Complex gold concentrate multi-element comprehensive recovery method

Abstract

The invention belongs to the technical field of metallurgy, and relates to a complex gold concentrate multi-element comprehensive recovery method which comprises the following steps: S1, complex gold concentrate and an auxiliary material I are subjected to bottom blowing smelting in a bottom blowing furnace in an oxygen-enriched air atmosphere to obtain matte, bottom slag and flue gas; s2, copper matte and an auxiliary material II are subjected to top-blowing blowing in a top-blowing furnace in the air atmosphere, and crude copper, flue gas and blowing slag are obtained; s3, refining the crude copper and carbonate in an air atmosphere, introducing methane gas after the reaction is finished, and reducing in an inert atmosphere to obtain a copper anode plate; wherein in the step S1, in the bottom blowing smelting process, a coking inhibitor composed of aluminum oxide, calcium oxide, sodium carbonate, ammonium molybdate, sodium borate and cerium dioxide is blown in an ascending flue of a bottom blowing furnace, the occurrence of the coking condition can be effectively reduced, and the complex gold concentrate recovery production start-up rate is increased.

Description

Technical Field

[0001] This invention belongs to the field of metallurgical technology, specifically relating to a method for comprehensive recovery of multiple elements from complex gold concentrates. Background Technology

[0002] As global gold demand continues to rise, the development and utilization of gold concentrate resources are accelerating. With the gradual depletion of high-quality gold concentrate reserves, complex gold concentrate resources have become a key part of the gold source for many countries. Although the development and refining technologies of these complex gold concentrates are more stringent, they contribute significantly to gold production and ensure the sustainability of gold supply.

[0003] Processing complex gold concentrates presents a significant challenge. These ores typically contain high levels of harmful substances such as arsenic, antimony, and bismuth. During the enrichment of gold using pyrometallurgical techniques, these elements can negatively impact the smelting process. The rising flue and waste heat boiler generate substantial amounts of coke during smelting, which disrupts normal combustion and can even clog the furnace outlet, forcing a shutdown and requiring regular cleaning. Furthermore, due to the presence of arsenic and antimony in the raw materials, the similar properties of arsenic and antimony cause the antimony in the flue gas to precipitate as antimony trioxide along with the arsenic trioxide during rapid cooling, resulting in the collected arsenic trioxide containing antimony trioxide impurities.

[0004] In existing technologies, treating or inhibiting coking typically involves adding coking inhibitors. However, due to variations in raw materials and the types and amounts of impurities, there is no single, universally applicable coking inhibitor. CN 119826567 A discloses a coking inhibitor and its preparation method. The coking inhibitor comprises a reducing agent, a sulfur-containing substance, a silicon-containing substance, and a calcium-containing substance, wherein the sulfur in the sulfur-containing substance has a oxidation state of -2 or zero. This coking inhibitor can lower the melting point of the coke, promoting its melting and thus inhibiting coking. However, this coking inhibitor can only convert high-melting-point compounds in the coke into low-melting-point compounds through reduction and sulfidation. Subsequent slag formation combines the silicon-containing and calcium-containing substances with the low-melting-point compounds to further promote the melting of the coke. Therefore, its coking inhibition effect is limited, and it is not suitable for coke with complex compositions and contents. Summary of the Invention

[0005] To address the shortcomings of the existing technologies, this invention provides a method for the comprehensive recovery of multiple elements from complex gold concentrates. This invention involves injecting coking inhibitors during the smelting and recovery of valuable elements from complex gold concentrates, which can reduce coking and increase production start-up rate.

[0006] The specific technical solutions are as follows:

[0007] A comprehensive method for recovering multiple elements from complex gold concentrates includes the following steps:

[0008] S1 involves bottom-blown smelting of complex gold concentrate and auxiliary material I in a bottom-blown furnace under an oxygen-enriched air atmosphere to obtain matte, bottom slag, and flue gas; the auxiliary material I is quartz and limestone.

[0009] S2 involves top-blowing copper matte and auxiliary material II in an air atmosphere in a top-blown furnace to obtain crude copper, flue gas, and smelting slag; the auxiliary material II is limestone.

[0010] S3 involves refining crude copper and carbonates in an air atmosphere. After the reaction is complete, methane gas is introduced for reduction in an inert atmosphere to obtain a copper anode plate.

[0011] In step S1, during the bottom-blown smelting process, a coking inhibitor is injected into the rising flue of the bottom-blown furnace; the coking inhibitor is composed of aluminum oxide, calcium oxide, sodium carbonate, ammonium molybdate, sodium borate, and cerium dioxide.

