Efficient separation and extraction method of high-sulfur and high-lead gold slurry by carbon slurry process

Abstract

The present application relates to a kind of high-efficiency separation extraction methods of high-sulfur high-lead gold mud of carbon slurry process, comprising the following steps: (1) high-temperature roasting;(2) gold mud slagging smelting;(3) liquid phase alloy oxidation refining;(4) alloy water quenching splashing bead;(5) alloy sheet lead leaching and filtration;(6) smelting slag oxidation lean smelting;(7) fume lead leaching and filtration;(8) lead-containing liquid precipitation recovers lead;(9) wet separation gold;(10) oxidation lean slag grinding cyanidation and magnetic separation;(11) mixed solution two-stage precipitation recovers fine gold.The present application can effectively remove sulfur, iron, lead and other impurities contained in gold mud, ensure that more than 90% of sheet alloy is obtained by water quenching splashing bead, and high-efficiency gold powder with color of 99% and above is obtained after wet separation, while also realizing the removal of lead, iron and the recovery of fine gold in mixed solution, and the effect of resource comprehensive recycling is remarkable.

Description

Technical Field

[0001] This invention belongs to the field of precious metal metallurgy technology, specifically relating to an efficient method for separating and extracting high-sulfur, high-lead gold mud using a carbon-in-pulp process. Background Technology

[0002] In recent years, with the decreasing availability of easily beneficiated metallurgical ores, the mining and use of low-gold, high-lead ores associated with magnetic and limonite ores, or those with high sulfur content in oxygen-sulfur mixed zones, has become increasingly difficult. These ores are characterized by being "poor, fine-grained, and impure." The main technical route for extracting gold, silver, iron, and other valuable metals from these ores is "grinding—cyanide leaching and activated carbon adsorption—magnetic separation—gold-loaded carbon desorption and electrolysis—gold mud purification and refining—alloyed gold." Due to the increased content of Pb, S, and Fe in the raw ore, the activated carbon in the carbon-in-pulp process adsorbs large amounts of Pb, S, and Fe along with Au and Ag. These adsorbents are then enriched in the crude gold mud product after desorption and electrolysis. The traditional process for separating and extracting Au and Ag from high-Pb, S, and Fe gold mud presents several problems, primarily manifested in the following ways: First, the liquid phase alloy is difficult to quench and bead in, and the alloy sheets are difficult to form into sheet-like shapes. In the traditional "slag-smelting-bead-nitric acid gold separation" process, the quality of the water-quenched beads is crucial to the efficiency of nitric acid gold separation and ensuring the color of the gold powder. However, the liquid phase alloy produced by slag smelting of high-S and Fe gold mud contains not only Au, Ag, and Cu, but also residual Fe that has not been separated by slag. Sometimes, matte (a eutectic of Cu2S-FeS) is also formed. When this metallic liquid phase is water-quenched and beaded, the gold and silver are encapsulated by these impurities Fe3O4 or matte. The water-quenched product is difficult to form alloy sheets with a large specific surface area, and 80% are spherical, which is not conducive to subsequent nitric acid gold separation. For gold mud smelting slag, the traditional induction furnace is generally used for secondary heating and depletion smelting, but the impurity Fe is still difficult to separate by slag and eutectic with the liquid phase. The water-quenched product of the high-temperature metallic eutectic is 90% spherical, which makes it impossible to obtain qualified gold powder in subsequent nitric acid gold separation.

[0003] Secondly, the gold powder impurity removal efficiency is low. Low Fe, Pb, S gold mud can generally obtain gold powder with a purity of over 99% by using the "slag-smelting-bead-nitric acid separation" process. However, when separating the spherical alloys coated with iron or copper matte with nitric acid, the separation time is more than 3 days and the nitric acid consumption is high (6-8 parts of 50% concentration nitric acid are consumed for one alloy sheet). Moreover, the purity of the gold powder is low, with Au content only reaching 85%-90% and Pb impurity content as high as 10%, making it impossible to obtain qualified gold powder.

[0004] Third, S, Pb, and Fe in the gold mud were not effectively removed and recovered. During the gold mud smelting process, matte (a eutectic of Cu2S-FeS) is generated between the slag layer and the metal layer, forming liquid droplets containing gold and silver. Fe is also difficult to separate in slag, and about 87% of Pb enters the wastewater and about 13% enters the smelting slag. The failure to effectively remove and recover S, Fe, and Pb resulted in metal loss.

[0005] Fourth, in traditional wet gold separation processes for high-sulfur, high-lead gold mud, fine gold particles are lost with the liquid phase during gold powder filtration. As the gold embedding characteristics in the source ore become finer, the gold particle size in deep-mined ores with pyrite and magnetite as carrier minerals or in gold-bearing ores in oxygen-sulfur mixed zones tends to be finer, mainly distributed below 10μm. The gold in the final secondary metallurgical process also exhibits the characteristic of fine embedding particle size. During the separation of gold by nitric acid and the filtration of gold powder, these fine gold particles easily penetrate the filter cloth and are lost with the liquid phase or with iron colloid suspension. Atomic absorption spectrometry analysis shows that the gold concentration in the liquid phase is generally as high as 0.08mg / L.

[0006] Fifth, the fumes from high-lead smelting pose a significant occupational health hazard to employees during the depletion smelting process. When the fumes generated from gold mud smelting are repeatedly depleted and smelted in an intermediate frequency furnace to recover Au, Ag, and Cu, the smelting temperature is generally between 1300℃ and 1600℃. At this temperature, Pb is more likely to volatilize in large quantities as the temperature increases, leading to a higher risk of occupational hazards for workers.

