Recycling process of vanadium titano-magnetite slag

Abstract

The invention discloses a recovery process of vanadium titano-magnetite slag, and relates to the technical field of mineral recovery. A recovery process of vanadium titano-magnetite slag comprises the following steps that a vanadium titano-magnetite sintering dedusting ash raw material is subjected to ball milling, and raw material powder is obtained; performing dry magnetic separation on the raw material powder in a high-voltage pulsed magnetic field to separate out magnetic iron minerals and dry tailings; mixing the dry tailings, NH4Cl and KCl, adding water, filtering and separating to obtain filtrate 1 and filter residue 1; filtering to obtain filtrate 2 and filter residue 2; adjusting the pH value of the filtrate 2, and filtering to separate filtrate 3 and filter residues 3; adjusting the pH value of the filtrate 3, adding K2S, and filtering to separate filtrate 4 and filter residues 4; filtering to obtain filtrate 5 and filter residue 5; a roasted product 1 and a roasted product 2 are obtained; filter residues 6 and filtrate 6 are obtained; and cooling and crystallizing the filtrate 6 at normal temperature, filtering, drying and crushing to obtain the potassium-nitrogen mixture. The problem that a systematic comprehensive separation and utilization method for various metal resources is lacked is solved.

Description

Technical Field

[0001] The present invention relates to the technical field of mineral recovery, and particularly to a recovery process for vanadium-titanium magnetite slag. Background Art

[0002] Vanadium-titanium resources in China are very rich. The proven titanium resource reserves (calculated as TiO2) are 720 million tons, accounting for about 1 / 3 of the world's total reserves. The vanadium resource reserves (calculated as V2O5) are 42.9 million tons, accounting for about 21% of the world's total reserves. In the mid-1960s, in order to make good use of the Panxi vanadium-titanium magnetite, the state organized large-scale industrial scientific experiments, solved the basic smelting process problems, and pioneered the technology of smelting vanadium-titanium magnetite in blast furnaces globally, laying a foundation for the development and utilization of resources in the Panzhihua-Xichang region. The relevant achievements won the national invention award. However, due to some important technical problems not being solved, such as foamy slag, sticking of molten iron in ladles, high iron loss, complex dedusting ash, low grade, large slag volume, etc., which have long troubled production, the smelting process and operation technology are still not completely mature, making the blast furnace indexes of Pangang still unable to compare with those of other domestic steel plants. In Panxi vanadium-titanium magnetite, in addition to iron, vanadium, and titanium, there are also considerable reserves of chromium, cobalt, nickel, potassium, copper, gallium, germanium, zinc, lead, as well as scandium, yttrium, rare earths, sulfur, tellurium, bismuth, platinum group, etc., which is a rare treasure house of rare and precious metal resources in China. At present, iron, vanadium, and titanium can be comprehensively recovered from Panxi vanadium-titanium magnetite, and its comprehensive market competitiveness is also much stronger than that of single iron ore resources. Vanadium-titanium magnetite sintering dedusting ash includes sintering machine head dedusting ash, sintering machine tail dedusting ash, ring cooler dedusting ash, screening dedusting ash, and ignition and breaking dedusting ash, etc., but its chemical composition and mineral structure are very different from those of ordinary iron ore sintering dedusting ash. Youlian Testing Technology Service Co., Ltd. in Jiangsu Province conducted a leaching toxicity analysis on the raw materials of vanadium-titanium magnetite sintering dedusting ash in November 2024. According to the "Identification Standard for Hazardous Wastes - Identification of Leaching Toxicity" (GB5085.3 - 2007), the test report shows that zinc and lead seriously exceed the standards, reaching 174 mg / L and 4000 mg / L respectively, far exceeding the limits of 100 mg / L and 5 mg / L, and nickel, chromium, beryllium, barium, silver, mercury, arsenic, fluoride, and cyanide all meet the standards.

