filter residue is recycled and / or backfilled, including: the dense underflow is dried in the shade and then backfilled into the mining pit; and / or, the dense underflow is dried in the shade and then sold as mortar; and / or, the dense underflow is filtered, and the obtained filtrate is returned to the preliminary flotation or closed-circuit flotation of step S2 to adjust the concentration of the flotation pulp, and the obtained filter residue is dried and ground in turn by using the return water, and then dried in a hot air furnace again, and then ground by a roller vertical mill to a specific surface area of ​​350-550m 2 / kg, the fine-grained tailings powder is blown up by the wind and collected by the dust bag, and the coarse-grained tailings not collected by the dust bag are returned to the roller mill for further grinding.

[0028] The beneficial effects of the present invention are as follows: the method of the present invention adopts the technical route of heap leaching tailings grinding-flotation-flotation gold concentrate roasting-roasting slag cyanidation leaching-leaching solution carbon adsorption-carbon desorption-desorption solution electrowinning-production of gold ingots, and the flotation tailings are recycled, and the gold concentrate cyanidation leaching slag is harmlessly treated and then backfilled, so as to realize the effective secondary recycling of cyanidation heap leaching tailings, make full use of the metals and non-metals in the tailings, increase the economic benefits, reduce the floor space occupied by the stockpiled tailings, and protect the ecological environment at the same time. BRIEF DESCRIPTION OF THE DRAWINGS

[0029] Figure 1 This is a flow chart of the method of Embodiment 1-2 of the present invention. DETAILED DESCRIPTION

[0030] The present invention will be further described below in conjunction with the accompanying drawings. It should be noted that this embodiment is based on the technical solution and provides a detailed implementation method and specific operation process, but the protection scope of the present invention is not limited to this embodiment.

[0031] Example 1

[0032] The gold content of cyanide heap leaching tailings in a gold mine is shown in Table 1. The valuable metals are Au and Ag, with Au content of 0.34 g / t, Ag content of 2.53 g / t, As content of 0.04%, Hg content of 0.53 g / t, CN - The content is 4.45g / t. The harmful elements are As and Hg.

[0033] Table 1 Composition of cyanide heap leaching tailings

[0034]

[0035] The method of this embodiment for secondary resource utilization of the cyanide heap leaching tailings is as follows: Figure 1 As shown, the specific steps include:

[0036] S1, grinding the cyanide heap leaching tailings to obtain a sample with a particle size of -75μm accounting for 70%;

[0037] S2. The sample obtained in step S1 is subjected to preliminary flotation to obtain a coarse gold concentrate and flotation tailings. The coarse gold concentrate is ground to a fineness of -20 μm, accounting for 60%. Then, amyl xanthate is used as a collector to perform closed-circuit flotation to obtain a gold concentrate yield of 2.0%, an Au grade of 20 g / t, a gold content of 0.08 g / t in the tailings, an Au recovery rate of 85%, and an enrichment rate of harmful elements As and Hg greater than 98%.

[0038] In this embodiment, the closed-circuit flotation adopts a process of one roughing, two scavenging, one concentrate, and three re-grinding, which is specifically as follows:

[0039] Primary roughing: First, the ore is roughly sorted to separate most of the useful minerals from the ore;

[0040] Double scavenging: The tailings after roughing are scavenged twice to recover the remaining useful minerals in the slurry1.

[0041] Regrinding of concentrate: The concentrate after roughing is subjected to a first concentration, and the concentrate obtained from the first concentration is regrinded to make the mineral particle size finer and improve the mineral processing efficiency.

[0042] Three-time concentration: The concentrate after regrinding is subjected to three rounds of concentration in order to further improve the purity of the concentrate.

[0043] This process is often used to increase the recovery and purity of useful minerals in ores.