[0007] Therefore, the traditional "slag-smelting-bead-nitric acid gold separation" process can no longer meet the requirements for the separation and extraction of gold mud with high Pb, S, and Fe content. Thus, this invention provides a highly efficient separation and extraction method for high-sulfur, high-lead gold mud to improve the separation and extraction effect of valuable metals in gold mud, efficiently separating gold while removing and recovering S, Pb, and Fe. Summary of the Invention

[0008] To address the above problems, this invention provides a highly efficient method for separating and extracting high-sulfur, high-lead gold mud using a carbon-in-pulp process.

[0009] The specific technical solution is: a highly efficient method for separating and extracting high-sulfur, high-lead gold mud using a carbon-in-pulp process, comprising the following steps: (1) High-temperature roasting: Saltpeter is added to the high-sulfur and high-lead gold mud with a water content of 50%-60% obtained by electrolytic desorption of gold-carrying carbon. After thorough mixing, the mixture is placed in an oven and heated to carry out dehydration, desulfurization, and decarbonization roasting reactions. The process first involves high-temperature decomposition to dehydrate, decarbonize, and remove cyanide ions (CN) from the gold mud. - When the moisture removal temperature reaches 400℃, the sulfur in the gold mud reacts with the nitrate to produce sulfur dioxide gas. Moreover, the desulfurization rate increases continuously with the increase of temperature. After roasting for 12-14 hours, most of the sulfur is removed, which can effectively reduce the amount of sulfur entering the smelting process and reduce the formation of smelted matte.

[0010] (2) Gold mud slag making and smelting: After the dry gold mud after high temperature roasting in step (1) is fully mixed with the slag-making agent, it is put into the high-purity graphite crucible of the medium frequency furnace for high temperature slag making and smelting. After high temperature smelting, the low-density calcium-silicon iron smelting slag and the liquid phase alloy containing copper, lead, gold and silver are separated into layers, completing the initial separation of smelting slag and valuable metals. Then, the upper layer of smelting slag is scraped into the slag tank, and the remaining liquid phase alloy layer (including 4-8cm smelting slag layer or matte layer and below liquid phase alloy layer) is transferred to the next process. This process utilizes the high efficiency of Cu and Pb in capturing Au and Ag, and through high-temperature slag smelting, Cu-Pb-Au-Ag is initially separated from impurities such as silicon, calcium, and iron.

[0011] (3) Liquid phase alloy oxidation refining: Pour the liquid phase alloy obtained in step (2) into a silicon carbide-graphite crucible, add slag-forming agent, and heat up and blow air to carry out oxidation refining. The copper matte is oxidized, and sulfur and some lead volatilize into the flue dust to obtain the upper layer of oxidation refining slag (including iron removal slag and low-density slag of iron silicate and calcium silicate) and the lower layer of liquid phase alloy with higher density. The purpose of oxidation refining is to deeply remove impurities such as iron and matte from the liquid alloy, creating favorable conditions for subsequent bead casting. The basic principle of this method is based on the different affinities of metals for oxygen. By introducing air, Fe, matte, and a small amount of Pb impurities are oxidized to form oxides insoluble in the Cu-Au-Ag matrix metal, which accumulate as slag on the surface of the melt. S is further separated in a gaseous state. Furthermore, by using slag-forming agents, a low-melting-point slag is formed, and sodium carbonate enhances the slag's fluidity, facilitating the flotation of impurities.

[0012] (4) Alloy water quenching beads: The upper layer of oxide refining slag obtained in step (3) is scraped into the slag tank. After the liquid phase alloy is cooled down to 900-1000℃, it is slowly and uniformly injected into the water tank. At the same time, when pouring in the high temperature liquid phase alloy, a high pressure water pipe (water pressure controlled at 0.15-0.2MPa) is used to flush and disperse the quenching beads. The water pipe outlet is 10cm away from the liquid phase alloy injection point. This continues until all the liquid phase alloy is poured in. Then, the copper-silver-gold-lead water quenched alloy sheet with a large specific surface area is obtained by filtration. This process has a significant impact on subsequent wet gold separation and gold powder coloration. Research shows that the Cu-Pb-Au-Ag liquid phase alloy exhibits the best bead-splashing effect when the temperature is 900-1000℃ and the water pressure is 0.15-0.2MPa, with most of the beads appearing as fine flakes and having a large specific surface area.

[0013] (5) Leaching and filtration of lead in alloy sheet: The water-quenched alloy sheet obtained in step (4) is loaded into a titanium reactor, industrial acetic acid is added, and the mixture is heated and allowed to react for 2 hours. Then, stirring is started and the reaction continues for 2 hours. After the reaction is completed, the lead-containing liquid and lead-free alloy sheet are obtained by filtration. This process utilizes the fact that gold, silver, and copper do not react with acetic acid, while lead does, to achieve lead removal and prevent excessive lead content in gold powder. The main chemical formula is: 2CH3COOH + Pb ==Pb(CH3COO)2 + H2↑.

[0014] (6) Oxidation and depletion smelting of smelting slag: The smelting slag and oxidative refining slag produced in steps (2) and (3) are crushed to less than 10 mm, mixed with slag-forming agent, and then loaded into an oxidation depletion furnace for heating and blowing gas to carry out depletion smelting. After the smelting is completed, a low-density upper layer of oxidation depletion slag (Au grade <50 g / t) and a lower layer of liquid phase recovery alloy containing a small amount of depletion slag are obtained. The upper layer of oxidation depletion slag is scraped to the slag tank, and the lower layer of liquid phase recovery alloy is returned to step (3) oxidation refining. The main purpose of this process is to further recover residual Au and Ag in the smelting slag. By rationally controlling the slag-forming agent, slag shape, and temperature, and by introducing air, the impurities Ca-Si-Fe are caused to form low-density lean slag that floats to the liquid surface and separates from Au-Ag. The introduction of air accelerates the oxidation and flotation of Ca-Si-Fe, sodium carbonate enhances the slag fluidity, and silica reacts with iron to form low-density, low-melting-point ferrosilicon slag. Because the material itself contains silicates, under these material conditions, a low-viscosity, well-flowing, weakly acidic slag is formed, and iron does not form iron(III) oxide (Fe3O4).

