A copper-lead-zinc mixed material smelting system and smelting method

Abstract

The application discloses a copper-lead-zinc mixed material smelting system and a smelting method, and particularly relates to the field of pyrometallurgy technology, and comprises an oxidation side-blown furnace, a reduction side-blown furnace and a fuming furnace, wherein the oxidation side-blown furnace, the reduction side-blown furnace and the fuming furnace are arranged in a triangular shape, and an electric heating front bed is arranged on the fuming furnace; the oxidation side-blown furnace, the reduction side-blown furnace and the electric heating front bed are connected through a chute, and the electric heating front bed and the fuming furnace are connected through a siphon slag discharge port. The bed area ratio of the oxidation side-blown furnace, the reduction side-blown furnace and the fuming furnace is A: 0.7A~0.8A: 0.5A~0.6A, and the three furnaces are arranged in a triangular shape, so that the slag production load, the metal reduction amount and the fuming and volatilization treatment amount of high-sulfur and low-sulfur copper-lead-zinc mixed materials under various working conditions can be accurately matched, and the defects of solidification and fixed material flow direction of the linear arrangement structure are avoided.

Description

Technical Field

[0001] The present invention relates to the technical field of pyrometallurgy, and specifically relates to a smelting system and smelting method for copper-lead-zinc miscellaneous materials. Background Art

[0002] In the copper, lead and zinc smelting process, various smelting miscellaneous materials with complex components and greatly different physical properties are generated, such as copper dust, copper slag, lead dust, lead slag, zinc leaching residue and iron slag, recycled waste lead batteries, lead paste, blast furnace dust produced by steel mills, etc. These miscellaneous materials generally contain valuable metals such as copper, lead and zinc, and also contain harmful impurities such as sulfur, arsenic and chlorine. If directly stockpiled, it will not only cause the loss of metal resources, but also cause serious environmental pollution.

[0003] At present, most of these miscellaneous materials are recovered by oxygen-enriched smelting methods, and the main processes include single-furnace smelting, double-furnace smelting, three-furnace smelting and the like. Among them, the single-furnace smelting adopts the "one furnace with three sections" mode, in which oxidation, reduction and fuming are completed in one smelting furnace. In this method, the flue gas in the oxidation section has a high SO₂ concentration and can be directly used for acid production, while the flue gas in the reduction and fuming sections is strongly reductive and has a low SO₂ concentration, which cannot be incorporated into the acid production system, resulting in a complicated treatment process and increased cost; the double-furnace smelting adopts the "one furnace with two sections" mode, in which oxidation and reduction are completed in one smelting furnace, and the smelting slag is treated in a matching fuming furnace. This method also has the problem that the low-sulfur flue gas in the reduction section is difficult to be directly used for acid production. The three-furnace smelting adopts two smelting furnaces and one fuming furnace arranged in a straight line in series, that is, the "three-connected furnace", in which oxidation, reduction and fuming are sequentially completed in the three furnaces respectively. Although separate collection of flue gas can be realized, the linear arrangement has a fixed structure and a fixed material flow direction, which can only realize single-sequence production and cannot flexibly adjust the process flow according to the composition of the miscellaneous materials.

[0004] In order to shorten the lead smelting process, Chinese Patent No. 201621416351.1 once disclosed a two-step lead smelting device with one bottom-blowing smelting furnace and two top-bottom combined blowing reduction furnaces arranged in a triangular layout. Lead concentrate is smelted in the bottom-blowing smelting furnace to produce part of crude lead and high-lead slag, and the high-lead slag alternately flows into the two parallel top-bottom combined blowing reduction furnaces through launders to further recover the remaining lead, thereby shortening the production cycle. However, materials cannot be exchanged between the two parallel reduction furnaces of the same specification in this triangular furnace device, which only serve as backups for each other for alternating production, and have no switchable multifunctionality. Using the reduction furnaces for lead reduction and high-lead slag depletion, that is, the "one furnace with two sections" mode, will cause a large amount of lead to volatilize into the dust, reduce the direct recovery rate of lead, and the treated materials are also relatively single, which limits its ability to widely treat complex copper-lead-zinc materials.

[0005] In summary, the above smelting devices and smelting methods all adopt the production mode of "one material, one furnace, one strategy", which cannot adapt to the collaborative treatment of multi-component miscellaneous materials with mixed high and low sulfur and large composition fluctuations, and have technical shortcomings such as weak working condition switching capability, high energy consumption, poor raw material adaptability, and low comprehensive recovery rate of valuable metals. Summary of the Invention

[0006] The purpose of this invention is to provide a smelting system and method for copper, lead, and zinc scraps to solve the problems mentioned in the background art.

[0007] To achieve the above objectives, the present invention provides the following technical solution: a smelting system for copper, lead, and zinc scraps, comprising: The furnace includes an oxidation side-blown furnace, a reduction side-blown furnace, and a fuming furnace, arranged in a triangular pattern. The fuming furnace is equipped with an electric heating front bed. The hearth area of ​​the side-blown oxidation furnace is Am. 2 The value of A ranges from 6 to 30, and the hearth area of ​​the reduction side-blown furnace is 0.7-0.8A m². 2 The furnace bed area of ​​the fuming furnace is 0.5-0.6 Å m². 2 The electric heating front bed is located at the feed front end of the fuming furnace, and the furnace bed area of ​​the electric heating front bed is 0.5-0.6 Å m². 2 ; The oxidation side-blown furnace, reduction side-blown furnace, and electric heating forecourt are connected by chutes, and the material conveying channels between the three furnace bodies are interconnected. Each chute is connected to the feed port and discharge port of the corresponding furnace, so that the metal oxide melt produced by oxidation smelting and the reducing slag or crude metal melt produced by reduction smelting can be selectively fed into the electric heating forecourt according to the smelting conditions. The oxidation side-blown furnace, reduction side-blown furnace, and fuming furnace are each equipped with independent flue gas collection pipelines; the flue gas collection pipeline of the oxidation side-blown furnace is connected to the acid production system after waste heat recovery and dust collection, and the flue gas collection pipelines of the reduction side-blown furnace and the fuming furnace are connected to the tail gas treatment system after waste heat recovery and dust collection. The electric heating forebed is connected to the fuming furnace via a siphon slag discharge port. The siphon slag discharge port is located at the lower part of the side wall of the electric heating forebed. The effective volume of the electric heating forebed is matched with the slag volume of one batch processed by the fuming furnace. The inlet height of the siphon slag discharge port is determined according to the molten slag flow rate and the processing cycle of the fuming furnace. The inlet height of the siphon slag discharge port is set such that when the newly introduced molten slag in the electric heating forebed causes the liquid level to reach the inlet height of the siphon slag discharge port, the amount of slag remaining in the forebed is exactly the amount of slag processed by one batch of the fuming furnace, and the liquid level difference generated by the newly fed slag is sufficient to drive the molten slag to flow into the fuming furnace continuously at a stable flow rate.

[0008] Furthermore, the oxidation side-blown furnace, reduction side-blown furnace, and fuming furnace all adopt a single-furnace-one-section independent operation mode, and the material conveying channels between the three furnaces are interconnected. The smelting system can be switched to any one of the following five operation modes: Mode 1: The oxidation side-blown furnace, reduction side-blown furnace, electric heating forecourt, and fuming furnace are connected in series. Mode 2: Two-stage operation: oxidation side-blown furnace and reduction side-blown furnace connected in series; Mode 3: The side-blown furnace, the electric heating forecourt, and the fuming furnace are connected in series in two stages; Mode 4: Two-stage operation in series: oxidation side-blown furnace, electric heating forecourt, and fuming furnace; Mode 5: The oxidation side-blown furnace and the reduction side-blown furnace are fed in parallel and independently. The melt or slag produced by the two furnaces is fed into the electrothermal forecourt, and after heat preservation and sedimentation, it is fed into the fuming furnace for parallel and coordinated operation.

