Efficient separation and extraction method for high-sulfur and high-lead gold mud in carbon slurry process
Through high-temperature roasting, slag smelting, liquid-phase alloy oxidation and refining and acetic acid leaching, the problem of difficult separation of impurities in high-sulfur and high-lead gold mud is solved, and the production of high-purity gold powder and efficient recycling of resources is achieved, and production costs and environmental risks are reduced.
Patent Information
- Application Number
- CN202510292656.3
- Authority / Receiving Office
- CN · China
- Patent Type
- Applications(China)
- Current Assignee / Owner
- Filing Date
- 2025-03-12
- Publication Date
- 2025-07-11
AI Technical Summary
Traditional processes are difficult to effectively separate and recover valuable metals in high-sulfur and high-lead gold mud, resulting in low purity of gold powder, difficult to remove impurities, and occupational health hazards, fine-grained gold is easy to lose, and it is difficult to deal with smoke and dust.
The steps of high-temperature roasting, slag smelting, liquid-phase alloy oxidation and refining, water-quenching beads, lead leaching of alloy sheets and acetic acid leaching, combined with acetic acid leaching and vulcanization precipitation processes, fine gold is recovered through multiple nitric acid separation and flocculation precipitation, so that the deep removal of impurities and efficient extraction of valuable metals can be achieved.
It has achieved efficient removal of impurities such as sulfur, iron, lead and other impurities in gold mud, and obtained high-purity gold powder, which has reduced occupational health hazards, recovered fine-grained gold and lead, improved resource utilization, and reduced production costs and environmental impact.
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Figure CN120290888A_ABST
Abstract
Description
Technical Field
[0001] The present invention belongs to the technical field of precious metal metallurgy, and particularly relates to an efficient separation and extraction method for high-sulfur and high-lead gold mud in a carbon-in-pulp process. Background Art
[0002] In recent years, with the increasing depletion of easily-selectable metallurgical ore resources, refractory and difficult-to-smelt ores such as those associated with low-grade gold and high lead in magnetite and limonite, or those with relatively high sulfur content in the oxygen-sulfur mixed zone, which are characterized by "poor, fine, and complex", are continuously mined and used. Such ores mainly adopt the technical route of "grinding - cyanidation leaching and activated carbon adsorption - magnetic separation - desorption and electrolysis of gold-loaded carbon - purification and refining of gold mud - crude gold" to extract valuable metals such as gold, silver, and iron. Due to the increase in the content of Pb, S, Fe, etc. in the raw ore, activated carbon in the carbon-in-pulp process will adsorb a large amount of Pb, S, and Fe while adsorbing Au and Ag. Subsequently, after desorption and electrolysis treatment, they are enriched in the crude gold mud product. There are many problems in separating and extracting Au and Ag from high-Pb, S, Fe gold mud using traditional processes, mainly manifested as follows: First, it is difficult to quench and splash the liquid-phase alloy into beads, and it is difficult to form alloy flakes. In the traditional "slag smelting - splashing beads - nitric acid gold parting" process, the quality of quenched and splashed beads is the key to the efficiency of nitric acid gold parting and ensuring the fineness of gold powder. In the liquid-phase alloy obtained by slag smelting of high-S, Fe gold mud, in addition to containing Au, Ag, and Cu, there will be residual Fe that has not been separated by slag formation, and sometimes matte (a eutectic melt of Cu2S - FeS) will be formed. When this metal liquid phase is quenched and splashed into beads, gold and silver will be wrapped by these impurities such as Fe3O4 or matte. The quenched product is difficult to form alloy flakes with a large specific surface area, and 80% is in the shape of round beads, which is not conducive to subsequent nitric acid gold parting; for gold mud smelting slag, traditional induction furnace secondary heating and impoverishment smelting are generally used, but it is still difficult to separate Fe by slag formation, and it will co-melt with the liquid phase. 90% of the quenched products of the high-temperature metal eutectic are in the shape of round beads, resulting in unqualified gold powder in subsequent nitric acid gold parting.
[0003] Second, the impurity removal efficiency of gold powder is low. Low-Fe, Pb, S gold mud generally can obtain gold powder with a purity of more than 99% by using the "slag smelting - splashing beads - nitric acid gold parting" process. However, when round-bead alloys wrapped by iron or matte are subjected to nitric acid gold parting, not only is the gold parting time greater than 3 days, the nitric acid unit consumption is high (one part of alloy flakes consumes 6 - 8 parts of 50% concentrated nitric acid), but also the gold powder purity is low, with only 85% - 90% Au content and up to 10% impurity Pb content, and qualified gold powder cannot be obtained.
[0004] Third, S, Pb, and Fe in the gold mud cannot be effectively removed and recovered. During the gold mud smelting process, matte (a eutectic melt of Cu2S - FeS) will be formed between the slag layer and the metal layer, forming entrained gold and silver liquid beads. Fe is also difficult to separate by slag formation. About 87% of Pb enters the wastewater and about 13% enters the smelting slag. S, Fe, and Pb cannot be effectively removed and recovered, resulting in metal loss.
[0005] Fourth, the traditional wet gold separation process treats high-sulfur and high-lead gold mud, and there is fine gold loss with the liquid phase during gold powder filtration. As the gold embedding characteristics in the source ore become finer, the gold embedding particle size of deep-mined ores with pyrite and magnetite as carrier minerals or gold-bearing ores in the oxygen-sulfur mixed zone tends to be finer, mainly distributed below 10μm. The gold in the final secondary metallurgical process also shows the characteristics of fine embedding particle size. In the process of nitric acid separation of gold and gold powder filtration, this part of fine gold easily penetrates the filter cloth and is lost with the liquid phase or with the iron colloid suspension. Atomic spectroscopy absorption analysis shows that the gold concentration in the liquid phase is generally as high as 0.08mg / L.
