Short-process synergetic smelting method for complex tin material

Abstract

The invention discloses a short-process collaborative smelting method for a complex tin material, and relates to the technical field of non-ferrous metallurgy. The smelting method comprises a raw material pretreatment stage, a smelting stage, a reduction stage and a slag phase treatment stage. The method is a smelting method based on stepped oxygen distribution and dynamic mass transfer, and is combined with an 'Ausmelt furnace-fuming furnace' collaborative smelting system, so that the oxygen concentration is increased in a segmented manner, the mass transfer efficiency of a molten pool is optimized, the slag type is stabilized, and the direct recovery rate of tin is efficiently increased; the fuming furnace is subjected to three-stage gradient oxygen reduction, so that the tin volatilization rate is improved, and meanwhile, the recovery rate of copper in the copper-iridium enrichment is also improved; and the high-temperature flue gas of the Ausmelt furnace is used for preheating the fuming furnace material, so that the comprehensive energy consumption is reduced by 25%, and the recycling efficiency of the high-sulfur high-arsenic high-iron complex tin material is remarkably improved.

Description

Technical Field

[0001] The present invention relates to the technical field of nonferrous metal metallurgy, and in particular to a short-process coordinated smelting method for complex tin materials. Background Art

[0002] The processing of complex tin materials containing high sulfur, high arsenic, and high iron content has always been a critical issue in the metallurgical industry. With increasing global demands for efficient resource utilization and environmental protection, traditional smelting processes face numerous challenges. Currently, the primary method for processing such complex tin materials is the Ausmelt furnace smelting technology. This technology has the following drawbacks: (1) Difficulty in controlling slag type: Due to the complex material composition, the Fe / SiO2 ratio fluctuates greatly (often deviating from the range of 1.0-1.8), resulting in low mass transfer efficiency in the molten pool and a tin direct recovery rate of less than 80%; (2) High energy consumption: The Ausmelt furnace and the fuming furnace operate independently, and additional heating is required for material transfer, with a total energy consumption of up to 1.2-1.5 tons of standard coal per ton of tin; (3) Low recovery rate of valuable metals: Copper, antimony and other associated metals have low separation efficiency (copper recovery rate <60%), and a large amount of them enter the waste slag, resulting in serious waste of resources; (4) Environmental issues: The arsenic content in the smelting slag is high (>0.5%), and the subsequent treatment cost is high and it is easy to cause environmental pollution.

[0003] Oxygen-enriched smelting technology has improved smelting efficiency to a certain extent, but it has not solved core problems such as coordinated control of dual furnaces and efficient separation of tin and copper, and still cannot meet the requirements of modern industry for comprehensive resource utilization and green environmental protection.

[0004] In view of this, the present invention is proposed. Summary of the Invention

[0005] The purpose of the present invention is to provide a short-process collaborative smelting method for complex tin materials to solve the above technical problems.

[0006] The present invention is achieved in that: In a first aspect, an embodiment of the present invention provides a short-process collaborative smelting method for complex tin materials, comprising the following steps: Raw material pretreatment stage, smelting stage, reduction stage and slag phase treatment stage; In the smelting stage, the complex tin material is added to the Ausmelt furnace, and the oxygen concentration is increased in stages using a step-by-step oxygen enrichment process to produce molten slag. In the initial stage, the oxygen concentration is 25%-28% and the temperature is 1200-1250°C; in the middle stage, it is increased to 33%-35% and the temperature is 1250-1300°C; in the final stage, it is increased to 38%-40% and the temperature is 1300-1400°C. Reduction stage: The obtained molten slag is introduced into a fuming furnace, and a step-by-step oxygen reduction process is adopted to complete tin volatilization and copper enrichment to obtain fumed slag; among them, the oxygen concentration in the first-level oxygen distribution zone is 14%-16%, and the temperature is 1200-1300℃; the oxygen concentration in the second-level oxygen distribution zone is 9%-11%, and the temperature is 1100-1150℃; the oxygen concentration in the third-level oxygen distribution zone is 4%-6%, and the temperature is 1000-1050℃.

