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 auxiliary materials are subjected to bottom blowing smelting in an oxygen-enriched air atmosphere to obtain matte, bottom slag and flue gas; s2, copper matte and auxiliary materials are subjected to top-blowing converting in the air atmosphere, and crude copper, flue gas and converting slag are obtained; and S3, the crude copper and carbonate are refined in the air atmosphere, after refining is finished, sulfur hexafluoride gas is introduced for a reaction in the air atmosphere, methane gas is introduced for reduction in the inert atmosphere after the reaction is finished, and the copper anode plate is obtained. According to the method, arsenic, antimony and bismuth in the crude copper are removed through two-step oxidation and impurity removal, and the phenomenon that arsenic, antimony and bismuth elements in the complex gold concentrate are enriched in the copper anode plate along with smelting and refining, and consequently the copper anode plate is passivated in the subsequent electrolysis process is avoided; and the recovery method is simple in process and high in recovery rate.

Description

Technical Field

[0001] The invention belongs to the technical field of metallurgy, and in particular relates to a method for comprehensive recovery of multiple elements from complex gold concentrates. Background Art

[0002] As global demand for gold continues to rise, the pace of development and utilization of gold concentrate resources has also accelerated. As the reserves of high-quality gold concentrates gradually decrease, complex gold concentrate resources have become a key part of the gold source for many countries. Although these complex gold concentrates have more stringent requirements for development and refining technology, they provide an important supplement to gold production and ensure the continuity of gold supply.

[0003] When processing complex gold concentrates, since these ores usually contain large amounts of harmful substances such as arsenic, antimony and bismuth, these elements will have a series of negative effects on smelting during the process of gold enrichment using pyrometallurgical technology, especially on the recovery rate and quality of the final product. There is also a problem of low recovery rate of valuable elements.

[0004] In the smelting process of complex gold concentrate, impurity removal is a crucial step. It is based on the principle that copper's affinity for oxygen is less than that of impurity elements. By injecting oxygen-rich air into the melt, copper is first oxidized to cuprous oxide, and then reacts with impurity elements to achieve the purpose of impurity removal. However, during the impurity removal process, it takes a relatively long time for cuprous oxide in the melt to reach saturation. During this period, some impurity elements are oxidized to oxides and volatilize into the flue gas, while the other part forms copper salts and dissolves in the melt. These copper salts will cause serious corrosion to the refractory materials of the furnace body at high temperatures, resulting in a shortened service life and increased consumption of the refractory materials of the furnace body.

[0005] In addition, during the electrolysis of high-arsenic antimony-bismuth copper anode plates, "passivation" is prone to occur, which will lead to copper ion depletion and increased impurity ion concentration during the electrolysis process. When the concentration of impurity ions in the electrolyte is high, the density of the electrolyte will increase, and the sedimentation rate of the anode mud will slow down, resulting in incomplete sedimentation of the anode mud. This incomplete sedimentation will increase the impurity content in the cathode copper; further, the increase in the concentration of impurity ions in the electrolyte will lead to an increase in the electrolyte resistance, which will in turn cause the cell voltage to rise. The increase in cell voltage will reduce the current efficiency and increase the residual pole rate, which will have an adverse effect on the entire electrolysis process.

[0006] To ensure the quality of subsequent products and improve the recovery rate, it is necessary to strictly control the contents of arsenic, bismuth, and antimony in copper anode plates. Generally, the measures taken include: First, through repeated oxidation-reduction processes, part of the arsenic and antimony are oxidized to As2O3 and Sb2O3 and volatilized to be removed. During the reduction process, the high-valent oxides of arsenic and antimony such as As2O5 and Sb2O5 and their salts are reduced to low-valent oxides and then volatilized and removed again. By repeating this process multiple times, most of the arsenic and antimony can be removed. Second, alkaline fluxes are added to react with arsenic and antimony to form arsenates and antimonates that are insoluble in copper, and they are removed by slag skimming. However, bismuth is mutually soluble with copper and is not easily formed into compounds or solid solutions. Bismuth is easily oxidized to Bi2O3, which forms compounds with the oxides of arsenic and antimony and dissolves in the copper liquid. Since Bi2O3 does not form slag with SiO2 or Na2CO3, it is difficult to remove bismuth during the pyrometallurgical refining process. In addition, the above two methods also have certain limitations for removing high contents of arsenic and antimony. Summary of the Invention

[0007] In view of the deficiencies of the above-mentioned existing technologies, the present invention provides a method for comprehensive recovery of multiple elements from complex gold concentrates, comprehensively recovering multiple elements in complex gold concentrates, and avoiding the "passivation" phenomenon of copper anode plates during subsequent electrolysis due to the enrichment of arsenic, antimony, and bismuth elements in complex gold concentrates in copper anode plates during smelting and refining.