[0006] Fifth, high-lead smelting smoke is a serious hazard to the occupational health of employees during the depletion smelting process. When the smoke generated by gold mud smelting is repeatedly depleted and smelted in a medium-frequency furnace to recover Au, Ag, and Cu, the smelting temperature is generally between 1300°C and 1600°C. At this time, as the temperature rises, Pb is more likely to volatilize in large quantities, resulting in a higher risk of occupational hazards for operators.
[0007] It can be seen that the traditional "slag smelting-bead pouring-nitric acid gold separation" process can no longer meet the separation and extraction requirements of high-Pb, S, and Fe gold mud. Therefore, the present invention provides an efficient separation and extraction method for high-sulfur and high-lead gold mud to improve the separation and extraction effect of valuable metals in gold mud, and remove and recover S, Pb, and Fe while efficiently separating gold. Summary of the invention
[0008] In view of the above problems, the present invention provides a method for efficiently separating and extracting high-sulfur and high-lead gold mud in a carbon slurry process.
[0009] The specific technical solution is: a method for efficiently separating and extracting high-sulfur and high-lead gold mud in a carbon slurry process, comprising the following steps: (1) High temperature roasting: Add saltpeter to the high-sulfur and high-lead gold mud with a water content of 50%-60% obtained by electrolytic desorption of gold-loaded carbon, mix thoroughly, and then put it into an oven to heat up for dehydration, desulfurization, and decarbonization roasting reactions; The process first dehydrates, decarbonizes and cyanides the gold mud by high temperature decomposition. - When the water removal temperature reaches 400℃, the sulfur in the gold mud reacts with saltpeter to generate sulfur dioxide gas, and the desulfurization rate increases with the increase of temperature. After roasting for 12-14h, most of the sulfur is removed, which can effectively reduce the sulfur entering the smelting process and reduce the generation of smelting matte.
[0010] (2)Smelting of gold mud to form slag: After the dried gold mud obtained from the high-temperature roasting in step (1) is fully mixed with the slag-forming agent, it is charged into a high-purity graphite crucible of an intermediate-frequency furnace for high-temperature slag-forming smelting. After high-temperature smelting, the low-density smelting slag containing calcium silico-ferrite and the liquid-phase alloy containing copper, lead, gold, and silver are layered, completing the preliminary separation of the smelting slag and valuable metals. Then, the upper-layer smelting slag is skimmed into the slag tank, and the remaining liquid-phase alloy layer (including a 4-8 cm smelting slag layer or matte layer and the liquid-phase alloy layer below) is transferred to the next process; This process utilizes the efficient capture of Au and Ag by Cu and Pb, and through high-temperature slag-forming smelting, Cu-Pb-Au-Ag is preliminarily separated from impurities such as silicon, calcium, and iron.
[0011] (3)Oxidative refining of liquid-phase alloy: The liquid-phase alloy obtained in step (2) is poured into a silicon carbide-graphite crucible, the slag-forming agent is added, and oxidation refining is carried out by heating and blowing air. The matte is oxidized, and sulfur and part of the lead volatilize into the fume and dust, obtaining the upper-layer oxidative refining slag (including iron-removing slag and low-density slag of calcium silicate and iron silicate) and the lower-layer liquid-phase alloy with a larger density; The purpose of oxidative refining is to deeply remove impurities such as iron and matte in the liquid-phase alloy, creating good conditions for subsequent bead splashing. The basic principle of this method is based on the different affinities of metals for oxygen. Air is introduced to oxidize Fe, matte, and a small amount of Pb impurity metals to form oxides that are insoluble in the main Cu-Au-Ag metals, aggregating on the surface of the melt in the form of slag, and S is further separated in a gaseous form. Moreover, by proportioning the slag-forming agent to form a low-melting-point slag, sodium carbonate can also enhance the fluidity of the slag, facilitating the floating of impurities.
[0012] (4)Water quenching and bead splashing of alloy: The upper-layer oxidative refining slag obtained in step (3) is skimmed into the slag tank. After the liquid-phase alloy is cooled to 900 - 1000 °C, it is slowly and evenly injected into the water tank. At the same time, when pouring the high-temperature liquid-phase alloy, a high-pressure water pipe (water pressure controlled at 0.15 - 0.2 MPa) is aimed at the injection point to disperse and quench the beads, and the water pipe outlet is at a position 10 cm away from the liquid-phase alloy injection point until all the liquid-phase alloy is poured. Then, a water-quenched alloy sheet with a larger specific surface area containing copper-silver-gold-lead is obtained by filtration; This process has a great influence on subsequent wet gold separation and the fineness of gold powder. Through research, when the temperature of the Cu-Pb-Au-Ag liquid-phase alloy is 900 - 1000 °C and the water pressure is 0.15 - 0.2 MPa, the bead-splashing effect is the best, and most of them are in fine sheet form with a large specific surface area.
[0013] (5)Leaching and filtration of lead from alloy sheet: The water-quenched alloy sheet obtained in step (4) is loaded into a titanium reaction kettle, industrial acetic acid is added, and after heating and waiting for the reaction for 2 h, stirring is started and the reaction continues for 2 h. After the reaction ends, a lead-containing solution and a lead-removed 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 react with acetic acid, to achieve the purpose of removing lead and prevent the lead content in the gold powder from exceeding the standard. The main chemical formula is: 2CH3COOH + Pb == Pb(CH3COO)2 + H2↑.