[0015] (7) Lead leaching and filtration of smoke dust: The smoke dust generated in steps (2), (3) and (6) is collected and poured into a titanium reactor. Acetic acid is added, and the static reaction is heated for 2 hours. Then, stirring is started and the reaction continues for 3 hours. After the reaction is completed, lead-containing liquid and lead-removed smoke dust are obtained by filtration. The lead-removed smoke dust is returned to step (6) for oxidation and depletion smelting. This process utilizes the fact that gold, silver, and copper do not react with acetic acid, while lead does, to achieve the purpose of leaching and separating lead from the flue dust in advance. The main chemical formula is: 2CH3COOH + Pb ==Pb(CH3COO)2 + H2↑.

[0016] (8) Lead-containing liquid precipitation and lead recovery: The lead-containing liquid obtained in steps (5) and (7) is transported to the lead precipitation reactor, stirring is started, and sodium sulfide solution is slowly added to react for 0.5 hours. After the reaction is completed, the lead concentrate product and lead-removing liquid are obtained by filtration. The lead concentrate is sold externally, and the lead-removing liquid is sent to the wastewater treatment system. Lead ions in lead-containing solutions undergo a sulfide precipitation reaction with sodium sulfide solution. The sulfide precipitation method utilizes the reaction between heavy metal ions and sulfur in the solution. 2- The reaction produces M2S X The precipitate, based on the difference in solubility product of different metal sulfides, KspAg2S (1.6x10 -49 ) < KspCuS(8.5x10 -45 ) < KspPbS(83.4x10-28 )<KspZnS(1.2x10 -23 )<KspFeS(3.7x10 -19） To achieve the purpose of removing or recycling heavy metals.

[0017] (9) Wet gold separation: The lead-free alloy sheet obtained in step (5) is loaded into an acid-resistant and high-temperature resistant titanium reactor. First, a small amount of water is added to wet the alloy sheet. Stirring and heating are started, and the temperature is controlled at 60-80℃. Then, sulfuric acid and nitric acid are added in batches for 3-5 separations. In the first 1-2 separations, sulfuric acid solution and nitric acid solution with a total weight of 4 times the alloy sheet are used, and the concentration of sulfuric acid solution is 98% and the concentration of nitric acid solution is 20%-25%. In the third-5 separations, nitric acid solution with a total weight of 2-3 times the alloy sheet is used, and the concentration of nitric acid solution is 30%-45%. No sulfuric acid is added until the solution color changes from dark blue to clear and there is no gas on the liquid surface. The wet gold separation operation is then considered to be over. Then, gold powder and filtrate are obtained by filtration. The filtrate is a mixed solution of copper nitrate, silver nitrate and copper sulfate. This process utilizes the strong acidity of sulfuric acid to accelerate the dissolution of Fe, Ag, and Cu in the alloy sheet under the combined effects of sulfuric acid, nitric acid, high temperature, and stirring.

[0018] (10) Oxidized lean slag grinding, cyanidation and magnetic separation: The oxidized lean slag obtained in step (6) is ground and cyanidated to recover gold. The tailings are then fed into the magnetic separation process to recover iron and obtain magnetite concentrate. (11) Two-stage precipitation of mixed solution to recover fine gold particles: The copper nitrate-silver nitrate-copper sulfate mixed solution obtained in step (9) is sequentially introduced into a closed primary precipitation tank and a two-stage precipitation tank. The bottom of the two precipitation tanks is conical and the top is filled with resin balls. When recovering fine gold particles, flocculant is added to the mixed solution below the resin balls. Under the action of flocculation and the interception of the resin balls, the fine gold particles agglomerate and precipitate at the bottom of the precipitation tank. After periodic discharge, filtration and smelting recovery, the remaining solution overflows from the top of the precipitation tank into the silver and copper recovery process.

[0019] Furthermore, in step (1), the amount of saltpeter used is 0.5%-1% of the amount of gold mud; the oven temperature is controlled at 700-800℃, and the static roasting reaction time is 12-14h.

[0020] Furthermore, the slag-forming agent mentioned in step (2) includes quartz, borax and sodium carbonate; in order to prevent lead volatilization, the temperature of the high-temperature slag-forming smelting should not be too high and should be controlled at 1200-1300℃; in order to prevent valuable metals such as gold and silver from being carried along with the slag during slag removal, a smelting slag layer of 4-8cm is reserved above the liquid phase alloy layer during slag removal.

[0021] Furthermore, the slag-forming agent in step (3) includes quartz, sodium carbonate, and borax, with the addition amounts of each component being 2%-3%, 1%-2%, and 0.5%-1.5% of the total melt weight, respectively. This ratio is beneficial for reducing slag viscosity, enhancing fluidity, and creating a reasonable slag shape. During refining, the melt temperature is maintained at 1100-1300℃. During air blowing, the air distribution pipe outlet is lowered to the middle position of the molten metal surface, and the air blowing oxidation time for each furnace is 0.5-1h (this time is crucial for de-ironizing and matte blowing), with an air volume of 0.2-0.3m³. 3 / min.

[0022] Furthermore, in step (4), the water pressure of the high-pressure water pipe is controlled at 0.15-0.2MPa, and the water pipe outlet is located 10cm away from the liquid phase alloy injection point. The incoming water flow must ensure that the bottom of the water quenching tank can form agitation and rotation.

[0023] Furthermore, in step (5), the acetic acid concentration is 50%-60%, the liquid-solid ratio of the acetic acid solution to the alloy sheet is 1:4-5, and the heating temperature is controlled at 60-80℃.

