Method for recovering metal elements in smoke dust of rotary hearth furnace

Abstract

The invention provides a method for recovering metal elements in smoke dust of a rotary hearth furnace, and relates to the technical field of comprehensive recovery of metals. The recovery method comprises the following steps: carrying out water leaching treatment and purification treatment on smoke dust of the rotary hearth furnace to obtain a potassium-sodium solution and a zinc-rich precipitate; the water leaching residues are subjected to first-time acid leaching treatment and second-time acid leaching treatment, and acid leaching liquid and acid leaching residues are obtained; ferric iron in the acid leaching solution is reduced into ferrous iron; and adjusting the pH value to precipitate and separate out bismuth, indium and germanium. According to the recovery method, through selective leaching, two-stage deep extraction and ingenious internal material circulation, efficient recovery of main metal such as potassium, sodium, zinc and lead is achieved, high-value scattered rare metal such as indium, germanium and bismuth is more successfully and comprehensively recovered, and comprehensive utilization of smoke dust resources of the rotary hearth furnace is achieved to the maximum extent.

Description

TECHNICAL FIELD

[0001] The present application relates to the technical field of resource recycling, in particular to a method for recovering metal elements from rotary hearth furnace dust. BACKGROUND

[0002] In the steel metallurgical process, a large amount of iron-containing dust slurry is produced. These dust slurry not only contains iron, calcium, magnesium and other components that can be recycled and utilized in the production process, but also enriches potassium, sodium, sulfur, zinc, lead and other elements that are harmful to steel production. Rotary hearth furnace process is one of the effective methods for treating such iron-containing dust slurry. It converts the iron oxides in the dust slurry into metallurgical pellets through high-temperature reduction reaction and returns them to steel production, while reducing zinc and other volatile metals into vapor form. These metal vapors are discharged with high-temperature flue gas and condensed into fine solid particles after cooling, which are captured by the dust removal system to form rotary hearth furnace dust. The composition of this dust is extremely complex. In addition to zinc oxide as the main component, which can contain more than 40% zinc, a large amount of potassium chloride, sodium chloride, and fluorides that are harmful to subsequent processing are also enriched due to differences in raw materials and processes. In addition, lead, bismuth, silver and scattered metals indium and germanium, as well as copper, cadmium, nickel, cobalt and other valuable metal elements associated with iron ore, will also migrate and enrich in the dust during the rotary hearth furnace process, making it a valuable secondary resource library.

[0003] Currently, the resource treatment of rotary hearth furnace dust and similar zinc-containing dust mainly adopts hydrometallurgical process, which is widely concerned due to its lower energy consumption, higher product purity and easier recovery of associated metals compared to pyrometallurgical process. The hydrometallurgical process mainly includes three technical routes: alkali leaching, ammonia leaching and acid leaching. The alkali leaching method uses sodium hydroxide solution to react with the material to make zinc generate soluble sodium zincate, but this method has problems such as harsh leaching conditions, long leaching time and low zinc leaching rate (about 80%). The ammonia leaching method uses the complexation of ammonia and ammonium chloride to leach zinc, with a zinc leaching rate of about 89%, but it also has the problem of strict process conditions, and the product scheme of the two methods is usually limited to the production of zinc oxide. The acid leaching method, especially the sulfuric acid method, is widely studied due to its high zinc leaching rate (more than 96%) and flexible product scheme (zinc oxide or electrolytic zinc). The typical sulfuric acid method process includes acid leaching, purification and electrodeposition steps. The purification process removes iron and zinc powder by oxidation and hydrolysis, and replaces copper by other operations, aiming to prepare high-purity zinc sulfate solution that meets the requirements of electrodeposition. In addition, some technical solutions also use water leaching to preferentially recover potassium and sodium salts in the dust, and then use extraction methods to recover zinc.

[0004] However, existing technologies generally suffer from a lack of focus on a single recovery target and low resource utilization. Whether using alkaline leaching, ammonia leaching, or acid leaching processes, the main research and application objectives are concentrated on the efficient extraction and purification of the primary metal, zinc. In these processes, other metals with high economic value present in the flue dust, such as rare metals indium and germanium, precious metal silver, and non-ferrous metals lead, bismuth, copper, cadmium, nickel, and cobalt, are typically considered impurities requiring removal. These are precipitated and separated during the purification process and then enter the slag, failing to achieve effective recovery and utilization. Even the few processes involving the recovery of multiple elements only achieve the recovery of a few elements such as potassium, sodium, and zinc. Existing literature reports very few comprehensive processes capable of separating and recovering almost all valuable metals from rotary hearth furnace flue dust in a cascade manner.

[0005] In summary, current technologies for treating rotary hearth furnace dust have significant limitations. While these technologies have achieved some success in recovering major metals (such as zinc, potassium, and sodium), they generally overlook the overall value of dust as a complex material containing multiple metals, failing to achieve comprehensive and integrated recovery of all valuable metals. Existing processes lack systematic separation and enrichment schemes for multiple elements such as indium, germanium, lead, silver, and bismuth, leading to the loss of these valuable resources. Furthermore, mature solutions are lacking for the efficient removal of harmful impurities such as fluorine and chlorine, as well as the integration and optimization of the overall process. Therefore, developing a clean and efficient technology capable of cascade separation and comprehensive recovery of multiple valuable metals from rotary hearth furnace dust is a crucial technical challenge that urgently needs to be addressed to maximize the value and achieve green disposal of this type of solid waste.

[0006] In view of this, the present invention is hereby proposed. Summary of the Invention

[0007] The purpose of this invention is to provide a method for recovering metal elements from rotary hearth furnace flue dust. This method not only achieves efficient recovery of major metals such as potassium, sodium, and zinc, but also successfully incorporates high-value rare and dispersed metals such as indium, germanium, and bismuth into the recovery system, maximizing the comprehensive utilization of resources.

[0008] In order to achieve the above-mentioned objectives of the present invention, the following technical solution is adopted: In a first aspect, the present invention provides a method for recovering metal elements from flue dust in a rotary hearth furnace, comprising: The flue dust from the rotary hearth furnace is subjected to water leaching treatment to obtain water leaching liquid and water leaching residue; and the water leaching liquid is purified to obtain purified potassium-sodium solution and zinc-rich precipitate. The water-leached residue was subjected to a first acidic leaching treatment using a first acidic leaching agent to obtain a zinc-rich neutral leachate and a neutral leaching residue. The neutral leaching residue is subjected to a second acidic leaching treatment using a second acidic leaching agent to obtain an acidic leaching solution and acidic leaching residue containing zinc, indium, germanium, bismuth and iron. The ferric iron in the acidic leachate is reduced to ferrous iron; and the zinc-rich precipitate is used as a neutralizing agent to adjust the pH value so as to precipitate and separate bismuth, indium and germanium.

[0009] In an optional embodiment, the water immersion treatment includes: The flue dust from the rotary hearth furnace is mixed with water, and the water-leached liquid and water-leached residue are obtained by solid-liquid separation. The temperature of the rotary hearth furnace flue dust is 100℃~200℃; and / or, the water is evaporative condensate; and / or, the water temperature is 50℃~55℃; and / or, the liquid-solid ratio of the rotary hearth furnace flue dust to water is (3mL~6mL):1g; and / or, the water immersion treatment time is 10 minutes~60 minutes.

[0010] In an optional implementation, the purification process includes: An impurity removal agent is added to the aqueous leaching solution to obtain the potassium-sodium solution and the zinc-rich precipitate; The impurity removal agent includes sodium carbonate; and / or, the temperature of the purification treatment is controlled at 25℃~60℃; and / or, in the purification treatment, the pH of the purification system after adding the impurity removal agent is 6.5~7.5.

[0011] In an optional embodiment, prior to the first acid leaching treatment, the following steps are also included: The water-leached residue is subjected to at least one fluoride and chlorine removal treatment; The fluorine and chlorine removal treatment includes: mixing the water leaching residue with sodium carbonate for leaching, followed by solid-liquid separation to obtain water leaching residue and fluorine-containing waste liquid after fluorine and chlorine removal treatment; The concentration of sodium carbonate is 2% to 5%; and / or the time for the fluoride and chlorine removal treatment is 30 minutes to 180 minutes; the temperature for the fluoride and chlorine removal treatment is 40°C to 95°C; and / or the liquid-solid ratio of sodium carbonate to the water leaching residue is (3 mL to 6 mL): 1 g.

[0012] In optional embodiments, the pH of the leaching system for the first acidic leaching treatment is 5.2 to 5.4; and / or, the first acidic leaching agent includes sulfuric acid, and the amount added is 1.0 to 1.05 times the theoretical amount of sulfuric acid calculated based on the zinc content in the water-leached residue; and / or, the leaching temperature for the first acidic leaching treatment is 40°C to 80°C; and / or, the liquid-to-solid ratio of the first acidic leaching agent to the water-leached residue is (4 mL to 8 mL): 1 g; and / or, the leaching time for the first acidic leaching treatment is 60 minutes to 180 minutes.