[0014] (6)Smelting slag oxidative impoverishment smelting: The smelting slag and oxidative refining slag produced in steps (2) and (3) are crushed to less than 10 mm, mixed with a slag-forming agent, and then charged into an oxidative impoverishment furnace for heating and blowing to carry out impoverishment smelting. After the smelting is completed, a low-density upper-layer oxidative impoverished slag (Au grade < 50 g / t) and a lower-layer liquid-phase recycled alloy containing a small amount of impoverished slag are obtained. The upper-layer oxidative impoverished slag is raked into the slag tank, and the lower-layer liquid-phase recycled alloy is returned to step (3) for oxidative refining; The main purpose of this process is to further recover the residual Au and Ag in the smelting slag. By reasonably controlling the slag-forming agent, slag type, and temperature, and introducing air, impurities such as Ca-Si-Fe are made to form a low-density impoverished slag that floats to the liquid surface to separate from Au-Ag. The introduction of air can accelerate the oxidation and floating of Ca-Si-Fe. Sodium carbonate can enhance the fluidity of the slag, and silicon dioxide can react with iron to form a low-density and low-melting-point slag of iron silicate. Due to the silicate contained in the material itself, under the condition of this ratio of materials, a weakly acidic slag with low viscosity and good fluidity is formed, and iron does not form magnetite.
[0015] (8)Lead leaching and filtration of soot: The soot produced in steps (2), (3), and (6) is collected and poured into a titanium reaction kettle, acetic acid is added, and after heating and static reaction for 2 h, stirring is started and the reaction continues for 3 h. After the reaction is completed, a lead-containing solution and lead-removed soot are obtained by filtration. The lead-removed soot is returned to step (6) for oxidative impoverishment smelting; This process utilizes the fact that gold, silver, and copper do not react with acetic acid, while lead does react with acetic acid, to achieve the purpose of pre-leaching and separating lead in the soot. The main chemical formula is: 2CH3COOH + Pb == Pb(CH3COO)2 + H2↑.
[0016] (8)Precipitation and recovery of lead from the lead-containing solution: The lead-containing solution obtained in steps (5) and (7) is transported to a lead precipitation reaction kettle, stirring is started, and a sodium sulfide solution is slowly added and reacted for 0.5 h. After the reaction is completed, lead concentrate products and lead-removed solution are obtained by filtration. The lead concentrate is sold externally, and the lead-removed solution enters the wastewater treatment system; The lead ions in the lead-containing solution undergo a sulfide precipitation reaction with the sodium sulfide solution. The sulfide precipitation method utilizes the reaction of heavy metal ions with S 2- to form M2S X precipitates. According to the differences in the solubility products 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 recovering heavy metals.
[0017] (9) Wet separation of gold: Load the de-leaded alloy sheet obtained in step (5) into an acid-resistant and high-temperature-resistant titanium reaction kettle. First, moisten the alloy sheet with a small amount of clear water, start stirring and heating, control the temperature at 60 - 80 °C, and then add sulfuric acid and nitric acid in portions for 3 - 5 separations; during the 1 - 2 separations, use sulfuric acid solution and nitric acid solution that are 4 times the total weight of the alloy sheet, and the concentration of the sulfuric acid solution is 98% and the concentration of the nitric acid solution is 20% - 25%; during the 3 - 5 separations, use nitric acid solution that is 2 - 3 times the total weight of the alloy sheet, and the nitric acid solution is 30% - 45%, without adding sulfuric acid, until the color of the solution changes from dark blue to clear and there is no gas on the liquid surface, then it is regarded as the end of the wet separation of gold operation; then filter to obtain gold powder and filtrate, where the filtrate is a mixed solution of copper nitrate - silver nitrate - copper sulfate; This process utilizes the strong acidic characteristic of sulfuric acid, and under the multiple actions of sulfuric acid + nitric acid + high temperature + stirring, etc., it accelerates the dissolution of Fe, Ag, and Cu in the alloy sheet.
[0018] (10) Oxidized lean slag grinding, cyanidation and magnetic separation: Grind the oxidized lean slag obtained in step (6), conduct cyanidation leaching to recover gold, and the tailings enter the magnetic separation process to recover iron to obtain magnetite concentrate; (11) Two-stage precipitation of the mixed solution to recover fine-grained gold: The copper nitrate - silver nitrate - copper sulfate mixed solution obtained in step (9) enters a closed first-stage precipitation tank and a two-stage precipitation tank in sequence. The bottoms of the two precipitation tanks are conical, and the tops are filled with resin balls. When recovering fine-grained gold, add a flocculant to the mixed solution below the resin balls. The fine-grained gold agglomerates and precipitates at the bottom of the precipitation tank under the action of flocculation and the interception of the resin balls, and is regularly discharged, filtered and then smelted for recovery. The remaining solution overflows from the upper part of the precipitation tank and enters the silver and copper recovery process.
[0019] Further, in step (1), the dosage of saltpeter is 0.5% - 1% of the amount of gold mud; the oven temperature is controlled at 700 - 800 °C, and the static roasting reaction time is 12 - 14 h.
[0020] Further, in step (2), the slag-making agent includes quartz, borax and sodium carbonate; to prevent the volatilization of lead, the temperature of the high-temperature slag-making smelting should not be too high and should be controlled at 1200 - 1300 °C; to prevent valuable metals such as gold and silver from being entrained with the slag during slag skimming, a 4 - 8 cm thick smelting slag layer is reserved above the liquid alloy layer during slag skimming.
[0021] Furthermore, the slag-forming agent in step (3) includes quartz, sodium carbonate and borax, and the addition amount of each component is 2%-3%, 1%-2% and 0.5%-1.5% of the total weight of the melt, respectively. This ratio is conducive to reducing the viscosity of the slag, enhancing the fluidity, and making a reasonable slag shape; during refining, the melt temperature is maintained at 1100-1300°C; during blowing, the outlet of the air distribution blowpipe is lowered to the middle position of the metal liquid surface, the blowing oxidation time of each furnace is 0.5-1h (this time is the key to deironing and matte blowing), and the air volume is 0.2-0.3m 3 / min.
[0022] Furthermore, in step (4), the water pressure of the high-pressure clean water pipe is controlled at 0.15-0.2 MPa, and the outlet of the water pipe is located 10 cm away from the injection point of the liquid alloy. The water flow entering the water pipe should ensure that the bottom of the water quenching tank can form a stirring rotation.