[0012] Specifically, the complex gold concentrate contains gold concentrate, silver concentrate and copper concentrate, and may also contain one or two of chalcopyrite and pyrite; among them, the copper concentrate contains CuS, Fe3O4, FeS2 and CuFeS2, the chalcopyrite contains CuFeS2, the pyrite contains FeS2, and the gold concentrate and silver concentrate contain FeS2, CuFeS2 and SiO2;

[0013] More specifically, the main components in the complex gold concentrate are as follows: Au 15~100 g / t, Ag 100~500 g / t, Cu 9.0 wt%~25 wt%, As ≤5.0 wt%, Sb ≤4.0 wt%, Bi ≤3.0 wt%, S 20 wt%~35 wt%, Fe 20 wt%~35 wt%, Te ≤1.0 wt%, Pd 0.02~5 g / t, and Pt 0.02~5 g / t.

[0014] Complex gold concentrates, due to their diverse mineral composition and the presence of numerous metallic and non-metallic impurities, exhibit significant liquefaction during high-temperature bottom-blown smelting, particularly in low-melting-point substances (such as lead and zinc). These molten substances rise with the flue gas flow. Upon reaching the inner walls of the flue and waste heat boiler, these liquid substances gradually adhere and deposit, simultaneously adsorbing fine solid particles carried by the flue gas. Over time, this adhesion and deposition accumulate, eventually forming dense aggregates. This process of gradually accumulating solid deposits on the inner walls of smelting equipment is collectively known as coking in metallurgical processes. Coking not only reduces heat transfer efficiency but also hinders normal flue gas flow, adversely affecting the normal operation of smelting equipment. Particularly inside the rising flue, thicker coke bodies tend to form, disrupting normal combustion conditions and, in severe cases, even blocking the furnace outlet, forcing a shutdown. Due to the different raw material composition, different coking inhibitors are used. The coking inhibitor of this invention transforms the coking material from a harmful substance with a low melting point, high viscosity, and dense hardness into a substance with a high melting point, good fluidity, and loose and porous structure that is easy to remove, thereby achieving the dual purpose of efficient inhibition and convenient cleaning.

[0015] Further, in step S1, the coking inhibitor is composed of 30-40 parts of aluminum oxide, 20-30 parts of calcium oxide, 10-20 parts of sodium carbonate, 10-20 parts of ammonium molybdate, 5-10 parts of sodium borate, and 1-3 parts of cerium dioxide, by mass fraction.

[0016] Among them, aluminum oxide, as a high-melting-point amphoteric oxide, is the framework material for constructing high-melting-point phases. It reacts with zinc ferrite to form zinc aluminum spinel, with FeO to form iron aluminum spinel, and with CaO to form calcium aluminate. These high-melting-point phases form a strong, non-sticky "skeleton" that encapsulates, divides, and isolates low-melting-point coking components (such as PbO), significantly increasing the softening temperature and hardness of the overall coking material, while making its structure brittle and easy to remove mechanically.

[0017] Calcium oxide, as a strong alkaline oxide, is the most critical component for arsenic fixation. It undergoes a high-temperature solid-phase reaction with arsenates to generate extremely stable calcium arsenate, which transforms the highly toxic and coking-promoting arsenic from an active form into an inert and stable form, fixing it in the slag. It also reacts synergistically with aluminum oxide to generate high-melting-point calcium aluminate, providing structural support for the entire system.

[0018] Sodium carbonate decomposes into sodium oxide (Na2O, which is strongly alkaline) and CO2 at high temperatures. Sodium oxide reacts with sulfides (PbS, CuS, Cu2S, etc.) to form sodium sulfate, which can significantly reduce the high-temperature viscosity of the entire system, improve the fluidity of the slag, and make the formed coke more easily washed away by the flue gas flow.

[0019] Ammonium molybdate decomposes into molybdenum trioxide and ammonia at high temperatures. Molybdenum trioxide embeds itself into the crystal structure of the growing coking material, destroying its dense and continuous structure. This causes the coking material to change from a hard, dense, and firmly attached state to a loose, porous, and brittle state. This "brittle" coking material is more likely to fall off when impacted by airflow or mechanically decoked.

[0020] Sodium borate has a low melting point (approximately 740°C). At this temperature, it melts to form a borate glass liquid phase. The molten borate glass liquid phase has excellent fluidity, which can wet and spread on the surface of unmelted coking particles and refractory materials, forming a physical barrier that isolates the contact between coking particles and the contact between coking and the reaction atmosphere (such as flue gas), thus inhibiting further adsorption and expansion of coking. At the same time, sodium borate and sodium carbonate work together to further reduce the viscosity of the system.

[0021] Cerium dioxide acts as an inhibitor of catalysis and a dispersant. Metals such as Cu and Fe, and their compounds, are catalysts for coking reactions. Cerium dioxide, through its unique electronic effect, strongly adsorbs onto these catalytically active sites, acting like a "seal" to block the contact between reactants and active sites, thereby inhibiting the activity of these catalysts and deactivating them. Furthermore, cerium dioxide particles can adhere to the surface of coking particles, preventing them from agglomerating and growing, thus playing a dispersing role.