[0009] Furthermore, the feed ends of the oxidation side-blown furnace and the reduction side-blown furnace are connected to a cylindrical granulator, and the feed end of the cylindrical granulator is connected to a batching silo. The dust collected by the dust collection devices of the oxidation side-blown furnace, the reduction side-blown furnace, the fuming furnace, and the electric heating front bed is transported back to the batching silo for recycling or sold directly.

[0010] A method for smelting copper-lead-zinc scrap using the above-mentioned copper-lead-zinc scrap smelting system includes the following steps: S1 Batching: Mix copper, lead, and zinc scraps with flux; the flux is at least one of limestone, quartz, iron filings, or iron sulfides; the copper, lead, and zinc scraps are divided into two categories: high-sulfur scraps and low-sulfur scraps. The high-sulfur scraps include copper dust, zinc leaching slag, lead-silver slag, oxygen leaching slag, and lead paste, with a sulfur content of 7% to 50%. The low-sulfur scraps include steel mill dust, lead dust, and lead refining slag, with a sulfur content of 2% to 4%. S2 Granulation: The prepared mixture is fed into a cylindrical granulator for mixing and granulation. The particle size of the granulated material is controlled to be 8~10mm and the moisture content is 8%~10%. S3 Oxidation Smelting: High-sulfur scrap is fed into an oxidation side-blown furnace, and granular coal with a particle size of 5-10mm and accounting for 6%-14% of the mass of the high-sulfur scrap is added. Oxygen-enriched air with an oxygen volume concentration of 70%-80% is introduced, and the smelting temperature is controlled at 1200-1250℃ and the smelting time is 2-2.5h. The high-sulfur dust-containing flue gas produced by oxidation is sent to the acid production system after waste heat recovery and dust collection, while metal oxide melt is produced at the same time. S4 Reduction Smelting: Desulfurized metal oxide melt or low-sulfur impurities are fed into a reduction side-blown furnace, and granular coal with a particle size of 5-10mm and accounting for 14%-16% of the corresponding material mass is added. Oxygen-enriched air with an oxygen volume concentration of 45%-55% is introduced, and the temperature is controlled at 1250-1300℃ for 2-2.5 hours. The low-sulfur dust-containing flue gas produced by reduction is sent to the tail gas absorption treatment system after waste heat recovery and dust collection. At the same time, crude metal or lead matte and reduction smelting slag are produced. S5 Electrothermal Forebed Precipitation: Oxidized or reduced smelting slag is fed into the electrothermal forebed and precipitated at 1250~1300℃ for 1.5~2.0h to separate valuable metals; the electrothermal forebed retains the amount of smelting slag to be processed in one furnace. When the next furnace smelting slag is added, the existing slag in the forebed is squeezed by the new feed slag and continuously discharged into the fuming furnace through the siphon slag discharge port; S6 continuous blowing of fuming material: 18% to 20% of pulverized coal by mass of slag is injected into the fuming furnace, the temperature is controlled at 1250 to 1300℃ and the blowing time is 1.5 to 2.0 hours; under a strong reducing atmosphere, lead and zinc metal oxides are reduced and volatilized into the flue gas, and the flue gas is sent to the tail gas treatment system after waste heat recovery and dust collection. The collected flue gas is returned to the batching process or sold externally, and the slag is finally sold externally after water quenching. The five operating modes can be freely switched according to the sulfur and lead content of the impurities to be processed.

[0011] Furthermore, the switching logic for the five operating modes is as follows: When the sulfur content of the impurities is ≥7% and the lead content is ≥30%, mode one is adopted, and steps S3, S4, S5 and S6 are executed in sequence, that is, the oxidation side-blown furnace, the reduction side-blown furnace, the electric heating forebed and the fuming furnace are connected in series. When the sulfur content of the scrap is ≥7% and no fuming recovery is required, mode 2 is adopted, and steps S3 and S4 are executed in sequence. The reduced smelting slag is directly water-quenched and sold externally, that is, the oxidation side-blown furnace and the reduction side-blown furnace are connected in series. When the sulfur content of the impurities is <7% and the lead content is ≥30%, mode 3 is adopted, and steps S4, S5 and S6 are executed in sequence, that is, the two-stage operation of the reduction side-blown furnace, the electric heating forebed and the fuming furnace is connected in series. When the sulfur content of the scrap is ≥7% and the lead content is <30%, mode four is adopted, and steps S3, S5 and S6 are executed in sequence, namely, the oxidation side-blown furnace, the electric heating forebed and the fuming furnace are connected in series. When high-sulfur and low-sulfur materials need to be processed simultaneously, mode five is adopted. High-sulfur materials are processed in step S3 and low-sulfur materials are processed in step S4. The melt or slag produced by the two streams are fed into the electric heating front bed through their respective chutes and then enter the fuming furnace together. That is, the oxidation side-blown furnace and the reduction side-blown furnace are fed independently and operate in parallel.

[0012] Furthermore, when operating in Mode 1, the lead recovery rate is >87%, the zinc recovery rate is >80%, and the copper recovery rate is >82% after processing.

[0013] Furthermore, when using the second mode of operation, it is suitable for high-sulfur copper dust or lead paste materials, and the copper recovery rate can reach 94.4% when treating high-sulfur copper dust.

[0014] Furthermore, when using the aforementioned Mode 3 operation, it is suitable for low-sulfur lead-zinc mixtures or steel mill dust. When treating steel mill dust alone, the zinc recovery rate can reach 92.4%.

[0015] Furthermore, when using the aforementioned Mode 4 operation, the zinc recovery rate can reach 93.1% when treating zinc leaching residue.

[0016] Furthermore, when operating in Mode 5, the steel plant's flue gas is directly fed into the fuming furnace, where the residual heat of the slag is used to complete the melting and reduction volatilization.

[0017] The technical effects and advantages provided by the present invention in the above technical solution are as follows: 1. This invention breaks through the rigid production mode of "one material, one furnace, one policy" in the traditional way. It can freely switch at least five smelting modes according to the sulfur content of the material, the type of metal and the recycling target. High sulfur impurities are removed through the whole process, and the process is simplified for low sulfur and lead-zinc impurities, realizing "one furnace for multiple uses and production on demand", and completely solving the defects of poor versatility of existing technologies. 2. In this invention, the furnace bed area ratio of the oxidation side-blown furnace, the reduction side-blown furnace, and the fuming furnace is A: 0.7A~0.8A: 0.5A~0.6A, and the three furnaces are arranged in a triangular pattern. Compared with the straight arrangement of the traditional three-furnace, it can accurately match the slag load of high and low sulfur copper, lead, zinc and other miscellaneous materials, the metal reduction amount and the fuming volatilization treatment amount. It avoids the defects of the rigid structure and fixed material flow direction of the straight arrangement, and also overcomes the limitation of the existing triangular furnace where two reduction furnaces of the same specifications serve as backups for each other and there is no material exchange. 3. In this invention, the oxidation side-blown furnace, the reduction side-blown furnace, and the fuming furnace are each equipped with independent flue gas collection pipelines. The high-sulfur flue gas produced by the oxidation side-blown furnace is directly used for acid production, while the low-sulfur reducing flue gas produced by the reduction side-blown furnace and the fuming furnace is uniformly sent to the tail absorption treatment system. This not only improves the utilization rate of sulfur dioxide resources but also reduces the cost of environmental protection and solves the industry problem that low-sulfur flue gas is difficult to incorporate into the acid production system in the traditional "one furnace, multiple stages" process. 4. The present invention retains the amount of smelting slag to be processed in one fuming furnace in the electric heating front bed. By utilizing the liquid level difference of the newly fed slag, the slag retained in the front bed is continuously pressed into the fuming furnace through the siphon slag discharge port, forming a cycle operation of "electric heating front bed heat preservation and sedimentation, siphon continuous slag discharge, and continuous fuming blowing", which significantly improves production efficiency, reduces energy consumption, and reduces equipment investment. Attached Figure Description

[0018] To more clearly illustrate the technical solutions in the embodiments of this application or the prior art, the drawings used in the embodiments will be briefly introduced below. Obviously, the drawings described below are only some embodiments recorded in this invention. For those skilled in the art, other drawings can be obtained based on these drawings.