[0007] The present invention has the following beneficial effects: The short-process collaborative smelting method for complex tin materials provided by an embodiment of the present invention is a smelting method based on cascade oxygen distribution and dynamic mass transfer. It combines the "Ausmelt furnace-fuming furnace" collaborative smelting system to increase the oxygen concentration in stages, optimize the mass transfer efficiency of the molten pool, stabilize the slag shape, and effectively improve the direct recovery rate of tin; the three-stage gradient oxygen reduction in the fuming furnace achieves an increase in the tin volatilization rate while also improving the copper recovery rate in copper-iridium enrichment; the high-temperature flue gas (1100-1300°C) of the Ausmelt furnace is used to preheat the fuming furnace material, reducing the comprehensive energy consumption by 25%, significantly improving the resource utilization efficiency of complex tin materials with high sulfur, high arsenic, and high iron. DETAILED DESCRIPTION

[0008] To make the purpose, technical solutions and advantages of the embodiments of the present invention clearer, the technical solutions in the embodiments of the present invention are described clearly and completely below. Where specific conditions are not specified in the embodiments, conventional conditions or conditions recommended by the manufacturer are used. Where the manufacturer of the reagents or instruments is not specified, all are conventional products that can be purchased commercially.

[0009] Raw material pretreatment stage, smelting stage, reduction stage and slag phase treatment stage; In the smelting stage, the complex tin material is added to the Ausmelt furnace, and the oxygen concentration is increased in stages using a step-by-step oxygen enrichment process to produce molten slag. In the initial stage, the oxygen concentration is 25%-28% and the temperature is 1200-1250°C; in the middle stage, it is increased to 33%-35% and the temperature is 1250-1300°C; in the final stage, it is increased to 38%-40% and the temperature is 1300-1400°C. Reduction stage: The obtained molten slag is introduced into a fuming furnace, and a step-by-step oxygen reduction process is adopted to complete tin volatilization and copper enrichment to obtain fumed slag; among them, the oxygen concentration in the first-level oxygen distribution zone is 14%-16%, and the temperature is 1200-1300℃; the oxygen concentration in the second-level oxygen distribution zone is 9%-11%, and the temperature is 1100-1150℃; the oxygen concentration in the third-level oxygen distribution zone is 4%-6%, and the temperature is 1000-1050℃.

[0010] It should be noted that the present invention, through the innovative design of an "Ausmelt furnace-fuming furnace" synergistic smelting system, increases the direct tin recovery rate from 75%-80% in traditional processes to ≥90%, while copper and antimony recoveries reach 85% and over 75%, respectively. By coupling the waste heat of the two furnaces and streamlining the process, overall energy consumption is reduced by 25% (energy consumption per ton of tin ≤ 1.0 ton of standard coal) and material transfer costs are reduced by 20%. Furthermore, the arsenic content of the smelting slag is consistently ≤ 0.3%, leaching toxicity is below the national standard limit (0.05 mg / L), and flue gas CO2 emissions are reduced by 18%-22%. This approach achieves optimal slag shape control, efficient tin and copper separation, low energy consumption, and resource utilization, meeting modern industrial requirements for comprehensive resource utilization and environmental protection.

[0011] Osmel furnace segmented oxygen distribution: oxygen concentration is regulated in three stages, which is beneficial to optimize metal recovery rate, accurately control slag shape to improve separation efficiency of slag and metal, reduce energy consumption and equipment loss, and enhance the process adaptability to multiple materials.

[0012] The Ausmelt furnace unit is also equipped with a segmented oxygen enrichment supply device, an adjustable depth spray gun, and an online melt pool composition monitoring module, which work together to achieve segmented oxygen distribution. The Ausmelt furnace segmented oxygen enrichment supply device includes three independently controlled oxygen supply modules, with an oxygen concentration adjustment accuracy of ±1%. Combined with dynamic adjustment of the spray gun air volume, this creates an alternating oxidation-reduction atmosphere, which helps to enhance melt pool turbulence and stabilize the slag shape.