[0013] In an optional embodiment, after the step of performing a first acid leaching treatment on the water-leached residue to obtain a zinc-rich neutral leachate and a neutral leaching residue, the method further includes: The neutral leachate was subjected to a copper-cadmium displacement reaction using elemental zinc to obtain copper-cadmium slag and a first purified liquid. Elemental zinc and antimony trioxide were added to the first purification solution to remove nickel and cobalt, resulting in a second purification solution and nickel-cobalt slag.

[0014] In an optional embodiment, in the copper-cadmium replacement reaction, the amount of elemental zinc added is 3 to 6 times the molar amount of copper and cadmium in the neutral leachate; and / or, the replacement temperature of the copper-cadmium replacement reaction is 45°C to 55°C; and / or, the reaction time of the copper-cadmium replacement reaction is 30 to 60 minutes; and / or, the mass concentration of antimony trioxide is [missing value] of the Co ion mass concentration. ; and / or, the method for calculating the mass concentration of elemental zinc in the nickel-cobalt removal process: ;in, This represents the mass concentration of elemental zinc. This represents the mass concentration of cobalt ions; The value of a represents the mass concentration of nickel ions; the value of a ranges from 30 to 100; the value of b ranges from 30 to 50; and / or the processing time for the nickel-cobalt removal treatment is 60 to 120 minutes; and / or the processing temperature for the nickel-cobalt removal treatment is 85°C to 95°C.

[0015] In an optional embodiment, after adding elemental zinc and antimony trioxide to the first purified liquid for nickel-cobalt removal treatment to obtain a second purified liquid and nickel-cobalt slag, the step further includes: The second purification solution is subjected to zinc electrowinning treatment to obtain electrowinning waste liquid and cathode zinc. The ratio of the electrowinning waste liquid to the second purified liquid is (10~20):1.

[0016] In an optional embodiment, both the first acidic leaching agent and the second acidic leaching agent are prepared by means of the electrowinning waste liquid and slag washing water generated from the zinc electrowinning process; The sulfuric acid concentration of the electrowinning waste liquid is 180 g / L to 220 g / L.

[0017] In an optional embodiment, the leaching temperature of the second acid leaching treatment is 70°C to 95°C; and / or, in the second acid leaching treatment, the liquid-to-solid ratio of the second acid leaching agent to the neutral leaching residue is (3 to 6) mL:1 g; and / or, the leaching time of the second acid leaching treatment is 180 minutes to 300 minutes.

[0018] In an optional embodiment, the step of reducing ferric iron in the acidic leachate to ferrous iron; and using the zinc-rich precipitate as a neutralizing agent to adjust the pH value to precipitate and separate bismuth, indium, and germanium, includes: The acidic leachate is passed through a sulfur dioxide mixture to carry out a reduction reaction, thereby reducing the ferric iron in the acidic leachate to ferrous iron, and obtaining a reduced reaction solution after the reduction reaction. The reduction reaction solution is subjected to a first neutralization treatment using the zinc-containing precipitate obtained from the purification process to obtain bismuth-rich material and bismuth-precipitated liquid. The bismuth-precipitated liquid was subjected to a second neutralization treatment using the zinc-containing precipitate to obtain an indium-germanium enrichment and an indium-germanium-precipitated liquid. The indium-germanium precipitate solution is subjected to a catalytic oxidation reaction of iron precipitation using the sulfur dioxide mixed gas, and the zinc-rich precipitate is separated into solid and liquid phases to obtain iron oxide precipitate and iron-removed solution. In the reduction reaction, the amount of sulfur dioxide mixed gas used is 2 to 5 times the theoretical molar amount of ferric iron to be reduced; and / or, the sulfur dioxide mixed gas is an air mixture containing 1.3% to 1.8% SO2; and / or, the reaction temperature of the reduction reaction is 40℃ to 80℃; and / or, the reaction time of the reduction reaction is 60 minutes to 180 minutes; and / or, in the first neutralization treatment, the neutralization temperature is 50℃ to 70℃; and / or, in the first neutralization treatment, the neutralization time is 20 minutes to 60 minutes; and / or, the bismuth content in the bismuth-precipitated liquid is not greater than 0.3%. mg / L; and / or, in the second neutralization treatment, the neutralization temperature is 50℃~70℃; and / or, in the second neutralization treatment, the neutralization time is 20 minutes~60 minutes; and / or, in the second neutralization treatment, the pH is neutralized to 4~4.2; and / or, the indium content in the indium-germanium precipitation solution is not greater than 0.15 mg / L; and / or, the germanium content in the indium-germanium precipitation solution is not greater than 0.15 mg / L; and / or, the iron content in the iron removal solution is not greater than 20 mg / L.

[0019] This application provides a method for recovering metal elements from rotary hearth furnace flue dust. Firstly, this method possesses high selectivity and staged separation capabilities. The initial water leaching treatment gently and effectively separates readily soluble potassium and sodium compounds from other water-insoluble metal components, achieving preliminary grouping in the first step and simplifying the complexity of subsequent processes. Next, a two-stage acid leaching strategy is employed. The first stage preferentially leaches the main metal, zinc, under relatively mild conditions, efficiently obtaining a relatively pure zinc solution that is easy to process later. The second stage, under stronger acidic conditions, specifically dissolves the insoluble valuable metals that were not leached in the first stage, such as indium, germanium, bismuth, and complex zinc-iron compounds. This step-by-step, staged treatment method not only ensures a high recovery rate of the main metal, zinc, but also ensures that valuable rare and dispersed metals such as indium and germanium are effectively extracted into the solution, avoiding their loss in the waste residue, thereby greatly improving the comprehensive recovery level of various valuable elements in the raw material.

[0020] Secondly, this method demonstrates outstanding economic efficiency and ingenious design in terms of material recycling and resource utilization. In the water leaching purification stage, not only are high-purity potassium and sodium products obtained, but the byproduct "zinc-rich precipitate" is not waste; instead, it is creatively used as a key chemical reagent in subsequent processes. Specifically, this zinc-rich precipitate is used as a neutralizing agent to adjust the pH of the acidic leachate, enabling the precipitation and recovery of metals such as bismuth, indium, and germanium. This design constructs an internal closed-loop material cycle, redirecting and utilizing the zinc resources dispersed in the initial steps in the core recovery stage, significantly reducing reliance on purchased chemicals (such as neutralizing agents), thereby lowering production costs and reducing waste generation.

[0021] Furthermore, the method's chemical reaction steps are meticulously designed, ensuring efficient recovery of high-value trace elements. When treating acidic leachates containing multiple metal ions, pre-reducing the reactive ferric iron (Fe3+) to the relatively stable ferrous iron (Fe2+) is a crucial prerequisite for achieving precise subsequent separation. This pretreatment step effectively prevents co-precipitation of iron with the target products (bismuth, indium, and germanium) during subsequent neutralization, creating the necessary conditions for the selective and stepwise precipitation separation of high-value bismuth, indium, and germanium through precise pH control. This ensures the efficient enrichment and recovery of these trace elements.

[0022] In summary, the organic combination of a series of operations in this method makes it possible to recover multiple metals from the complex composition of rotary hearth furnace flue gas. Through selective leaching, two-stage deep extraction, and ingenious internal material recycling, this method not only achieves efficient recovery of major metals such as potassium, sodium, and zinc, but also successfully incorporates high-value rare and dispersed metals such as indium, germanium, and bismuth into the recovery system, maximizing the comprehensive utilization of resources. Attached Figure Description

[0023] To more clearly illustrate the specific embodiments of the present invention or the technical solutions in the prior art, the drawings used in the description of the specific embodiments or the prior art will be briefly introduced below. Obviously, the drawings described below are some embodiments of the present invention. For those skilled in the art, other drawings can be obtained from these drawings without creative effort.

[0024] Figure 1 This is a schematic flowchart of the method for recovering metal elements from rotary hearth furnace flue dust provided in the embodiments of this application; Figure 2 This is a schematic diagram of the overall process for the method of recovering metal elements from flue dust in a rotary hearth furnace used in the embodiments of this application. Detailed Implementation

[0025] The embodiments of the present invention will be described in detail below with reference to examples. However, those skilled in the art will understand that the following examples are for illustrative purposes only and should not be considered as limiting the scope of the invention. Unless otherwise specified in the examples, conventional conditions or conditions recommended by the manufacturer are followed. Reagents or instruments whose manufacturers are not specified are all commercially available conventional products.

[0026] refer to Figure 1 This application provides a method for recovering elemental substances from rotary hearth furnace flue dust, comprising: Step S1: The flue dust from the rotary hearth furnace is subjected to water leaching treatment to obtain water leaching liquid and water leaching residue; and the water leaching liquid is purified to obtain purified potassium-sodium solution and zinc-rich precipitate.

[0027] In the above steps, the rotary hearth furnace flue gas used as raw material is mixed and stirred with water. Utilizing the principle of water dissolution, a preliminary separation is performed. This is a solid-liquid separation operation aimed at selectively dissolving water-soluble components in the flue gas. After processing, two substances are obtained through solid-liquid separation methods such as filtration: (1) Aqueous extract: A clear liquid in which all water-soluble substances in the dust are dissolved (such as potassium chloride, sodium chloride and a small amount of soluble zinc salts).