[0023] Furthermore, in step (5), the concentration of acetic acid is 50%-60%, the liquid-to-solid ratio of the acetic acid solution to the alloy sheet is 1:4-5; and the heating temperature is controlled at 60-80°C.
[0024] Furthermore, the slag-forming agent in step (6) comprises cullet, sodium carbonate and borax, and the addition amount of each component is 5%-8%, 10%-15% and 2%-5% of the total weight of the slag, respectively, wherein the particle size of the cullet is less than 20 mm; the smelting temperature is controlled at 1300-1500°C; when blowing, the outlet of the air distribution blowpipe is lowered to the middle position of the slag, and each furnace is blown for 1 hour, and the air volume is 0.4-1m 3 / min.
[0025] Furthermore, in step (7), the concentration of acetic acid is 50%-70%, the liquid-to-solid ratio of the acetic acid solution to the smoke is 1:6-8; and the heating temperature is controlled at 60-80°C.
[0026] Furthermore, in step (8), the concentration of the sodium sulfide solution is 10%-15%, and the amount of the sodium sulfide solution used is 1.2-1.5 times the amount of lead metal in the lead-containing liquid.
[0027] Furthermore, in step (10), the grinding particle size is -200 mesh accounting for 90%-92%; the tailings magnetic separation concentration is 30%-38%, and the magnetic field strength is 0.5T.
[0028] Furthermore, in step (11), the diameter of the resin ball is 1.5 cm; the amount of the flocculant is 0.1-0.2 g / m 3 .
[0029] Advantages of the present invention: The present invention can effectively remove impurities such as sulfur, iron, and lead contained in gold slime, ensure that more than 90% of the water-quenched and splashed beads are flaky alloys, and efficiently obtain gold powder with a fineness of 99% or more after wet separation. At the same time, it also realizes the removal of lead, iron, and the recovery of fine-grained gold in the mixed solution. The comprehensive resource recovery and utilization effect is remarkable, providing an economical, environmentally friendly and efficient method for impurity removal and gold and silver extraction from gold slime in the carbon-in-pulp process of multi-metal refractory gold mines, and having broad application prospects. The technical advantages of the present invention are specifically embodied as follows: (1) Before slag-making smelting, the present invention first removes most of the sulfur impurities in gold slime through high-temperature roasting to reduce the formation of matte when sulfur enters the smelting process. Secondly, the liquid-phase alloy obtained from slag-making smelting is subjected to oxidative refining to deeply remove impurities such as S, Fe, Pb, and matte. Further, by reasonably controlling the splashing temperature and water pressure of the beads, the problem that the liquid-phase alloy beads are seriously wrapped by impurities such as matte and present a spherical shape is completely solved, and flaky alloys with a proportion of more than 90% are produced. Its large specific surface area creates favorable conditions for the full reaction of subsequent nitric acid gold parting.
[0030] (2) Before wet gold parting, a lead leaching link for alloy flakes is added, and acetic acid is used to leach lead in the alloy flakes in advance, solving the problem that lead and lead oxides are tightly wrapped with gold and do not react with nitric acid, resulting in high lead content in the gold powder product. In addition, in the wet gold parting link, by controlling the concentration and addition amount of sulfuric acid and nitric acid and separating through multiple reactions, gold powder products with a fineness of up to 99% are separated while reducing the consumption of reagents.
[0031] (3) For lead in soot and alloy flakes, through the "acetic acid leaching + sulfide precipitation" process, not only the removal of lead is realized, but also the recovery of lead is further achieved, obtaining lead concentrate products that can be sold externally, creating value while also reducing subsequent treatment problems and environmental problems caused by lead, and having good economic and social benefits.
[0032] (4) The present invention fully recovers the intermediate products - liquid-phase recovery alloy, oxidative refining slag, and lead-removed soot by returning them to the corresponding links for cyclic treatment; for the intermediate product - depleted slag, the residual gold is also recovered by grinding and cyanidation, and magnetite in the cyanidation tailings is further recovered by magnetic separation; especially for fine-grained gold, a "two-stage flocculation precipitation and resin ball adsorption process" is set up for deep recovery, all of which fully reflect the advantages of comprehensive resource recovery and utilization of the present invention. Description of the Drawings
[0033] Figure 1 is a flow chart of an efficient separation and extraction method for high-sulfur and high-lead gold slime in a carbon-in-pulp process of the present invention. Detailed Embodiments
[0034] In order to make the technical problems and technical solutions solved by the present invention clearer and more understandable, 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 only used to explain the present invention and are not used to limit the present invention. Example 1
[0035] The gold mud of a polymetallic gold mine in southern China is obtained by desorbing and electrolyzing the 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% respectively.
[0036] As Figure 1 shown, the method of the present invention is used to process the gold mud, and the specific steps are as follows: (1) High-temperature roasting: Potassium nitrate is added to the gold mud, and the amount of potassium nitrate is 0.6% of the amount of gold mud. After fully mixing, it is loaded into an oven, and the temperature is raised to 750°C for dehydration, desulfurization, and decarbonization roasting reactions. The roasting duration is 12 hours, and the desulfurization rate reaches 85.6%; (2) Gold mud slag-making smelting: The dry gold mud after high-temperature roasting in step (1) is fully mixed with a slag-making agent (quartz + borax + sodium carbonate), and then loaded into a high-purity graphite crucible of an intermediate-frequency furnace for high-temperature slag-making smelting. The smelting temperature is controlled at 1200 - 1250°C, and the smelting time is 0.5 hours. The low-density calcium-containing ferrosilicon smelting slag and the liquid alloy containing copper, lead, gold, and silver are layered, completing the preliminary separation of the smelting slag and valuable metals. Among them, the yield of the smelting slag is 15%, the yield of the liquid alloy is 84%, and it contains Fe 2.3% and Pb 8.5%.