[0024] Furthermore, the slag-forming agent mentioned in step (6) includes crushed glass, sodium carbonate, and borax, and the addition amounts of each component are 5%-8%, 10%-15%, and 2%-5% of the total weight of the molten slag, respectively, wherein the particle size of the crushed glass is less than 20mm; the smelting temperature is controlled at 1300-1500℃; during air blowing, the outlet of the air distribution pipe is lowered to the middle position of the molten slag, and each furnace is smelted by air blowing for 1 hour, with an air volume of 0.4-1m³. 3 / min.

[0025] Furthermore, in step (7), the acetic acid concentration is 50%-70%, the liquid-solid ratio of the acetic acid solution to the dust is 1:6-8, and the heating temperature is controlled at 60-80℃.

[0026] Furthermore, in step (8), the concentration of sodium sulfide solution is 10%-15%, and the amount of sodium sulfide solution used is 1.2-1.5 times the amount of lead metal in the lead-containing solution.

[0027] Furthermore, in step (10), the grinding particle size is -200 mesh, accounting for 90%-92%; the magnetic separation concentration of tailings is 30%-38%, and the magnetic field strength is 0.5T.

[0028] Furthermore, the diameter of the resin balls in step (11) is 1.5 cm; the dosage of the flocculant is 0.1-0.2 g / m³. 3 .

[0029] The beneficial effects of this invention are as follows: This invention can effectively remove impurities such as sulfur, iron, and lead from gold mud, ensuring that water-quenched beads yield more than 90% flake alloy. After wet separation, gold powder with a purity of 99% or higher is obtained efficiently. Simultaneously, it also achieves the removal of lead and iron, as well as the recovery of fine gold particles from mixed solutions. The comprehensive resource recovery and utilization effect is significant. This invention provides an economical, environmentally friendly, and efficient method for gold mud impurity removal and gold and silver extraction in the carbon-in-pulp process of polymetallic refractory gold mines, and has broad application prospects. The specific technical advantages of this invention are as follows: (1) Before slag smelting, the present invention first removes most of the sulfur impurities in the gold mud by high-temperature roasting to reduce the sulfur from entering the smelting process to generate matte. Then, the liquid phase alloy obtained by slag smelting is oxidized and refined to deeply remove impurities such as S, Fe, Pb and matte. Furthermore, by reasonably controlling the bead temperature and water pressure, the problem of the liquid phase alloy beads being severely encased by impurities such as matte and appearing as round beads is completely solved, producing a sheet alloy with a content of more than 90%. Its large specific surface area creates favorable conditions for the subsequent nitric acid gold separation reaction.

[0030] (2) Before the wet gold separation process, a lead leaching step was added to the alloy sheet. Acetic acid was used to leach the lead from the alloy sheet in advance, which solved the problem that the lead and lead oxides were tightly wrapped with gold and did not react with nitric acid, resulting in high lead content in the gold powder product. In addition, by controlling the concentration and amount of sulfuric acid and nitric acid in the wet gold separation process, the gold powder product with a purity of up to 99% was separated in multiple reactions while reducing the amount of reagents consumed.

[0031] (3) The process of “acetic acid leaching + sulfide precipitation” for lead in dust and alloy sheets not only removes lead but also recovers it, resulting in lead concentrate products that can be sold. This process creates value while reducing subsequent treatment and environmental problems caused by lead, and has good economic and social benefits.

[0032] (4) The present invention fully recovers intermediate products—liquid phase recovery alloy, oxidized refining slag and lead removal dust by returning them to the corresponding recycling process; the intermediate product—depleted slag also recovers residual gold again by grinding and cyanidation, and further recovers magnetite from cyanide tailings by magnetic separation; especially for fine gold particles, a "two-stage flocculation sedimentation and resin ball adsorption process" is set up to achieve deep recovery. All of these fully demonstrate the advantages of the present invention in comprehensive resource recovery and utilization. Attached Figure Description

[0033] Figure 1 This is a flowchart of a highly efficient separation and extraction method for high-sulfur and high-lead gold mud using a carbon-in-pulp process according to the present invention. Detailed Implementation

[0034] To make the technical problems and solutions solved by the present invention clearer, the present invention will be further described in detail below with reference to the accompanying drawings and embodiments. It should be understood that the specific embodiments described herein are merely for explaining the present invention and are not intended to limit the present invention. Example 1

[0035] Gold mud from a polymetallic gold mine in southern my country was obtained through desorption and electrolysis of gold-loaded carbon in a carbon-in-pulp plant. Its main components and contents are: Au 6.1%, Ag 37.2%, Cu 28.07%, Pb 7.78%, Fe 2%, and S 2.24%.

[0036] like Figure 1 As shown, the specific steps for processing the gold mud using the method described in this invention are as follows: (1) High-temperature roasting: Saltpeter was added to the gold mud, the amount of saltpeter being 0.6% of the gold mud amount. After thorough mixing, the mixture was placed in an oven and heated to 750℃ for dehydration, desulfurization, and decarbonization roasting reaction. The roasting time was 12 hours, achieving a desulfurization rate of 85.6%. (2) Gold mud slag making and smelting: After the dry gold mud after high temperature roasting in step (1) is fully mixed with the slag-making agent (quartz + borax + sodium carbonate), it is loaded into a high-purity graphite crucible in a medium frequency furnace for high-temperature slag making and smelting. The smelting temperature is controlled at 1200-1250℃ and smelting is carried out for 0.5h. The low-density calcium-containing ferrosilicon smelting slag and the liquid phase alloy containing copper, lead, gold and silver are separated into layers, completing the initial separation of smelting slag and valuable metals. The yield of smelting slag is 15%, the yield of liquid phase alloy is 84%, and it contains 2.3% Fe and 8.5% Pb.