[0022] This invention involves injecting a coking inhibitor into the rising flue of the bottom-blown furnace during the bottom-blown smelting process of complex gold concentrate. This transforms the coking material from a low-melting-point, high-viscosity, dense, and hard harmful substance into a high-melting-point, fluid, loose, and porous substance that is easy to remove. This achieves the dual purpose of highly efficient coking inhibition and convenient cleaning, reducing the generation of coking material in the rising flue of the bottom-blown furnace. Furthermore, it extends the cleaning frequency of coking material flowing with the flue gas to the subsequent waste heat boiler from once every 20-25 days to once every 45-50 days. The cleaning time for each cleaning session is reduced from 6-12 hours to 4-8 hours, reducing labor intensity and improving production efficiency.

[0023] In this invention, after injecting a coking inhibitor, the main coking phases generated in the waste heat boiler are zinc ferrite, copper sulfide, cuprous sulfide, ferrous oxide, lead sulfide, lead oxide, arsenates (ferric arsenate, bismuth arsenate, antimony arsenate, etc.), cuprous oxide, etc., with the following main component contents: Cu 10wt%~25wt%, Fe 5wt%~20wt%, S 5wt%~10wt%, Pb 5wt%~20wt%, Zn 3wt%~10wt%, As 5wt%~15wt%, Sb 0.1wt%~1wt%, Bi 0.1wt%~1wt%.

[0024] Furthermore, in step S1, the total amount of the coking inhibitor is 0.01wt% to 0.05wt% of the total amount of complex gold concentrate.

[0025] Furthermore, in step S1, the particle size of the coking inhibitor is no greater than 74 μm.

[0026] Preferably, in step S1, the bottom blowing smelting temperature is 1150~1250℃, and the S / Cu mass ratio in the complex gold concentrate is ≥1.3; during the bottom blowing smelting process, continuous feeding is adopted, and bottom slag is discharged once every 25~35 minutes, and matte is discharged once every 1.3~1.6 hours.

[0027] Preferably, in step S1, the amounts of auxiliary materials quartz and limestone are calculated based on SiO2 and CaO in the bottom slag, respectively. The mass ratio of Fe / SiO2 in the bottom slag is (1.2~2.0):1, and the mass ratio of CaO / SiO2 in the bottom slag is (0.12~0.16):1. Quartz is used for slag formation, and CaO generated from the reaction of limestone is used to adjust the viscosity and fluidity of the bottom slag.

[0028] Preferably, in step S1, the oxygen volume fraction of the oxygen-enriched air is 70% to 85%.

[0029] In step S1, the bottom slag is returned to the slag beneficiation system to recover copper concentrate and iron concentrate. The copper concentrate can be returned to step S1 for bottom blowing smelting.

[0030] In step S1, the flue gas is sequentially returned to the waste heat recovery system, the electrostatic precipitator system, the arsenic recovery system, and the sulfuric acid recovery system.

[0031] Preferably, the arsenic recovery system uses a rapid cooling method to recover crude arsenic trioxide solid.

[0032] During the rapid cooling process of arsenic recovery, due to the similar properties of arsenic and antimony, the antimony in the flue gas settles together with the arsenic trioxide as antimony trioxide during the cooling process, resulting in the presence of antimony trioxide impurities in the collected arsenic trioxide. This invention utilizes the characteristic that arsenic trioxide is sparingly soluble in a saturated dilute hydrochloric acid solution, while antimony trioxide is soluble in solution and forms soluble chlorides. By acid leaching the antimony-containing arsenic trioxide (in a saturated dilute hydrochloric acid solution), the antimony impurities are effectively removed, yielding arsenic trioxide with a purity of over 99.5%. The main reaction is as follows:

[0033] Sb₂O₃ + 6HCl → 2SbCl₃ + 3H₂O.

[0034] Specifically, a saturated aqueous solution of arsenic trioxide is mixed with crude arsenic trioxide at a liquid-to-solid mass ratio of (4~8):1, and then 12 mol / L analytical grade hydrochloric acid is added. The mixture is stirred at 5~25℃ for 4~8 hours, and then filtered and dried to obtain refined arsenic trioxide with a purity of 99.5% or higher. Preferably, the molar ratio of hydrochloric acid to the total amount of antimony in crude arsenic trioxide is (5~20):1.

[0035] Furthermore, in step S2, the top blowing temperature is 1150~1250℃. During the top blowing process, the material is fed once every 1.3~1.6 hours, the slag is released once every 3.5~4.5 hours, and the crude copper is released once every 7.5~8.5 hours.