[0008] The specific technical solution is as follows: A method for comprehensive recovery of multiple elements from complex gold concentrates, comprising the following steps: S1 Bottom-blow smelting the complex gold concentrates and auxiliary materials in an oxygen-enriched air atmosphere to obtain matte, bottom slag, and flue gas; the auxiliary materials are quartzite and limestone; S2 Top-blow converting the matte and auxiliary materials in an air atmosphere to obtain blister copper, flue gas, and converting slag; the auxiliary material is limestone; S3 Refining the blister copper with carbonate in an air atmosphere, after completion, introducing sulfur hexafluoride gas to react in an air atmosphere, and after the reaction is completed, introducing methane gas to carry out reduction in an inert atmosphere to obtain copper anode plates.

[0009] Specifically, the complex gold concentrates contain gold concentrates, silver concentrates, and copper concentrates, and may also contain one or two of chalcopyrite and pyrite; among them, the copper concentrates contain CuS, Fe3O4, FeS2, CuFeS2, the chalcopyrite contains CuFeS2; the pyrite contains FeS2; the gold concentrates and silver concentrates contain FeS2, CuFeS2, and SiO2.

[0010] More specifically, the main contents of the complex gold concentrate are as follows: the Au content is 15 - 100 g / t, the Ag content is 100 - 500 g / t, the Cu content is 9.0 wt% - 25 wt%, the S content is 20 wt% - 35 wt%, the Fe content is 20 wt% - 35 wt%, the As content is ≤ 5.0 wt%, the Sb content is ≤ 4.0 wt%, the Bi content is ≤ 3.0 wt%, the Te content is ≤ 1.0 wt%, the Pd content is 0.02 - 5 g / t, and the Pt content is 0.02 - 5 g / t.

[0011] The reaction mechanism of the present invention is as follows: In the present invention, in order to recover valuable elements from the complex gold concentrate, the complex gold concentrate is successively subjected to bottom-blown smelting and top-blown converting; subsequently, the obtained crude copper is first refined in the presence of carbonate and oxygen to preliminarily oxidize and remove arsenic and antimony, and the arsenic and antimony therein react with the basic flux to form arsenates and antimonates insoluble in copper and are removed by slag skimming; then sulfur hexafluoride gas is introduced again for reaction to further oxidize and remove arsenic and antimony, especially to remove bismuth; finally, methane gas is used to reduce cuprous oxide therein to elemental copper, and copper anode plates are obtained by casting.

[0012] In step S1, the complex gold concentrate and auxiliary materials undergo physical and chemical reactions with the oxygen-enriched air blown into the furnace under a high-temperature environment to generate matte, bottom slag, and flue gas. The main reactions are as follows: FeS2 → FeS + S; S + O2 → SO2; 4CuFeS2 + 5O2 → 2(Cu2S·FeS) + 2FeO + 4SO2; 2CuS + O2 → Cu2S + SO2; 2FeS + 3O2 → 2FeO + 2SO2; 2FeO + SiO2 → 2FeO·SiO2; FeS + 3Fe3O4 + 5SiO2 → 5(2FeO·SiO2) + SO2; CaCO3 → CaO + CO2.

[0013] The content of impurity elements in complex gold concentrates is relatively high, including arsenic, antimony, bismuth, etc., among which arsenic is particularly corrosive to the furnace body during the smelting process; the present invention controls the process flow during the smelting process and is equipped with bottom blowing smelting, so that most of the arsenic enters the flue gas in gaseous form and is finally recovered in the form of arsenic trioxide. 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; the control of the S / Cu mass ratio is on the one hand to provide heat for the exothermic combustion of S and the physical and chemical reactions of the melting of the ore powder, and on the other hand to generate matte. Preferably, in the bottom blowing smelting process, continuous feeding is adopted, and the bottom slag is released once every 25~35min, and the matte is released once every 1.3~1.6h. Preferably, in step S1, the amounts of auxiliary materials quartz stone 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 stone is used for slag making, and the CaO generated by the reaction of limestone is used to adjust the viscosity and fluidity of the bottom slag.