[0037] (3) Oxidative refining of the liquid alloy: The liquid alloy obtained in step (2) is poured into a silicon carbide-graphite crucible, and a slag-making agent is added (including quartz, borax, and sodium carbonate, and the addition amounts of each component are 2%, 1%, and 0.5% of the total weight of the melt respectively); the temperature is raised to 1100 - 1200°C, and the outlet of the air blowing pipe is lowered to the middle position of the metal liquid surface, and air is blown for oxidation for 0.5 hours, and the air volume is 0.25 m 3 / min, so that Fe, matte, and a small amount of Pb impurity metals are oxidized to form oxides insoluble in the main Cu-Au-Ag metals, and are aggregated on the surface of the melt in the form of slag, obtaining a liquid alloy containing Fe 1% and Pb 7%; (4) Alloy water quenching and bead splashing: After the liquid alloy obtained in step (3) is cooled to 900°C, it is slowly and evenly injected into a water tank. At the same time, when pouring the high-temperature liquid alloy, a high-pressure water pipe (water pressure controlled at 0.16 MPa) is used to aim at and disperse the quenched beads at a position 10 cm away from the liquid alloy injection point until all the liquid alloy is poured. Then, filter to obtain water-quenched alloy flakes containing copper, silver, gold, and lead with a flake shape and a particle size less than 2 mm, accounting for 92%; (5)Leaching of lead from alloy flakes: The water-quenched alloy flakes obtained in step (4) are loaded into a titanium reactor, and acetic acid solution with a concentration of 52% is added. The liquid-solid ratio of the acetic acid solution to the alloy flakes is 1:4. It is heated to 60 - 70 °C and left to react for 2 h. Then, stirring is started and the reaction continues for another 2 h. After the reaction ends, a lead-containing solution and lead-depleted alloy flakes are obtained by filtration, and the lead leaching rate is 98.1%. (6)Oxidative impoverishment 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 agents (including cullet, sodium carbonate, and borax, and the addition amounts of each component are 5%, 12%, and 3% of the total weight of the molten slag respectively), and then loaded into an oxidative impoverishment furnace. The temperature is raised and air is blown in for impoverishment smelting. The smelting temperature is controlled at 1300 - 1400 °C. The outlet of the air blowing pipe is lowered to the middle position of the molten slag, and air is blown for smelting for 1 h. The air volume is 0.45 m 3 / min, obtaining upper-layer oxidative impoverished slag with low density, containing Au < 40 g / t and mFe 1%. The produced fume contains Pb 6.5%, and a lower-layer liquid-phase recycled alloy containing a small amount of impoverished slag is returned to step (3) for oxidative refining. (7)Leaching and filtration of lead from fume: The fume produced in steps (2), (3), and (6) is collected and poured into a titanium reactor. Acetic acid solution with a concentration of 70% is added. The liquid-solid ratio of the acetic acid solution to the fume is 1:8. It is heated to 75 °C and left to react statically for 2 h. Then, stirring is started and the reaction continues for 3 h. After filtration, a lead-containing solution and lead-depleted fume are obtained. The lead-depleted fume is returned to step (6) for oxidative impoverishment smelting; the lead leaching rate reaches 94.2%.
[0038] (8)Precipitation and recovery of lead from the lead-containing solution: The lead-containing solutions obtained in steps (5) and (7) are transported to a lead precipitation reactor. Stirring is started, and a sodium sulfide solution with a concentration of 14% is slowly added. The dosage of the sodium sulfide solution is 1.4 times the amount of lead metal in the lead-containing solution. After reacting for 0.5 h, filtration is carried out to obtain lead concentrate products and lead-depleted solution. The lead precipitation recovery rate is 99%, and the lead concentrate contains 60% lead.
[0039] (9)Wet gold separation: Load the de-leaded alloy sheets obtained in step (5) into an acid-resistant and high-temperature-resistant titanium reactor. First, add a small amount of clear water to moisten the alloy sheets, start stirring and heating, control the temperature at 60 - 70 °C, and then add sulfuric acid and nitric acid in portions for 3 - 5 separations; during the 1 - 2 separations, use sulfuric acid solution and nitric acid solution that are 4 times the total weight of the alloy sheets, with the sulfuric acid solution concentration being 98% and the nitric acid solution concentration being 20% - 25%; during the 3 - 5 separations, use nitric acid solution that is 2 times the total weight of the alloy sheets, and the nitric acid solution is 30% - 45%, without adding sulfuric acid until the solution color changes from dark blue to clear and there is no gas on the liquid surface, then it is regarded as the end of the wet gold separation operation; then filter to obtain gold powder with a fineness of 99.5% and filtrate, where the filtrate is a mixed solution of copper nitrate - silver nitrate - copper sulfate. During this process, the single consumption of nitric acid corresponding to the alloy sheets drops from the original 3 kg / kg alloy sheets to 2 kg / kg alloy sheets, and the production efficiency drops from the original 3 days / batch to 2 days / batch.
[0040] (10)Oxidized lean slag grinding, cyanidation and magnetic separation: Grind the oxidized lean slag obtained in step (6) to 92% passing 200 mesh, conduct cyanidation leaching to recover gold, and the tailings enter the magnetic separation process at a concentration of 31% to recover iron, with a magnetic field intensity of 0.5 T. mFe drops from 1% to 0.5%, and the magnetite concentrate recovery rate is 50%.
[0041] (11)Two-stage precipitation of the mixed solution to recover fine gold: Feed the copper nitrate - silver nitrate - copper sulfate mixed solution obtained in step (9) into a closed first-stage precipitation tank and two-stage precipitation tank in sequence, and add 0.1 g / m 3 of flocculant to each of the two precipitation tanks. After secondary flocculation precipitation, fine gold is recovered, and the Au content in the final overflow liquid is 0.01 mg / L, showing a significant decrease compared to 0.08 mg / L of the traditional method. Example 2
[0042] The gold mud from a polymetallic gold mine in the south of China, which is obtained by desorbing and electrolyzing the gold-loaded carbon in a carbon-in-pulp plant. It mainly contains 5.20% Au, 20.53% Ag, 32.16% Cu, 10% Pb, 2.2% Fe, and 5.36% S.