[0036] Preferably, in step S2, the amount of limestone used as an auxiliary material is based on the CaO in the blowing slag, and the mass ratio of CaO / Fe2O3 in the blowing slag is (0.8~1.0):1; the CaO generated by the reaction of limestone is used for slag formation.

[0037] In step S2, the flue gas is sequentially returned to the waste heat recovery system, the electrostatic precipitator system, and the sulfuric acid recovery system to recover boiler ash, electrostatic precipitator ash, and sulfuric acid; the boiler ash can be returned to step S1 for bottom blowing smelting, and the electrostatic precipitator ash is sent to the flue gas recovery system to recover ferroarsenic alloy.

[0038] In step S2, the blown slag is returned to the slag beneficiation system to recover copper concentrate and iron concentrate, and the copper concentrate can be returned to step S1 for bottom blowing smelting.

[0039] Preferably, in step S3, the carbonate is sodium carbonate or potassium carbonate.

[0040] Preferably, in step S3, the mass ratio of carbonate to the total amount of arsenic and antimony in crude copper is (1.5~5):1, the refining temperature is 1150~1200℃, and the refining time is 4~6h; the molar ratio of methane to the total amount of copper in crude copper is (1~2):4, the reduction temperature is 1200~1300℃, and the reduction time is 2~4h.

[0041] Furthermore, in step S3, after refining, slag can be obtained by removing the slag, and the slag can be returned to step S2 for top blowing refining.

[0042] Furthermore, the copper anode plate in step S3 is electrolyzed to obtain cathode copper and copper anode sludge, and the copper anode sludge is returned to the rare and precious metals recovery system to recover valuable metals such as gold, silver, tellurium, bismuth, palladium, platinum and antimony trioxide.

[0043] Compared with the prior art, the present invention has the following beneficial effects:

[0044] This invention involves injecting a coking inhibitor composed of aluminum oxide, calcium oxide, sodium carbonate, ammonium molybdate, sodium borate, and cerium dioxide during the smelting and recovery of valuable elements from complex gold concentrates. This effectively reduces coking and increases the production start-up rate of complex gold concentrate smelting and recovery of valuable elements. Attached Figure Description

[0045] Figure 1 This is a process flow diagram of the multi-element comprehensive recovery method for complex gold concentrates in a specific implementation. Detailed Implementation

[0046] The embodiments of the present invention will be described in further detail below with reference to the accompanying drawings and examples. The following examples are for illustrative purposes only and should not be construed as limiting the scope of the invention. Unless otherwise specified, the experimental methods used in the following examples are conventional methods. Unless otherwise specified, the materials and reagents used in the following examples are commercially available.

[0047] Example 1

[0048] A comprehensive method for multi-element recovery from complex gold concentrates, in which...

[0049] The complex gold concentrate contains chalcopyrite, pyrite, gold concentrate, silver concentrate, and copper concentrate. The main contents of the complex gold concentrate are: Au 15.21 g / t, Ag 101.24 g / t, Cu 24.58 wt%, S 34.25 wt%, Fe 20.58 wt%, As 0.25 wt%, Sb 0.14 wt%, Bi 0.57 wt%, Te 0.14 wt%, Pd 0.02 g / t, and Pt 0.02 g / t. The S / Cu mass ratio in the complex gold concentrate is 1.39.

[0050] The steps are as follows:

[0051] S1 mixes complex gold concentrate, quartz, and limestone, and continuously conveys the mixture to a bottom-blown furnace via belt conveyor. Oxygen-enriched air (85% oxygen by volume) is continuously blown into the furnace through bottom spray guns. Continuous bottom-blown smelting is carried out at 1150°C. During the bottom-blown smelting process, a coking inhibitor with a particle size not exceeding 74μm is injected into the rising flue of the bottom-blown furnace. The coking inhibitor is composed of 30 parts alumina, 20 parts calcium oxide, 10 parts sodium carbonate, 10 parts ammonium molybdate, 5 parts sodium borate, and 1 part aluminum oxide by mass fraction. The composition consists of 1 part cerium chloride. The total amount of coking inhibitor injected is 0.05 wt% of the total amount of complex gold concentrate. Bottom ash is discharged from the slag discharge port every 25 minutes and can flow out automatically from the slag discharge port. Copper matte is discharged from the copper discharge port every 1.3 hours and can flow out automatically from the copper discharge port. The total Si and Ca element content in the bottom ash is detected by X-ray fluorescence spectrometry and converted into SiO2 and CaO mass. The Fe / SiO2 mass ratio in the bottom ash is 1.2:1, and the CaO / S content in the bottom ash is... The iO2 mass ratio is 0.12:1; the flue gas obtained during the bottom-blown smelting process is sequentially returned to the waste heat recovery system (waste heat boiler), the electrostatic precipitator system (electrostatic precipitator), the arsenic recovery system, and the sulfuric acid recovery system to recover boiler ash, electrostatic precipitator ash, crude arsenic trioxide, and sulfuric acid. The boiler ash is sent to the bottom-blown furnace for bottom-blown smelting, and the electrostatic precipitator ash is sent to the flue gas recovery system to recover arsenic-ferroalloy. The steam obtained in the waste heat recovery system (waste heat boiler) can be used for power generation using waste heat; the bottom ash is returned to the slag beneficiation system. Copper concentrate and iron concentrate are recovered. The copper concentrate can be sent to a bottom-blown furnace for bottom-blown smelting. The arsenic recovery system uses a rapid cooling method to recover crude arsenic trioxide solid. Then, a saturated arsenic trioxide aqueous solution is mixed with the crude arsenic trioxide at a liquid-to-solid mass ratio of 4:1. Then, 12 mol / L analytical grade hydrochloric acid is added. The molar ratio of hydrochloric acid to the total antimony in the crude arsenic trioxide is 5:1. The mixture is stirred at 25°C for 6 hours. After filtration and drying, refined arsenic trioxide with a purity of 99.58% is obtained.