[0044] S3. The gold concentrate obtained in step S2 is roasted in stages, with the first stage at a temperature of 550°C and a roasting time of 3 hours, and the second stage at a temperature of 650°C and a roasting time of 3 hours. The final roasting slag has a desulfurization rate and a demercuration rate greater than 99.5%, and a dearsenication rate greater than 80%. The flue gas generated during the roasting process is absorbed by an alkaline solution with a NaOH concentration of 20g / L to achieve desulfurization, dearsenication, and demercuration treatment, and is discharged after meeting the standards;

[0045] S4, cyanide leaching the roasted slag obtained in step S3, adjusting the pH value to 10.5 with sodium hydroxide, adding sodium cyanide until the mass concentration of sodium cyanide is 1‰, the cyanide leaching time is 24h, and the gold leaching rate is greater than 95%;

[0046] S5, the leached slurry obtained after cyanide leaching is subjected to belt filtration to obtain a leachate and a leached residue, the leachate enters step S6, and the leached residue is transferred to step S11 for cyanide breaking treatment;

[0047] S6, the leaching solution obtained in step S5 is subjected to carbon adsorption to obtain gold-loaded carbon and adsorption liquid, the gold-loaded carbon is desorbed to obtain gold-free carbon and desorption liquid, and the adsorption liquid is supplemented with sodium cyanide and sodium hydroxide and then returned to step S4 to participate in cyanide leaching; the desorption temperature is 150°C, the desorption column pressure is 0.6MPa, and the desorbent flow rate is 6m 3 / h, the desorbent is 4g / L sodium hydroxide solution, and the desorption time is 20h;

[0048] S7, using a magnetic pump to drive the desorption solution obtained in step S6 to continuously circulate through the electrowinning tank in a closed system, and when the desorption solution containing gold complex ions flows through the electrowinning tank, an electrowinning reaction occurs, so that the gold complex ions are converted into gold single substance and precipitated on the cathode, thereby obtaining gold mud. The voltage of the electrowinning tank is 4V, and the gold mud contains about 50% gold.

[0049] S8. The gold-removed carbon obtained in step S6 is regenerated by pyrolysis and then returned for carbon adsorption. Specifically, the gold-removed carbon is first dried at 120° C. in a regeneration kiln, carbonized at 650° C. for 2 h, and then quenched with water and returned for carbon adsorption.

[0050] S9. The gold mud obtained in step S7 is sequentially pickled with nitric acid and washed with water to remove acid, and then dried. The dried gold mud enters step S10 for smelting. The pickling liquid obtained by nitric acid pickling is precipitated. The precipitated residue obtained by the precipitation is smelted to recover silver. Na2S is added to the precipitated liquid obtained by the precipitation to precipitate copper, and the obtained copper precipitated residue is sold.

[0051] S10. Add flux (80% of the weight of the gold mud, mass ratio of borax: potassium nitrate = 1) to the dried gold mud obtained in step S9, and put it into a medium frequency furnace for smelting at a smelting temperature of 1250° C. for 2 hours to obtain alloy gold with a gold content of 85% for external sale.

[0052] S11, the pH value of the leached slag slurry obtained in step S5 is adjusted to 8 with lime, and then sodium pyrosulfite and copper sulfate are added to break the cyanide, wherein 200 mg of copper sulfate is added per liter of leached slag slurry, and 3 kg of sodium pyrosulfite is added per ton of leached slag, the cyanide-breaking slag is filtered and dried in the shade, and the filter residue obtained by filtration is tested for CN by HJ / T 299 method - When the content is less than 5 mg / L, the mining pit is backfilled and the filtrate obtained by filtration is returned to the cyanide residue cyanide breaking section for recycling.

[0053] S12, the flotation tailings obtained in step S2 are thickened and subjected to 6-stage countercurrent thickening washing (CCD), the thickened overflow is returned to participate in flotation, and the thickened underflow is backfilled into the mining pit after being dried in the shade.

[0054] In this embodiment, the mining pit is an open pit where ore has been mined. According to local environmental protection requirements, an anti-seepage geomembrane can be laid in the open pit.