[0019] Figure 1 This is a schematic diagram of the triangular arrangement of the smelting system of the present invention; Figure 2 This is a schematic diagram of the five operating modes of the smelting system of the present invention; Figure 3 This is a process flow diagram of the smelting method of the present invention. Detailed Implementation

[0020] To enable those skilled in the art to better understand the technical solution of the present invention, the present invention will be further described in detail below with reference to the accompanying drawings.

[0021] This invention provides, for example Figures 1 to 2 The smelting system shown includes: The furnace includes an oxidation side-blown furnace, a reduction side-blown furnace, and a fuming furnace, arranged in a triangular pattern. The fuming furnace is equipped with an electric heating front bed. The hearth area of ​​the side-blown oxidation furnace is Am. 2 The value of A ranges from 6 to 30, and the hearth area of ​​the reduction side-blown furnace is 0.7-0.8A m². 2 Preferably 0.75A m 2 ; The hearth area of ​​the fuming furnace is 0.5-0.6 Å m². 2 Preferably 0.55A m 2 ; The electric heating front bed is located at the feed front end of the fuming furnace, and the furnace bed area of ​​the electric heating front bed is 0.5-0.6 Å m². 2 Preferably 0.55A m 2 .

[0022] The above area ratio is designed based on the following: the slag load of high and low sulfur copper, lead and zinc scraps, the metal reduction amount and the fuming and volatilization treatment amount are matched with each other in an oxidizing side-blown furnace, a reducing side-blown furnace and a fuming furnace; the volume of the electric heating forecourt is matched with the processing capacity of one furnace of the fuming furnace to ensure the stability of continuous blowing.

[0023] The oxidation side-blown furnace, the reduction side-blown furnace, and the electric heating forecourt are connected by chutes, and the material conveying channels between the three furnaces are interconnected, allowing the smelting system to switch to any of the following material flow directions: The molten metal oxide produced by the oxidation side-blown furnace is fed into the reduction side-blown furnace via a chute, and the reducing slag produced by the reduction side-blown furnace is fed into the electric heating forecourt via a chute. The molten metal oxide produced by the side-blown oxidation furnace is directly fed into the electric heating forecourt via a chute. The reduction slag produced from the low-sulfur mixed materials processed in the reduction side-blown furnace is fed into the electric heating forecourt via a chute; The molten metal oxide produced by the oxidation side-blown furnace and the reducing slag produced by the reduction side-blown furnace are simultaneously fed into the electric heating forebed via their respective chutes.

[0024] Each of the chutes is connected to the feed inlet and discharge outlet of the corresponding furnace.

[0025] The oxidation side-blown furnace, reduction side-blown furnace, and fuming furnace are each equipped with independent flue gas collection pipelines. The high-concentration SO2 flue gas produced by the oxidation side-blown furnace undergoes waste heat recovery in a waste heat boiler and dust collection by an electrostatic precipitator or bag filter before being sent to the acid production system through independent pipelines to produce sulfuric acid. The low-sulfur reducing flue gas produced by the reduction side-blown furnace and the low-sulfur reducing flue gas produced by the fuming furnace undergo their respective waste heat recovery and dust collection devices before being merged into the same tail-end treatment system for desulfurization and valuable metal recovery. The electric heating forebed is connected to the fuming furnace via a siphon slag discharge port. The siphon slag discharge port is located at the lower part of the side wall of the electric heating forebed. The effective volume of the electric heating forebed is matched with the slag volume of one batch processed by the fuming furnace. The inlet height of the siphon slag discharge port is determined according to the molten slag flow rate and the processing cycle of the fuming furnace. The inlet height of the siphon slag discharge port is set such that when the newly introduced molten slag in the electric heating forebed causes the liquid level to reach the inlet height of the siphon slag discharge port, the amount of slag remaining in the forebed is exactly the amount of slag processed by one batch of the fuming furnace, and the liquid level difference generated by the newly fed slag is sufficient to drive the molten slag to flow into the fuming furnace continuously at a stable flow rate.

[0026] The electric heating forebed is equipped with an electrode heating device to keep the molten slag at a temperature of 1250~1300℃. The operation mode of the electric heating forebed is as follows: first, the amount of molten slag to be processed in one batch of fuming furnace is injected into the bed. When the molten slag of the next batch is added to the electric heating forebed through the chute, the newly added high-temperature molten slag raises the liquid level. The static pressure generated by the molten slag level difference is used to continuously press the previously retained molten slag into the fuming furnace from bottom to top through the siphon slag discharge port, thereby realizing the continuous operation of the electric heating forebed and the fuming furnace.

[0027] The feed ends of the oxidation side-blown furnace and the reduction side-blown furnace are connected to a cylindrical granulator, and the feed end of the cylindrical granulator is connected to a batching silo. The dust collected by the dust collection devices of the oxidation side-blown furnace, the reduction side-blown furnace, the fuming furnace and the electric heating front bed is transported back to the batching silo for recycling or sold directly.

[0028] The oxidation side-blown furnace, reduction side-blown furnace, and fuming furnace all adopt a single-furnace-one-section independent operation mode, and the material conveying channels between the three furnaces are interconnected. The smelting system can be switched to any one of the following five operation modes: Mode 1: The oxidation side-blown furnace, reduction side-blown furnace, electric heating forecourt, and fuming furnace are connected in series. Mode 2: Two-stage operation: oxidation side-blown furnace and reduction side-blown furnace connected in series; Mode 3: The side-blown furnace, the electric heating forecourt, and the fuming furnace are connected in series in two stages; Mode 4: Two-stage operation in series: oxidation side-blown furnace, electric heating forecourt, and fuming furnace; Mode 5: The oxidation side-blown furnace and the reduction side-blown furnace are fed in parallel and independently. The melt or slag produced by the two furnaces is fed into the electrothermal forecourt, and after heat preservation and sedimentation, it is fed into the fuming furnace for parallel and coordinated operation.

[0029] The smelting method of the present invention can freely switch between five operating modes according to the sulfur and lead content of the impurities to be processed. In each mode, a single furnace body is operated in a "one furnace, one section" manner.

[0030] The miscellaneous materials are divided into two categories: High-sulfur mixed materials, with a sulfur content of 7% to 50%, include: Copper dust: Cu 6~18%, Pb 6~12%, Zn 3~8%, S 8~10%; Zinc leaching residue: Zn 3~16%, Pb 4~19%, S 7~9%; Lead-silver slag: Zn 3~6%, Pb 26~36%, S 8~9%; Oxygen leaching residue: Zn 2~3%, Pb 8~35%, S 10~50%; Lead paste: Pb 60~70%, S 7~9%.

[0031] Low-sulfur mixed materials, with a sulfur content of 2% to 4%, include: Steel mill dust: Zn 10~15%, S 3~4%; Lead fumes: Pb 30~40%, Zn 8~10%, S 3~4%; Lead refining slag: Pb 25~60%, Cu 5~7%, S 2~3%.

[0032] The smelting method includes the following steps: S1 Batching: Mix one or more of the above-mentioned miscellaneous materials with flux (at least one of limestone, quartz, iron filings or iron sulfides) in a batching bin in proportion.