[0052] S2: The matte and limestone obtained in step S1 are mixed and fed directly into the top-blown furnace via a chute. Feeding occurs every 1.3 hours, while air is continuously blown into the furnace through the top spray gun. Top-blown smelting is carried out at 1150℃. Smelting slag is discharged from the slag discharge port every 3.5 hours, and the smelting slag flows out automatically from the slag discharge port. Crude copper is discharged from the copper discharge port every 7.5 hours, and the crude copper flows out automatically from the copper discharge port. X-ray fluorescence spectrometry is used to detect the total Ca and Fe element content in the smelting slag, and these are converted into the mass of CaO and Fe2O3. The mass ratio of CaO / Fe2O3 is 0.9:1. The flue gas obtained during the top-blown smelting process is sequentially returned to the waste heat recovery system (waste heat boiler), the electrostatic precipitator system (electrostatic precipitator), and the sulfuric acid recovery system to recover boiler ash, electrostatic precipitator ash, and sulfuric acid. The boiler ash is sent to the bottom-blown furnace for bottom-blown smelting, and the electrostatic precipitator ash is sent to the flue gas recovery system to recover ferroarsenic alloy. The steam obtained in the waste heat recovery system can be used for power generation using waste heat. The smelting slag is returned to the slag beneficiation system to recover copper concentrate and iron concentrate. The copper concentrate can be sent to the bottom-blown furnace for bottom-blown smelting.

[0053] S3. The crude copper obtained in step S2 is transferred to the anode furnace. Sodium carbonate is added at a mass ratio of 1.5:1 to the total mass of arsenic and antimony in the crude copper. Air is continuously blown in, and the refining is carried out at 1150℃ for 4 hours. After the refining is completed, slag is obtained by skimming and returned to step S2 for top blowing refining. After the reaction is completed, methane gas is introduced at a molar ratio of 1:4 to copper in the crude copper. The refining is carried out at 1200℃ in a nitrogen atmosphere for 4 hours. Finally, copper anode plates are obtained by casting. The copper anode plates are sent to the electrolysis system for electrolysis to obtain cathode copper and copper anode mud. The copper anode mud is returned to the rare and precious metals recovery system to recover gold ingots, silver ingots, tellurium ingots, bismuth ingots, sponge palladium, sponge platinum, and antimony trioxide.

[0054] Example 2

[0055] A comprehensive method for multi-element recovery from complex gold concentrates, in which...

[0056] The complex gold concentrate contains chalcopyrite, pyrite, gold concentrate, silver concentrate, and copper concentrate. The main contents of the complex gold concentrate are: Au 99.84 g / t, Ag 499.21 g / t, Cu 9.21 wt%, S 20.47 wt%, Fe 34.27 wt%, As 4.81 wt%, Sb 3.87 wt%, Bi 2.54 wt%, Te 0.94 wt%, Pd 4.85 g / t, and Pt 4.97 g / t. The S / Cu mass ratio in the complex gold concentrate is 2.22.