[0055] Example 2

[0056] The gold content of cyanide heap leaching tailings in a gold mine is shown in Table 2. The valuable metals are Au and Ag, with Au content of 0.55 g / t, Ag content of 4.33 g / t, As content of 0.08%, Hg content of 0.22 g / t, CN - The content is 3.41g / t. The harmful elements are As and Hg.

[0057] Table 2 Composition of cyanide heap leaching tailings

[0058]

[0059] The method of this embodiment for secondary resource utilization of the cyanide heap leaching tailings is as follows: Figure 1 As shown, the specific steps include:

[0060] S1, grinding the cyanide heap leaching tailings to obtain a sample with a particle size of -75μm accounting for 65%;

[0061] S2. Preliminary flotation is performed on the sample obtained in step S1 to obtain a coarse gold concentrate and flotation tailings. The coarse gold concentrate is re-ground to a fineness of -20 μm accounting for 70%, and then amyl xanthate is used as a collector to perform closed-circuit flotation to obtain a gold concentrate yield of 3.2%, an Au grade of 28 g / t, a gold content of less than 0.15 g / t in the tailings, an Au recovery rate of 78%, and an enrichment rate of harmful elements As and Hg greater than 98%;

[0062] This embodiment also adopts the process of one roughing, two scavenging, one concentrate, re-grinding and three concentrating to carry out closed-circuit flotation, and the specific operation is the same as that of Embodiment 1.

[0063] S3. The gold concentrate obtained in step S2 is roasted in stages, with the first stage at a temperature of 650°C and a roasting time of 0.5h, and the second stage at a temperature of 750°C and a roasting time of 0.5h. The obtained roasting slag has a desulfurization rate and a demercuration rate greater than 99.5%, and a dearsenication rate greater than 80%. The flue gas generated during the roasting process is treated with an alkaline solution with a NaOH concentration of 35g / L for desulfurization, dearsenication, and demercuration, and is discharged after meeting the standards;

[0064] S4, cyanide leaching is performed on the roasted slag obtained in step S3, the pH value is adjusted to 11 with sodium hydroxide, and sodium cyanide is added until the mass concentration of sodium cyanide is 0.5‰, the time is 48 hours, and the leaching rate of gold is greater than 94%;

[0065] S5, the leached slurry obtained after cyanide leaching is filtered by a thickener to obtain a leachate and a leached residue, the leachate enters step S6, and the leached residue is transferred to step S11 for cyanide breaking treatment;

[0066] S6, the leaching solution obtained in step S5 is subjected to carbon adsorption to obtain gold-loaded carbon and adsorption liquid, the gold-loaded carbon is desorbed to obtain gold-free carbon and desorption liquid, and the adsorption liquid is supplemented with sodium cyanide and sodium hydroxide and then returned to step S4 to participate in cyanide leaching; the desorption temperature is 150°C, the desorption column pressure is 0.7MPa, and the desorbent flow rate is 8m 3 / h, the desorbent is 5g / L sodium hydroxide solution, and the desorption time is 30h.

[0067] S7, using a magnetic pump to drive the desorption solution obtained in step S6 to continuously circulate through the electrowinning tank in a closed system, and when the desorption solution containing gold complex ions flows through the electrowinning tank, an electrowinning reaction occurs, so that the gold complex ions are converted into gold single substance and precipitated on the cathode, thereby obtaining gold mud. The voltage of the electrowinning tank is 4V, and the gold mud contains about 65% gold.

[0068] S8. The gold-removed carbon obtained in step S6 is regenerated by pyrolysis and then returned for carbon adsorption. Specifically, the gold-removed carbon is first dried at 150° C. in a regeneration kiln, then carbonized at 750° C. for 1 hour, and then returned for carbon adsorption after water quenching.