[0028] (2) Water leaching residue: The solid residue remaining after filtration mainly contains substances insoluble in water from the flue dust (such as zinc oxide, zinc iron spinel, lead silver bismuth compounds, etc.).

[0029] This step effectively groups and separates water-soluble salts (such as potassium and sodium salts) with vastly different chemical properties from insoluble metal oxides and other main components, greatly simplifying the processing difficulty of their respective subsequent steps.

[0030] The purification process involves chemically treating the "aqueous extract" produced in the previous step to remove impurity ions, primarily purifying the potassium and sodium salt solutions. This is typically achieved by adding a chemical reagent that causes the target impurity ions to react and precipitate from the liquid. After this treatment, a solid-liquid separation is performed again to obtain two substances: (1) Purified potassium and sodium solution: A purer potassium and sodium salt solution with impurities removed, which can be used for subsequent crystallization to produce high-quality potassium and sodium products.

[0031] (2) Zinc-rich precipitate: The separated solid is mainly composed of a certain insoluble compound of zinc, which captures the zinc that was originally dissolved in the aqueous solution.

[0032] This step achieves two goals at once. First, it purifies the potassium and sodium solution, laying the foundation for producing high-quality potassium and sodium products. Second, it converts and enriches the zinc, which was originally dispersed in the aqueous solution, into a solid form, preventing the loss of this zinc resource and making it a valuable intermediate material that can be used in subsequent processes, embodying the closed-loop concept of resource utilization.

[0033] Step S2: The water leaching residue is subjected to a first acidic leaching treatment using a first acidic leaching agent to obtain a zinc-rich neutral leaching solution and a neutral leaching residue.

[0034] In the above steps, the "water-leached residue" separated in step S1 is treated to dissolve the main valuable metal, zinc, using acid. This "first acid leaching" is carried out under relatively mild acidic conditions, with the pH value controlled at the reaction endpoint within a weakly acidic to near-neutral range. After treatment, solid-liquid separation is performed to obtain two substances: (1) Neutral leachate: Zinc sulfate solution rich in zinc ions. Because the leaching conditions are mild, the content of other impurity metal ions in the solution is relatively low.

[0035] (2) Neutral leaching residue: The solid residue remaining after the first acid leaching, which contains substances that are more difficult to dissolve by acid, such as zinc iron spinel and other valuable metals such as indium, germanium, lead, silver, and bismuth.

[0036] The advantage of this step lies in its selectivity. By controlling the acidity to a mild level, the most reactive zinc oxide can be dissolved efficiently and preferentially, resulting in a relatively pure zinc solution and reducing the difficulty of subsequent zinc purification. Simultaneously, the more difficult-to-process components are concentrated in the "neutral leaching residue," preparing the material for the next stage of advanced treatment.

[0037] Step S3: The neutral leaching residue is subjected to a second acidic leaching treatment using a second acidic leaching agent to obtain an acidic leaching solution and acidic leaching residue containing zinc, indium, germanium, bismuth and iron.

[0038] This step involves further treatment of the "neutral leaching residue" produced in the previous step, with the aim of dissolving all remaining valuable metals. The "second acid leaching" here is carried out under more strongly acidic conditions. After treatment, solid-liquid separation is performed to obtain two substances: (1) Acidic leachate: A complex strongly acidic solution in which zinc (from zinc-iron spinel, etc.) that was not dissolved in the previous round is dissolved, as well as various metals such as indium, germanium, bismuth, and iron.

[0039] (2) Acid leaching residue: The final solid residue, which is mainly enriched with substances that are not soluble in strong acids, such as compounds of lead and silver.

[0040] This step, by employing stronger leaching conditions, ensures that valuable but difficult-to-dissolve rare metals such as indium, germanium, and bismuth are transferred to the liquid phase to the maximum extent, making their recovery possible. This achieves full utilization of the raw material value and greatly improves the overall resource recovery rate.

[0041] Step S4: Reduce the ferric iron in the acidic leachate to ferrous iron; and use the zinc-rich precipitate as a neutralizing agent to adjust the pH value to precipitate and separate bismuth, indium and germanium.

[0042] The above steps consist of two interconnected parts: purifying the complex "acidic leachate" obtained in the previous step and recovering valuable metals. First, the solution undergoes chemical reduction treatment to remove the trivalent ferric ions (Fe3+) that are harmful to subsequent separation. 3+ ) transforms into the more stable divalent iron ion (Fe) 2+ Then, the "zinc-rich precipitate" obtained in the second step is used as a neutralizing agent and slowly added to the acid solution. By gradually increasing the pH value of the solution, different metal ions are precipitated sequentially at different pH values, thereby achieving separation.

[0043] Through filtration and separation, solid products such as bismuth-rich precipitate and indium-germanium enrichment, as well as the separated liquid, can be obtained sequentially.

[0044] This step ensures the accuracy of subsequent separations. By pre-reducing ferric iron, it effectively prevents it from precipitating along with target products such as bismuth, indium, and germanium during the subsequent neutralization process. This is a crucial prerequisite for achieving selective separation. A closed-loop material circulation is achieved: using the internal byproduct "zinc-rich precipitate" as a neutralizing agent replaces externally purchased lime and soda ash, reducing production costs and allowing the zinc separated in the previous steps to return to the main process flow, thus achieving resource recycling. Through precise pH control, elements such as bismuth, indium, and germanium, which are present in small amounts but are of high value, can be effectively separated as concentrates, achieving the final recovery goal.

[0045] In some embodiments, step S1, the water immersion treatment, includes: The flue gas from the rotary hearth furnace is mixed with water, and the water-leached liquid and the water-leached residue are obtained through solid-liquid separation.

[0046] In some embodiments, the temperature of the flue gas from the rotary hearth furnace is 100°C to 200°C. For example, it can be 100°C, 120°C, 140°C, 160°C, 180°C, 200°C, etc.

[0047] In some implementations, the water is evaporative condensate.

[0048] In some implementations, the water temperature is 50°C to 55°C. For example, it can be 50°C, 51°C, 52°C, 53°C, 54°C, 55°C, etc.

[0049] In some embodiments, the liquid-to-solid ratio of the rotary hearth furnace dust to water is (3 mL to 6 mL): 1 g. For example, it can be 3 mL: 1 g, 4 mL: 1 g, 5 mL: 1 g, 6 mL: 1 g, etc.

[0050] In some embodiments, the water immersion treatment time is 10 to 60 minutes. For example, it can be 10 minutes, 20 minutes, 30 minutes, 40 minutes, 50 minutes, 60 minutes, etc.

[0051] In some embodiments, step S1, the purification process, includes: An impurity removal agent is added to the aqueous leaching solution to obtain the potassium-sodium solution and the zinc-rich precipitate; In some embodiments, the impurity removal agent includes sodium carbonate.

[0052] In some embodiments, the temperature of the purification process is controlled between 25°C and 60°C. For example, it can be 25°C, 28°C, 30°C, 35°C, 40°C, 45°C, 50°C, 55°C, 60°C, etc.

[0053] In some embodiments, the pH of the purification system after adding the impurity removal agent is 6.5 to 7.5. For example, it can be 6.5, 6.6, 6.7, 6.8, 6.9, 7.0, 7.1, 7.2, 7.3, 7.4, 7.5, etc.

[0054] In some embodiments, prior to the first acid leaching treatment, the process further includes: The water-leached residue is subjected to at least one fluoride and chlorine removal treatment.

[0055] The fundamental purpose of this step is to purify the solid material, "water-leached residue," with the core objective being the selective removal of fluorine (F) and chlorine (Cl) elements that are harmful to subsequent processes. Since rotary hearth furnace flue dust may contain fluorides and chlorides, if these substances are not removed beforehand, they may corrode equipment or adversely affect product quality during subsequent acid leaching and electrowinning processes.

[0056] The fluorine and chlorine removal treatment includes: mixing the water leaching residue with sodium carbonate for leaching, followed by solid-liquid separation to obtain water leaching residue and fluorine-containing waste liquid after fluorine and chlorine removal treatment; This step employs a chemical leaching method. Specifically, the process involves mixing the solid "water-leached residue" with a liquid "sodium carbonate solution" and stirring for a period of time. During this process, certain fluorides and chlorides contained in the water-leached residue react chemically with the sodium carbonate to form water-soluble fluoride and chloride salts, thereby transferring them from the solid "water-leached residue" to the liquid phase.

[0057] After mixed leaching and solid-liquid separation operations such as filtration, this step yields two products: (1) Water leaching residue after fluorine and chlorine removal treatment: This is the main product, namely the "purified" solid. Compared with the original product, the content of fluorine and chlorine is significantly reduced, making it a "cleaner" raw material that can be sent to the next step of "first acid leaching treatment".

[0058] (2) Fluorine and chlorine-containing waste liquid: This is a byproduct, namely the liquid separated after leaching. It contains dissolved fluorine and chlorine-containing salts transferred from the water leaching residue and requires separate wastewater treatment.