[0037] (3) Liquid phase alloy oxidation refining: Pour the liquid phase alloy obtained in step (2) into a silicon carbide-graphite crucible, add slag-forming agents (including quartz, borax and sodium carbonate, with the addition amounts of each component being 2%, 1%, and 0.5% of the total weight of the melt, respectively); raise the temperature to 1100-1200℃, lower the outlet of the air blowing pipe to the middle position of the metal liquid surface, and blow air for oxidation for 0.5h, with an air volume of 0.25m³. 3 / min, which oxidizes Fe, matte and a small amount of Pb impurity metal to form oxides that are insoluble in the Cu-Au-Ag main metal, which accumulate on the surface of the melt in the form of slag, to obtain a liquid phase alloy containing 1% Fe and 7% Pb. (4) Alloy water quenching beads: After cooling the liquid phase alloy obtained in step (3) to 900℃, slowly and uniformly inject it into a clean water tank. At the same time, when pouring in the high-temperature liquid phase alloy, use a high-pressure clean water pipe (water pressure controlled at 0.16MPa) to flush and disperse the quenching beads. The water pipe outlet is 10cm away from the liquid phase alloy injection point. Continue until all the liquid phase alloy is poured in. Then filter to obtain flaky, copper, silver, gold and lead-containing water quenched alloy flakes with a particle size of less than 2mm, accounting for 92%. (5) Lead leaching of alloy sheet: The water-quenched alloy sheet obtained in step (4) is placed in a titanium reactor, and a 52% acetic acid solution is added. The liquid-solid ratio of the acetic acid solution to the alloy sheet is 1:4. The mixture is heated to 60-70℃ and allowed to stand for 2 hours. Then, stirring is started and the reaction continues for another 2 hours. After the reaction is completed, the lead-containing liquid and lead-free alloy sheet are obtained by filtration. The lead leaching rate is 98.1%. (6) Oxidation and depletion smelting of smelting slag: The smelting slag and oxidized refining slag produced in steps (2) and (3) are crushed to less than 10 mm, mixed with slag-forming agents (including crushed glass, sodium carbonate and borax, with the addition amounts of each component being 5%, 12% and 3% of the total weight of the slag, respectively), and then loaded into an oxidation depletion furnace for heating and blowing air for depletion smelting. The smelting temperature is controlled at 1300-1400℃, and the outlet of the air blowing pipe is lowered to the middle position of the slag. The blowing smelting is carried out for 1 hour with an air volume of 0.45 m³ / h. 3 / min, to obtain a low-density upper layer of oxidized lean slag containing Au<40g / t, mFe1%, the generated flue dust containing Pb6.5%, and a lower layer of liquid phase recovery alloy containing a small amount of lean slag, which is returned to step (3) oxidative refining; (7) Lead leaching and filtration of flue dust: The flue dust generated in steps (2), (3), and (6) is collected and poured into a titanium reactor. A 70% acetic acid solution is added, with a liquid-to-solid ratio of acetic acid solution to flue dust of 1:8. The mixture is heated to 75°C and statically reacted for 2 hours. Then, stirring is started and the reaction continues for 3 hours. The mixture is filtered to obtain lead-containing liquid and lead-removed flue dust. The lead-removed flue dust is returned to step (6) for oxidation and smelting. The lead leaching rate reaches 94.2%.

[0038] (8) Lead recovery from lead-containing liquid precipitation: The lead-containing liquid obtained in steps (5) and (7) is transported to the lead precipitation reactor, stirring is started, and a sodium sulfide solution with a concentration of 14% is slowly added. The amount of sodium sulfide solution used is 1.4 times the amount of lead metal in the lead-containing liquid. After reacting for 0.5 hours, the solution is filtered to obtain lead concentrate and lead-removing liquid. The lead precipitation recovery rate is 99%, and the lead concentrate contains 60% lead.

[0039] (9) Wet gold separation: The lead-free alloy sheet obtained in step (5) is loaded into an acid-resistant and high-temperature-resistant titanium reactor. A small amount of water is added to wet the alloy sheet. Stirring and heating are started, and the temperature is controlled at 60-70℃. Sulfuric acid and nitric acid are added in batches for 3-5 separations. In the first 1-2 separations, sulfuric acid solution and nitric acid solution with a total weight of 4 times the alloy sheet are used, and the concentration of sulfuric acid solution is 98% and the concentration of nitric acid solution is 20%-25%. In the third-5 separations, nitric acid solution with a total weight of 2 times the alloy sheet is used, and the concentration of nitric acid solution is 30%-45%. No sulfuric acid is added until the solution color changes from dark blue to clear and there is no gas on the liquid surface. The wet gold separation operation is then considered to be over. Then, the gold powder with a purity of 99.5% and the filtrate are obtained by filtration. The filtrate is a mixed solution of copper nitrate, silver nitrate and copper sulfate. In this process, the consumption of nitric acid per alloy sheet is reduced from 3 kg / kg alloy sheet to 2 kg / kg alloy sheet, and the production efficiency is reduced from 3 days / batch to 2 days / batch.

[0040] (10) Oxidized lean slag grinding, cyanidation and magnetic separation: The oxidized lean slag obtained in step (6) is ground to 200 mesh (92%), cyanidation is performed to recover gold, and the tailings are fed into the magnetic separation process at a concentration of 31% to recover iron. The magnetic field strength is 0.5T. mFe is reduced from 1% to 0.5%, and the magnetite concentrate recovery rate is 50%.

[0041] (11) Two-stage precipitation of mixed solution to recover fine gold particles: The copper nitrate-silver nitrate-copper sulfate mixed solution obtained in step (9) is sequentially fed into a closed primary precipitation tank and a two-stage precipitation tank, and flocculant 0.1 g / m³ is added to each of the two precipitation tanks. 3 After secondary flocculation and sedimentation, fine gold particles were recovered, and the final Au content in the overflow was 0.01 mg / L, which is a significant reduction compared to 0.08 mg / L in the traditional method. Example 2

[0042] The gold mud from a polymetallic gold mine in southern my country was obtained through desorption and electrolysis of gold-loaded carbon in a carbon-in-pulp plant. Its main components are Au 5.20%, Ag 20.53%, Cu 32.16%, Pb 10%, Fe 2.2%, and S 5.36%.