[0013] During the reduction stage, the fuming furnace unit is equipped with a three-stage oxygen distribution zone, a temperature gradient control device, and a CO / CO2 ratio sensor. The oxygen concentration in the three-stage oxygen distribution zone is dynamically controlled by a gas mixer, with an oxygen distribution error of ≤0.5%. The fuming furnace's cascade oxygen reduction process, through precise, staged oxygen control, effectively inhibits secondary solidification of metal oxides and slag encapsulation, reducing the formation of difficult-to-volatilize substances such as Fe3O4. This stage achieves efficient tin volatilization (SnO2 volatilization rate ≥95%) and targeted enrichment of copper and iridium (copper content in the slag ≥12wt%).

[0014] In an optional embodiment, during the raw material pretreatment stage, SiO2 or Fe3O4 is added to adjust the slag shape and control the Fe / SiO2 in the molten pool to be 1.2-1.5.

[0015] It should be noted that SiO2 is a typical acidic oxide and acts as an "acidic component" in the slag. Adding SiO2 can reduce the alkalinity of the slag and avoid excessive erosion of the furnace lining by high-alkalinity slag.

[0016] Fe₃O₄ decomposes in the slag or reacts with other slag components to form FeO (and even some Fe₂O₃), which is considered an "iron oxide component." Although FeO does not directly contribute to basicity calculations, the addition of Fe₃O₄ increases the total iron (TFe) content in the slag, indirectly affecting its oxidizability. It may also modulate reactivity by altering the ionic structure of the slag (e.g., by increasing low-valent iron ions).

[0017] Controlling the Fe / SiO2 ratio in the molten pool helps control the melting point and viscosity of the slag, optimize reaction kinetics, balance iron oxidation and metal recovery, stabilize the slag's oxidizability, avoid excessive metal oxidation, and optimize slag-iron separation. If the ratio is too high, iron loss will increase, while if it is too low, impurities may not be fully oxidized, affecting the final metal recovery rate.

[0018] In an optional embodiment, during the smelting stage, the turbulence intensity of the molten pool is ≥0.8 m / s, and the smelting time is 2-4 hours.

[0019] It should be noted that turbulence intensity refers to the intensity of the flow of molten metal in the molten pool. Reasonable control of turbulence intensity is conducive to promoting the uniformity of temperature and composition in the molten pool, accelerating the floating or dissolution of impurities (such as desulfurization and deoxidation reactions), reducing segregation; enhancing the renewal of reaction interfaces (such as slag-gold reaction), shortening refining time, and improving the recovery rate of alloy elements.

[0020] Reasonable control of smelting time is conducive to sufficient reaction and homogenization. If the time is too long, fuel / electricity consumption will increase directly and the yield of certain metals may also decrease. If the time is too short, the target material cannot be effectively recovered, which will also cause waste of resources.

[0021] In an optional embodiment, the air volume of the spray gun is adjusted to 0.8-1.5 Nm 3 / min·t, the spray gun insertion depth is 0.5-1.2m.

[0022] It should be noted that bubbles are formed after the gas is sprayed into the molten pool. During the rising process of the bubbles, friction with the melt produces a strong stirring effect. The air volume of the spray gun determines the gas input rate in the molten pool. By changing the number of bubbles, the rising speed and the contact area with the melt, it significantly affects the mass transfer, heat transfer and chemical reaction dynamics. When the air volume increases, the number of bubbles increases, the flow rate accelerates, the stirring intensity increases, the temperature field and composition field of the melt tend to be uniform, which is conducive to heat transfer. If the air volume is too small and the stirring is insufficient, the melt may be "stratified" (such as incomplete separation of the metal layer and the slag layer), resulting in local low temperature or incomplete oxidation of impurities (such as Fe, Pb), affecting the metal recovery or slag stability, and further separating the hardness slag and metal phases.