[0059] Fluorine and chloride ions (especially chloride ions) are highly corrosive in acidic environments. Pre-treatment effectively protects reactors, pipelines, and electrode plates in subsequent acid leaching and electrowinning processes, extending equipment lifespan and ensuring production stability. If fluorine and chloride ions enter the electrowinning solution, they may affect the purity and quality of the final cathode zinc product. Pre-removing these harmful impurities helps produce higher purity cathode zinc products. By implementing a dedicated pretreatment step to centrally treat harmful components, interference with subsequent complex processes is avoided, making the subsequent metal leaching and recovery processes more efficient and controllable.

[0060] In some embodiments, the concentration of sodium carbonate is 2% to 5%; for example, it can be 2%, 3%, 4%, 5%, etc.

[0061] In some embodiments, the fluoride and chlorine removal treatment time is 30 minutes to 180 minutes; for example, it can be 30 minutes, 40 minutes, 50 minutes, 60 minutes, 80 minutes, 100 minutes, 120 minutes, 140 minutes, 160 minutes, 180 minutes, etc.

[0062] In some embodiments, the temperature of the fluorine and chlorine removal treatment is 40°C to 95°C; for example, it can be 40°C, 50°C, 60°C, 80°C, 90°C, 95°C, etc.

[0063] In some embodiments, the liquid-to-solid ratio of sodium carbonate to the water-leached residue is (3 mL to 6 mL): 1 g. For example, it can be 3 mL: 1 g, 4 mL: 1 g, 5 mL: 1 g, 6 mL: 1 g, etc.

[0064] In some embodiments, the pH of the leaching system for the first acidic leaching treatment is 5.2 to 5.4; and / or, In some embodiments, the first acidic leaching agent includes sulfuric acid, and the amount added is 1.0 to 1.05 times the theoretical amount of sulfuric acid calculated based on the zinc content in the water leaching residue; for example, it can be 1.00 times, 1.01 times, 1.02 times, 1.03 times, 1.04 times, 1.05 times, etc.

[0065] In some embodiments, the leaching temperature of the first acid leaching treatment is 40°C to 80°C; for example, it can be 40°C, 50°C, 60°C, 70°C, 80°C, etc.

[0066] In some embodiments, the liquid-to-solid ratio of the first acidic leaching agent to the water-leached residue is (4 mL to 8 mL): 1 g; for example, it can be 4 mL: 1 g, 5 mL: 1 g, 6 mL: 1 g, 7 mL: 1 g, 8 mL: 1 g, etc.

[0067] In some embodiments, the leaching time of the first acid leaching treatment is 60 to 180 minutes. For example, it can be 60 minutes, 80 minutes, 100 minutes, 120 minutes, 140 minutes, 160 minutes, 180 minutes, etc.

[0068] In an optional embodiment, after performing a first acid leaching treatment on the water-leached residue to obtain a zinc-rich neutral leachate and a neutral leaching residue in step S2, the method further includes: Step S5: Use elemental zinc to perform a copper-cadmium replacement reaction on the neutral leachate to obtain copper-cadmium slag and a first purification solution.

[0069] This embodiment provides a two-step deep purification process. Its fundamental purpose is to almost completely remove trace impurity elements—copper (Cu), cadmium (Cd), nickel (Ni), and cobalt (Co)—present in the "neutral leachate" that pose a serious threat to the subsequent electrowinning zinc refining process, thereby obtaining a high-purity zinc sulfate solution that meets the requirements of electrowinning production.

[0070] This step utilizes a metal displacement reaction (also known as a cementing reaction) in chemistry. Elemental zinc (usually zinc powder, but other reactive forms of elemental zinc can also be added) with higher chemical reactivity to a neutral leachate containing copper and cadmium ions. Because zinc is more reactive than copper and cadmium, it preferentially dissolves into the solution, displacing copper and cadmium ions and transforming them into water-insoluble elemental metal particles. After the reaction, filtration yields a solid copper-cadmium slag (mainly composed of displaced copper and cadmium, and a small amount of remaining zinc powder) and a purified solution from which copper and cadmium have been removed.

[0071] The reaction equation for the substitution of copper in the above reaction can be: Zn + Cu 2+ →Zn 2+ +Cu↓; Solid zinc powder dissolves into the solution, becoming zinc ions, while copper ions in the solution gain electrons and precipitate as insoluble metallic copper.

[0072] The reaction equation for cadmium displacement can be: Zn + Cd 2+ →Zn 2+ +Cd↓; Similar to the principle of copper displacement, zinc replaces cadmium ions in the solution with metallic cadmium precipitate.

[0073] Step S6: Add elemental zinc and antimony trioxide to the first purification liquid to remove nickel and cobalt, and obtain a second purification liquid and nickel-cobalt slag.

[0074] This step also utilizes the principle of displacement reaction, but the process is more complex. Because the chemical reactivity of nickel and cobalt is similar to that of zinc, simply using zinc powder for displacement is inefficient. Therefore, when treating the "first purification solution," elemental zinc and antimony trioxide (Sb₂O₃) as an activator need to be added simultaneously. The presence of antimony trioxide significantly improves the reaction rate and completeness of zinc powder displacement of nickel and cobalt.

[0075] After the reaction, the product is filtered to obtain solid "nickel-cobalt slag" (mainly the displaced metallic nickel, cobalt, and antimony) and the final "second purification liquid" (i.e., a high-purity zinc sulfate solution).

[0076] The principle of this step is similar to that of step S5, also utilizing the displacement effect of zinc. However, because nickel (Ni) and cobalt (Co) have chemical reactivity more similar to zinc, the displacement reaction is more difficult to occur. Therefore, antimony trioxide (Sb₂O₃) is added as an activator to promote the reaction. The main displacement reactions include: (1) The reaction formula for the substitution of nickel can be: Zn + Ni 2+ →Zn 2+ +Ni↓; (2) The reaction formula for cobalt substitution can be: Zn + Co 2+ →Zn 2+ +Co↓.

[0077] Antimony trioxide itself does not directly participate in the replacement of nickel and cobalt, but plays a catalytic or activating role.

[0078] This two-step purification process has significant advantages in ensuring the quality of core products, improving resource utilization efficiency, and optimizing process economics. Its primary value lies in the deep removal of impurities such as copper, cadmium, nickel, and cobalt, which are extremely harmful to subsequent electrowinning processes. This effectively avoids problems such as decreased current efficiency, energy waste, and substandard purity of the final cathode zinc product caused by impurity interference, and is a prerequisite for ensuring the quality of core products and production economics.

[0079] Furthermore, this method enriches and concentrates these impurity elements into marketable copper-cadmium slag and nickel-cobalt slag through displacement reactions, achieving resource recovery of harmful impurities and further improving the overall economic efficiency of the process. From a process design perspective, stepwise treatment is also more efficient and economical: the first step removes the easier-to-treat copper and cadmium under relatively mild conditions, while the second step can centrally treat the more chemically stable nickel and cobalt under more optimized and demanding conditions. This targeted design avoids redundant consumption of energy and materials, which is superior to the "one-pot" purification mode.

[0080] In some embodiments, in the copper-cadmium replacement reaction, the amount of elemental zinc added is 3 to 6 times the molar amount of copper and cadmium in the neutral leachate. For example, it can be 3 times, 4 times, 5 times, 6 times, etc.

[0081] In some embodiments, the replacement temperature of the copper-cadmium replacement reaction is 45°C to 55°C. For example, it can be 45°C, 46°C, 47°C, 48°C, 49°C, 50°C, 51°C, 52°C, 53°C, 54°C, 55°C, etc.

[0082] In some embodiments, the reaction time for the copper-cadmium replacement reaction is 30 to 60 minutes. For example, it can be 30 minutes, 40 minutes, 50 minutes, 60 minutes, etc.

[0083] In some embodiments, the antimony trioxide mass concentration is equal to the Co ion mass concentration. For example, it can be 1 / 25, 1 / 30, 1 / 35, 1 / 40, 1 / 45, 1 / 50, etc.

[0084] In some embodiments, the method for calculating the mass concentration of elemental zinc in the nickel-cobalt removal process is as follows: ; in, This represents the mass concentration of elemental zinc. This represents the mass concentration of cobalt ions; This represents the mass concentration of nickel ions; the value of a ranges from 30 to 100 (for example, it can be 30, 40, 50, 60, 70, 80, 90, 100, etc.); the value of b ranges from 30 to 50 (for example, it can be 30, 40, 50, etc.).

[0085] In some embodiments, the processing time for the nickel-cobalt removal treatment is 60 to 120 minutes. For example, it can be 60 minutes, 70 minutes, 80 minutes, 90 minutes, 100 minutes, 110 minutes, 120 minutes, etc.

[0086] In some embodiments, the processing temperature for the nickel-cobalt removal treatment is 85°C to 95°C. For example, it can be 85°C, 86°C, 87°C, 88°C, 89°C, 90°C, 91°C, 92°C, 93°C, 94°C, 95°C, etc.

[0087] In some embodiments, after adding elemental zinc and antimony trioxide to the first purified liquid for nickel-cobalt removal treatment to obtain a second purified liquid and nickel-cobalt slag, the step further includes: The second purification solution is subjected to zinc electrowinning treatment to obtain electrowinning waste liquid and cathode zinc.