[0014] Among them, the grade of matte is 45wt%~75wt%, the arsenic content in matte is ≤0.35wt%, and the copper content in bottom slag is ≤4.0wt%.

[0015] Preferably, in step S1, the oxygen volume fraction of the oxygen-enriched air is 70% to 85%.

[0016] Among them, in step S1, the bottom slag is returned to the slag selection system to recover copper concentrate and iron ore concentrate, and the copper concentrate can be returned to step S1 for bottom blowing smelting.

[0017] Among them, in step S1, the flue gas returns to the waste heat recovery system, the electrostatic precipitator system, the arsenic recovery system and the sulfuric acid recovery system in sequence to recover boiler ash, electrostatic precipitator ash, arsenic trioxide 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 smoke recovery system to recover arsenic iron alloy.

[0018] In step S2, the matte and auxiliary materials react physically and chemically with the air blown into the furnace under high temperature to obtain crude copper, flue gas and blowing slag. The main reactions are: CaCO3→CaO+CO2; 3FeS+5O2→Fe3O4+3SO2; 4FeS+7O2→2Fe2O3+4SO2; 2CaO+Fe2O3→2CaO·Fe2O3; 2Cu2S+3O2→2Cu2O+2SO2; Cu2S+2Cu2O→6Cu+SO2.

[0019] Preferably, in step S2, the top-blowing smelting temperature is 1150°C to 1250°C; during the top-blowing smelting process, the feed is carried out once every 1.3 to 1.6 hours, the blowing slag is discharged once every 3.5 to 4.5 hours, and the blister copper is discharged once every 7.5 to 8.5 hours.

[0020] Preferably, in step S2, the dosage of the auxiliary material limestone is calculated based on CaO in the blowing slag, and the mass ratio of CaO / Fe2O3 in the blowing slag is (0.8 to 1.0):1; the CaO generated by the reaction of limestone is used for slag making.

[0021] Among them, the blister copper grade is ≥96.5 wt%, and the copper content in the blowing slag is ≤4.5 wt%.

[0022] Among them, in step S2, the flue gas is successively returned to the waste heat recovery system, the electrostatic precipitation system, and the sulfuric acid recovery system to recover boiler ash, electrostatic precipitation ash, and sulfuric acid; the boiler ash can be returned to step S1 for bottom-blowing smelting, and the electrostatic precipitation ash is sent to the dust recovery system to recover arsenic-iron alloy.

[0023] Among them, in step S2, the blowing slag is returned to the slag separation system to recover copper concentrate and iron concentrate powder, and the copper concentrate can be returned to step S1 for bottom-blowing smelting.

[0024] In step S3, the blister copper is subjected to two-step oxidation and deimpurity removal and reduction to remove arsenic, antimony, and bismuth contained in the blister copper, and to avoid the enrichment of arsenic, antimony, and bismuth elements in the copper anode plate to form a high-arsenic-antimony-bismuth copper anode plate, resulting in the "passivation" phenomenon during the subsequent electrolysis process; preferably, the carbonate is sodium carbonate or potassium carbonate, and more preferably, the carbonate is sodium carbonate; the main reactions are: 4Cu3As + 5O2 + 2Na2CO3 → 2(Na2O·As2O5) + 2CO2 + 12Cu; 4Cu3Sb + 5O2 + 2Na2CO3 → 2(Na2O·Sb2O5) + 2CO2 + 12Cu; 2Cu3As + O2 + SF6 → 2AsF3 + SO2 + 6Cu; 2Cu3Sb + O2 + SF6 → 2SbF3 + SO2 + 6Cu; 2Bi2O3 + 2SF6 → 4BiF3 + 2SO2 + O2; 4Cu2O + CH4 → 8Cu + CO2 + 2H2O.

[0025] Preferably, in step S3, the mass ratio of carbonate to the total amount of arsenic and antimony in crude copper is (1.5 - 5.0):1, the refining temperature is 1150 - 1200 °C, and the refining time is 4 - 6 h; the mass ratio of sulfur hexafluoride to the total amount of bismuth in crude copper is (0.4 - 1.0):1, the reaction time is 2 - 4 h, and the reaction temperature is 1150 - 1200 °C; the molar ratio of methane to the total amount of copper in crude copper is (1 - 2):4, the reduction temperature is 1200 - 1300 °C, and the reduction time is 2 - 4 h.