[0043] (1)High-temperature roasting: Add saltpeter to the gold mud, with the amount of saltpeter being 1% of the amount of gold mud. After fully mixing, load it into an oven, raise the temperature to 800 °C for dehydration, desulfurization, and decarbonization roasting reactions, and the roasting duration is 14 h to achieve a desulfurization rate of 60%; (2)Smelting of gold mud to form slag: After fully mixing the dried gold mud obtained from the high-temperature roasting in step (1) with a slag-forming agent (quartz + borax + sodium carbonate), it is charged into a high-purity graphite crucible in an intermediate-frequency furnace for high-temperature slag-forming smelting. The smelting temperature is controlled at 1250 - 1300 °C, and smelting is carried out for 0.52 h. After smelting, a low-density smelting slag containing calcium, silicon, and iron and a liquid-phase alloy layer containing copper, lead, gold, and silver are formed. Moreover, a grayish-white matte layer with a thickness of about 2 - 3 cm is generated between the slag layer and the liquid-phase layer. Among them, the yield of the smelting slag is 15%, the yield of the liquid-phase alloy is 82%, and it contains 2.5% Fe and 8.5% Pb.
[0044] (3)Oxidative refining of the liquid-phase alloy: Pour the liquid-phase alloy obtained in step (2) into a silicon carbide - graphite crucible, and add a slag-forming agent (including quartz, borax, and sodium carbonate, and the addition amounts of each component are 2%, 2%, and 1.5% of the total weight of the melt respectively); heat up to 1100 - 1200 °C, lower the outlet of the air blowing pipe to the middle position of the metal liquid surface, and blow air for oxidation for 1 h, with the air volume being 0.3 m 3 / min, so that Fe, matte, and a small amount of Pb impurity metals are oxidized to form oxides that are insoluble in the main Cu - Au - Ag metals and aggregate on the surface of the melt in the form of slag, resulting in a liquid-phase alloy containing Fe < 1%, Pb 6%, and S < 1%; (4)Water quenching and bead splashing of the alloy: After cooling and reducing the temperature of the liquid-phase alloy obtained in step (3) to 910 °C, slowly and uniformly pour it into a water tank. At the same time, when pouring the high-temperature liquid-phase alloy, use a high-pressure water pipe (with the water pressure controlled at 0.2 MPa) to aim at and disperse the quenched beads at a position 10 cm away from the injection point of the liquid-phase alloy until all the liquid-phase alloy is poured. Then, filter to obtain water-quenched alloy flakes containing copper, silver, gold, and lead, with a particle size less than 2 mm, accounting for 90%; (5)Leaching of lead from the alloy flakes: Load the water-quenched alloy flakes obtained in step (4) into a titanium reaction kettle, add acetic acid solution with a concentration of 60%. The liquid-solid ratio of the acetic acid solution to the alloy flakes is 1:5. Heat to 70 - 80 °C, wait for the reaction for 2 h, then start stirring and continue the reaction for 2 h. After the reaction ends, filter to obtain a lead-containing solution and lead-removed alloy flakes, and the lead leaching rate is 96.5%; (6)Oxidative depletion smelting of the smelting slag: Crush the smelting slag and oxidative refining slag generated in steps (2) and (3) to less than 10 mm, mix them with a slag-forming agent (including broken glass, sodium carbonate, and borax, and the addition amounts of each component are 5%, 15%, and 5% of the total weight of the molten slag respectively), then charge them into an oxidative depletion furnace, heat up and blow air for depletion smelting. The smelting temperature is controlled at 1300 - 1350 °C, lower the outlet of the air blowing pipe to the middle position of the molten slag, and blow air for smelting for 1 h, with the air volume being 0.8 m 3 / min to obtain a low-density upper oxidized lean slag containing Au < 45 g / t and mFe 2%, the generated fume contains Pb 10.5%, and a lower-layer liquid-phase recycled alloy containing a small amount of depleted slag is returned to step (3) for oxidative refining; (7) Lead leaching and filtration of fume: The fume generated in steps (2), (3), and (6) is collected and poured into a titanium reaction kettle, and acetic acid solution with a concentration of 70% is added. The liquid-solid ratio of the acetic acid solution to the fume is 1:8. After heating to 80 °C and static reaction for 2 h, stirring is started and the reaction continues for 3 h. After filtration, a lead-containing solution and lead-depleted fume are obtained, and the lead-depleted fume is returned to step (6) for oxidative depletion smelting; the lead leaching rate reaches 95.5%.
[0045] (8) Precipitation and recovery of lead from the lead-containing solution: The lead-containing solution obtained in steps (5) and (7) is transported to a lead precipitation reaction kettle, stirring is started, and a sodium sulfide solution with a concentration of 15% is slowly added. The dosage of the sodium sulfide solution is 1.2 times the amount of lead metal in the lead-containing solution. After reacting for 0.5 h, filtration is carried out to obtain lead concentrate products and lead-depleted solution. The precipitation recovery rate of lead is 99%, and the lead concentrate contains 62% lead.