[0088] The ratio of the electrowinning waste liquid to the second purified liquid is (10~20):1. For example, it can be 10:1, 11:1, 12:1, 13:1, 14:1, 15:1, 16:1, 17:1, 18:1, 19:1, or 20:1.

[0089] This step means removing zinc ions (Zn) dissolved in the "second purification solution". 2+ It is converted into high-purity elemental zinc (Zn) through electrochemical methods.

[0090] Zinc electrowinning is an electrolytic process that uses direct current to extract metallic zinc from a pure zinc sulfate solution. The basic process is as follows: A "second purification solution" (also called "new solution") is mixed with a large amount of "electrowinning waste solution" at a volume ratio of 10 to 20. That is, for every part of new solution added, 10 to 20 parts of waste solution are mixed in. This is done to maintain the conductivity and acidity of the solution in the electrolytic cell within a stable and suitable range. The mixed electrolyte is then fed into an electrolytic cell containing a cathode (usually an aluminum plate) and an anode (usually a lead-silver alloy plate). Direct current is applied across the electrolytic cell. Under the influence of the electric field, positively charged zinc ions move towards the cathode and undergo a chemical reaction on the cathode surface.

[0091] After electrowinning, two products are obtained: (1) Cathode Zinc: Zinc ions in the solution gain electrons on the cathode plate and are reduced to a pure layer of metallic zinc. After a period of electrolysis, this zinc layer reaches a certain thickness and is peeled off from the cathode plate to become the final product. (2) Electrowinning Waste Liquid: The liquid remaining in the electrolytic cell. Because a large number of zinc ions are consumed and new sulfuric acid is produced by the anode reaction, this waste liquid is characterized by "zinc deficiency and acid enrichment".

[0092] The electrodeposition process mainly involves two reactions that occur at the cathode and anode.

[0093] (1) Cathode reaction: At the cathode (negative electrode), zinc ions (Zn) in the solution 2+ The electrons are gained, reduced to metallic zinc (Zn), and deposited on the cathode plate.

[0094] Reaction formula: Zn 2+ +2e - →Zn(s) (2) Anode Reaction: At the anode (positive electrode), water molecules (H2O) lose electrons (i.e., are oxidized), decomposing to produce oxygen (O2) and hydrogen ions (H+). + ).

[0095] Reaction formula: 2H₂O - 4e - →O2(g)+4H + ; Electrowinning is a standard method for producing high-purity metals. This step produces cathode zinc with very high purity (e.g., above 99.9%), a high-value product that can be sold directly as a commodity. The electrowinning process not only produces zinc but also regenerates sulfuric acid. The resulting "electrowinning waste liquid" is essentially a highly concentrated sulfuric acid solution. This waste liquid can be directly returned to the upstream of the process, used as the "first acid leaching agent" and "second acid leaching agent" mentioned in the aforementioned embodiments. This establishes a complete acid cycle system, eliminating the need to purchase large quantities of sulfuric acid, significantly reducing production costs and environmental impact.

[0096] In some embodiments, both the first acidic leaching agent and the second acidic leaching agent are prepared by means of the electrowinning waste liquid and slag washing water generated from the zinc electrowinning process; In the recycling method provided in this embodiment, the two main acidic reagents (first and second acidic leaching agents) used to dissolve the metal can be prepared without relying on commercially available sulfuric acid purchased from outside, but by utilizing the by-product → "electrowinning waste liquid" generated in the "zinc electrowinning" stage within this process.

[0097] This method's acid recycling design delivers significant economic benefits, environmental value, and high process integration. By utilizing the sulfuric acid regenerated during the zinc electrowinning step and recycling it as a leaching agent, this method virtually eliminates the need to purchase commercial sulfuric acid, a major consumable, thereby drastically reducing raw material procurement costs. Simultaneously, this closed-loop recycling model, transforming traditional "waste liquid" into valuable chemical raw materials, minimizes waste liquid emissions and avoids the environmental footprint associated with new acid production and transportation, fully aligning with the principles of green and sustainable production. More importantly, this feature demonstrates the completeness and ingenuity of the method design, closely linking the endpoint (electrowinning) and the starting point (leaching) of the zinc recovery process, forming a highly efficient, compact, and self-sufficient acid recycling system.

[0098] The sulfuric acid concentration of the electrowinning waste liquid is 180 g / L to 220 g / L. For example, it can be 180 g / L, 190 g / L, 200 g / L, 210 g / L, 220 g / L, etc.

[0099] In some embodiments, the leaching temperature of the second acid leaching treatment is 70°C to 95°C; for example, it can be 70°C, 75°C, 80°C, 85°C, 90°C, 95°C, etc.

[0100] In the second acid leaching treatment, the liquid-to-solid ratio of the second acid leaching agent to the neutral leaching residue is (3~6) mL:1g; for example, it can be 3 mL:1 g, 4 mL:1 g, 5 mL:1 g, 6 mL:1 g, etc.

[0101] The leaching time for the second acid leaching treatment is 180 minutes to 300 minutes. For example, it can be 180 minutes, 190 minutes, 200 minutes, 220 minutes, 230 minutes, 250 minutes, 260 minutes, 270 minutes, 280 minutes, 300 minutes, etc.

[0102] In some embodiments, step S4, which involves reducing ferric iron in the acidic leachate to ferrous iron, and using the zinc-rich precipitate as a neutralizing agent to adjust the pH value to precipitate and separate bismuth, indium, and germanium, includes: Step S41: A sulfur dioxide mixture is introduced into the acidic leachate to carry out a reduction reaction, so as to reduce the ferric iron in the acidic leachate to ferrous iron, and obtain the reduced reaction solution after the reduction reaction.

[0103] This step is a crucial pretreatment for all subsequent separation steps. Due to the presence of ferric ions (Fe3+)... 3+ During subsequent neutralization, ferrous ions (Fe2+) tend to precipitate along with bismuth and indium, causing interference and making separation difficult. Therefore, it is necessary to first remove the more chemically stable ferrous ions (Fe2+). 2+ The treatment method involves introducing a mixture of sulfur dioxide (SO2) gas into the "acidic leachate" to initiate a chemical reduction reaction, thereby obtaining a "reduction reaction solution" in which the iron element mainly exists in the form of ferrous ions. The reaction formula can be: 2Fe 3+ +SO2+2H2O→2Fe 2+ +SO4 2- +4H + .

[0104] In the above reaction, when the sulfur dioxide mixture is passed into an acidic aqueous solution containing ferric ions, the sulfur dioxide in the mixture is oxidized to sulfate ions (SO42-). 2- At the same time, it reduces ferric ions to ferrous ions (Fe2+). 2+ ), and in the process, hydrogen ions (H) are produced. + This further enhances the acidity of the solution.

[0105] Specifically, in the reduction reaction, the amount of sulfur dioxide gas used can be 2 to 5 times the theoretical molar amount required to reduce ferric iron (for example, 2 times, 3 times, 4 times, 5 times, etc.), and the reaction is carried out at a temperature of 40℃ to 80℃ (for example, 40℃, 50℃, 60℃, 70℃, 80℃, etc.) for 60 minutes to 180 minutes (for example, 60 minutes, 70 minutes, 80 minutes, 90 minutes, 100 minutes, 120 minutes, 140 minutes, 160 minutes, 180 minutes, etc.).

[0106] In addition, the sulfur dioxide mixture can be an air mixture containing 1.3% to 1.8% SO2.

[0107] Step S42: The reduction reaction solution is subjected to a first neutralization treatment using the zinc-rich precipitate obtained from the purification process to obtain bismuth-rich material and bismuth-precipitated liquid.

[0108] The purpose of the above steps is to selectively precipitate and separate bismuth. The treatment method involves using the "zinc-rich precipitate" (mainly zinc carbonate) obtained in the purification step of the aforementioned embodiments as a neutralizing agent, which is slowly added to the "reduction reaction solution" to precisely raise the pH value of the solution to a specific, extremely low range. At this specific pH value, bismuth ions preferentially precipitate. Through solid-liquid separation, a solid "bismuth-rich substance" and a liquid "bismuth-precipitated liquid" with most of the bismuth removed are obtained.

[0109] The neutralization reaction can be carried out at a temperature of 50°C to 70°C (e.g., 50°C, 60°C, 70°C, etc.) for 20 to 60 minutes (e.g., 20 minutes, 30 minutes, 40 minutes, 50 minutes, 60 minutes, etc.). The ultimate goal is to ensure that the bismuth content in the "post-bismuth precipitation solution" is no more than 0.3 mg / L, which typically corresponds to neutralizing the pH to a range of 0.1 to 0.5.

[0110] Step S43: The bismuth-precipitated liquid is subjected to a second neutralization treatment using the zinc-rich precipitate to obtain an indium-germanium enrichment and an indium-germanium-precipitated liquid.

[0111] The purpose of the above steps is to precipitate and separate indium and germanium. Zinc-rich precipitate is then added to the "bismuth-precipitated solution" as a neutralizing agent to further increase the pH of the solution. When the pH reaches a suitable range, indium and germanium will precipitate. Through solid-liquid separation, a solid "indium-germanium concentrate" and a liquid "indium-germanium-precipitated solution" with indium and germanium removed are obtained.