[0023] The depth of the spray gun insertion determines the contact point between the gas jet and the melt (such as the surface, middle, or bottom of the melt pool) and the depth of gas penetration into the melt, which directly affects the thermodynamic environment and kinetic efficiency of the reaction zone. If the insertion depth is small, the spray gun muzzle is close to the melt pool surface, and the gas jet acts primarily on the surface of the melt pool. The resulting bubbles have a short rising path, and stirring is concentrated on the surface, resulting in heat loss and possible insufficient reaction of the deeper melt. If the insertion depth is large, the spray gun muzzle penetrates deep into the melt pool, the gas jet penetrates the entire melt pool, the bubble rising path is long, stirring is more uniform, and heat is concentrated in the center of the melt pool. However, it should be noted that high-temperature melt will intensify erosion of the spray gun, increasing the frequency of spray gun replacement.

[0024] The embodiment of the present invention dynamically adjusts the depth of the spray gun according to the viscosity of the molten pool, effectively avoiding fluid splashing and improving oxygen utilization.

[0025] In an optional embodiment, during the smelting stage, the Fe content in the molten slag is 35%-45%, and the SiO2 content is 25%-30%.

[0026] In an optional embodiment, during the reduction stage, the residence time in the primary oxygen distribution zone is 30-50 min; the residence time in the secondary oxygen distribution zone is 20-30 min; and the residence time in the tertiary oxygen distribution zone is 10-20 min.

[0027] It should be noted that properly controlling the residence time is beneficial for optimizing reaction kinetics and mass transfer efficiency. If the residence time is too short, the metal oxides may not be fully reduced to metal, resulting in a reduced recovery rate. Alternatively, the metal vapor generated by the reduction process may be reoxidized to fine oxide particles by residual O₂ in the exhaust gas, reducing the recovery rate and increasing the dust treatment load. If the residence time is too long, some metals may be over-reduced, forming non-volatile alloys or agglomerates, which may block the flue or reduce slag fluidity.

[0028] In an optional embodiment, during the reduction stage, the CO / CO2 ratio in the reducing atmosphere is controlled to be 0.5-1.2, and the amount of reducing agent added is 4wt%-6wt% of the molten slag.

[0029] It should be noted that the combustion of CO to produce CO2 releases a significant amount of heat, providing a heat source for the molten pool or furnace. Properly increasing the CO / CO2 ratio (increasing CO concentration) can enhance combustion heat release, maintain furnace temperature, and promote the reduction of refractory components. CO2 acts as a "coolant," reacting with carbon (C) to produce CO, which absorbs heat and lowers furnace temperature. Excessively high CO2 ratios (e.g., >20%) can lead to insufficient furnace temperature and reduced reaction rates.

[0030] If the amount of reducing agent added is excessive, the excess reducing agent may participate in side reactions, hinder the reduction reaction, and significantly increase energy consumption; if the reducing agent is insufficient, the metal oxide cannot be completely reduced, resulting in metal loss; or secondary reduction may be required, increasing process costs.

[0031] The coordinated regulation of the CO / CO2 ratio and the amount of reducing agent added is conducive to ensuring complete reduction, improving efficiency and reducing costs.

[0032] The reducing agent is selected from at least one of coke, coal powder, charcoal, coke powder and anthracite, and the particle size of the reducing agent is ≤100 mesh.

[0033] It should be noted that the type of reducing agent can be reasonably selected according to actual needs.

[0034] The particle size of the reducing agent directly affects its contact area with the reactants, mass and heat transfer efficiency, furnace permeability, and process stability. Smaller particle sizes can significantly increase reaction rates and shorten reaction times by increasing specific surface area and reducing internal diffusion resistance. However, if the particle size is too large, the small specific surface area slows the reaction rate, potentially leading to uneven temperature distribution within the furnace (localized undercooling) and prolonged smelting cycles.

[0035] The use of a reasonable particle size of reducing agent is conducive to balancing the utilization rate of reducing agent to reduce energy consumption, inhibiting side reactions to improve product purity, and ultimately achieving the reduction process goal of "high efficiency, low consumption, and high quality".