[0046] (9) Wet separation of gold: The lead-depleted alloy sheets obtained in step (5) are loaded into an acid-resistant and high-temperature-resistant titanium reaction kettle. First, add a small amount of clear water to moisten the alloy sheets, start stirring and heating, control the temperature at 70 - 80 °C, and then add sulfuric acid and nitric acid in portions for 3 - 5 separations; during the 1 - 2 separations, sulfuric acid solution and nitric acid solution with a total weight 4 times that of the alloy sheets are used, and the concentration of the sulfuric acid solution is 98% and the concentration of the nitric acid solution is 20% - 25%; during the 3 - 5 separations, nitric acid solution with a total weight 3 times that of the alloy sheets is used, and the nitric acid solution is 30% - 45%, without adding sulfuric acid. When the color of the solution changes from dark blue to clear and there is no gas on the liquid surface, it is considered that the wet separation of gold operation is completed; then filtration is carried out to obtain gold powder with a fineness of 99% and a filtrate, and the filtrate is a mixed solution of copper nitrate - silver nitrate - copper sulfate. During this process, the single consumption of nitric acid corresponding to the alloy sheets is reduced from the original 3 kg / kg alloy sheets to 2.2 kg / kg alloy sheets, and the production efficiency is reduced from the original 3 days / batch to 1.5 days / batch.
[0047] (10) Grinding, cyanidation and magnetic separation of oxidized lean slag: The oxidized lean slag obtained in step (6) is ground to 91% passing 200 mesh, and gold is recovered by cyanidation leaching. The tailings enter the magnetic separation process at a concentration of 33% to recover iron, and the magnetic field intensity is 0.5 T. mFe is reduced from 2% to 0.43%, and the recovery rate of magnetite concentrate is 78.5%.
[0048] (11) Two-stage precipitation of the mixed solution to recover fine-grained gold: The copper nitrate - silver nitrate - copper sulfate mixed solution obtained in step (9) enters a closed first-stage precipitation tank and two-stage precipitation tank in sequence, and 0.15 g / m of flocculant is added to each of the two precipitation tanks 3, after secondary flocculation precipitation, fine-grained gold is recovered. The Au content in the final overflow liquid is 0.02 mg / L, showing a significant decrease compared with 0.08 mg / L of the traditional method.
[0049] As can be seen from the above embodiments, when using the method of the present invention to treat gold slime, impurities such as sulfur, iron, and lead contained in the gold slime can be effectively removed through processes such as high-temperature roasting, oxidative refining, and lead leaching, obtaining gold powder with a fineness of up to 99% and above. At the same time, the recovered lead and iron are achieved through the "acetic acid leaching + sulfide precipitation" and magnetic separation processes, obtaining lead concentrate and magnetite concentrate, and the recovery rates are 99% and over 50% respectively. In addition, through the "two-stage flocculation precipitation and resin ball adsorption process", the deep recovery of fine-grained gold in the mixed solution is further realized, reducing the proportion of fine-grained gold lost with the liquid phase. It can be seen that the present invention has significant advantages in the comprehensive recovery and utilization of resources, providing an economic, environmentally friendly and efficient method for impurity removal and gold and silver extraction from carbon-in-pulp process gold slime in multi-metal refractory gold mines, and having broad application prospects.
[0050] The present invention has been described in detail through specific and preferred embodiments above. However, those skilled in the art should understand that the present invention is not limited to the above embodiments. Any modifications, equivalent replacements, etc. made within the spirit and principle of the present invention shall be included within the protection scope of the present invention.
Claims
1. An efficient separation and extraction method for high-sulfur and high-lead gold mud in the carbon-in-pulp process, characterized in that, It includes the following steps: (1) High-temperature roasting: Potassium nitrate is added to the high-sulfur and high-lead gold mud with a water content of 50%-60% obtained by electrolytic desorption of gold-loaded carbon. After thorough mixing, it is loaded into an oven, and the temperature is raised for dehydration, desulfurization, and decarbonization roasting reactions; (2) Slag-making smelting of gold mud: After the dry gold mud obtained from the high-temperature roasting in step (1) is thoroughly mixed with a slag-making agent, it is loaded into a high-purity graphite crucible of an intermediate-frequency furnace for high-temperature slag-making smelting. After high-temperature smelting, the low-density calcium-containing silicon iron smelting slag and the copper-lead-gold-silver-containing liquid alloy are layered, completing the preliminary separation of the smelting slag and valuable metals. Then, the upper smelting slag is skimmed into the slag tank, and the remaining liquid alloy layer is transferred to the next process; (3) Oxidative refining of liquid alloy: The liquid alloy obtained in step (2) is poured into a silicon carbide-graphite crucible, a slag-making agent is added, and the temperature is raised and blown with air for oxidative refining. The matte is oxidized, and sulfur and part of the lead volatilize into the flue dust, obtaining the upper oxidative refining slag and the lower liquid alloy with a higher density; (4) Water quenching and bead splashing of alloy: The upper oxidative refining slag obtained in step (3) is skimmed into the slag tank. After the liquid alloy is cooled to 900-1000 °C, it is slowly and evenly poured into a water tank. At the same time, when pouring the high-temperature liquid alloy, a high-pressure water pipe is used to aim at and disperse the quenched beads until all the liquid alloy is poured. Then, a water-quenched alloy sheet with a larger specific surface area containing copper-silver-gold-lead is obtained by filtration; (5) Lead leaching and filtration of alloy sheet: The water-quenched alloy sheet obtained in step (4) is loaded into a titanium reaction kettle, industrial acetic acid is added, and after heating and waiting for the reaction for 2 h, stirring is started and the reaction continues for 2 h. After the reaction ends, a lead-containing solution and a lead-removed alloy sheet are obtained by filtration; (6) Oxidative impoverishment smelting of smelting slag: The smelting slag and oxidative refining slag generated in steps (2) and (3) are crushed to less than 10 mm, mixed with a slag-making agent, and then loaded into an oxidative impoverishment furnace. The temperature is raised and blown with air for impoverishment smelting. After