[0112] Specifically, the neutralization reaction can also be carried out at a temperature of 50℃~70℃ (e.g., for 50 minutes, 60 minutes, 70 minutes, etc.) for 20~60 minutes (e.g., for 20 minutes, 30 minutes, 40 minutes, 50 minutes, 60 minutes, etc.), ultimately neutralizing the pH of the solution to 4.0~4.2 (e.g., 4, 4.1, 4.2, etc.). The goal is to ensure that the indium content in the "indium-germanium precipitation solution" is no greater than 0.15 mg / L and the germanium content is no greater than 0.15 mg / L.

[0113] Step S44: The indium-germanium precipitate solution is subjected to a catalytic oxidation reaction of iron precipitation using the sulfur dioxide mixed gas, and the zinc-rich precipitate is separated into solid and liquid phases to obtain iron oxide precipitate and iron-removed solution.

[0114] The purpose of the above steps is to remove iron from the solution. The treatment method includes, firstly, introducing a mixture of air and sulfur dioxide as a catalyst into the "indium-germanium precipitation solution" to remove stable ferrous ions (Fe²⁺) from the solution. 2+ ) Re-oxidized back to ferric ions (Fe 3+ Then, the "zinc-rich precipitate" is used as a neutralizing agent to maintain the pH within a specific range, allowing the newly generated ferric ions to hydrolyze and precipitate in the form of stable iron oxides. The sulfur dioxide mixed gas can be an air mixture containing 1.3% to 1.8% SO2. The sulfur dioxide mixed gas used here can be the same as or different from the mixed gas used in the reduction reaction.

[0115] Through solid-liquid separation, a solid "iron oxide precipitate" (i.e., iron slag) and a "removed iron liquid" that is essentially free of iron are obtained. The goal of this step is to ensure that the iron content in the final "removed iron liquid" is no more than 20 mg / L. This iron slag can be returned to the rotary hearth furnace system for iron recovery, while the removed iron liquid can be returned to the neutral leaching solution purification stage for further recovery of residual metals such as zinc.

[0116] The aforementioned multi-step separation process demonstrates extremely high separation precision and selectivity. Through precise stepwise control of the redox state of iron and the pH value of the solution, it achieves the individual and highly selective separation of components such as bismuth, indium, germanium, and iron from a complex mixed system. This precise control aims to ensure near-complete recovery of these valuable rare metals. The extremely low target residual concentrations (e.g., bismuth no more than 0.3 mg / L, indium and germanium no more than 0.15 mg / L) reflect the high recovery rate, effectively avoiding resource waste. More importantly, this process achieves a complete closed-loop resource cycle. Not only are the target metals recovered, but the final products, "iron oxide precipitate" and "iron-removed liquid," are also returned to the upstream of the process for reuse, maximizing resource utilization and embodying the "zero-waste" circular economy concept.

[0117] In summary, the recycling method provided in this application embodiment has significant comprehensive advantages over traditional recycling technologies in terms of raw material adaptability, resource utilization efficiency, and internal circular economy.

[0118] First, this method is highly adaptable to rotary hearth furnace flue dust with complex compositions, effectively handling the differences in valuable metal content caused by different iron ore raw materials, especially materials with zinc content ranging from 10% to 55%. Second, through a tiered recovery design, this method can maximize the comprehensive recovery and utilization of various valuable metals contained in the flue dust, such as potassium, sodium, zinc, indium, germanium, lead, silver, bismuth, copper, cadmium, nickel, and cobalt. The leaching rates of potassium and sodium can reach over 99%, zinc over 99%, and indium and germanium over 98%.

[0119] These advantages are achieved through its highly integrated and closed-loop process design. First, the process achieves closed-loop recovery of zinc metal. The small amount of zinc dissolved in the initial water leaching step is purified with sodium carbonate to generate a zinc-rich precipitate (zinc carbonate). This precipitate is then creatively used as a neutralizing agent in subsequent acid leaching steps to neutralize and precipitate indium and germanium, preventing zinc resource loss and reducing reagent costs. Second, the process achieves closed-loop recycling of sulfur dioxide (SO2) gas. After reducing ferric iron with SO2 gas, a trace amount escapes and can be directly mixed into the air for subsequent oxidation of ferrous iron and precipitation of iron resources, improving gas utilization efficiency. Third, the method achieves the regeneration and recycling of sulfuric acid. The sulfuric acid required for acid leaching in the zinc recovery system comes entirely from the waste electrowinning solution generated in the final zinc electrowinning process, eliminating the need for external purchases. This achieves closed-loop acid recycling while producing cathode zinc, significantly reducing production costs.

[0120] The present invention will be further illustrated below with specific embodiments. However, it should be understood that these embodiments are merely for the purpose of more detailed illustration and should not be construed as limiting the present invention in any way.

[0121] In this embodiment, the specific composition of the rotary hearth furnace flue dust is shown in Table 1: Table 1. Composition of Flue Dust from Rotary Hearth Furnace

[0122] Table 1 shows that the rotary hearth furnace flue dust has a high potassium and sodium content. The main phases in the rotary hearth furnace flue dust are zinc oxide (ZnO), followed by sodium chloride (NaCl), potassium chloride (KCl), and potassium zinc tetrachloride (K2ZnCl4), as well as small amounts of lead fluoride chloride (PbClF), zinc iron spinel (ZnFe2O4), and trace amounts of calcium iron silicates, magnesium oxide, and magnesium salts.

[0123] Example 1 In this embodiment, potassium, sodium, zinc, indium, germanium, lead, silver, and bismuth are recovered from the flue dust of the rotary hearth furnace.

[0124] Experimental methods (reference) Figure 2 ): Step 1: Mix sensible heat 100℃ rotary hearth furnace dust with 50℃ evaporative condensate (initially, laboratory tap water is used), control the liquid-solid ratio at 6mL:1g, and leach at 80℃ (in the laboratory, a heating and insulation device is used) for 30 minutes. Filter to obtain water leaching solution and water leaching residue. At this time, the potassium leaching rate is 99%, the sodium leaching rate is 99%, and the zinc leaching rate is 7%.

[0125] Step 2: Add Na2CO3 solid to the above water extract to adjust the pH of the system to 7.5, and remove zinc, calcium and magnesium impurities at a temperature of 40℃ for a reaction time of 30 minutes to obtain purified potassium-sodium solution and zinc-rich precipitate (zinc carbonate); the purified potassium-sodium solution is then concentrated by evaporation and cooled by crystallization to obtain the corresponding products.

[0126] Step 3: Mix the water-leaching residue with a 5% sodium carbonate solution at a liquid-to-solid ratio of 4 mL:1 g. Leach at 80°C for 60 minutes with stirring to remove fluoride and chlorine. Filter to obtain defluorinated chlorine residue and fluorine-containing chlorine wastewater. The fluoride content in the defluorinated chlorine residue is ≤0.015%, and the chloride content is ≤0.1%. The secondary defluorinated chlorine wastewater can be reused after adding sodium carbonate. The fluorine-containing chlorine wastewater is treated as wastewater.

[0127] Step 4: Mix the defluorination and chlorination residue with sulfuric acid of a certain concentration. The amount of sulfuric acid used is 1.05 times the theoretical amount calculated based on the zinc content in the defluorination and chlorination residue, with a liquid-to-solid ratio of 6 mL:1 g. Perform neutral sulfuric acid leaching of zinc, with a leaching system pH of 5.2. Leach at 80°C with stirring for 180 minutes. Filter to obtain neutral leaching residue and neutral leaching solution. Waste electrolyte from zinc electrowinning (H2SO4 concentration approximately 180-220 g / L) can be used as the sulfuric acid leaching agent, mixed with the acid leaching residue wash water, and adjusted to a suitable sulfuric acid concentration.

[0128] Step 5: Add a certain amount of zinc powder to the neutral leachate for displacement to remove copper and cadmium. The amount of zinc powder is 5 times the molar amount of copper and cadmium. The displacement temperature is about 50℃, and the time is 30 minutes. Filter to obtain copper-cadmium slag and the solution after copper and cadmium removal. Analyze the copper and cadmium content in the solution after copper and cadmium removal. It meets the requirements for copper and cadmium content in zinc electrowinning, i.e., Cu≤0.5mg / L, Cd≤2mg / L. Continue to add a certain amount of zinc powder and antimony trioxide powder to the solution after copper and cadmium removal for displacement to remove nickel and cobalt. The antimony trioxide powder (analytical grade reagent Sb2O3>99.5%) is 1 / 25 of [Co 2+ ] g / L, zinc powder is 1+100[Co 2+ ]+50[Ni 2+[g / LL], temperature approximately 90℃, time 90 minutes. Filter to obtain purified liquid and cobalt-nickel slag. Analyze the nickel and cobalt content in the purified liquid to ensure it meets the requirements for nickel and cobalt content in zinc electrowinning, i.e., Ni≤1.5mg / L, Co≤2mg / L. The purified liquid is used as new liquid and mixed with electrowinning waste liquid at a certain mixing ratio to produce cathode zinc products that meet national standards through electrowinning. Copper-cadmium slag and nickel-cobalt slag can be sold at a price.