[0026] Among them, the As content in the copper anode plate is ≤ 0.1 wt%, the Sb content is ≤ 0.1 wt%, and the Bi content is < 0.01 wt%.

[0027] Furthermore, in step S3, slag can be obtained by skimming after refining, and the slag can be returned to step S2 for top-blown smelting.

[0028] Furthermore, the copper anode plate in step S3 is electrolyzed to obtain cathode copper and copper anode slime, and the copper anode slime is returned to the precious metal recovery system to recover gold, silver, tellurium, bismuth, palladium, platinum, and antimony trioxide.

[0029] The beneficial effects of the present invention are as follows: The present invention separates arsenic and iron in complex gold concentrate from copper, gold, silver and other elements therein through bottom-blown smelting and top-blown smelting, and the latter coexist in the form of crude copper; subsequently, arsenic, antimony and bismuth in crude copper are removed by two-step oxidation and impurity removal, avoiding the phenomenon of "passivation" of copper anode plates during subsequent electrolysis due to the enrichment of arsenic, antimony and bismuth elements in complex gold concentrate in copper anode plates; and the recovery method of the present invention has a simple process and a high recovery rate. Description of the Drawings

[0030] Figure 1 It is a process flow chart of the method for comprehensive recovery of multiple elements in complex gold concentrate in the specific embodiment. Specific Embodiment

[0031] The principles and features of the present invention are described below in conjunction with examples. The examples given are only for explaining the present invention and are not intended to limit the scope of the present invention. The experimental methods used in the following examples are all conventional methods unless otherwise specified. The materials, reagents, etc. used in the following examples can be obtained from commercial sources unless otherwise specified. Example 1

[0032] A method for comprehensive recovery of multiple elements in complex gold concentrate, wherein, The complex gold concentrate contains chalcopyrite, pyrite, gold concentrate, silver concentrate and copper concentrate; the main contents of the complex gold concentrate are 15.21 g / t of Au, 101.24 g / t of Ag, 24.58 wt% of Cu, 34.25 wt% of S, 20.58 wt% of Fe, 0.25 wt% of As, 0.14 wt% of Sb, 0.57 wt% of Bi, 0.14 wt% of Te, 0.02 g / t of Pd, and 0.02 g / t of Pt; among them, the S / Cu mass ratio in the complex gold concentrate is 1.39.