[0057] The steps are as follows:

[0058] S1 mixes complex gold concentrate, quartz, and limestone, and continuously conveys the mixture to a bottom-blown furnace via belt conveyor. Oxygen-enriched air (70% oxygen by volume) is continuously blown into the furnace through bottom spray guns. Continuous bottom-blown smelting is carried out at 1250℃. During the bottom-blown smelting process, a coking inhibitor with a particle size not exceeding 74μm is injected into the rising flue of the bottom-blown furnace. The coking inhibitor is composed of 35 parts alumina, 25 parts calcium oxide, 15 parts sodium carbonate, 20 parts ammonium molybdate, and 10 parts sodium borate by mass fraction. The composition consists of 3 parts cerium oxide. The total amount of coking inhibitor injected is 0.03 wt% of the total amount of complex gold concentrate. Bottom ash is discharged from the slag discharge port every 35 minutes and can flow out automatically from the slag discharge port. Copper matte is discharged from the copper discharge port every 1.6 hours and can flow out automatically from the copper discharge port. The total Si and Ca element content in the bottom ash is detected by X-ray fluorescence spectrometry and converted into SiO2 and CaO mass. The Fe / SiO2 mass ratio in the bottom ash is 2.0:1, and the CaO / The SiO2 mass ratio is 0.16:1. The flue gas obtained during bottom-blown smelting is sequentially returned to the waste heat recovery system (waste heat boiler), the electrostatic precipitator system (electrostatic precipitator), the arsenic recovery system, and the sulfuric acid recovery system to recover boiler ash, electrostatic precipitator ash, crude arsenic trioxide, and sulfuric acid. The boiler ash is then sent to the bottom-blown furnace for bottom-blown smelting, and the electrostatic precipitator ash is sent to the flue gas recovery system to recover arsenic-ferroalloy. The steam obtained in the waste heat recovery system (waste heat boiler) can be used for power generation. The bottom ash is returned to the slag beneficiation system. Copper concentrate and iron concentrate are recovered. The copper concentrate can be sent to a bottom-blown furnace for bottom-blown smelting. The arsenic recovery system uses a rapid cooling method to recover crude arsenic trioxide solid. Then, a saturated arsenic trioxide aqueous solution is mixed with the crude arsenic trioxide at a liquid-to-solid mass ratio of 8:1. Then, 12 mol / L analytical grade hydrochloric acid is added. The molar ratio of hydrochloric acid to the total antimony in the crude arsenic trioxide is 20:1. The mixture is stirred at 5°C for 8 hours. After filtration and drying, refined arsenic trioxide with a purity of 99.61% is obtained.

[0059] S2: The matte and limestone obtained in step S1 are mixed and fed directly into the top-blown furnace via a chute. Feeding occurs every 1.6 hours, and air is continuously blown into the furnace through the top spray gun. Top-blown smelting is carried out at 1250℃. Smelting slag is discharged from the slag discharge port every 4.5 hours, and the smelting slag flows out automatically from the slag discharge port. Crude copper is discharged from the copper discharge port every 8.5 hours, and the crude copper flows out automatically from the copper discharge port. X-ray fluorescence spectrometry is used to detect the total Ca and Fe element content in the smelting slag, and these are converted into the mass of CaO and Fe2O3. The mass ratio of CaO / Fe2O3 is 0.8:1. The flue gas obtained during the top-blown smelting process is sequentially returned to the waste heat recovery system (i.e., waste heat boiler), the electrostatic precipitator system (i.e., electrostatic precipitator), and the sulfuric acid recovery system to recover boiler ash, electrostatic precipitator ash, and sulfuric acid. The boiler ash is sent to the bottom-blown furnace for bottom-blown smelting, and the electrostatic precipitator ash is sent to the flue gas recovery system to recover ferroarsenic alloy. The steam obtained in the waste heat recovery system can be used for power generation using waste heat. The smelting slag is returned to the slag beneficiation system to recover copper concentrate and iron concentrate. The copper concentrate can be sent to the bottom-blown furnace for bottom-blown smelting.

[0060] S3. The crude copper obtained in step S2 is transferred to the anode furnace. Sodium carbonate is added at a mass ratio of 5:1 (sodium carbonate to the total amount of arsenic and antimony in the crude copper). Air is continuously blown in, and the mixture is refined at 1200℃ for 6 hours. After the refinement, slag is obtained by skimming and returned to step S2 for top-blown refining. After the reaction, methane gas is introduced at a molar ratio of 2:4 (methane to copper in the crude copper). The mixture is reduced at 1250℃ in a nitrogen atmosphere for 2 hours. Finally, copper anode plates are obtained by casting. The copper anode plates are sent to the electrolysis system for electrolysis to obtain cathode copper and copper anode mud. The copper anode mud is returned to the rare and precious metals recovery system to recover gold ingots, silver ingots, tellurium ingots, bismuth ingots, sponge palladium, sponge platinum, and antimony trioxide.

[0061] Example 3

[0062] A comprehensive method for multi-element recovery from complex gold concentrates, in which...