the smelting ends, a low-density upper oxidative impoverished slag and a lower liquid-phase recycled alloy containing a small amount of impoverished slag are obtained. The upper oxidative impoverished slag is skimmed into the slag tank, and the lower liquid-phase recycled alloy is returned to step (3) for 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 reaction kettle, acetic acid is added, and after heating and static reaction for 2 h, stirring is started and the reaction continues for 3 h. After the reaction ends, a lead-containing solution and lead-removed flue dust are obtained by filtration. The lead-removed flue dust is returned to step (6) for oxidative impoverishment smelting; (8) Precipitation and recovery of lead from lead-containing solution: The lead-containing solution obtained in steps (5) and (7) is transported to a lead precipitation reaction kettle, stirring is started, and a sodium sulfide solution is slowly added for reaction for 0.5 h. After the reaction ends, lead concentrate products and lead-removed solution are obtained by filtration. The lead concentrate is sold externally, and the lead-removed solution enters the wastewater treatment system; (9)Wet gold separation: Load the de-leaded alloy sheets obtained in step (5) into an acid-resistant and high-temperature-resistant titanium reactor. First, moisten the alloy sheets with a small amount of clear water, start stirring and heating, control the temperature at 60 - 80 °C, and then add sulfuric acid and nitric acid in portions for 3 - 5 separations. During the 1 - 2 separations, use sulfuric acid solution and nitric acid solution that are 4 times the total weight of the alloy sheets, with the sulfuric acid solution concentration being 98% and the nitric acid solution concentration being 20% - 25%. During the 3 - 5 separations, use nitric acid solution that is 2 - 3 times the total weight of the alloy sheets, with the nitric acid solution concentration being 30% - 45%, and do not add sulfuric acid until the solution color changes from dark blue to clear and there is no gas on the liquid surface, which is regarded as the end of the wet gold separation operation. Then, filter to obtain gold powder and filtrate, where the filtrate is a mixed solution of copper nitrate - silver nitrate - copper sulfate; (10)Oxidized lean slag grinding, cyanidation and magnetic separation: Grind and cyanide leach the oxidized lean slag obtained in step (6) to recover gold, and the tailings enter the magnetic separation process to recover iron to obtain magnetite concentrate; (11)Two-stage precipitation of the mixed solution to recover fine gold: The copper nitrate - silver nitrate - copper sulfate mixed solution obtained in step (9) enters a closed first-stage precipitation tank and a two-stage precipitation tank in sequence. The bottoms of the two precipitation tanks are conical, and the tops are filled with resin balls. When recovering fine gold, add a flocculant to the mixed solution below the resin balls. The fine gold aggregates and precipitates at the bottom of the precipitation tank under the action of flocculation and the interception of the resin balls, and is regularly discharged, filtered and then smelted for recovery. The remaining solution overflows from the upper part of the precipitation tank and enters the silver and copper recovery process.
2. The efficient separation and extraction method of high-sulfur and high-lead gold mud in a carbon-in-pulp process 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 °C, and the static roasting reaction time is 12 - 14 h.
3. The high-efficiency separation and extraction method of high-sulfur and high-lead gold mud in a carbon-in-pulp process according to claim 1, characterized in that, In step (2), the slag-forming agent includes quartz, borax and sodium carbonate; the temperature of the high-temperature slag-forming smelting is controlled at 1200 - 1300 °C; when skimming the slag, leave a smelting slag layer of 4 - 8 cm above the liquid alloy layer.
4. The high-efficiency separation and extraction method of high-sulfur and high-lead gold mud in a carbon-in-pulp process according to claim 1, characterized in that, The slag-making agent described in step (3) includes quartz, sodium carbonate and borax, and the addition amounts of each component are 2%-3%, 1%-2%, 0.5%-1.5% of the total weight of the melt respectively; during refining, the melt temperature is maintained at 1100-1300 °C; when blowing air, the outlet of the air distribution pipe is lowered to the middle position of the metal liquid surface, the blowing oxidation time for each furnace is 0.5-1 h, and the air volume is 0.2-0.3 m 3 / min.
5. The high-efficiency separation and extraction method of high-sulfur and high-lead gold slime by carbon-in-pulp process according to claim 1, characterized in that, In step (5), the concentration of the acetic acid solution is 50% - 60%, and the liquid-solid ratio of the acetic acid solution to the alloy sheets is 1:4 - 5; the heating temperature is controlled at 60 - 80 °C.
6. The high-efficiency separation and extraction method of high-sulfur and high-lead gold slime in a carbon-in-pulp process according to claim 1, characterized in that The slag-making agent described in step (6) includes cullet, 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, where the particle size of the cullet is less than 20 mm; the smelting temperature is controlled at 1300-1500 °C; when blowing air, the outlet of the air distribution pipe is lowered to the middle position of the molten slag, and each furnace is blown and smelted for 1 h, and the air volume is 0.4-1 m 3 / min.
7. The efficient separation and extraction method of high-sulfur and high-lead gold slime in a carbon-in-pulp process according to claim 1, characterized in that, In step (7), the acetic acid concentration is 50% - 70%, and the liquid-solid ratio of the acetic acid solution to the soot is 1:6 - 8; the heating temperature is controlled at 60 - 80 °C.
8. The efficient separation and extraction method of high-sulfur and high-lead gold mud in a carbon-in-pulp process according to claim 1, characterized in that, In step (8), the concentration of the sodium sulfide solution is 10% - 15%, and the amount of the sodium sulfide solution used is 1.2 - 1.5 times the amount of lead metal in the lead-containing solution.
9. The high-efficiency separation and extraction method of high-sulfur and high-lead gold slime in a carbon-in-pulp process according to claim 1, characterized in that, In step (10), the grinding particle size is 90% - 92% of -200 mesh; the magnetic separation concentration of the tailings is 30% - 38%, and the magnetic field intensity is 0.5 T.
10. The high-efficiency separation and extraction method of high-sulfur and high-lead gold slime in a carbon-in-pulp process according to claim 1, characterized in that, The diameter of the resin balls described in step (11) is 1.5 cm; the dosage of the flocculant is 0.1 - 0.2 g / m 3 .