[0033] It includes the following steps: S1 Mix the complex gold concentrate, quartz stone and limestone, and continuously convey them to the bottom-blown furnace through a belt. Continuously blow oxygen-rich air with an oxygen volume fraction of 85% into the furnace through the bottom lance, and conduct continuous bottom-blown smelting at 1150 °C. Discharge the bottom slag from the slag discharge port once every 25 minutes, and the bottom slag can flow out automatically from the slag discharge port. Discharge the matte from the copper discharge port once every 1.3 hours, and the matte can flow out automatically from the copper discharge port; among them, the matte grade is 45.21 wt%, the arsenic content in the matte is 0.20 wt%, and the copper content in the bottom slag is 3.95 wt%. Use an X-ray fluorescence spectrometer to detect the total Si and Ca element contents in the bottom slag respectively, and convert them into the masses of SiO2 and CaO. The Fe / SiO2 mass ratio in the bottom slag is 1.2:1, and the CaO / SiO2 mass ratio in the bottom slag is 0.12:1. The flue gas obtained during the bottom-blown smelting process is successively returned to the waste heat recovery system, the electrostatic precipitation system, the arsenic recovery system and the sulfuric acid recovery system to recover boiler ash, electrostatic precipitation ash, arsenic trioxide and sulfuric acid, and send the boiler ash to the bottom-blown furnace for bottom-blown smelting. Send the electrostatic precipitation ash to the dust recovery system to recover arsenical iron alloy. The steam obtained in the waste heat recovery system can be used to generate electricity using its waste heat; return the bottom slag to the slag separation system to recover copper concentrate and iron concentrate powder, and the copper concentrate can be sent to the bottom-blown furnace for bottom-blown smelting; S2 Mix the matte obtained in step S1 with limestone and directly pour it into the top-blown converter through a chute. Feed once every 1.3 h, and continuously blow air into the furnace through the top lance. Conduct top-blown smelting at 1150 °C. Discharge the smelting slag from the slag tapping hole once every 3.5 h. The smelting slag can flow out automatically from the slag tapping hole. Discharge the blister copper from the copper tapping hole once every 7.5 h. The blister copper can flow out automatically from the copper tapping hole. Among them, the grade of the blister copper is 96.65 wt%. The copper content in the smelting slag is 4.48 wt%. Use an X-ray fluorescence spectrometer to detect the total Ca and Fe element contents in the smelting slag respectively, and convert them into the masses of CaO and Fe2O3. The mass ratio of CaO / Fe2O3 in the smelting slag is 0.9:1. The flue gas obtained during the top-blown smelting process is successively returned to the waste heat recovery system, the electrostatic precipitation system, and the sulfuric acid recovery system to recover boiler ash, electrostatic precipitation ash, and sulfuric acid. Send the boiler ash to the bottom-blown converter for bottom-blown smelting. Send the electrostatic precipitation ash to the dust recovery system to recover arsenic-iron alloy. The steam obtained in the waste heat recovery system can be used to generate electricity with its waste heat. Return the smelting slag to the slag separation system to recover copper concentrate and iron concentrate. The copper concentrate can be sent to the bottom-blown converter for bottom-blown smelting; S3 Transfer the blister copper obtained in step S2 to the anode furnace. Add sodium carbonate according to the mass ratio of sodium carbonate to the total amount of arsenic and antimony in the blister copper being 1.5:1. Continuously blow air and refine at 1150 °C for 4 h. After completion, obtain slag by skimming the slag, and return the slag to step S2 for top-blown smelting. Then, introduce sulfur hexafluoride gas according to the mass ratio of sulfur hexafluoride to the total amount of bismuth in the blister copper being 0.4:1, and react at 1150 °C in an air atmosphere for 2 h. After the reaction, introduce methane gas according to the molar ratio of methane to copper in the blister copper being 1:4, and reduce in a nitrogen atmosphere at 1200 °C for 4 h. Finally, obtain copper anode plates by casting. Send the copper anode plates to the electrolysis system for electrolysis to obtain cathode copper and copper anode slime, and return the copper anode slime to the precious and rare metal recovery system to recover gold ingots, silver ingots, tellurium ingots, bismuth ingots, sponge palladium, sponge platinum, and antimony trioxide. Example 2

[0034] A method for comprehensive recovery of multiple elements from complex gold concentrate, wherein, The complex gold concentrate contains chalcopyrite, pyrite, gold concentrate, silver concentrate, and copper concentrate; the main contents of the complex gold concentrate are Au content of 99.84 g / t, Ag content of 499.21 g / t, Cu content of 9.21 wt%, S content of 20.47 wt%, Fe content of 34.27 wt%, As content of 4.81 wt%, Sb content of 3.87 wt%, Bi content of 2.54 wt%, Te content of 0.94 wt%, Pd content of 4.85 g / t, Pt content of 4.97 g / t; among them, the S / Cu mass ratio in the complex gold concentrate is 2.22.