[0033] S2 Granulation: The mixed materials are fed into a cylindrical granulator for mixing and granulation. The particle size of the granulated material is controlled to be 8~10mm and the moisture content is 8%~10% to ensure that the material is mixed evenly and fully wetted, and to prevent dust from flying after entering the furnace.

[0034] S3 Oxidation Smelting: High-sulfur scrap is fed into an oxidation side-blown furnace, and granular coal with a particle size of 5-10 mm, accounting for 6%-14% of the high-sulfur scrap mass, is added as fuel. Oxygen-enriched air with an oxygen volume concentration of 70%-80% is introduced, and the smelting temperature is controlled at 1200-1250℃ for 2-2.5 hours. Under the oxidizing atmosphere, metal sulfides are oxidized to metal oxides, and sulfur enters the flue gas in the form of SO2. The high-sulfur, dust-laden flue gas produced by oxidation is sent to the acid production system after waste heat recovery and dust collection, and the produced molten metal oxides are sent to the next process via a chute.

[0035] S4 Reduction Smelting: Desulfurized molten metal oxides or low-sulfur impurities are fed into a reduction side-blown furnace. Coal granules (5-10 mm in diameter) accounting for 14%-16% of the corresponding material mass are added, and oxygen-enriched air with a volume concentration of 45%-55% is introduced. The temperature is controlled at 1250-1300℃, and the smelting time is 2-2.5 hours. Under the reducing atmosphere, the metal oxides are reduced by CO to crude metal or lead matte. The low-sulfur, dust-laden flue gas produced from the reduction process is sent to the tail-end treatment system after waste heat recovery and dust collection. The crude metal or lead matte produced is enriched and recovered, and the reduction smelting slag is fed into the electrothermal forecourt.

[0036] S5 Electrothermal Forebed Precipitation: Oxidized or reduced smelting slag is fed into the electrothermal forebed and precipitated at 1250~1300℃ for 1.5~2.0h. During the precipitation process, denser valuable metals such as lead matte settle to the bottom of the bed and are periodically discharged and recovered from the bottom outlet; less dense smelting slag floats to the top and is continuously discharged into the fuming furnace through the siphon slag discharge port. The electrothermal forebed retains the amount of smelting slag to be processed for one fuming furnace cycle. When the next batch of smelting slag is added, the existing slag in the forebed is squeezed by the new feed slag and continuously discharged into the fuming furnace through the siphon slag discharge port.

[0037] S6 continuous blowing furnace: Pulverized coal, accounting for 18%~20% of the slag mass, is injected into the furnace as a reducing agent and fuel. The temperature is controlled at 1250~1300℃, and the blowing time is 1.5~2.0h. Under a high-temperature, strong reducing atmosphere, lead and zinc metal oxides are reduced by CO to metal vapor, which then volatilizes. The metal vapor enters the subsequent cooling and dust collection system with the flue gas, where it is oxidized into metal oxide dust during the cooling process and collected. After waste heat recovery and dust collection, the flue gas is sent to the tail-end treatment system. The dust collected is returned to the batching process or sold externally, and the produced slag is sold externally after water quenching.

[0038] The system can freely switch between five operating modes based on the sulfur and lead content of the impurities to be processed. The switching logic for the five operating modes is as follows: When the sulfur content of the scrap is ≥7% and the lead content is ≥30%, mode one is adopted, and steps S3, S4, S5 and S6 are executed in sequence, that is, the oxidation side-blown furnace, the reduction side-blown furnace, the electric heating forebed and the fuming furnace are connected in series. When the sulfur content of the scrap is ≥7% and no fuming recovery is required, mode 2 is adopted, and steps S3 and S4 are executed in sequence. The reduced smelting slag is directly water-quenched and sold externally, that is, the oxidation side-blown furnace and the reduction side-blown furnace are connected in series. When the sulfur content of the impurities is <7% and the lead content is ≥30%, mode 3 is adopted, and steps S4, S5 and S6 are executed in sequence, that is, the two-stage operation of the reduction side-blown furnace, the electric heating forebed and the fuming furnace is connected in series. When the sulfur content of the scrap is ≥7% and the lead content is <30%, mode four is adopted, and steps S3, S5 and S6 are executed in sequence, namely, the oxidation side-blown furnace, the electric heating forebed and the fuming furnace are connected in series. When it is necessary to process high-sulfur and low-sulfur impurities at the same time, mode five is adopted. High-sulfur impurities are processed in step S3 and low-sulfur impurities are processed in step S4. The melt or slag produced by both streams is fed into the electric heating front bed and then enters the fuming furnace in a unified manner. That is, the oxidation side-blown furnace and the reduction side-blown furnace are fed independently and operate in parallel and in a coordinated manner.

[0039] The main chemical reactions in the oxidation smelting stage are: MeS + 1.5O2 → MeO + SO2 MeSO4 → MeO + SO3 MeSO4 + MeS → 2Me + 2SO2 The main reaction occurring during the reduction smelting stage is: MeO + CO → Me + CO2 During the fuming and blowing stage, metal oxides are reduced to metal vapor and volatilized. MeO + CO → Me + CO2 The volatilized metal vapors are oxidized into metal oxide dust during flue gas cooling and dust collection. Me + 0.5O2 → MeO Me represents at least one metal selected from Cu, Pb, and Zn.

[0040] In the following examples, unless otherwise specified, all material components are expressed as mass percentages (wt%).

[0041] Examples 1 to 6 below adopt the Mode 1 operation method, that is, the oxidation side-blown furnace, reduction side-blown furnace, electric heating forecourt, and fuming furnace are connected in series, and steps S3, S4, S5, and S6 are executed sequentially. This mode is suitable for mixed materials with high sulfur and lead content ≥30%. It requires first entering the oxidation side-blown furnace for desulfurization and acid production, then entering the reduction side-blown furnace to recover crude lead or lead matte, and the reduction slag entering the fuming furnace to recover lead and zinc.

[0042] Example 1: Lead paste (Pb 65%, S 8%), zinc leaching residue (Zn 4%, Pb 16%, S 7%), and copper fume (Cu 6%, Pb 8%, Zn 4%, S 8%) were mixed in a weight ratio of 45%, 45%, and 10% to obtain a mixture with the composition of Pb 37.3%, Zn 2.2%, Cu 0.6%, and S 7.6%.

[0043] The mixture is granulated by adding fluxes such as limestone and quartz, and then 13% (based on the mass of the high-sulfur mixed materials) of granulated coal is added to the side-blown oxidation furnace (hearth area A=12 m²). 2 The furnace is smelted by introducing oxygen-enriched air (70 vol%) at a controlled temperature of 1200℃ for 2.2 hours. The resulting dust-laden flue gas is collected using waste heat and then sent to acid production. The resulting molten metal oxides are sent to a reduction side-blown furnace (hearth area 9 m²). 2 15% granulated coal was added, and oxygen-enriched air with an oxygen concentration of 45 vol% was introduced for reduction smelting, controlled at 1250℃ for 2.2 h. The dust-laden flue gas produced was collected by waste heat recovery and then sent to the tail end of the furnace for reduction to produce crude lead and lead matte. The resulting reduction slag was fed into the electric heating forecourt (furnace area 6.5 m²). 2 The furnace is kept at 1250℃ for 1.7 hours, and slag is continuously discharged into the fuming furnace (furnace bed area 6.5 m²). 2 The furnace is continuously smelted by injecting 18% pulverized coal into it, maintaining a temperature of 1250℃ for 1.7 hours. The resulting dust-laden flue gas is collected using waste heat recovery and then sent to the tail-end suction duct. The collected dust is returned to the feedstock, and the produced slag is water-quenched and sold. The final recovery rates are: Pb 97.3%, Zn 80.2%, and Cu 82.2%.