[0003] How to dispose of a large amount of sintering ash generated by iron and steel plants has always been a major problem faced by iron and steel plants. If the sintering ash is directly returned to the sintering cycle as a secondary raw material, with the enrichment of alkali metal elements potassium and sodium in it, it is extremely easy to corrode the blast furnace wall, affecting the service life of the blast furnace and the production quality of steel. At the same time, it will also reduce the dust removal efficiency and operation stability of the sintering electrostatic precipitator, increase the operation energy consumption of the device, cause the dust concentration in the discharged flue gas to exceed the standard, and pollute the environment and a series of other problems. Therefore, even though the sintering ash contains a large amount of iron, it is not suitable for direct recycling. However, the stacking treatment of sintering ash requires a large amount of land, wasting a lot of manpower and financial resources, and the soluble salts contained in it will also pollute the environment with water penetration, causing soil salinization and surface water halogenation, which poses a great harm to the ecological environment. In recent years, the resource utilization and development of sintering ash have been increasingly emphasized at home and abroad, and great progress has been made in the technical research of recovering metals such as K and Zn, creating a foundation for improving resource utilization rate theoretically and practically. In terms of the recovery of metallic iron, sintering soot contains a large amount of iron elements. Usually, the carbon in the soot can be used as a reducing agent, and after mixing it with iron ore, the iron in the soot and iron ore can be reduced. In terms of potassium salt extraction, developing the technology for extracting potassium from sintering soot not only helps to alleviate the tight trend of domestic potassium resources, but also can achieve resource recycling, and has become a research hotspot in recent years.

[0004] To sum up, at the present stage, great research progress has been made in the comprehensive utilization of the main elements in sintering soot at home and abroad. However, the different sources of sintering soot have a great impact on its composition, and the content of some rare metals is not high, which brings great difficulties to the reasonable utilization of soot. At present, the following problems in the research and industrial utilization of sintering soot need to be further improved: 1) Lack of a systematic solution. Researchers more often simply treat sintering ash as a solid waste and are limited to the utilization of a single metal resource, lacking comprehensive consideration of all valuable elements; 2) The separation technology still needs to be improved, especially the separation of some rare elements. Due to their low content and high separation cost, it is difficult to achieve industrialization; 3) The amount of secondary by-products is still large. Summary of the Invention

[0005] The technical problem to be solved by the present invention is the lack of a systematic method for the comprehensive separation and utilization of multiple metal resources. The purpose is to provide a recovery process for vanadium-titanium magnetite slag, which solves the problem of the lack of a systematic method for the comprehensive separation and utilization of multiple metal resources.

[0006] The present invention is achieved through the following technical solutions:

[0007] A recovery process for vanadium-titanium magnetite slag includes the following steps:

[0008] Ball-mill the raw material of vanadium-titanium magnetite sintering dust to obtain raw material powder;

[0009] The raw material powder is separated from gangue minerals, weakly magnetic iron minerals, medium magnetic iron minerals, and strongly magnetic iron minerals by dry magnetic separation under a high-voltage pulsed magnetic field. The residue other than magnetic iron minerals is the dry tailings;

[0010] Mix the dry tailings, NH4Cl, and KCl, add water to obtain a mixed solution, stir and react, filter and separate to obtain filtrate 1 and residue 1; cool filtrate 1 and perform two cooling precipitations, filter and separate to obtain filtrate 2 and residue, combine the residues precipitated twice to obtain residue 2; adjust the pH value of filtrate 2 to 9-10, filter and separate to obtain filtrate 3 and residue 3; adjust the pH value of filtrate 3 to 2-3, then add a saturated K2S solution, stir, and filter and separate to obtain filtrate 4 and residue 4; add a saturated potassium carbonate solution to residue 2, stir, and filter and separate to obtain filtrate 5 and residue 5; roast residue 4 and residue 5 respectively to obtain roasted product 1 and roasted product 2; mix filtrate 5 and filtrate 4, introduce CO2, stop introducing CO2 when the pH value reaches 7-8, then filter and separate to obtain residue 6 and filtrate 6; perform three-stage cyclic vacuum concentration - normal temperature cooling crystallization on filtrate 6, filter, dry, and pulverize to obtain a potassium-nitrogen mixture.