[0036] In an optional embodiment, during the reduction stage, the copper-enriched slag has a copper content of ≥12% and a particle size range of 50-100 μm.

[0037] In an optional embodiment, the slag phase treatment includes water quenching the fumed slag to produce water-quenched slag; Among them, the water-slag ratio is (1-5): (2.5-3), the moisture content of the water-quenched slag is ≤5%, and the arsenic leaching toxicity is less than 0.05 mg / L.

[0038] It should be noted that the water-slag ratio can be reasonably set according to the actual amount of material processed, and can be selected from any one of 1:2.5, 1:3, 2:2.8, 3:2.5, 4:2.7, 5:2.5 and 5:3, or other values ​​within the range of (1-5): (2.5-3).

[0039] The slag treatment unit, connected to the fuming furnace outlet, includes a water quenching device and a copper-iridium flotation recovery system. Arsenic leaching toxicity is ≤ 0.05mg / L, allowing for direct and safe storage.

[0040] Water quenching involves rapidly placing high-temperature slag into a large volume of circulating water, utilizing the water's latent heat of vaporization to achieve "rapid cooling." This creates a dramatic temperature gradient within the slag, resulting in a sudden cooling of the surface while maintaining a high temperature within. This alters the slag's physical and chemical properties. This effectively inhibits secondary metal oxidation, improving the recovery rate of valuable metals; reduces thermal damage to equipment, extending its service life; and suppresses dust and harmful gas emissions, enhancing environmental benefits.

[0041] Properly setting the water-slag ratio optimizes cooling and ensures equipment safety; it also regulates slag structure and improves subsequent processing efficiency. A low water-slag ratio (low water, high slag) results in slow cooling, allowing minerals within the slag ample time to crystallize and grow, forming a dense, blocky structure (high hardness, low brittleness), which in turn increases energy consumption for subsequent crushing. A high water-slag ratio (high water, low slag) results in rapid cooling (even "quenching"), resulting in numerous microcracks within the slag due to the dramatic temperature gradient. This prevents minerals from crystallizing and results in a glassy or loose, porous structure (low hardness, high brittleness), making it easier to crush and separate.

[0042] In an optional embodiment, the slag phase treatment also includes flotation of the water-quenched slag, using butyl xanthate as the collector and pine oil as the frother. Flotation of the water-quenched slag achieves a copper recovery rate of 95%, with a final copper concentrate grade of ≥25%.

[0043] In an optional embodiment, the smelting method further includes cooling the flue gas in the smelting stage to 850-950° C. and then introducing it into the reduction stage, while controlling the arsenic content in the fumed slag to ≤0.3%.

[0044] It should be noted that the high-temperature flue gas (1100-1300°C) generated by the Ausmelt furnace is cooled to 850-950°C in a waste heat boiler and then directly introduced into the fuming furnace. This process uses the sensible heat of the flue gas to preheat the materials, reducing the fuming furnace's fuel consumption by ≥20%. This process involves a heat coupling unit, including a waste heat boiler, high-temperature flue gas ducting, and a fuming furnace preheating chamber, enabling direct utilization of the Ausmelt furnace's flue gas waste heat. In the heat coupling unit, the high-temperature flue gas ducting is lined with refractory material with a temperature resistance of ≥1400°C, and heat loss during flue gas transmission is ≤5%.

[0045] In summary, the short-process collaborative smelting method for complex tin materials according to the embodiment of the present invention includes the following steps: (1) Raw materials and pretreatment Add SiO2 or Fe3O4 to adjust the slag shape and control the Fe / SiO2 in the molten pool to 1.2-1.5.

[0046] (2) Melting stage An Ausmelt furnace is used for step-by-step oxygen distribution to increase the oxygen concentration in stages to produce molten slag; during this process, parameters such as oxygen concentration, temperature, spray gun air volume and spray gun insertion depth are set in the initial stage, middle stage and final stage.