[0069] S9. The gold mud obtained in step S7 is sequentially pickled with nitric acid and washed with water to remove acid, and then dried. The dried gold mud enters step S10 for smelting. The pickling liquid obtained by nitric acid pickling is precipitated. The precipitated residue obtained by the precipitation is smelted to recover silver. Na2S is added to the precipitated liquid obtained by the precipitation to precipitate copper, and the obtained copper precipitated residue is sold.

[0070] S10. Add flux (100% of the weight of the gold mud, mass ratio of borax: potassium nitrate = 0.8) to the dried gold mud obtained in step S9, and put it into a medium frequency furnace for smelting at a smelting temperature of 1250° C. for 2.5 hours to obtain alloy gold with a gold content of 90% for external sale.

[0071] S11, the pH value of the leached slag slurry obtained in step S5 is adjusted to 8 with lime, and then sodium pyrosulfite and copper sulfate are added to break the cyanide, wherein 100 mg of copper sulfate is added per liter of leached slag, and 1 kg of sodium pyrosulfite is added per ton of leached slag, the cyanide-breaking slag is filtered and dried in the shade, and the filter residue obtained by filtration is tested for CN by HJ / T 299 method - When the content is less than 5 mg / L, the mining pit is backfilled and the filtrate obtained by filtration is returned to the cyanide residue cyanide breaking section for recycling.

[0072] S12, the flotation tailings obtained in step S2 are thickened and washed with 5-stage countercurrent thickening (CCD), the thickened overflow is returned to participate in flotation, part of the thickened underflow is dried in the shade and backfilled into the mining pit and made into mortar for sale, and the other part of the thickened underflow is filtered to further remove moisture, dried, ground, and then dried in a hot air furnace, and ground by a roller vertical mill to a specific surface area of ​​350m 2 / kg, the fine-grained tailings powder is blown up by the wind to the dust bag for collection, and the coarse-grained particles return to the roller mill for further grinding.

[0073] In this embodiment, the mining pit is an underground mining pit where ore has been mined.

[0074] For those skilled in the art, various corresponding changes and modifications can be made according to the above technical solutions and concepts, and all of these changes and modifications should be included in the protection scope of the claims of the present invention.

Claims

1. A method for secondary resource utilization of cyanide heap leaching tailings in the gold industry, characterized in that: The steps include: S1. Grinding the cyanide heap leaching tailings to obtain a ground sample; S2, performing preliminary flotation on the sample obtained in step S1 to obtain a coarse gold concentrate and flotation tailings, and the flotation tailings are transferred to step S12 for treatment; the coarse gold concentrate is regrinded and closed-circuit flotation is performed using amyl xanthate as a collector to obtain a gold concentrate; S3, roasting the gold concentrate obtained in step S2 in stages, absorbing the flue gas generated during the staged roasting with an alkaline solution to achieve desulfurization, dearsenicization and demercuration, and discharging the tail gas after it meets the standards; S4, adding sodium cyanide to the roasted slag finally obtained by staged roasting in step S3 for cyanide leaching, and adjusting the pH value with alkali; S5, performing solid-liquid separation on the leached slurry obtained after cyanide leaching in step S4 to obtain a leaching solution and a leaching residue, the leaching solution enters step S6 for treatment, and the leaching residue enters step S11 for treatment; S6, the leaching solution obtained in step S5 is subjected to carbon adsorption to obtain gold-loaded carbon and adsorption liquid, the gold-loaded carbon is desorbed by the electrodeposition lean solution in step S7 to obtain desorption liquid and gold-removed carbon, the desorption liquid is transferred to step S7 for treatment, and the gold-removed carbon is transferred to step S8 for treatment; sodium cyanide and alkali are added to the adsorption liquid and then returned to step S4 to participate in cyanide leaching; S7, the desorption liquid obtained in step S6 is continuously circulated and flows through the electrolytic cell, where an electrolytic reaction occurs, and the gold complex ions in the desorption liquid are converted into gold element and precipitated on the cathode, thereby obtaining gold mud, which is transferred to step S9 for treatment, and the electrolytic lean liquid enters step S6; S8, the gold-free carbon obtained in step S6 is calcined and regenerated and then returned to step S6 for carbon adsorption; S9: The gold mud obtained in step S7 is successively washed with nitric acid and washed with water for deacidification, and then dried, and then enters the process of step S10; the pickling liquid obtained by nitric acid washing is precipitated, and the precipitated slag obtained by the precipitation is smelted to recover silver, and Na2S is added to the precipitated liquid obtained by the precipitation to precipitate copper to obtain copper precipitated slag; S10, adding flux to the dried gold mud obtained in step S9, and then entering the medium frequency furnace for smelting to obtain alloy gold; S11, the pH value of the leached residue obtained in step S5 is adjusted with lime, the pH value is maintained above 8, and then sodium pyrosulfite and copper sulfate are added to break the cyanide, the broken cyanide residue is filtered and dried in the shade, and the CN of the filter residue is - After the content reaches the standard, the mining pit is backfilled, and the filtrate is returned to the cyanide breaking section for recycling, and the residual sodium cyanide and alkali in the filtrate are recycled; S12, thickening the flotation tailings obtained in step S2 to obtain a thickened underflow and a thickened overflow, the thickened underflow is recycled and / or returned to the mining pit, and the thickened overflow is returned to the preliminary flotation and / or closed-circuit flotation of step S2 to adjust the flotation pulp concentration and utilize the return water.