[0035] It includes the following steps: S1 Mix the complex gold concentrate, quartzite and limestone, and continuously convey them into the bottom-blown furnace through a belt. Continuously blow oxygen-rich air with an oxygen volume fraction of 70% into the furnace through the bottom lance, and conduct continuous bottom-blown smelting at 1250 °C. Discharge the bottom slag from the slag discharge port once every 35 minutes, and the bottom slag can flow out automatically from the slag discharge port. Discharge the matte from the copper discharge port once every 1.6 hours, and the matte can flow out automatically from the copper discharge port; among them, the matte grade is 74.65 wt%, the arsenic content in the matte is 0.35 wt%, and the copper content in the bottom slag is 3.24 wt%. Use an X-ray fluorescence spectrometer to detect the total Si and Ca element contents in the bottom slag respectively, and convert them into the masses of SiO2 and CaO. The mass ratio of Fe / SiO2 in the bottom slag is 2.0:1, and the mass ratio of CaO / SiO2 in the bottom slag is 0.16:1; the flue gas obtained during the bottom-blown smelting process is successively returned to the waste heat recovery system, the electrostatic precipitation system, the arsenic recovery system and the sulfuric acid recovery system to recover boiler ash, electrostatic precipitation ash, arsenic trioxide and sulfuric acid, and send the boiler ash to the bottom-blown furnace for bottom-blown smelting, and send the electrostatic precipitation ash to the dust recovery system to recover arsenic-iron alloy. The steam obtained in the waste heat recovery system can be used to generate electricity using its waste heat; the bottom slag is returned to the slag separation system to recover copper concentrate and iron concentrate, and the copper concentrate can be sent to the bottom-blown furnace for bottom-blown smelting; S2 Mix the matte obtained in step S1 and limestone, and directly flow them into the top-blown furnace through a chute. Feed once every 1.6 hours, and continuously blow air into the furnace through the top lance. Conduct top-blown smelting at 1250 °C. Discharge the smelting slag from the slag discharge port once every 4.5 hours, and the smelting slag can flow out automatically from the slag discharge port. Discharge the blister copper from the copper discharge port once every 8.5 hours, and the blister copper can flow out automatically from the copper discharge port; among them, the blister copper grade is 97.24 wt%, and the copper content in the smelting slag is 3.58 wt%. Use an X-ray fluorescence spectrometer to detect the total Ca and Fe element contents in the smelting slag respectively, and convert them into the masses of CaO and Fe2O3. The mass ratio of CaO / Fe2O3 in the smelting slag is 0.8:1; the flue gas obtained during the top-blown smelting process is successively returned to the waste heat recovery system, the electrostatic precipitation system and the sulfuric acid recovery system to recover boiler ash, electrostatic precipitation ash and sulfuric acid, and send the boiler ash to the bottom-blown furnace for bottom-blown smelting, and send the electrostatic precipitation ash to the dust recovery system to recover arsenic-iron alloy. The steam obtained in the waste heat recovery system can be used to generate electricity using its waste heat; the smelting slag is returned to the slag separation system to recover copper concentrate and iron concentrate, and the copper concentrate can be sent to the bottom-blown furnace for bottom-blown smelting; S3 transfers the crude copper obtained in step S2 into an anode furnace, adds sodium carbonate at a mass ratio of sodium carbonate to the total amount of arsenic and antimony in the crude copper of 5.0:1, continuously blows air, and refines at 1200 °C for 6 h. After completion, slag is removed to obtain slag, and the slag is returned to step S2 for top-blown smelting; then sulfur hexafluoride gas is introduced at a mass ratio of sulfur hexafluoride to the total amount of bismuth in the crude copper of 1.0:1, and reacts at 1200 °C for 4 h in an air atmosphere; after the reaction is completed, methane gas is introduced at a molar ratio of methane to copper in the crude copper of 2:4, and reduces in a nitrogen atmosphere at 1250 °C for 2 h. Finally, a copper anode plate is obtained by casting; the copper anode plate is sent to an electrolysis system for electrolysis to obtain cathode copper and copper anode slime, and the copper anode slime is returned to the precious and rare metal recovery system to recover gold ingots, silver ingots, tellurium ingots, bismuth ingots, sponge palladium, sponge platinum and antimony trioxide. Example 3

[0036] A method for comprehensive recovery of multiple elements from complex gold concentrate, wherein, The complex gold concentrate contains chalcopyrite, pyrite, gold concentrate, silver concentrate and copper concentrate; the main contents of the complex gold concentrate are 60.21 g / t of Au content, 280.14 g / t of Ag content, 14.21 wt% of Cu content, 26.57 wt% of S content, 29.32 wt% of Fe content, 2.31 wt% of As content, 1.58 wt% of Sb content, 1.74 wt% of Bi content, 0.54 wt% of Te content, 1.21 g / t of Pd content, 1.09 g / t of Pt content; wherein, the S / Cu mass ratio in the complex gold concentrate is 1.87.