[0043] (1) High-temperature roasting: Saltpeter is added to the gold mud, the amount of saltpeter is 1% of the amount of gold mud, and after being thoroughly mixed, it is put into an oven and heated to 800℃ to carry out dehydration, desulfurization and decarbonization roasting reaction. The roasting time is 14h, and the desulfurization rate is 60%; (2) Gold mud slag smelting: After the dry gold mud after high temperature roasting in step (1) is fully mixed with the slag-forming agent (quartz + borax + sodium carbonate), it is loaded into a high-purity graphite crucible in a medium frequency furnace for high-temperature slag smelting. The smelting temperature is controlled at 1250-1300℃ and smelting is carried out for 0.52h. After smelting, a low-density smelting slag containing calcium and silicon iron and a liquid phase alloy layer containing copper, lead, gold and silver are formed. A gray-white matte layer with a thickness of about 2-3cm is generated between the slag layer and the liquid phase layer. The smelting slag yield is 15%, the liquid phase alloy yield is 82%, and it contains 2.5% Fe and 8.5% Pb.

[0044] (3) Liquid phase alloy oxidation refining: Pour the liquid phase alloy obtained in step (2) into a silicon carbide-graphite crucible, add slag-forming agents (including quartz, borax and sodium carbonate, with the addition amounts of each component being 2%, 2%, and 1.5% of the total weight of the melt, respectively); raise the temperature to 1100-1200℃, lower the outlet of the air blowing pipe to the middle position of the metal liquid surface, and blow air for oxidation for 1 hour, with an air volume of 0.3 m³ / h. 3 / min, which oxidizes Fe, matte and a small amount of Pb impurity metal to form oxides that are insoluble in the Cu-Au-Ag main metal, which accumulate on the surface of the melt in the form of slag, to form a liquid phase alloy containing Fe <1%, Pb 6% and S <1%. (4) Alloy water quenching beads: After cooling the liquid phase alloy obtained in step (3) to 910℃, slowly and uniformly inject it into a clean water tank. At the same time, when pouring in the high-temperature liquid phase alloy, use a high-pressure clean water pipe (water pressure controlled at 0.2MPa) to flush away the quenching beads. The water pipe outlet is 10cm away from the liquid phase alloy injection point. Continue until all the liquid phase alloy is poured in. Then filter to obtain flaky, copper, silver, gold and lead-containing water quenched alloy flakes with a particle size of less than 2mm, accounting for 90%. (5) Lead leaching of alloy sheet: The water-quenched alloy sheet obtained in step (4) is placed in a titanium reactor, and a 60% acetic acid solution is added. The liquid-solid ratio of the acetic acid solution to the alloy sheet is 1:5. The mixture is heated to 70-80℃ and allowed to stand for 2 hours. Then, stirring is started and the reaction continues for another 2 hours. After the reaction is completed, the lead-containing liquid and lead-free alloy sheet are obtained by filtration. The lead leaching rate is 96.5%. (6) Oxidation and depletion smelting of smelting slag: The smelting slag and oxidized refining slag produced in steps (2) and (3) are crushed to less than 10 mm, mixed with slag-forming agents (including crushed glass, sodium carbonate and borax, with the addition amounts of each component being 5%, 15%, and 5% of the total weight of the slag, respectively), and then loaded into an oxidation and depletion furnace for heating and blowing air for depletion smelting. The smelting temperature is controlled at 1300-1350℃, and the outlet of the air blowing pipe is lowered to the middle position of the slag. The blowing smelting is carried out for 1 hour with an air volume of 0.8 m³ / h. 3 / min, to obtain a low-density upper layer of oxidized lean slag containing Au<45g / t, mFe2%, the generated flue dust containing Pb10.5%, and a lower layer of liquid phase recovery alloy containing a small amount of lean slag, return to step (3) oxidative refining; (7) Lead leaching and filtration of flue dust: The flue dust generated in steps (2), (3), and (6) is collected and poured into a titanium reactor. A 70% acetic acid solution is added, with a liquid-to-solid ratio of acetic acid solution to flue dust of 1:8. The mixture is heated to 80°C and statically reacted for 2 hours. Then, stirring is started and the reaction continues for 3 hours. The mixture is filtered to obtain lead-containing liquid and lead-removed flue dust. The lead-removed flue dust is returned to step (6) for oxidation and smelting. The lead leaching rate reaches 95.5%.

[0045] (8) Lead recovery from lead-containing liquid precipitation: The lead-containing liquid obtained in steps (5) and (7) is transported to the lead precipitation reactor, stirring is started, and a 15% sodium sulfide solution is slowly added. The amount of sodium sulfide solution used is 1.2 times the amount of lead metal in the lead-containing liquid. After reacting for 0.5 hours, the solution is filtered to obtain lead concentrate and lead-removing liquid. The lead precipitation recovery rate is 99%, and the lead concentrate contains 62% lead.

[0046] (9) Wet gold separation: The lead-free alloy sheet obtained in step (5) is loaded into an acid-resistant and high-temperature-resistant titanium reactor. A small amount of water is added to wet the alloy sheet. Stirring and heating are started, and the temperature is controlled at 70-80℃. Sulfuric acid and nitric acid are added in batches for 3-5 separations. In the first 1-2 separations, sulfuric acid solution and nitric acid solution with a total weight of 4 times the alloy sheet are used, and the concentration of sulfuric acid solution is 98% and the concentration of nitric acid solution is 20%-25%. In the third-5 separations, nitric acid solution with a total weight of 3 times the alloy sheet is used, and the concentration of nitric acid solution is 30%-45%. No sulfuric acid is added until the solution color changes from dark blue to clear and there is no gas on the liquid surface. The wet gold separation operation is then considered to be over. Then, the gold powder with a purity of 99% and the filtrate are obtained by filtration. The filtrate is a mixed solution of copper nitrate, silver nitrate and copper sulfate. In this process, the consumption of nitric acid per alloy sheet is reduced from 3 kg / kg alloy sheet to 2.2 kg / kg alloy sheet, and the production efficiency is reduced from 3 days / batch to 1.5 days / batch.