[0011] As a possible design, the average particle size of the above raw materials is 0.13mm - 0.17mm.

[0012] As a possible design, the above ball milling is dry ball milling; the average particle size of the raw material powder is 300 - 450 mesh.

[0013] As a possible design, the above dry magnetic separation under a high-voltage pulsed magnetic field is specifically to perform three-stage magnetic separation on the raw material powder: first, magnetic separation is performed under a magnetic field of 0.3 - 0.7T to obtain gangue minerals and strongly magnetic minerals, then magnetic separation is performed under a magnetic field of 0.9 - 1.3T to obtain medium magnetic minerals, and finally magnetic separation is performed under a magnetic field of 1.5 - 1.7T to obtain weakly magnetic minerals. The residue other than magnetic iron minerals is the dry tailings.

[0014] As a possible design, the molar concentration of KCl in the above mixed solution is 4 - 8mol / L, the molar concentration of NH4Cl is 2 - 2.4mol / L, the solid-liquid ratio is 1:(1.5 - 2.5), and the pH value is 4 - 5; the stirring of the above mixed solution is specifically to stir and react at 70 - 80°C for 60 - 120min.

[0015] As a possible design, the above filtrate 1 is cooled to 15 - 35°C and then two cooling precipitations are performed. The first precipitation time is 80 - 100min, and the second precipitation time is 110 - 130min.

[0016] As a possible design, the pH value of the above-mentioned filtrate 2 is specifically adjusted by concentrated ammonia water, and after adjusting the pH value, stirring is continued for 20 to 40 minutes.

[0017] As a possible design, the molar amount of K2S in the saturated K2S solution added to the above-mentioned filtrate 3 and the molar amount of Zn 2+ are in a ratio of 1:(0.9 to 1.1). After adding the saturated K2S solution, stirring is continued for 20 to 40 minutes.

[0018] As a possible design, after adding the saturated potassium carbonate solution to the above-mentioned filter residue 2, stirring is carried out for 10 to 30 minutes; the molar amount of saturated potassium carbonate and the molar amount of lead chloride in the filter residue 2 are in a ratio of 1:(0.9 to 1.1).

[0019] As a possible design, the above-mentioned filter residue 4 is calcined at 800 to 1000 °C for 1 to 2 hours; the above-mentioned filter residue 5 is calcined at 660 to 720 °C for 2 to 3 hours.

[0020] Compared with the prior art, the present invention has the following advantages and beneficial effects:

[0021] The present invention first obtains fine particles through ball milling, which is convenient for subsequent separation; then iron is separated by magnetic separation, and the recovery rate of iron in this step is as high as 93.89%; then the separated dry tailings are added to react with NH4Cl and a small amount of KCl, and NH4Cl provides H + by hydrolysis to dissolve metal oxides, and uses the complexing effect of Cl- to stabilize the dissolved metal ions (Zn 2+ , Pb 2+ ), and finally forms soluble complexes ([ZnCl4] 2- , [PbCl4] 2- ). This process proceeds efficiently under weak acidic and high Cl- concentration conditions, gently dissolving zinc oxide and lead oxide; the filtrate 1 is subjected to cooling crystallization + cyclic precipitation to precipitate PbCl2; in the filtrate 2, the pH value is adjusted by concentrated ammonia water, and ammonia water will react with Zn 2+ , Al 3+ , Fe 2+ and Mg 2+ to form precipitates; in the filtrate 3, by adjusting the pH of the filtrate 3 to 2-3, the [Zn(NH3)4] 2+ complex is destroyed, releasing , adding K2S to react to separate out ZnS precipitate; in the filter residue 2, PbCO3 solid is formed by potassium carbonate; then the filter residue 4 and the filter residue 5 are respectively calcined to obtain ZnO product and PbO product; calcium carbonate and potassium-nitrogen mixture are separated after mixing the filtrate 5 and the filtrate 4. The whole step can achieve the effect of synergistically recovering iron, lead, zinc, and potassium in the sintering dust removal ash of vanadium-titanium magnetite.