[0047] (3) Reduction stage Use a fuming furnace to perform step-by-step oxygen reduction to complete tin volatilization and copper enrichment to produce fumed slag; During this period, the oxygen concentration, temperature, residence time, CO / CO2 ratio in the reducing atmosphere, amount of reducing agent added and particle size of the reducing agent in each oxygen distribution zone are controlled.

[0048] (4) Thermal coupling and environmental protection treatment Waste heat utilization: the flue gas from the Ausmelt furnace is cooled to 850-950℃ by the waste heat boiler and then passed into the fuming furnace for preheating treatment; Slag treatment: the fumed slag is water quenched, and the water-quenched slag is flotation treated to recover copper.

[0049] The smelting system of the embodiment of the present invention for realizing a short-process coordinated smelting method of complex tin materials includes: an Ausmelt furnace unit, equipped with a segmented oxygen enrichment supply device, an adjustable depth spray gun and an online monitoring module for the composition of the molten pool; a fuming furnace unit, with a built-in three-stage oxygen distribution zone, a temperature gradient control device and a CO / CO2 ratio sensor; a heat coupling unit, including a waste heat boiler, a high-temperature flue gas pipeline and a fuming furnace preheating chamber, to realize the direct utilization of the waste heat of the Ausmelt furnace flue gas; a slag treatment unit, connected to the slag outlet of the fuming furnace, including a water quenching device and a copper-iridium flotation recovery system.

[0050] The features and performance of the present invention are further described in detail below with reference to the embodiments.

[0051] Example 1 This embodiment provides a short-process collaborative smelting method for complex tin materials, including the following steps: (1) Raw materials and pretreatment Material composition: Tin concentrate contains Sn 15.2%, Cu 3.1%, As 2.3%, Fe 28%, S 8%, particle size ≤5mm.

[0052] Ore blending: Add SiO2 to adjust the slag shape so that the Fe / SiO2 ratio in the mixture is 1.3.

[0053] (2) During the smelting stage, the oxygen distribution parameters of the Ausmelt furnace are: Initial stage: oxygen concentration 26%, spray gun air volume 1.0 Nm 3 / min·t, the spray gun insertion depth is 0.8m, and the molten pool temperature is 1250℃; Mid-stage: oxygen concentration 34%, air volume 1.2 Nm 3 / min·t, spray gun insertion depth 1.0m, temperature 1300℃; Final stage: oxygen concentration 39%, air volume 1.5 Nm 3 / min·t, spray gun depth 0.6m, temperature 1380℃.

[0054] Melting effect: The smelting time was 3 hours, the slag shape was stable (Fe / SiO2 ratio 1.3), the tin direct recovery rate was 91.5%, copper was initially enriched in the slag (Cu content 5.8%), the Fe content in the molten slag was 35%, and the SiO2 content was 25%.

[0055] (3) In the reduction stage, the three-stage gradient oxygen reduction parameters of the fuming furnace are: Primary oxygen distribution zone: oxygen concentration 15%, temperature 1250℃, residence time 40min, CO / CO2 ratio 0.8; Secondary oxygen distribution zone: oxygen concentration 10%, temperature 1130℃, residence time 25min, CO / CO2 ratio 1.0; The third-stage oxygen distribution zone: oxygen concentration 5%, temperature 1020℃, residence time 15min, CO / CO2 ratio 1.2; Reducing agent: The addition amount is 5wt% of the molten slag, and the reducing agent is coal powder (particle size ≤ 100 mesh).

[0056] Separation effect: Tin volatilization rate 96.2% (SnO2 content in flue gas 12.5g / Nm 3 ); Copper is enriched in the slag in the form of copper-iridium (Cu-Fe alloy) (copper content 12.3%) with a particle size of 60-90μm.

[0057] (4) Thermal coupling and environmental protection treatment Waste heat utilization: The flue gas from the Ausmelt furnace (initial temperature 1200°C) is cooled to 900°C by the waste heat boiler and then passed into the fuming furnace, reducing fuel consumption by 23%.