[0044] Example 2: Lead paste (Pb 65%, S 8%), lead-silver slag (Zn 5%, Pb 30%, S 8%), and copper fumes (Cu 6%, Pb 8%, Zn 4%, S 8%) were mixed at 40%, 40%, and 20% respectively to obtain Pb 39.6%, Zn 2.8%, Cu 1.2%, and S 8%.

[0045] The mixture of raw materials and flux is granulated, and 13% coal granules are added to an oxidation side-blown furnace. Oxidation smelting is carried out using oxygen-enriched air (72 vol%), with the temperature controlled at 1220°C for 2.3 hours. The resulting dust-laden flue gas is collected using waste heat and then sent to acid production. The resulting molten metal oxide is sent to a reduction side-blown furnace, where 15% coal granules are added, and reduction smelting is carried out using oxygen-enriched air (46 vol%), with the temperature controlled at 1260°C for 2.3 hours. The resulting dust-laden flue gas... After waste heat recovery and dust collection, the slag is sent to the tail end of the furnace for reduction to produce crude lead and lead matte. The resulting reduction slag is fed into an electric heating forecourt and held at 1260°C for 1.8 hours. The slag is then continuously discharged into a fuming furnace for blowing. 18% pulverized coal is injected into the furnace for continuous fuming blowing, with the temperature controlled at 1260°C for 1.8 hours. The resulting dust-laden flue gas is then sent to the tail end of the furnace after waste heat recovery and dust collection. The collected dust is returned to the batching process. The resulting slag is water-quenched and sold. The final recovery rates are 97.4% for Pb, 80.3% for Zn, and 82.8% for Cu.

[0046] Example 3: Lead-silver slag (Zn 3%, Pb 35%, S 9%), zinc leaching slag (Zn 4%, Pb 18%, S 8%), and copper fume (Cu 6%, Pb 8%, Zn 4%, S 8%) were mixed at 80%, 10%, and 10% respectively to obtain Pb 30.6%, Zn 3.2%, Cu 0.6%, and S 8.8%. The mixture of raw materials and flux is granulated, and 12% coal granules are added to an oxidation side-blown furnace. Oxidation smelting is carried out using oxygen-enriched air (74 vol%), with the temperature controlled at 1210°C for 2.2 hours. The resulting dust-laden flue gas is collected using waste heat and then sent to acid production. The resulting molten metal oxide is sent to a reduction side-blown furnace, where 14% coal granules are added, and reduction smelting is carried out using oxygen-enriched air (48 vol%), with the temperature controlled at 1250°C for 2.2 hours. The resulting dust-laden flue gas... After waste heat recovery and dust collection, the slag is sent to the tail end of the furnace for reduction to produce crude lead and lead matte. The resulting reduction slag is fed into an electric heating forecourt and held at 1250°C for 1.7 hours. The slag is then continuously discharged into a fuming furnace for blowing. 18% pulverized coal is injected into the furnace for continuous fuming blowing, with the temperature controlled at 1250°C for 1.7 hours. The resulting dust-laden flue gas is then sent to the tail end of the furnace after waste heat recovery and dust collection. The collected dust is returned to the batching process. The resulting slag is water-quenched and sold. The final recovery rates are 97.0% for Pb, 80.2% for Zn, and 82.1% for Cu.

[0047] Example 4: Oxygen leaching residue (Zn 2%, Pb 34%, S 36%) and copper dust (Cu 6%, Pb 8%, Zn 4%, S 8%) were mixed at 90% and 10% respectively to obtain Pb 31.4%, Zn 2.2%, Cu 0.6%, and S 33.2%. The mixture of raw materials and flux is granulated, and 8% coal granules are added to an oxidation side-blown furnace. Oxidation smelting is carried out using oxygen-enriched air (78 vol%), with the temperature controlled at 1240°C for 2.4 hours. The resulting dust-laden flue gas is collected using waste heat and then sent to acid production. The resulting molten metal oxide is sent to a reduction side-blown furnace, where 15% coal granules are added, and reduction smelting is carried out using oxygen-enriched air (48 vol%), with the temperature controlled at 1280°C for 2.4 hours. The resulting dust-laden flue gas... After waste heat recovery and dust collection, the slag is sent to the tail end of the furnace for reduction to produce crude lead and lead matte. The resulting reduction slag is fed into an electric heating forecourt and held at 1280°C for 1.9 hours. The slag is then continuously discharged into a fuming furnace for blowing. 18% pulverized coal is injected into the furnace for continuous fuming blowing, with the temperature controlled at 1280°C for 1.9 hours. The resulting dust-laden flue gas is then sent to the tail end of the furnace after waste heat recovery and dust collection. The dust collected is returned to the batching process. The resulting slag is water-quenched and sold. The final recovery rates are 97.5% for Pb, 80.1% for Zn, and 82.1% for Cu.

[0048] Example 5: Lead paste (Pb 65%, S 8%) and oxygen leaching residue (Zn 3%, Pb 10%, S 30%) were mixed at 50% and 50% respectively to obtain Pb 37.5%, Zn 1.5%, and S 19%. The mixture of raw materials and flux is granulated, and 9% coal granules are added to an oxidation side-blown furnace. Oxidation smelting is carried out using oxygen-enriched air (75 vol%), with the temperature controlled at 1230°C for 2.1 hours. The resulting dust-laden flue gas is collected using waste heat and then sent to acid production. The resulting molten metal oxide is sent to a reduction side-blown furnace, where 16% coal granules are added, and reduction smelting is carried out using oxygen-enriched air (50 vol%), with the temperature controlled at 1270°C for 2.1 hours. The dust-laden flue gas is collected by waste heat recovery and then sent to the tail end of the furnace for reduction to produce crude lead. The resulting reduction slag is fed into an electric heating forecourt and held at 1270°C for 1.6 hours. The slag is then continuously discharged into a fuming furnace for blowing. 19% pulverized coal is injected into the furnace for continuous fuming blowing, with the temperature controlled at 1270°C for 1.6 hours. The resulting dust-laden flue gas is collected by waste heat recovery and then sent to the tail end of the furnace. The collected dust is returned to the batching process. The resulting slag is water-quenched and then sold. The final recovery rates are 97.3% for Pb and 80.1% for Zn.

[0049] Example 6: Lead-silver slag (Zn 3%, Pb 35%, S 9%), zinc leaching slag (Zn 4%, Pb 18%, S 8%), and oxygen leaching slag (Zn 2%, Pb 34%, S 36%) were mixed at 40%, 20%, and 40% respectively to obtain Pb 31.2%, Zn 2.8%, and S 19.6%. The mixture of raw materials and flux is granulated, and 9% coal granules are added to an oxidation side-blown furnace. Oxidation smelting is carried out using oxygen-enriched air (77 vol%), with the temperature controlled at 1230°C for 2.2 hours. The resulting dust-laden flue gas is collected using waste heat and then sent to acid production. The resulting molten metal oxide is sent to a reduction side-blown furnace, where 15% coal granules are added, and reduction smelting is carried out using oxygen-enriched air (47 vol%), with the temperature controlled at 1260°C for 2.2 hours. Dust-laden flue gas is collected by waste heat recovery and then sent to the tail end of the furnace for reduction to produce crude lead. The resulting reduction slag is fed into an electric heating forecourt and held at 1260°C for 1.7 hours. The slag is then continuously discharged into a fuming furnace for blowing. 19% pulverized coal is injected into the furnace for continuous fuming blowing, with the temperature controlled at 1260°C for 1.7 hours. The resulting dust-laden flue gas is collected by waste heat recovery and then sent to the tail end of the furnace. The collected dust is returned to the batching process. The resulting slag is water-quenched and then sold. The final recovery rates are 97.1% for Pb and 80.2% for Zn.