[0037] It includes the following steps: S1 Mix the complex gold concentrate, quartz stone and limestone, and continuously convey them into the bottom-blown furnace through a belt. Continuously blow oxygen-enriched air with an oxygen volume fraction of 78.25% into the furnace through the bottom lance, and carry out continuous bottom-blown smelting at 1200 °C. Discharge the bottom slag from the slag discharge port once every 30 minutes. The bottom slag can flow out automatically from the slag discharge port. Discharge the matte from the copper discharge port once every 1.5 hours. The matte can flow out automatically from the copper discharge port. Among them, the matte grade is 58.24 wt%, the arsenic content in the matte is 0.15 wt%, and the copper content in the bottom slag is 3.87 wt%. Use an X-ray fluorescence spectrometer to detect the total Si and Ca element contents in the bottom slag respectively, and convert them into the masses of SiO2 and CaO. The mass ratio of Fe / SiO2 in the bottom slag is 1.6:1, and the mass ratio of CaO / SiO2 in the bottom slag is 0.14:1. The flue gas obtained during the bottom-blown smelting process is successively returned to the waste heat recovery system, the electrostatic precipitator system, the arsenic recovery system, and the sulfuric acid recovery system to recover boiler ash, electrostatic precipitator ash, arsenic trioxide, and sulfuric acid. Send the boiler ash to the bottom-blown furnace for bottom-blown smelting, send the electrostatic precipitator ash to the dust recovery system to recover ferric arsenide alloy, and the steam obtained in the waste heat recovery system can be used to generate electricity with its waste heat. Return the bottom slag to the slag separation system to recover copper concentrate and iron concentrate powder, and the copper concentrate can be sent to the bottom-blown furnace for bottom-blown smelting; S2 Mix the matte obtained in step S1 and limestone, and directly flow them into the top-blown furnace through a chute. Feed once every 1.5 hours, and continuously blow air into the furnace through the top lance. Carry out top-blown converting at 1200 °C. Discharge the converting slag from the slag discharge port once every 4.0 hours. The converting slag can flow out automatically from the slag discharge port. Discharge the blister copper from the copper discharge port once every 8.0 hours. The blister copper can flow out automatically from the copper discharge port. Among them, the blister copper grade is 97.29 wt%, and the copper content in the converting slag is 3.69 wt%. Use an X-ray fluorescence spectrometer to detect the total Ca and Fe element contents in the converting slag respectively, and convert them into the masses of CaO and Fe2O3. The mass ratio of CaO / Fe2O3 in the converting slag is 1:1. The flue gas obtained during the top-blown converting process is successively 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. Send the boiler ash to the bottom-blown furnace for bottom-blown smelting, send the electrostatic precipitator ash to the dust recovery system to recover ferric arsenide alloy, and the steam obtained in the waste heat recovery system can be used to generate electricity with its waste heat. Return the converting slag to the slag separation system to recover copper concentrate and iron concentrate powder, and the copper concentrate can be sent to the bottom-blown furnace for bottom-blown smelting; S3 transfers the crude copper obtained in step S2 into an anode furnace, adds sodium carbonate at a mass ratio of 2.5:1 to the total amount of arsenic and antimony in the crude copper, continuously blows air, and refines it at 1175 °C for 5 h. After completion, slag is obtained by slag skimming, and the slag is returned to step S2 for top-blown smelting; then sulfur hexafluoride gas is introduced at a mass ratio of 0.6:1 to the total amount of bismuth in the crude copper, and reacts at 1175 °C for 3 h in an air atmosphere; after the reaction is completed, methane gas is introduced at a molar ratio of 1.5:4 to the copper in the crude copper, and is reduced in a nitrogen atmosphere at 1300 °C for 3 h. Finally, copper anode plates are obtained by casting; the copper anode plates are sent to an electrolysis system for electrolysis to obtain cathode copper and copper anode slime, and the copper anode slime is returned to the precious and rare metal recovery system to recover ingots of gold, silver, tellurium, bismuth, sponge palladium, sponge platinum, and antimony trioxide. Comparative Example 1

[0038] The specific steps refer to Example 1, the difference is that: sodium carbonate is not added in step S3. Comparative Example 2

[0039] The specific steps refer to Example 1, the difference is that: sulfur hexafluoride gas is not introduced in step S3. Test

[0040] Detect the contents of arsenic, antimony, and bismuth in the copper anode plates obtained in Examples 1 to 3 and Comparative Examples 1 to 2. The results are shown in Table 1; calculate the recovery rates of various valuable metals in Examples 1 to 3 and Comparative Examples 1 to 2. The results are shown in Table 2.

[0041] Among them, the contents of arsenic, antimony, and bismuth in the copper anode plates are determined according to the method for chemical analysis of crude copper - Part 6: Determination of arsenic, antimony, bismuth, lead, zinc, and nickel contents - Inductively coupled plasma atomic emission spectrometry (YS / T 521.6 - 2009).