[0047] (10) Oxidized lean slag grinding, cyanidation and magnetic separation: The oxidized lean slag obtained in step (6) is ground to 200 mesh (91%), cyanidation is performed to recover gold, and the tailings are fed into the magnetic separation process at a concentration of 33% to recover iron. The magnetic field strength is 0.5T. mFe decreases from 2% to 0.43%, and the magnetite concentrate recovery rate is 78.5%.

[0048] (11) Two-stage precipitation of mixed solution to recover fine gold particles: The copper nitrate-silver nitrate-copper sulfate mixed solution obtained in step (9) is sequentially fed into a closed primary precipitation tank and a two-stage precipitation tank, and flocculant 0.15 g / m³ is added to each of the two precipitation tanks. 3After secondary flocculation and sedimentation, fine gold particles were recovered, and the final Au content in the overflow was 0.02 mg / L, which is a significant reduction compared to 0.08 mg / L in the traditional method.

[0049] As can be seen from the above embodiments, the method described in this invention effectively removes impurities such as sulfur, iron, and lead from gold mud through high-temperature roasting, oxidation refining, and lead leaching, yielding gold powder with a purity of 99% or higher. Simultaneously, the removed lead and iron are recovered through acetic acid leaching + sulfide precipitation and magnetic separation processes, yielding lead concentrate and magnetite concentrate with recovery rates of 99% and 50% or higher, respectively. Furthermore, the "two-stage flocculation sedimentation and resin ball adsorption process" further achieves deep recovery of fine gold particles from the mixed solution, reducing the proportion of fine gold particles lost with the liquid phase. Therefore, this invention has significant advantages in comprehensive resource recovery and utilization, providing an economical, environmentally friendly, and efficient method for gold mud impurity removal and gold and silver extraction in the carbon-in-pulp process of polymetallic refractory gold mines, with broad application prospects.

[0050] The present invention has been described in detail above through specific and preferred embodiments. However, those skilled in the art should understand that the present invention is not limited to the embodiments described above. Any modifications, equivalent substitutions, etc., made within the spirit and principles of the present invention should be included within the protection scope of the present invention.

Claims

1. A highly efficient method for separating and extracting high-sulfur, high-lead gold mud using a carbon-in-pulp process, characterized in that, Includes the following steps: (1) High-temperature roasting: Saltpeter is added to the high-sulfur and high-lead gold mud with a water content of 50%-60% obtained by electrolytic desorption of gold-carrying carbon. After thorough mixing, the mixture is placed in an oven and heated to carry out dehydration, desulfurization, and decarbonization roasting reactions. (2) Gold mud slag making and smelting: After the dry gold mud after high temperature roasting in step (1) is fully mixed with the slag-making agent, it is put into the high-purity graphite crucible of the medium frequency furnace for high temperature slag making and smelting. After high temperature smelting, the low-density calcium-containing ferrosilicon smelting slag and the liquid phase alloy containing copper, lead, gold and silver are separated into layers, completing the initial separation of smelting slag and valuable metals. Then the upper layer of smelting slag is scraped into the slag tank, and the remaining liquid phase alloy layer is transferred to the next process. (3) Liquid phase alloy oxidation refining: Pour the liquid phase alloy obtained in step (2) into a silicon carbide-graphite crucible, add slag-forming agent, and heat up and blow gas to carry out oxidation refining. Copper matte is oxidized, and sulfur and some lead volatilize into the flue dust to obtain the upper layer of oxidation refining slag and the lower layer of liquid phase alloy with higher density. (4) Alloy water quenching beads: The upper layer of oxide refining slag obtained in step (3) is scraped into the slag tank. After the liquid phase alloy is cooled down to 900-1000℃, it is slowly and uniformly injected into the water tank. At the same time, when pouring in the high temperature liquid phase alloy, the high pressure water pipe is used to flush and disperse the quenching beads until all the liquid phase alloy is poured in. Then, the copper-silver-gold-lead water quenching alloy sheet with a large specific surface area is obtained by filtration. (5) Leaching and filtration of lead in alloy sheet: The water-quenched alloy sheet obtained in step (4) is loaded into a titanium reactor, industrial acetic acid is added, and the mixture is heated and allowed to react for 2 hours. Then, stirring is started and the reaction continues for 2 hours. After the reaction is completed, the lead-containing liquid and lead-free alloy sheet are obtained by filtration. (6) Oxidation and depletion smelting of smelting slag: The smelting slag and oxidative refining slag produced in steps (2) and (3) are crushed to less than 10 mm, mixed with slag-forming agent, and then loaded into an oxidation depletion furnace for heating and blowing gas to carry out depletion smelting. After the smelting is completed, a low-density upper layer of oxidation depletion slag and a lower layer of liquid phase recovery alloy containing a small amount of depletion slag are obtained. The upper layer of oxidation depletion slag is scraped to the slag tank, and the lower layer of liquid phase recovery alloy is returned to step (3) oxidation refining. (7) Lead leaching and filtration of smoke dust: The smoke dust generated in steps (2), (3) and (6) is collected and poured into a titanium reactor. Acetic acid is added, and the static reaction is heated for 2 hours. Then, stirring is started and the reaction continues for 3 hours. After the reaction is completed, lead-containing liquid and lead-removed smoke dust are obtained by filtration. The lead-removed smoke dust is returned to step (6) for oxidation and depletion smelting. (8) Lead-containing liquid precipitation and lead recovery: The lead-containing liquid obtained in steps (5) and (7) is transported to the lead precipitation reactor, stirring is started, and sodium sulfide solution is slowly added to react for 0.5 hours. After the reaction is completed, the lead concentrate product and lead-removing liquid are obtained by filtration. The lead concentrate is sold externally, and the lead-removing liquid is sent to the wastewater treatment system. (9) Wet gold separation: The lead-free alloy sheet obtained in step (5) is loaded into an acid-resistant and high-temperature resistant titanium reactor. First, a small amount of water is added to wet the alloy sheet. Stirring and heating are started, and the temperature is controlled at 60-80℃. Then, sulfuric acid and nitric acid are added in batches for 3-5 separations. In the first 1-2 separations, sulfuric acid solution and nitric acid solution with a total weight of 4 times the alloy sheet are used, and the concentration of sulfuric acid solution is 98% and the concentration of nitric acid solution is 20%-25%. In the third-5 separations, nitric acid solution with a total weight of 2-3 times the alloy sheet is used, and the concentration of nitric acid solution is 30%-45%. No sulfuric acid is added until the solution color changes from dark blue to clear and there is no gas on the liquid surface. The wet gold separation operation is then considered to be over. Then, gold powder and filtrate are obtained by filtration. The filtrate is a mixed solution of copper nitrate, silver nitrate and copper sulfate. (10) Oxidized lean slag grinding, cyanidation and magnetic separation: The oxidized lean slag obtained in step (6) is ground and cyanidated to recover gold. The tailings are then fed into the magnetic separation process to recover iron and obtain magnetite concentrate. (11) Two-stage precipitation of mixed solution to recover fine gold particles: The copper nitrate-silver nitrate-copper sulfate mixed solution obtained in step (9) is sequentially introduced into a closed primary precipitation tank and a two-stage precipitation tank. The bottom of the two precipitation tanks is conical and the top is filled with resin balls. When recovering fine gold particles, flocculant is added to the mixed solution below the resin balls. Under the action of flocculation and the interception of the resin balls, the fine gold particles agglomerate and precipitate at the bottom of the precipitation tank. After periodic discharge, filtration and smelting recovery, the remaining solution overflows from the top of the precipitation tank into the silver and copper recovery process.