[0129] Step 6: Mix the neutral leaching residue with sulfuric acid of a certain concentration for acidic leaching. The sulfuric acid concentration is 180 g / L. Waste electrolyte from zinc electrowinning (H2SO4 concentration approximately 180-220 g / L) can be used as the sulfuric acid leaching agent. Mix the acid leaching residue wash water to adjust the sulfuric acid concentration to a suitable level. The leaching temperature is 90℃, the liquid-to-solid ratio is 5 mL:1 g, and the leaching time is 240 minutes. Filter to obtain acidic leaching residue and acidic leachate. Wash the acidic leaching residue to obtain acidic leaching residue wash water and lead-silver slag rich in lead, silver, and bismuth. The lead-silver slag can be sold at a price.

[0130] Step 7: For the acidic leachate, SO2 gas is introduced to reduce the ferric iron in the leachate to ferrous iron. The amount of sulfur dioxide used is 2.5 times the theoretical molar amount of ferric iron to be reduced. The reduction temperature is 60℃ and the reduction time is 90 minutes. The resulting reduced acidic leachate is neutralized with the zinc-rich slag (zinc carbonate) obtained from the water leachate purification at a temperature of 50℃ for 30 minutes, until the pH value reaches 0.3 to precipitate bismuth. The precipitate is filtered to separate the bismuth-rich material and the bismuth-precipitated liquid. The bismuth content in the bismuth-precipitated liquid is ≤0.3 mg / L. The bismuth-precipitated liquid is further neutralized to around pH 4 to precipitate indium and germanium. The precipitate is filtered to separate the indium and germanium-rich material and the indium and germanium-precipitated liquid. The indium content in the indium and germanium-precipitated liquid is ≤0.15 mg / L and the germanium content is ≤0.15 mg / L.

[0131] Step 8: For the indium-germanium precipitate solution, use an air mixture containing 1.5% SO2. The SO2 can be recovered and reused from the trace amount that escaped during the reduction of ferric iron. The reaction temperature is 60℃, allowing Fe... 2+ Catalytically oxidized to Fe 3+ After neutralization with zinc-rich precipitate (zinc carbonate) and stirring at pH 4.0 for 120 minutes, the generated Fe... 3+ Hydrolysis precipitation produces an iron oxide precipitate similar to goethite. Filtration yields a de-ironized liquid and iron slag. Analysis of the iron content in the de-ironized liquid confirms it meets the requirements for zinc electrowinning (Fe ≤ 20 mg / L). The de-ironized liquid is then mixed with a neutral leaching solution for copper, cadmium, nickel, and cobalt removal and purification treatment, followed by zinc recovery through electrowinning. The iron slag is returned to the rotary hearth furnace feeding system for further iron recovery.

[0132] Example 2 like Figure 1As shown, this embodiment provides a method for the cascade recovery of potassium, sodium, zinc, indium, germanium, lead, silver, and bismuth from rotary hearth furnace flue dust, specifically including the following steps: Step 1: Mix sensible heat 100℃ rotary hearth furnace dust with 50℃ evaporative condensate (initially, laboratory tap water is used), control the liquid-solid ratio at 5mL:1g, and leach at 80℃ (in the laboratory, a heating and insulation device is used) for 60 minutes. Filter to obtain water leaching solution and water leaching residue. At this time, the potassium leaching rate is 99%, the sodium leaching rate is 99%, and the zinc leaching rate is 8%.

[0133] Step 2: Add Na2CO3 solid to the above water extract to adjust the pH of the system to 7.4. Remove zinc, calcium and magnesium impurities at 40℃ for 30 minutes to obtain purified potassium-sodium solution and zinc-rich precipitate (zinc carbonate). The purified potassium-sodium solution is then concentrated by evaporation and cooled by crystallization to obtain the corresponding products.

[0134] Step 3: Mix the water-leached residue with a 3% sodium carbonate solution at a liquid-solid ratio of 4 mL:1 g. Leach at 80°C for 60 minutes with stirring to remove fluoride and chlorine. Filter to obtain defluorinated chlorine residue and fluorine-containing chlorine wastewater. Alternatively, mix the primary defluorinated chlorine residue with a 3% sodium carbonate solution at a liquid-solid ratio of 4 mL:1 g. Leach at 80°C for 120 minutes with stirring to remove fluoride and chlorine. Filter to obtain defluorinated chlorine residue and fluorine-containing chlorine wastewater. The defluorinated chlorine residue has an F content ≤0.015% and a Cl content ≤0.1%. The fluorine-containing chlorine wastewater is treated as wastewater.

[0135] Step 4: Mix the defluorination and chlorination residue with sulfuric acid of a certain concentration. The amount of sulfuric acid used is 1.03 times the theoretical amount calculated based on the zinc content in the defluorination and chlorination residue, with a liquid-to-solid ratio of 7 mL:1 g. Perform neutral sulfuric acid leaching of zinc, with the pH of the leaching system at 5.0. Leach at a leaching temperature of 60°C with stirring for 120 minutes. Filter to obtain neutral leaching residue and neutral leaching solution. Waste electrolyte from zinc electrowinning (H2SO4 concentration approximately 180-220 g / L) can be used as the sulfuric acid leaching agent, mixed with the acid leaching residue wash water, and adjusted to a suitable sulfuric acid concentration.

[0136] Step 5: Add a certain amount of zinc powder to the neutral leachate for displacement to remove copper and cadmium. The amount of zinc powder is 4.5 times the molar amount of copper and cadmium. The displacement temperature is about 50℃, and the time is 30 minutes. Filter to obtain copper-cadmium slag and the solution after copper and cadmium removal. Analyze the copper and cadmium content in the solution after copper and cadmium removal. It meets the requirements for copper and cadmium content in zinc electrowinning, i.e., Cu≤0.5mg / L, Cd≤2mg / L. Continue to add a certain amount of zinc powder and antimony trioxide powder to the solution after copper and cadmium removal for displacement to remove nickel and cobalt. The antimony trioxide powder (analytical grade reagent Sb2O3>99.5%) is 1 / 25 of [Co 2+ ] g / L, zinc powder is 1+100[Co 2+ ]+50[Ni 2+The concentration of nickel and cobalt in the solution was [g / L], the temperature was approximately 90℃, and the time was 90 minutes. Filtration yielded purified liquid and cobalt-nickel slag. Analysis of the nickel and cobalt content in the purified liquid confirmed that it met the requirements for nickel and cobalt content in zinc electrowinning, i.e., Ni ≤ 1.5 mg / L, Co ≤ 2 mg / L. The purified liquid was used as new liquid and mixed with electrowinning waste liquid at a certain mixing ratio to produce cathode zinc products that meet national standards through electrowinning. Copper-cadmium slag and nickel-cobalt slag can be sold at a price.

[0137] Step 6: Mix the neutral leaching residue with sulfuric acid of a certain concentration for acidic leaching. The sulfuric acid concentration is 180 g / L. Waste electrolyte from zinc electrowinning (H2SO4 concentration approximately 180 g / L~220 g / L) can be used as the sulfuric acid leaching agent. Mix the acid leaching residue wash water to adjust the sulfuric acid concentration to a suitable level. The leaching temperature is 90℃, the liquid-to-solid ratio is 5 mL:1 g, and the leaching time is 240 minutes. Filter to obtain acidic leaching residue and acidic leachate. Wash the acidic leaching residue to obtain acidic leaching residue wash water and lead-silver slag rich in lead, silver, and bismuth. The lead-silver slag can be sold at a price.

[0138] Step 7: For the acidic leachate, SO2 gas is introduced to reduce the ferric iron in the leachate to ferrous iron. The amount of sulfur dioxide used is 5 times the theoretical molar amount of ferric iron to be reduced. The reduction temperature is 60℃ and the reduction time is 80 minutes. The resulting reduced acidic leachate is neutralized with the zinc-rich slag (zinc carbonate) obtained from the water leachate purification at a temperature of 50℃ for 30 minutes, until the pH value reaches 0.3 to precipitate bismuth. The precipitate is filtered to separate the bismuth-rich material and the bismuth-precipitated liquid. The bismuth content in the bismuth-precipitated liquid is ≤0.3 mg / L. The bismuth-precipitated liquid is further neutralized to around pH 4 to precipitate indium and germanium. The precipitate is filtered to separate the indium and germanium-rich material and the indium and germanium-precipitated liquid. The indium content in the indium and germanium-precipitated liquid is ≤0.15 mg / L and the germanium content is ≤0.15 mg / L.