[0042] Table 1 Contents of arsenic, antimony, and bismuth in copper anode plates Content (wt%) Example 1 Example 2 Example 3 Comparative Example 1 Comparative Example 2 As 0.097 0.089 0.094 0.12 0.14 Sb 0.081 0.092 0.078 0.11 0.13 Bi 0.005 0.008 0.006 0.010 0.16

[0043] Table 2 Recovery rates of valuable metals Total recovery rate (%) Example 1 Example 2 Example 3 Comparative Example 1 Comparative Example 2 Au 98.58 98.57 98.59 98.69 98.57 Ag 98.61 98.59 98.64 98.54 98.61 Cu 98.52 98.57 98.51 95.36 95.02 Te 85.65 85.34 85.94 85.21 86.21 Pd 88.68 89.24 88.74 88.69 88.27 Pt 89.24 89.17 88.95 89.01 88.65 Bi 92.15 91.87 92.02 91.59 91.26 Sb 93.25 94.21 93.88 93.58 93.61 As shown in Table 1 and Table 2, after adding sodium carbonate and sulfur hexafluoride as impurity removal agents during the refining of crude copper smelted from complex gold concentrates, the contents of arsenic, antimony, and bismuth in the copper anode plates obtained in Examples 1 to 3 are significantly lower than those in Comparative Examples 1 to 2, and the copper recovery rates in Examples 1 to 3 are significantly higher than those in Comparative Examples 1 to 2, all by more than 3%. This shows that using sodium carbonate and sulfur hexafluoride as impurity removal agents can effectively improve the removal rates of impurity elements arsenic, antimony, and bismuth during the refining process of complex gold concentrates smelting.

[0044] The foregoing are only preferred embodiments of the present invention and are not intended to limit the present invention. Any modifications, equivalent replacements, improvements, etc. made within the spirit and principles of the present invention shall 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: The steps include: S1. Bottom-blowing smelting of complex gold concentrate and auxiliary materials in an oxygen-rich air atmosphere to obtain matte, bottom slag and flue gas; the auxiliary materials are quartz and limestone; S2 top-blowing the matte and auxiliary materials in an air atmosphere to obtain blister copper, flue gas and blowing slag; the auxiliary material is limestone; S3: The crude copper and carbonate are refined in an air atmosphere. After the refining is completed, sulfur hexafluoride gas is introduced to react in an air atmosphere. After the reaction is completed, methane gas is introduced to reduce the reaction in an inert atmosphere to obtain a copper anode plate.

2. The method according to claim 1, characterized in that In step S1, the bottom blowing smelting temperature is 1150°C-1250°C, and the S / Cu mass ratio in the complex gold concentrate is ≥1.

3.

3. The method according to claim 1, characterized in that In step S1, during the bottom blowing smelting process, continuous feeding is adopted, with bottom slag added once every 25 to 35 minutes and matte added once every 1.3 to 1.6 hours.

4. The method according to claim 1, characterized in that: In step S1, the amounts of auxiliary materials quartz stone 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.

5. The method according to claim 1, characterized in that In step S2, the top blowing temperature is 1150°C to 1250°C.

6. The method according to claim 1, characterized in that In step S2, during the top blowing process, feed is added once every 1.3-1.6 hours, blowing slag is added once every 3.5-4.5 hours, and crude copper is added once every 7.5-8.5 hours.

7. The method according to claim 1, characterized in that In step S2, the amount of the auxiliary material limestone is calculated based on CaO in the blowing slag, and the mass ratio of CaO / Fe2O3 in the blowing slag is (0.8~1.0):

1.

8. The method according to claim 1, characterized in that In step S3, the mass ratio of carbonate to the total amount of arsenic and antimony in crude copper is (1.5-5.0):1, the refining temperature is 1150-1200° C., and the refining time is 4-6 hours.

9. The method according to claim 1, characterized in that: In step S3, the mass ratio of sulfur hexafluoride to the total amount of bismuth in the crude copper is (0.4-1.0):1, the reaction time is 2-4 hours, and the reaction temperature is 1150-1200°C.

10. The method according to claim 1, characterized in that In step S3, the molar ratio of methane to the total amount of copper in the crude copper is (1-2):4, the reduction temperature is 1200-1300° C., and the reduction time is 2-4 hours.

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