2. The efficient separation and extraction method for high-sulfur and high-lead gold mud using carbon-in-pulp technology according to claim 1, characterized in that, In step (1), the amount of saltpeter used is 0.5%-1% of the amount of gold mud; the oven temperature is controlled at 700-800℃, and the static roasting reaction time is 12-14h.

3. The efficient separation and extraction method for high-sulfur and high-lead gold mud using carbon-in-pulp technology according to claim 1, characterized in that, The slag-forming agent mentioned in step (2) includes quartz, borax and sodium carbonate; the temperature of the high-temperature slag-forming smelting is controlled at 1200-1300℃; a smelting slag layer of 4-8cm is reserved above the liquid phase alloy layer when removing slag.

4. The efficient separation and extraction method for high-sulfur and high-lead gold mud using carbon-in-pulp technology according to claim 1, characterized in that, The slag-forming agent in step (3) includes quartz, sodium carbonate, and borax, with the addition amounts of each component being 2%-3%, 1%-2%, and 0.5%-1.5% of the total melt weight, respectively. During refining, the melt temperature is maintained at 1100-1300℃. During air blowing, the air distribution pipe outlet is lowered to the middle position of the molten metal surface, and the air blowing oxidation time per furnace is 0.5-1 hour, with an air volume of 0.2-0.3 m³ / h. 3 / min.

5. The efficient separation and extraction method for high-sulfur, high-lead gold mud using carbon-in-pulp technology according to claim 1, characterized in that, In step (5), the concentration of acetic acid solution is 50%-60%, and the liquid-solid ratio of acetic acid solution to alloy sheet is 1:4-5; the heating temperature is controlled at 60-80℃.

6. The efficient separation and extraction method for high-sulfur and high-lead gold mud using carbon-in-pulp process according to claim 1, characterized in that, The slag-forming agent in step (6) includes crushed glass, sodium carbonate, and borax, with the addition amounts of each component being 5%-8%, 10%-15%, and 2%-5% of the total weight of the molten slag, respectively, wherein the particle size of the crushed glass is less than 20mm; the smelting temperature is controlled at 1300-1500℃; during air blowing, the outlet of the air distribution pipe is lowered to the middle position of the molten slag, and each furnace is smelted by air blowing for 1 hour, with an air volume of 0.4-1m³. 3 / min.

7. The efficient separation and extraction method for high-sulfur and high-lead gold mud using carbon-in-pulp process according to claim 1, characterized in that, In step (7), the acetic acid concentration is 50%-70%, the liquid-solid ratio of the acetic acid solution to the dust is 1:6-8, and the heating temperature is controlled at 60-80℃.

8. The efficient separation and extraction method for high-sulfur and high-lead gold mud using carbon-in-pulp technology according to claim 1, characterized in that, In step (8), the concentration of sodium sulfide solution is 10%-15%, and the amount of sodium sulfide solution used is 1.2-1.5 times the amount of lead metal in the lead-containing solution.

9. The efficient separation and extraction method for high-sulfur and high-lead gold mud using carbon-in-pulp process according to claim 1, characterized in that, In step (10), the grinding particle size is -200 mesh, accounting for 90%-92%; the magnetic separation concentration of tailings is 30%-38%, and the magnetic field strength is 0.5T.

10. The efficient separation and extraction method for high-sulfur and high-lead gold mud using carbon-in-pulp technology according to claim 1, characterized in that, The diameter of the resin balls in step (11) is 1.5 cm; the dosage of the flocculant is 0.1-0.2 g / m³. 3 .

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