[0139] Step 8: For the indium-germanium precipitate solution, use an air mixture containing 1.6% SO2. The SO2 can be recovered and reused from the trace amount that escaped during the reduction of ferric iron. The reaction temperature is 60℃, allowing Fe... 2+ Catalytically oxidized to Fe 3+ After neutralization with zinc-rich slag (zinc carbonate) and stirring for 240 minutes at pH 4.1, the generated Fe... 3+ Hydrolysis precipitation produces an iron oxide precipitate similar to goethite. Filtration yields a de-ironized liquid and iron slag. Analysis of the iron content in the de-ironized liquid confirms it meets the requirements for zinc electrowinning (Fe ≤ 20 mg / L). The de-ironized liquid is then mixed with a neutral leaching solution for copper, cadmium, nickel, and cobalt removal and purification treatment, followed by zinc recovery through electrowinning. The iron slag is returned to the rotary hearth furnace feeding system for further iron recovery. Finally, it should be noted that the above embodiments are only used to illustrate the technical solutions of the present invention, and not to limit them. Although the present invention has been described in detail with reference to the foregoing embodiments, those skilled in the art should understand that modifications can still be made to the technical solutions described in the foregoing embodiments, or equivalent substitutions can be made to some or all of the technical features therein. Such modifications or substitutions do not cause the essence of the corresponding technical solutions to deviate from the scope of the technical solutions of the embodiments of the present invention.

Claims

1. A method for recovering metal elements from a smelter dust in a rotary hearth furnace, characterized by, The application relates to a method for treating rotary hearth furnace dust. The method comprises the following steps: treating rotary hearth furnace dust by water immersion to obtain water immersion liquid and water immersion residue; purifying the water immersion liquid to obtain purified potassium-sodium liquid and zinc-rich precipitate; treating the water immersion residue by first acid immersion with a first acid immersion agent to obtain zinc-rich neutral immersion liquid and neutral immersion residue; treating the neutral immersion residue by second acid immersion with a second acid immersion agent to obtain acid immersion liquid containing zinc, indium, germanium, bismuth and iron and acid immersion residue; reducing trivalent iron in the acid immersion liquid to divalent iron; and adjusting the pH value with the zinc-rich precipitate as a neutralizing agent to precipitate and separate bismuth, indium and germanium. The water immersion treatment comprises the following steps: mixing the rotary hearth furnace dust with water to obtain the water immersion liquid and the water immersion residue through solid-liquid separation; the temperature of the rotary hearth furnace dust is 100-200 DEG C; the water is evaporated condensate water; the temperature of the water is 50-55 DEG C; the liquid-solid ratio of the rotary hearth furnace dust to water is (3-6 mL):1 g; and the water immersion treatment time is 10-60 minutes. The purification treatment comprises the following steps: adding an impurity removal agent into the water immersion liquid to obtain the potassium-sodium liquid and the zinc-rich precipitate; the impurity removal agent comprises sodium carbonate; the purification treatment temperature is controlled to be 25-60 DEG C; and the pH value of the purification system after the impurity removal agent is added is 6.5-7.

5. Before the first acid immersion treatment, the method further comprises the following steps: treating the water immersion residue by fluorine-chlorine removal at least once; the fluorine-chlorine removal treatment comprises the following steps: mixing the water immersion residue with sodium carbonate for immersion, and then performing solid-liquid separation to obtain fluorine-chlorine removal treated water immersion residue and fluorine-chlorine containing waste liquid; the concentration of the sodium carbonate is 2-5%; the fluorine-chlorine removal treatment time is 30-180 minutes; the fluorine-chlorine removal treatment temperature is 40-95 DEG C; and the liquid-solid ratio of the sodium carbonate to the water immersion residue is (3-6 mL):1 g.

2. The method of claim 1, wherein the metal elements in the rotary hearth furnace dust are recovered, The pH value of the leaching system of the first acid immersion treatment is 5.2-5.4; the first acid immersion agent comprises sulfuric acid, and the addition amount of the sulfuric acid is 1.0-1.05 times the theoretical sulfuric acid amount calculated according to the zinc content in the water immersion residue; the leaching temperature of the first acid immersion treatment is 40-80 DEG C; the liquid-solid ratio of the first acid immersion agent to the water immersion residue is (4-8 mL):1 g; and the leaching time of the first acid immersion treatment is 60-180 minutes. ​ After the step of treating the water immersion residue by first acid immersion to obtain zinc-rich neutral immersion liquid and neutral immersion residue, the method further comprises the following steps: performing copper-cadmium displacement reaction on the neutral immersion liquid with elemental zinc to obtain copper-cadmium residue and first purification liquid; adding elemental zinc and diantimony trioxide into the first purification liquid to remove nickel and cobalt, and then obtaining second purification liquid and nickel-cobalt residue. ​ 3. The method of claim 1, wherein the rotary hearth furnace dust is a dust generated in a process of producing a ferroalloy or a metal by using a rotary hearth furnace. ​ ​ ​ 4. The method for recovering metal elements from rotary hearth furnace flue dust as described in claim 1, characterized in that, ​ ​ ​ ​ 5. The method for recovering metal elements from rotary hearth furnace flue dust as described in claim 1, characterized in that, ​ ​ ​ ​ ​ 6. The method of claim 1, wherein the metal elements in the rotary hearth furnace dust are recovered, ​ ​ ​ ​ 7. The method for recovering metal elements from rotary hearth furnace flue dust as described in claim 6, characterized in that, The copper-cadmium displacement reaction, the amount of elemental zinc added is 3-6 times the molar amount of copper and cadmium in the neutral leaching solution; and / or, The displacement temperature of the copper-cadmium displacement reaction is 45-55℃; and / or, The reaction time of the copper-cadmium displacement reaction is 30-60 minutes; and / or, The mass concentration of the antimony trioxide is less than 1 / 10 of the mass concentration of the Co ions ; and / or, The calculation method of the mass concentration of elemental zinc in the nickel-cobalt removal treatment: ； wherein, a mass concentration representing elemental zinc; a mass concentration representing cobalt ions; a mass concentration representing nickel ions; a has a value in the range of 30 to 100; b has a value in the range of 30 to 50; and / or, The treatment time of the nickel-cobalt removal treatment is 60-120 minutes; and / or, The treatment temperature of the nickel-cobalt removal treatment is 85-95℃.

8. The method for recovering metal elements from rotary hearth furnace flue dust as described in claim 6, characterized in that, The step of adding elemental zinc and diantimony trioxide in the first purified solution for nickel-cobalt removal treatment, obtaining second purified solution and nickel-cobalt residue, further comprises: The second purified solution is subjected to zinc electrowinning treatment to obtain electrowinning waste liquid and cathode zinc; The ratio of the electrowinning waste liquid to the second purified solution is (10-20):

1.

9. The method for recovering metal elements from rotary hearth furnace flue dust as described in claim 8, characterized in that, The first acidic leaching agent and the second acidic leaching agent are prepared by the electrowinning waste liquid and residue washing water generated by the zinc electrowinning treatment; The sulfuric acid concentration of the electrowinning waste liquid is 180-220g / L.

10. The method of claim 1, wherein the metal elements in the rotary hearth furnace dust are recovered, The leaching temperature of the second acidic leaching treatment is 70-95℃; and / or, ​ In the second acidic leaching treatment, the liquid-solid ratio of the second acidic leaching agent to the neutral leaching residue is (3-6)mL:1g; and / or, The leaching time of the second acidic leaching treatment is 180-300 minutes.

11. The method of claim 1, wherein the metal elements in the rotary hearth furnace dust are recovered, The step of reducing trivalent iron to divalent iron in the acidic leaching solution; and adjusting the pH value with the zinc-rich precipitate as a neutralizing agent to precipitate and separate bismuth, indium and germanium, comprises: ​ The acidic leaching solution is mixed with sulfur dioxide gas to perform a reduction reaction, so as to reduce trivalent iron in the acidic leaching solution to divalent iron, and obtain a reduction reaction liquid after reduction reaction; The reduction reaction liquid is subjected to a first neutralization treatment with the zinc-rich precipitate obtained by the purification treatment, to obtain a bismuth-rich substance and a post-bismuth precipitation liquid; The post-bismuth precipitation liquid is subjected to a second neutralization treatment with the zinc-rich precipitate, to obtain an indium-germanium enriched substance and a post-indium-germanium precipitation liquid; The post-indium-germanium precipitation liquid is subjected to a catalytic oxidation and iron precipitation reaction with the sulfur dioxide gas, and the zinc-rich precipitate is separated into an iron oxide precipitate and an iron-removed liquid; In the reduction reaction, the amount of sulfur dioxide gas in the sulfur dioxide gas mixture is 2-5 times the theoretical reduction molar amount of trivalent iron; and / or, The sulfur dioxide gas mixture is air mixed gas containing 1.3%-1.8% SO2; The reaction temperature of the reduction reaction is 40-80℃; and / or, The reaction time of the reduction reaction is 60-180 minutes; and / or, In the first neutralization treatment, the neutralization temperature is 50-70℃; and / or, In the first neutralization treatment, the neutralization time is 20-60 minutes; and / or, The bismuth content in the post-bismuth precipitation liquid is not more than 0.3mg / L; and / or, In the second neutralization treatment, the neutralization temperature is 50-70℃; and / or, The neutralization time in the second neutralization treatment is 20-60 minutes; and / or, The neutralization in the second neutralization treatment is to pH 4-4.2; and / or, The indium content in the indium-germanium precipitation post-solution is not more than 0.15 mg / L; and / or, The germanium content in the indium-germanium precipitation post-solution is not more than 0.15 mg / L; and / or, The iron content in the iron removal post-solution is not more than 20 mg / L.

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