[0050] Table 1 below shows the ingredient list for Examples 1-6:

[0051] Note: In Examples 1-6, the materials contain Pb ≥ 30% and S ≥ 7%, requiring initial desulfurization and acid production in an oxidation side-blown furnace, followed by recovery of crude lead or lead matte in a reduction side-blown furnace. The reduction slag is then fed into a fuming furnace to recover lead and zinc. If the material contains Pb < 25%, crude lead or lead matte cannot be produced, only smoke and dust.

[0052] Examples 7 and 8 below employ Mode 2 operation, which involves a two-stage operation of an oxidation side-blown furnace and a reduction side-blown furnace connected in series. Steps S3 and S4 are executed sequentially, and the reduction smelting slag is directly water-quenched and sold, omitting the fuming process. This mode is suitable for high-sulfur copper dust or pure lead paste materials, eliminating the need for fuming recovery.

[0053] Example 7: Copper dust (Cu 16%, Pb 10%, Zn 5%, S 10%) was granulated with flux and then added to an oxidation side-blown furnace with 14% granulated coal. Oxidation smelting was carried out using oxygen-enriched air (70 vol%), controlled at 1250℃ for 2.5 hours. The resulting dust-laden flue gas was collected using waste heat recovery and then sent to acid production. The resulting copper oxide melt was sent to a reduction smelting side-blown furnace, where 16% granulated coal was added and oxygen-enriched air (45 vol%) was introduced for reduction smelting, controlled at 1300℃ for 2.5 hours. The resulting dust-laden flue gas was collected using waste heat recovery and then sent to a tail gas extraction system for reduction to produce crude copper. The resulting reduction slag was water-quenched and sold. The final recovery rates were Cu 94.4%, Pb 88.7%, and Zn 81.9%.

[0054] Note: Copper flue dust with sulfur content ≥ 7% is desulfurized in an advanced oxidation side-blown furnace to obtain copper oxide melt, which is then reduced in a reduction side-blown furnace to produce crude copper. Lead and zinc are recovered in the reduction flue gas. Copper flue dust without iron oxides can be directly reduced to crude copper after desulfurization.

[0055] Example 8: Lead paste (Pb 65%, S 8%) was granulated with flux and mixed with 13% granulated coal. The mixture was then added to an oxidation side-blown furnace, and oxygen-enriched air with an oxygen concentration of 70 vol% was introduced for oxidation smelting. The temperature was controlled at 1200°C and the time was 2.2 hours. The dust-containing flue gas produced was collected by waste heat recovery and then sent to acid production. The resulting lead oxide melt was sent to a reduction side-blown furnace, where 15% granulated coal was added and oxygen-enriched air with an oxygen concentration of 45 vol% was introduced for reduction smelting. The temperature was controlled at 1250°C and the time was 2.2 hours. The dust-containing flue gas produced was collected by waste heat recovery and then sent to the tail gas extraction furnace. The reduction produced crude lead with a lead recovery rate of 97.9% was obtained. The resulting reduction slag was water-quenched and then sold.

[0056] Note: Lead paste has Pb≥30% and S≥7%. It is desulfurized in an advanced oxidation side-blown furnace to obtain lead oxide melt, which is then reduced in a reduction side-blown furnace to produce crude lead.

[0057] Examples 9 and 10 below employ Mode 3 operation, which involves a two-stage series operation of the reduction side-blown furnace, the electric heating forecourt, and the fuming furnace, executing steps S4, S5, and S6 sequentially, omitting the oxidation smelting process in step S3. This mode is suitable for low-sulfur materials with a sulfur content of <7%, eliminating the need for desulfurization via the oxidation side-blown furnace.

[0058] Example 9: Lead dust (Pb 32%, Zn 8%, S 3%) and lead refining slag (Pb 50%, Cu 6%, S 3%) were mixed in a 50% weight ratio to obtain a mixture (Pb 41%, Zn 4%, Cu 3%, S 3%). 3%), the mixture is granulated with flux, and 15% granulated coal is added to the reducing side-blown furnace. Oxygen-enriched air with an oxygen concentration of 45 vol% is introduced for reduction smelting, controlled at 1250°C for 2.2 hours. The resulting dust-laden flue gas is collected by waste heat recovery and then sent to the tail end of the furnace. The reduction produces crude lead and lead matte. The resulting reduction slag is fed into an electric heating forecourt and held at 1250°C for 1.7 hours. The slag is continuously discharged into a fuming furnace for smelting. 18% pulverized coal is injected into the furnace for continuous fuming smelting, controlled at 1250°C for 1.7 hours. The resulting dust-laden flue gas is collected by waste heat recovery and then sent to the tail end of the furnace. The collected dust is returned to the batching process. The resulting slag is water-quenched and sold. The final recovery rates are Pb 97.4%, Zn 81.1%, and Cu 83.4%.

[0059] Note: If the mixture contains Pb≥30% and S<7%, it does not need to be desulfurized in the oxidation side-blown furnace. It can be directly fed into the reduction side-blown furnace to produce crude lead and lead matte, and then fed into the fuming furnace to volatilize and recover lead and zinc.

[0060] Example 10: Steel plant flue dust (12% Zn, 3% S) was granulated with flux and mixed with 15% granulated coal. The mixture was then added to a reducing side-blown furnace, and oxygen-enriched air with an oxygen concentration of 45 vol% was introduced for slag smelting. The temperature was controlled at 1250°C for 1 hour. The dust-laden flue gas was collected by waste heat recovery and then sent to the tail end of the furnace. The produced melt was fed into an electric heating forecourt and held at 1250°C for 1.7 hours. The slag was continuously discharged into a fuming furnace for blowing. 18% pulverized coal was injected into the furnace for continuous fuming blowing. The temperature was controlled at 1250°C for 1.7 hours. The dust-laden flue gas was collected by waste heat recovery and then sent to the tail end of the furnace. The collected flue dust was sold externally, and the produced slag was water-quenched and then sold externally. The final Zn recovery rate was 92.4%.

[0061] Note: Steel plant flue dust S < 7%, no need to go through oxidation side-blown furnace for desulfurization, advanced reduction side-blown furnace for slag formation, and then into fuming furnace for zinc volatilization and recovery.

[0062] Examples 11 and 12 below employ Mode 4 operation, which involves a two-stage series operation of an oxidation side-blown furnace, an electric heating forecourt, and a fuming furnace, executing steps S3, S5, and S6 sequentially, omitting the reduction smelting process in step S4. This mode is suitable for high-sulfur impurities with a sulfur content ≥7% and a lead content <30%. The first step is desulfurization in the oxidation side-blown furnace to obtain lead-zinc oxide melt. Due to the low lead content, crude lead or lead matte cannot be produced; the melt is directly fed into the fuming furnace for reduction and recovery of lead and zinc.

[0063] Example 11: Oxygen leaching slag (Zn 3%, Pb 10%, S 30%) was granulated with flux and mixed with 6% granulated coal. The slag was then added to an oxidation side-blown furnace, and oxygen-enriched air (70 vol%) was introduced for oxidation smelting. The temperature was controlled at 1200°C for 2.2 hours. The resulting dust-laden flue gas was collected using waste heat and then sent to acid production. The resulting lead-zinc oxide melt was fed into an electric heating forecourt and held at 1250°C for 1.7 hours. The slag was continuously discharged into a fuming furnace for blowing. 18% pulverized coal was injected into the furnace for continuous fuming blowing, with the temperature controlled at 1250°C for 1.7 hours. The resulting dust-laden flue gas was collected using waste heat and then sent to the tail end of the furnace. The collected dust was sold externally, and the resulting slag was water-quenched and then sold externally. The final recovery rates were 89.3% for Pb and 80.1% for Zn.