[0058] Slag treatment: The fumed slag was quenched in water (water-slag ratio 1:3), with a moisture content of 4.5%, an arsenic content of 0.22%, and a leaching toxicity of 0.03 mg / L (lower than the national standard limit of 0.05 mg / L).

[0059] Copper-iridium recovery: Water-quenched slag is subjected to flotation (collector butyl xanthate, frother pine oil), with a copper recovery rate of 95% and a final copper concentrate grade of ≥25%.

[0060] Comprehensive technical indicators: Metal recovery rate: Sn 91.5%, Cu 87%, Sb 76%.

[0061] Energy consumption: The comprehensive energy consumption per ton of tin is 0.92 tons of standard coal (1.25 tons for traditional process).

[0062] Environmental protection: Arsenic emissions are reduced by 65% ​​and CO2 emissions are reduced by 21%.

[0063] Example 2 This embodiment provides a short-process collaborative smelting method for complex tin materials, which is a secondary resource recovery method for complex tin slag, and includes the following steps: (1) Raw materials and pretreatment Material composition: smelting waste slag contains Sn 8.5%, Cu 2.6%, As 1.8%, Fe 35%, SiO220%, particle size ≤10mm.

[0064] Ore blending: Add Fe3O4 to adjust the Fe / SiO2 ratio to 1.4.

[0065] (2) In the furnace smelting stage, the oxygen distribution parameters of the Ausmelt furnace are: Initial stage: oxygen concentration 25%, air volume 0.9 Nm 3 / min·t, spray gun depth 1.0m, temperature 1220℃; Mid-stage: oxygen concentration 35%, air volume 1.3 Nm 3 / min·t, spray gun depth 0.7m, temperature 1280℃; Final stage: oxygen concentration 40%, air volume 1.4 Nm 3 / min·t, spray gun depth 0.5m, temperature 1300℃.

[0066] Melting effect: The smelting time is 4 hours, the slag Fe / SiO2 ratio is 1.4, the tin direct recovery rate is 89.8%, the copper enrichment is 6.2%, the Fe content in the molten slag is 40%, and the SiO2 content is 27%.

[0067] (3) In the reduction stage, the three-stage gradient oxygen reduction parameters of the fuming furnace are: Primary oxygen distribution zone: oxygen concentration 14%, temperature 1230℃, residence time 45min, CO / CO2 ratio 0.8; Secondary oxygen distribution zone: oxygen concentration 9%, temperature 1120℃, residence time 30min, CO / CO2 ratio 1.0; The third-stage oxygen distribution zone: oxygen concentration 4%, temperature 1010℃, residence time 20min, CO / CO2 ratio 1.2; Reducing agent: The addition amount is 4.5wt% of the molten slag, and the reducing agent is coke powder (particle size ≤ 80 mesh).

[0068] Separation effect: The tin volatilization rate is 94.5%, the copper content in the copper-iridium slag is 11.8%, and the particle size is 50-80μm.

[0069] (4) Thermal coupling and economic benefits Waste heat utilization: Flue gas waste heat reduces the fuel consumption of the fuming furnace by 21%.

[0070] Metal recovery rate: Sn 90.68%, Cu 86.16%, Sb 73.69%.

[0071] Cost analysis: 10,000 tons of waste slag were processed, saving 800 tons of standard coal and reducing arsenic treatment costs by RMB 1.2 million.

[0072] In summary, the present invention significantly improves the resource utilization efficiency of complex tin materials with high sulfur, high arsenic, and high iron content through the use of cascade oxygen distribution and dynamic mass transfer optimization technology, combined with the "Ausmelt furnace-fuming furnace" collaborative smelting system: the direct recovery rate of tin is increased from 75%-80% in the traditional process to ≥85%, and the recovery rates of copper and antimony reach 85% and 70% respectively; through the coupling of double furnace waste heat and process simplification, the comprehensive energy consumption is reduced by 25% (energy consumption per ton of tin ≤1.0 ton of standard coal), and the material transfer cost is reduced by 20%; in terms of environmental protection, the arsenic content of the smelting slag is stable at ≤0.3%, the leaching toxicity is lower than the national standard limit (0.05 mg / L), and the flue gas CO2 emissions are reduced by 18%-22%.