[0022] The present invention provides a process technology for simultaneously recovering iron, potassium, lead, and zinc from sintering dedusting ash of vanadium-titanium magnetite, obtaining four industrial by-products, namely iron concentrate (59.01%), zinc oxide (97.2%), lead oxide (99.85%), and potassium-nitrogen compound fertilizer (KCl 98.6%, NH4Cl 99.3%), all of which meet national standards. This turns hazardous waste into solid waste, maximally comprehensively recovers valuable elements, and realizes closed-loop recycling of wastewater.

[0023] The present invention simultaneously considers the mutual interference of multiple elements and the collaborative coupling of multiple processes, and integrally integrates multi-process and full-process recovery and disposal technologies. Basically, no excessive additional chemical raw materials are introduced, but the existing elements in the dedusting ash are fully utilized.

[0024] The present invention adopts the method of first recovering lead and zinc and then recovering potassium. On the one hand, it fully utilizes the potassium and chlorine elements in the sintering dedusting ash as leaching agents for leaching lead and zinc in the ammonium salt complexation step, avoiding excessive addition of additional leaching agents. On the other hand, when finally extracting potassium salts, almost all impurity elements are removed, making the preparation process of potassium-nitrogen compound fertilizer very simple and with high grade. In contrast, existing process technologies all recover potassium salts first and then extract other valuable metals, making the potassium salt recovery process very complex, with low grade, and affecting the extraction of subsequent valuable elements.

[0025] The process technology proposed by the present invention simultaneously adopts dry magnetic separation, high-voltage pulsed magnetic field, and three-stage magnetic separation for iron selection, greatly reducing water consumption, with a relatively high grade of iron concentrate and significantly reduced iron selection cost. In contrast, existing process technologies mostly use wet magnetic separation and ordinary magnetic fields, resulting in high water consumption and high grinding cost.

[0026] The process technology proposed by the present invention adopts the ammonium salt complexation - fractional crystallization method for lead extraction, avoiding high-temperature / strong acid or strong base operations, greatly protecting the environment and extending the service life of equipment. In contrast, existing process technologies mostly involve pyrometallurgical (high-temperature) or strong acid or strong base operations.

[0027] The process technology proposed by the present invention, by adopting carbonization to remove calcium technology, unexpectedly obtains industrial-grade calcium carbonate by-product without initially considering specifically recovering calcium elements, improving economic benefits, and this process is very environmentally friendly. Existing process technologies have not been found to extract calcium from sintering dedusting ash.

[0028] The process technology proposed by the present invention, by adopting ammonium salt complexation technology, unexpectedly obtains potassium-nitrogen compound fertilizer by-product without initially considering preparing nitrogen fertilizer, providing a new variety of compound fertilizer for the market and significantly improving the economy of this project. Existing process technologies have not been found to prepare potassium-nitrogen compound fertilizer from sintering dedusting ash.

[0029] The process technology proposed by the present invention has a relatively high purity of the demineralized water finally obtained, and all of it can be recycled to the two processes of "ammonium salt complexation" and "stepwise crystallization for lead extraction", achieving zero wastewater discharge. However, the secondary by-products of the existing process technology are relatively serious, and it is difficult to solve.

[0030] Therefore, the process technology of "iron-priority dry separation - mild synergistic leaching of lead, zinc, and potassium" of the present invention has remarkable creativity in terms of the synergistic coupling of the overall process, the parameter configuration of local processes, the ingenious selection of leaching materials, etc., and has great industrialization prospects. BRIEF DESCRIPTION OF THE DRAWINGS

[0031] In order to more clearly illustrate the technical solutions of the exemplary embodiments of the present invention, the drawings required for use in the embodiments will be briefly introduced below. It should be understood that the following drawings only show some embodiments of the present invention, and thus should not be regarded as limiting the scope. For those of ordinary skill in the art, other related drawings can be obtained based on these drawings without creative efforts. In the drawings:

[0032] Figure 1 It is a graph showing the relationship between the grinding fineness of the dust removal ash of raw material powder No. 1 in Experimental Example 1 and the iron recovery rate;

[0033] Figure 2 It is a graph showing the relationship between the grinding fineness of the dust removal ash of raw material powder No. 2 in Experimental Example 1 and the iron recovery rate;

[0034] Figure 3 It is a flow chart of high-voltage pulsed magnetic field dry magnetic separation for raw material powder No. 1 in Experimental Example 1;

[0035] Figure 4 It is a graph showing the relationship between the ammonium chloride concentration and the lead leaching rate in Experimental Example 2;

[0036] Figure 5 It is a graph showing the relationship between the potassium chloride concentration and the lead leaching rate in Experimental Example 2;

[0037] Figure 6 It is a graph showing the relationship between the temperature and the lead leaching rate in Experimental Example 2;

[0038] Figure 7 It is a graph showing the relationship between the reaction time and the lead leaching rate in Experimental Example 2;

[0039] Figure 8 It is a graph showing the relationship between the solid-liquid ratio and the lead leaching rate in Experimental Example 2;

[0040] Figure 9 It is a graph showing the relationship between the volume ratio of the mother liquor to water and the lead chloride precipitation rate in Experimental Example 3. DETAILED DESCRIPTION OF THE EMBODIMENTS

[0041] To make the objectives, technical solutions, and advantages of the present invention more clearly understood, the present invention will be further described in detail below with reference to embodiments and the accompanying drawings. The illustrative embodiments and descriptions thereof of the present invention are only used to explain the present invention and are not intended to limit the present invention. For those conditions not specified in the embodiments, they are carried out according to conventional conditions or the conditions recommended by the manufacturer. For reagents or instruments without indicating the manufacturer, they are all conventional products that can be obtained through commercial purchase.

[0042] A recovery process for vanadium-titanium magnetite slag comprises the following steps:

[0043] S1. Ball-mill the raw material of vanadium-titanium magnetite sintering dust to obtain raw material powder.

[0044] In some embodiments of the present invention, the average particle size of the above raw material is 0.13 mm to 0.17 mm.

[0045] In some embodiments of the present invention, the above ball-milling is dry ball-milling.

[0046] In some embodiments of the present invention, the average particle size of the above raw material powder is 300 to 450 mesh. The raw material powder in this range can adapt to subsequent magnetic separation for iron removal and accelerate the dry metallurgy kinetics process, without using overly fine raw material powder (such as 500 mesh - 600 mesh). Overly fine powder will increase the working load of the ball mill and consume too much energy, while overly coarse particle size (such as 200 mesh - 300 mesh) will seriously reduce the iron recovery rate of magnetic separation for iron removal and inhibit the chemical reaction kinetics.

[0047] S2. Separate gangue minerals, weakly magnetic iron minerals, medium magnetic iron minerals, and strongly magnetic iron minerals from the raw material powder by high-pressure pulsed magnetic field dry magnetic separation. The residue other than the magnetic iron minerals is the dry tailings.

[0048] In some embodiments of the present invention, the above high-pressure pulsed magnetic field dry magnetic separation specifically performs three-stage magnetic separation on the raw material powder: first, magnetic separation is carried out at a magnetic field of 0.3 - 0.7 T to obtain gangue minerals and strongly magnetic minerals, then magnetic separation is carried out at a magnetic field of 0.9 - 1.3 T to obtain medium magnetic minerals, and finally magnetic separation is carried out at a magnetic field of 1.5 - 1.7 T to obtain weakly magnetic minerals. The residue other than the magnetic iron minerals is the dry tailings. This type of hierarchical magnetic separation method can achieve an iron recovery effect of more than 93.89%, while the experimental data of the traditional wet magnetic separation iron recovery rate is usually <80%, which is significantly lower than this method. The wet magnetic separation requires a large amount of water resources, and the sewage treatment is difficult and the quantity is huge.