2. The method according to claim 1, characterized in that In step S1, the grinding particle size is -75 μm, accounting for 60%-70%; in step S2, the coarse gold concentrate is re-ground to a fineness of -20 μm, accounting for 60%-80%.

3. The method according to claim 1, characterized in that In step S3, the staged roasting includes a first stage roasting and a second stage roasting; the first stage roasting temperature is 550-650°C, and the first stage roasting time is 0.5-3h; the second stage roasting temperature is 650-750°C, and the second stage roasting time is 0.5-3h.

4. The method according to claim 1 or 3, characterized in that: In step S4, the alkali is sodium hydroxide, the pH value of cyanide leaching is 9.5-11, the mass concentration of sodium cyanide is 0.5‰-1‰, and the time is 24-48h.

5. The method according to claim 1, characterized in that: In step S6, the desorbent is 4-8 g / L sodium hydroxide solution, the desorption temperature is 120-150°C, the desorption column pressure is 0.6-0.8 MPa, and the desorbent flow rate is 6-8 m 3 / h, the desorption time is 20-30h.

6. The method according to claim 1, characterized in that In step S7, the voltage of the electrowinning cell is 4V; the mass content of gold in the gold mud finally obtained is 40%-70%.

7. The method according to claim 1, characterized in that In step S8, the specific process of the gold-removed carbon undergoing pyrometallurgical regeneration is as follows: the gold-removed carbon is dried at 100-150° C. in a regeneration kiln, then carbonized at 650-750° C. for 1-2 hours, and then water quenched.

8. The method according to claim 1, characterized in that: In step S10, the solvent is borax and potassium nitrate, the smelting temperature is 1250° C., the smelting time is 2 to 2.5 hours, and the gold mass content of the alloy gold is 75-95%.

9. The method according to claim 1, characterized in that: In step S12, the filter residue is recycled and / or backfilled, including: the dense underflow is dried in the shade and then backfilled into the mining pit; and / or, the dense underflow is dried in the shade and then sold as mortar; and / or, the dense underflow is filtered, and the obtained filtrate is returned to the preliminary flotation or closed-circuit flotation in step S2 to adjust the concentration of the flotation pulp. The filter residue is dried and ground in turn by using the return water, and then dried in a hot air furnace again, and then ground by a roller vertical mill to a specific surface area of ​​350-550m 2 / kg, the fine-grained tailings powder is blown up by the wind and collected by the dust bag, and the coarse-grained tailings not collected by the dust bag are returned to the roller mill for further grinding.

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