[0063] The complex gold concentrate contains chalcopyrite, pyrite, gold concentrate, silver concentrate, and copper concentrate. The main contents of the complex gold concentrate are: Au 60.21 g / t, Ag 280.14 g / t, Cu 14.21 wt%, S 26.57 wt%, Fe 29.32 wt%, As 2.31 wt%, Sb 1.58 wt%, Bi 1.74 wt%, Te 0.54 wt%, Pd 1.21 g / t, and Pt 1.09 g / t. The S / Cu mass ratio in the complex gold concentrate is 1.87.

[0064] The steps are as follows:

[0065] S1 mixes complex gold concentrate, quartz, and limestone, and continuously conveys the mixture to a bottom-blown furnace via belt conveyor. Oxygen-enriched air with a volume fraction of 78.25% is continuously blown into the furnace through bottom spray guns. Continuous bottom-blown smelting is carried out at 1200℃. During the bottom-blown smelting process, a coking inhibitor with a particle size not exceeding 74μm is injected into the rising flue of the bottom-blown furnace. The coking inhibitor is composed of 40 parts alumina, 30 parts calcium oxide, 20 parts sodium carbonate, 15 parts ammonium molybdate, and 7 parts sodium borate by mass fraction. The composition consists of 2 parts cerium dioxide. The total amount of coking inhibitor injected is 0.01 wt% of the total amount of complex gold concentrate. Bottom ash is discharged from the slag discharge port every 30 minutes and can flow out automatically from the slag discharge port. Copper matte is discharged from the copper discharge port every 1.5 hours and can flow out automatically from the copper discharge port. The total Si and Ca element content in the bottom ash is detected by X-ray fluorescence spectrometry and converted into SiO2 and CaO mass. The Fe / SiO2 mass ratio in the bottom ash is 1.6:1, and the CaO / The SiO2 mass ratio is 0.14:1. The flue gas obtained during bottom-blown smelting is sequentially returned to the waste heat recovery system (waste heat boiler), the electrostatic precipitator system (electrostatic precipitator), the arsenic recovery system, and the sulfuric acid recovery system to recover boiler ash, electrostatic precipitator ash, crude arsenic trioxide, and sulfuric acid. The boiler ash is then sent to the bottom-blown furnace for bottom-blown smelting, and the electrostatic precipitator ash is sent to the flue gas recovery system to recover arsenic-ferroalloy. The steam obtained in the waste heat recovery system (waste heat boiler) can be used for power generation. The bottom ash is returned to the slag beneficiation system. Copper concentrate and iron concentrate are collected. The copper concentrate can be sent to a bottom-blown furnace for bottom-blown smelting. The arsenic recovery system uses a rapid cooling method to recover crude arsenic trioxide solid. Then, a saturated arsenic trioxide aqueous solution is mixed with the crude arsenic trioxide at a liquid-to-solid mass ratio of 6:1. Then, 12 mol / L analytical grade hydrochloric acid is added. The molar ratio of hydrochloric acid to the total antimony in the crude arsenic trioxide is 10:1. The mixture is stirred at 15°C for 4 hours. After filtration and drying, refined arsenic trioxide with a purity of 99.64% is obtained.

[0066] S2: The matte and limestone obtained in step S1 are mixed and fed directly into the top-blown furnace via a chute. Feeding occurs every 1.5 hours, while air is continuously blown into the furnace through the top spray gun. Top-blown smelting is carried out at 1200℃. Smelting slag is discharged from the slag discharge port every 4.0 hours, and blister copper is discharged from the copper discharge port every 8 hours, also automatically flowing out. X-ray fluorescence spectrometry is used to detect the total Ca and Fe element content in the smelting slag, and these are converted into CaO and Fe2O3 masses. The mass ratio of CaO / Fe2O3 is 1:1. The flue gas obtained during the top-blown smelting process is sequentially returned to the waste heat recovery system (waste heat boiler), the electrostatic precipitator system (electrostatic precipitator), and the sulfuric acid recovery system to recover boiler ash, electrostatic precipitator ash, and sulfuric acid. The boiler ash is sent to the bottom-blown furnace for bottom-blown smelting, and the electrostatic precipitator ash is sent to the flue gas recovery system to recover ferroarsenic alloy. The steam obtained in the waste heat recovery system can be used for power generation using waste heat. The smelting slag is returned to the slag beneficiation system to recover copper concentrate and iron concentrate. The copper concentrate can be sent to the bottom-blown furnace for bottom-blown smelting.

[0067] S3. The crude copper obtained in step S2 is transferred to the anode furnace. Sodium carbonate is added at a mass ratio of 2.5:1 (sodium carbonate to the total amount of arsenic and antimony in the crude copper). Air is continuously blown in, and the mixture is refined at 1175°C for 5 hours. After the refinement, slag is obtained by skimming and returned to step S2 for top-blown refining. After the reaction is complete, methane gas is introduced at a molar ratio of 1.5:4 (methane to copper in the crude copper). The mixture is reduced at 1300°C in a nitrogen atmosphere for 3 hours. Finally, copper anode plates are obtained by casting. The copper anode plates are sent to the electrolysis system for electrolysis to obtain cathode copper and copper anode mud. The copper anode mud is returned to the rare and precious metals recovery system to recover gold ingots, silver ingots, tellurium ingots, bismuth ingots, sponge palladium, sponge platinum, and antimony trioxide.