[0064] Note: The oxygen leaching residue contains S≥7%. The lead-zinc oxide melt is obtained by desulfurization in an advanced oxidation side-blown furnace. Since Pb<25%, it is directly fed into the fuming furnace for reduction and recovery of lead and zinc.

[0065] Example 12: Zinc leaching residue (Zn 15%, Pb 4%, S 7%) was granulated with flux and mixed with 13% granulated coal. The mixture was then added to an oxidation side-blown furnace, and oxygen-enriched air (70 vol%) was introduced for oxidation smelting. The temperature was controlled at 1200°C for 2.2 hours. The resulting dust-laden flue gas was collected using waste heat and then sent to acid production. The resulting lead-zinc oxide melt was fed into an electric heating forecourt and held at 1250°C for 1.7 hours. The slag was continuously discharged into a fuming furnace for smelting. 18% pulverized coal was injected into the furnace for continuous fuming smelting, with the temperature controlled at 1250°C for 1.7 hours. The resulting dust-laden flue gas was collected using waste heat and then sent to the tail end of the furnace. The collected dust was sold externally, and the resulting slag was water-quenched and then sold externally. The final recovery rates were 87.1% for Pb and 93.1% for Zn.

[0066] Note: The zinc leaching residue contains S≥7%. The lead-zinc oxide melt is obtained by desulfurization in an advanced oxidation side-blown furnace. Since Pb<25%, it is directly fed into the fuming furnace for reduction and recovery of lead and zinc.

[0067] Examples 13 and 14 below employ Mode 5 operation, where the oxidation side-blown furnace and the reduction side-blown furnace are fed independently in parallel. The melt or slag produced by both is fed into the electrothermal forecourt, and after heat preservation and sedimentation, it is uniformly fed into the fuming furnace for parallel and coordinated operation. This mode is used to simultaneously process high-sulfur and low-sulfur impurities, realizing the combined parallel operation of the oxidation side-blown furnace-fuming furnace (including the electrothermal forecourt) and the reduction side-blown furnace-fuming furnace (including the electrothermal forecourt).

[0068] Example 13: In Examples 9, 10, 11, and 12 above, the specific combination is as follows: Combined Example 11 + Example 9: High-sulfur oxygen leaching slag (Zn 3%, Pb 10%, S 30%) and a mixture of low-sulfur lead dust and lead refining slag (Pb 41%, Zn 4%, Cu 3%, S 3%) are fed into an oxidation side-blown furnace and a reduction side-blown furnace for smelting, respectively. The oxygen leaching slag undergoes oxidation smelting and desulfurization in the oxidation side-blown furnace to produce lead-zinc oxide melt, which is then fed into the electrothermal forecourt via a chute. The mixture of lead dust and lead refining slag undergoes reduction smelting in the reduction side-blown furnace to produce crude lead, lead matte, and reduction slag, which is then fed into the same electrothermal forecourt via a chute. Both melt / slag streams are held at 1250~1300℃ in the electrothermal forecourt for precipitation for 1.5~2.0 h, clarified, and then continuously discharged into a fuming furnace via a siphon slag discharge port for fuming blowing to recover lead and zinc.

[0069] Combined Example 11 + Example 10: High-sulfur oxygen-leaching slag (Zn 3%, Pb 10%, S 30%) and low-sulfur steel mill dust (Zn 12%, S 3%) are fed into an oxidation side-blown furnace and a reduction side-blown furnace for smelting, respectively. The oxygen-leaching slag, after oxidation smelting, produces lead-zinc oxide melt which enters the electrothermal forecroom; the steel mill dust, after slag formation in the reduction side-blown furnace, produces slag melt which enters the same electrothermal forecroom. After being held and clarified in the electrothermal forecroom, both melts are fed into a fuming furnace for continuous blowing to recover lead and zinc.

[0070] Combined Example 12 + Example 9: High-sulfur zinc leaching residue (Zn 15%, Pb 4%, S 7%) and low-sulfur lead dust and lead refining slag mixture (Pb 41%, Zn 4%, Cu 3%, S 3%) are fed into an oxidation side-blown furnace and a reduction side-blown furnace for smelting, respectively. The resulting lead-zinc oxide melt and reduction slag are fed into an electric heating forecourt via chutes, and after heat preservation and clarification, they are fed into a fuming furnace for continuous blowing to recover lead and zinc.

[0071] Combined Example 12 + Example 10: High-sulfur zinc leaching slag (Zn 15%, Pb 4%, S 7%) and low-sulfur steel mill dust (Zn 12%, S 3%) are fed into an oxidation side-blown furnace and a reduction side-blown furnace for smelting, respectively. The resulting lead-zinc oxide melt and slag melt are fed into an electric heating forecourt via chutes, and after heat preservation and clarification, they are fed into a fuming furnace for continuous blowing to recover lead and zinc.

[0072] Note: For slag with Pb < 25% and S ≥ 7%, the slag is desulfurized in an advanced oxidation side-blown furnace and then directly fed into a fuming furnace (including an electric heating forehearth) for lead and zinc volatilization and recovery. For slag with Pb ≥ 30% and S < 7%, the slag is reduced in an advanced reduction side-blown furnace to produce crude lead or lead matte, and then fed into a fuming furnace (including an electric heating forehearth) for lead and zinc volatilization and recovery. For steel plant flue dust with S < 7%, slag is only processed in a reduction side-blown furnace and then fed into a fuming furnace (including an electric heating forehearth) for zinc volatilization and recovery. In this case, the processes from the oxidation side-blown furnace to the fuming furnace (including an electric heating forehearth) and the reduction side-blown furnace to the fuming furnace (including an electric heating forehearth) can be combined into one.

[0073] Example 14: In Examples 1-6, 9, and 11-12 above, steel plant dust can be directly added to the fuming furnace for blowing. The calorific value of the high-temperature smelting slag produced by the oxidation furnace or reduction furnace is used to melt the steel plant dust, so that it enters the fuming blowing stage together with the smelting slag to recover valuable metals such as zinc.

[0074] The foregoing has only described certain exemplary embodiments of the present invention by way of illustration. Undoubtedly, those skilled in the art can modify the described embodiments in various ways without departing from the spirit and scope of the present invention. Therefore, the foregoing drawings and descriptions are illustrative in nature and should not be construed as limiting the scope of protection of the claims of the present invention.

Claims

1. A smelting system for copper, lead, and zinc scraps, characterized in that, include: The furnace includes an oxidation side-blown furnace, a reduction side-blown furnace, and a fuming furnace, arranged in a triangular pattern. The fuming furnace is equipped with an electric heating front bed. The hearth area of ​​the side-blown oxidation furnace is Am. 2 The value of A ranges from 6 to 30, and the hearth area of ​​the reduction side-blown furnace is 0.7-0.8A m². 2 The furnace bed area of ​​the fuming furnace is 0.5-0.6 Å m². 2 ; The electric heating front bed is located at the feed front end of the fuming furnace, and the furnace bed area of ​​the electric heating front bed is 0.5-0.6 Å m². 2 ; The oxidation side-blown furnace, the reduction side-blown furnace, and the electric heating front bed are connected by chutes, and the material conveying channels between the three furnace bodies are interconnected. Each of the chutes is connected to the feed inlet and discharge outlet of the corresponding furnace. The oxidation side-blown furnace, reduction side-blown furnace, and fuming furnace are each equipped with independent flue gas collection pipelines; the flue gas collection pipeline of the oxidation side-blown furnace is connected to the acid production system after waste heat recovery and dust collection, and the flue gas collection pipelines of the reduction side-blown furnace and the fuming furnace are connected to the tail gas treatment system after waste heat recovery and dust collection. The electric heating forebed is connected to the fuming furnace via a siphon slag discharge port. The siphon slag discharge port is located at the lower part of the side wall of the electric heating forebed. The effective volume of the electric heating forebed is matched with the slag volume of one batch processed by the fuming furnace. The inlet height of the siphon slag discharge port is determined according to the molten slag flow rate and the processing cycle of the fuming furnace. The inlet height of the siphon slag discharge port is set such that when the newly introduced molten slag in the electric heating forebed causes the liquid level to reach the inlet height of the siphon slag discharge port, the amount of slag remaining in the forebed is exactly the amount of slag processed by one batch of the fuming furnace, and the liquid level difference generated by the newly fed slag is sufficient to drive the molten slag to flow into the fuming furnace continuously at a stable flow rate.