[0073] The foregoing description is merely a preferred embodiment of the present invention and is not intended to limit the present invention. Those skilled in the art will readily appreciate that various modifications and variations of the present invention are possible. Any modifications, equivalent substitutions, or improvements made within the spirit and principles of the present invention are intended to be within the scope of protection of the present invention.

Claims

1. A short-process collaborative smelting method for complex tin materials, characterized in that: The steps include: Raw material pretreatment stage, smelting stage, reduction stage and slag phase treatment stage; In the smelting stage, the complex tin material is added to the Ausmelt furnace, and the oxygen concentration is increased in stages using a step-by-step oxygen enrichment process to produce molten slag. In the initial stage, the oxygen concentration is 25%-28% and the temperature is 1200-1250°C; in the middle stage, it is increased to 33%-35% and the temperature is 1250-1300°C; in the final stage, it is increased to 38%-40% and the temperature is 1300-1400°C. Reduction stage: The obtained molten slag is introduced into a fuming furnace, and a step-by-step oxygen reduction process is adopted to complete tin volatilization and copper enrichment to obtain fumed slag; among them, the oxygen concentration in the first-level oxygen distribution zone is 14%-16%, and the temperature is 1200-1300℃; the oxygen concentration in the second-level oxygen distribution zone is 9%-11%, and the temperature is 1100-1150℃; the oxygen concentration in the third-level oxygen distribution zone is 4%-6%, and the temperature is 1000-1050℃.

2. The smelting method according to claim 1, characterized in that: During the raw material pretreatment stage, SiO2 or Fe3O4 is added to adjust the slag shape and control the Fe / SiO2 in the molten pool to be 1.2-1.

5.

3. The smelting method according to claim 1, characterized in that: During the smelting stage, the turbulence intensity of the molten pool is ≥0.8m / s, and the smelting time is 2-4 hours; And / or, adjust the air volume of the spray gun to 0.8-1.5 Nm 3 / min·t, the spray gun insertion depth is 0.5-1.2m.

4. The smelting method according to claim 1, characterized in that: During the smelting stage, the Fe content in the molten slag is 35%-45%, and the SiO2 content is 25%-30%.

5. The smelting method according to claim 1, characterized in that: In the reduction stage, the residence time of the primary oxygen distribution zone is 30-50 minutes; the residence time of the secondary oxygen distribution zone is 20-30 minutes; and the residence time of the tertiary oxygen distribution zone is 10-20 minutes.

6. The smelting method according to claim 1, characterized in that: In the reduction stage, the CO / CO2 ratio in the reducing atmosphere is controlled to be 0.5-1.2, and the amount of reducing agent added is 4wt%-6wt% of the molten slag; The reducing agent is selected from at least one of coke, coal powder, charcoal, coke powder and anthracite, and the particle size of the reducing agent is ≤100 mesh.

7. The smelting method according to claim 1, characterized in that: In the reduction stage, the copper content in the copper-enriched slag is ≥11%, and the particle size range is 50-100 μm.

8. The smelting method according to claim 1, characterized in that: The slag phase treatment includes water quenching the fumed slag to obtain water-quenched slag; Among them, the water-slag ratio is (1-5): (2.5-3), the moisture content of the water-quenched slag is ≤5%, and the arsenic leaching toxicity is less than 0.05 mg / L.

9. The smelting method according to claim 8, characterized in that: The slag phase treatment also includes flotation treatment of the water-quenched slag, wherein the collector in the flotation stage is butyl xanthate and the foaming agent is pine oil.

10. The smelting method according to claim 1, characterized in that: The smelting method further includes cooling the flue gas from the smelting stage to 850-950° C. and then introducing the flue gas into the reduction stage, while controlling the arsenic content in the fumed slag to ≤0.3%.