[0068] Comparative Example 1

[0069] For specific steps, please refer to Example 1. The difference is that in step S1, no coking inhibitor is sprayed during the bottom blowing smelting process.

[0070] Comparative Example 2

[0071] For specific steps, please refer to Example 2. The difference is that in step S1, no coking inhibitor is sprayed during the bottom blowing smelting process.

[0072] Since coking inhibitors were injected into the rising flue of the bottom-blown furnace in Examples 1-3, the coking that occurred with the flue gas flow mainly settled in the subsequent waste heat recovery system, i.e., the waste heat boiler. The coking cleaned at this time was the coking in the waste heat boiler. The coking in the waste heat boiler was cleaned when the hourly steam output of the bottom-blown smelting process decreased by 5%. The coking cleaning cycle and time for the waste heat boilers of Examples 1-3 and Comparative Examples 1-2 are shown in Table 1; the coking component content in the waste heat boilers of Examples 1-3 is shown in Table 2.

[0073] Table 1. Coking removal cycle and time consumption.

[0074]

[0075] Table 2. Content of coking components

[0076]

[0077] As shown in Table 1, compared with Comparative Examples 1 and 2, the coking of the waste heat boiler in Examples 1-3 was extended from once every 20-25 days to once every 45-50 days; the cleaning time for each cleaning was reduced from 6-12 hours to 4-8 hours, indicating that this coking inhibitor has a better effect on inhibiting coking of flue gas generated by oxygen-enriched bottom blowing smelting of complex gold concentrates. The coking component content in the waste heat boiler of Examples 1-3 is shown in Table 2. The coking mainly contains Cu, Fe, S, Pb, Zn, As, etc.

[0078] The above description is only a preferred embodiment of the present invention and is not intended to limit the present invention. Any modifications, equivalent substitutions, improvements, etc., made within the spirit and principles of the present invention should be included within the protection scope of the present invention.

Claims

1. A method for comprehensive recovery of multiple elements from complex gold concentrates, characterized in that, Includes the following steps: S1 involves bottom-blown smelting of complex gold concentrate and auxiliary material I in a bottom-blown furnace under an oxygen-enriched air atmosphere to obtain matte, bottom slag, and flue gas; the auxiliary material I is quartz and limestone. S2 involves top-blowing copper matte and auxiliary material II in an air atmosphere in a top-blown furnace to obtain crude copper, flue gas, and smelting slag; the auxiliary material II is limestone. S3 involves refining crude copper and carbonates in an air atmosphere. After the reaction is complete, methane gas is introduced for reduction in an inert atmosphere to obtain a copper anode plate. In step S1, during the bottom-blown smelting process, a coking inhibitor is injected into the rising flue of the bottom-blown furnace; the coking inhibitor is composed of aluminum oxide, calcium oxide, sodium carbonate, ammonium molybdate, sodium borate, and cerium dioxide.

2. The method according to claim 1, characterized in that, In step S1, the total amount of coking inhibitor is 0.01wt% to 0.05wt% of the total amount of complex gold concentrate.

3. The method according to claim 1, characterized in that, In step S1, the coking inhibitor is composed of 30-40 parts of aluminum oxide, 20-30 parts of calcium oxide, 10-20 parts of sodium carbonate, 10-20 parts of ammonium molybdate, 5-10 parts of sodium borate, and 1-3 parts of cerium dioxide, by mass fraction.

4. The method according to claim 1, characterized in that, In step S1, the particle size of the coking inhibitor is no greater than 74 μm.

5. The method according to claim 1, characterized in that, In step S1, the flue gas is returned sequentially to the waste heat recovery system, the electrostatic precipitator system, the arsenic recovery system, and the sulfuric acid recovery system.

6. The method according to claim 5, characterized in that, The arsenic recovery system uses a rapid cooling method to recover crude arsenic trioxide.

7. The method according to claim 6, characterized in that, The crude arsenic trioxide was mixed with a saturated aqueous solution of arsenic trioxide, and then hydrochloric acid was added. After the reaction was completed, refined arsenic trioxide was obtained.

8. The method according to claim 7, characterized in that, The solid-to-mass ratio of the saturated arsenic trioxide aqueous solution to the crude arsenic trioxide mixture is (4~8):

1.

9. The method according to claim 7, characterized in that, The molar ratio of hydrochloric acid to the total amount of antimony in crude arsenic trioxide is (5~20):1.

Images