2. The smelting system for copper, lead, and zinc scraps according to claim 1, characterized in that: The oxidation side-blown furnace, reduction side-blown furnace, and fuming furnace all adopt a single-furnace-one-section independent operation mode, and the material conveying channels between the three furnaces are interconnected. The smelting system can be switched to any one of the following five operation modes: Mode 1: The oxidation side-blown furnace, reduction side-blown furnace, electric heating forecourt, and fuming furnace are connected in series. Mode 2: Two-stage operation: oxidation side-blown furnace and reduction side-blown furnace connected in series; Mode 3: The side-blown furnace, the electric heating forecourt, and the fuming furnace are connected in series in two stages; Mode 4: Two-stage operation in series: oxidation side-blown furnace, electric heating forecourt, and fuming furnace; Mode 5: The oxidation side-blown furnace and the reduction side-blown furnace are fed in parallel and independently. The melt or slag produced by the two furnaces is fed into the electrothermal forecourt, and after heat preservation and sedimentation, it is fed into the fuming furnace for parallel and coordinated operation.

3. The smelting system for copper, lead, and zinc scraps according to claim 1, characterized in that: The feed ends of the oxidation side-blown furnace and the reduction side-blown furnace are connected to a cylindrical granulator, and the feed end of the cylindrical granulator is connected to a batching silo. The dust collected by the dust collection devices of the oxidation side-blown furnace, the reduction side-blown furnace, the fuming furnace and the electric heating front bed is transported back to the batching silo for recycling or sold directly.

4. A method for smelting copper-lead-zinc scrap using the smelting system of any one of claims 1 to 3, characterized in that, Includes the following steps: S1 Batching: Mix copper, lead, and zinc scraps with flux; the flux is at least one of limestone, quartz, iron filings, or iron sulfides; the copper, lead, and zinc scraps are divided into two categories: high-sulfur scraps and low-sulfur scraps. The high-sulfur scraps include copper dust, zinc leaching slag, lead-silver slag, oxygen leaching slag, and lead paste, with a sulfur content of 7% to 50%. The low-sulfur scraps include steel mill dust, lead dust, and lead refining slag, with a sulfur content of 2% to 4%. S2 Granulation: The prepared mixture is fed into a cylindrical granulator for mixing and granulation. The particle size of the granulated material is controlled to be 8~10mm and the moisture content is 8%~10%. S3 Oxidation Smelting: High-sulfur scrap is fed into an oxidation side-blown furnace, and granular coal with a particle size of 5-10mm and accounting for 6%-14% of the mass of the high-sulfur scrap is added. Oxygen-enriched air with an oxygen volume concentration of 70%-80% is introduced, and the smelting temperature is controlled at 1200-1250℃ and the smelting time is 2-2.5h. The high-sulfur dust-containing flue gas produced by oxidation is sent to the acid production system after waste heat recovery and dust collection, while metal oxide melt is produced at the same time. S4 Reduction Smelting: Desulfurized metal oxide melt or low-sulfur impurities are fed into a reduction side-blown furnace, and granular coal with a particle size of 5-10mm and accounting for 14%-16% of the corresponding material mass is added. Oxygen-enriched air with an oxygen volume concentration of 45%-55% is introduced, and the temperature is controlled at 1250-1300℃ for 2-2.5 hours. The low-sulfur dust-containing flue gas produced by reduction is sent to the tail gas absorption treatment system after waste heat recovery and dust collection. At the same time, crude metal or lead matte and reduction smelting slag are produced. S5 Electrothermal Forebed Precipitation: Oxidized or reduced smelting slag is fed into the electrothermal forebed and precipitated at 1250~1300℃ for 1.5~2.0h to separate valuable metals; the electrothermal forebed retains the amount of smelting slag to be processed in one furnace. When the next furnace smelting slag is added, the existing slag in the forebed is squeezed by the new feed slag and continuously discharged into the fuming furnace through the siphon slag discharge port; S6 continuous blowing of fuming material: 18% to 20% of pulverized coal by mass of slag is injected into the fuming furnace, the temperature is controlled at 1250 to 1300℃ and the blowing time is 1.5 to 2.0 hours; under a strong reducing atmosphere, lead and zinc metal oxides are reduced and volatilized into the flue gas, and the flue gas is sent to the tail gas treatment system after waste heat recovery and dust collection. The collected flue gas is returned to the batching process or sold externally, and the slag is finally sold externally after water quenching. The five operating modes can be freely switched according to the sulfur and lead content of the impurities to be processed.

5. The method for smelting copper, lead, and zinc scraps according to claim 4, characterized in that: The switching logic for the five operating modes is as follows: When the sulfur content of the scrap is ≥7% and the lead content is ≥30%, mode one is adopted, and steps S3, S4, S5 and S6 are executed in sequence, that is, the oxidation side-blown furnace, the reduction side-blown furnace, the electric heating forebed and the fuming furnace are connected in series. When the sulfur content of the scrap is ≥7% and no fuming recovery is required, mode 2 is adopted, and steps S3 and S4 are executed in sequence. The reduced smelting slag is directly water-quenched and sold externally, that is, the oxidation side-blown furnace and the reduction side-blown furnace are connected in series. When the sulfur content of the impurities is <7% and the lead content is ≥30%, mode 3 is adopted, and steps S4, S5 and S6 are executed in sequence, that is, the two-stage operation of the reduction side-blown furnace, the electric heating forebed and the fuming furnace is connected in series. When the sulfur content of the scrap is ≥7% and the lead content is <30%, mode four is adopted, and steps S3, S5 and S6 are executed in sequence, namely, the oxidation side-blown furnace, the electric heating forebed and the fuming furnace are connected in series. When high-sulfur and low-sulfur materials need to be processed simultaneously, mode five is adopted. High-sulfur materials are processed in step S3 and low-sulfur materials are processed in step S4. The melt or slag produced by the two streams are fed into the electric heating front bed through their respective chutes and then enter the fuming furnace together. That is, the oxidation side-blown furnace and the reduction side-blown furnace are fed independently and operate in parallel.

6. The method for smelting copper, lead, and zinc scraps according to claim 5, characterized in that: When operating in Mode 1, the recovery rates of lead, zinc, and copper after processing are >87%, >80%, and >82%, respectively.

7. The method for smelting copper, lead, and zinc scraps according to claim 5, characterized in that: When operating in Mode 2, it is suitable for high-sulfur copper dust or lead paste materials, and the copper recovery rate can reach 94.4% when treating high-sulfur copper dust.

8. The method for smelting copper, lead, and zinc scraps according to claim 5, characterized in that: When operating in Mode 3, it is suitable for low-sulfur lead-zinc mixtures or steel mill dust. When treating steel mill dust alone, the zinc recovery rate can reach 92.4%.

9. The method for smelting copper, lead, and zinc scraps according to claim 5, characterized in that: When using the aforementioned Mode 4 operation, the zinc recovery rate can reach 93.1% when processing zinc leaching residue.

10. The method for smelting copper, lead, and zinc scraps according to claim 5, characterized in that: When operating under Mode 5, the steel plant's flue gas is directly fed into the fuming furnace, where it is melted and reduced by utilizing the residual heat of the slag.

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