[0049] S3. Mix dry tailings, NH4Cl, and KCl, add water to obtain a mixed solution, stir and react, then filter and separate to obtain filtrate 1 and residue 1; cool the filtrate and perform two-stage cooling precipitation, then filter and separate to obtain filtrate 2 and residues. Combine the residues precipitated in the two stages to obtain residue 2; adjust the pH value of filtrate 2 to 9 - 10, then filter and separate to obtain filtrate 3 and residue 3; adjust the pH value of filtrate 3 to 2 - 3, add saturated K2S solution, stir, and then filter and separate to obtain filtrate 4 and residue 4; add saturated potassium carbonate solution to residue 2, stir, and then filter and separate to obtain filtrate 5 and residue 5; roast residue 4 to obtain roasted product 1; roast residue 5 to obtain roasted product 2; mix filtrate 5 and filtrate 4, introduce CO2, stop introducing CO2 when the pH value reaches 7 - 8, then filter and separate to obtain residue 6 and filtrate 6; perform conventional three-stage cyclic vacuum concentration - normal temperature cooling crystallization on filtrate 6, filter, dry, and pulverize to obtain a potassium-nitrogen mixture.

[0050] In the whole process of S3, there are actually only two types of waste by-products, namely residue 1 and residue 3. Residue 1 contains 6 types of compounds (elements), namely SiO2, magnesium chloride, titanium dioxide, vanadium pentoxide, C, and S. Residue 3 contains 3 types of compounds, namely Al(OH)3↓, Fe(OH)3↓, and Mg(OH)2↓. Currently, neither of these two solid residues has recycling value. Filtrate 1 contains zinc chloride, lead chloride, potassium chloride, calcium chloride, aluminum chloride, ferrous chloride, magnesium chloride, and NH4Cl.

[0051] In some embodiments of the present invention, the molar concentration of KCl in the above-mentioned mixed solution is 4 - 8 mol / L, the molar concentration of NH4Cl is 2 - 2.4 mol / L, the solid-liquid ratio is 1:(1.5 - 2.5), and the pH value is 4 - 5. The inventor of the present invention has studied and found that when the molar concentration of the NH4Cl solution is controlled at 2 - 2.4 mol / L, and the NH4Cl solution is at 70 - 80 °C, the lead leaching rate at this time reaches more than 98.5%. If the concentration of NH4Cl is further increased, the increase in the lead leaching rate is not significant; when the molar concentration of the KCl solution is 4 - 8 mol / L, the lead leaching rate reaches more than 95%. If the concentration of potassium chloride is further increased, the increase in the lead leaching rate is not significant; the inventor has found that the lead leaching rate is the best when the solid-liquid ratio is 1:(1.5 - 2.5). As the solid-liquid ratio increases, the lead leaching rate first increases and then decreases. When the solid-liquid ratio is 1:2, the lead leaching rate is the highest at 97.6%. The essence of the influence of the solid-liquid ratio is due to the change in the amount of potassium chloride solution, which causes changes in the ammonium chloride concentration (equivalent to the amount of ammonium chloride) and the total amount of chloride ions in the leaching system (equivalent to the potassium chloride concentration). Therefore, the influence of the solid-liquid ratio on the lead leaching rate is the result of the comprehensive effect of the amount of ammonium chloride and the potassium chloride concentration on the lead leaching rate. And because the complexation of ammonium salts with zinc is easier than with lead, as long as the lead leaching rate meets the requirements, then the zinc leaching rate will also meet the requirements.

[0052] In some embodiments of the present invention, the stirring of the above-mentioned mixed solution is specifically carried out by stirring and reacting at 70-80 °C for 60-120 min. When the leaching temperature is 70-80 °C, the leaching rate is the best. It has been found through research that when the temperature is 76 °C, the leaching rate reaches more than 95.5%; thereafter, as the leaching temperature rises, the leaching rate of lead decreases instead. The inventor of the present invention has found that this is because the temperature leaching causes serious evaporation of water in the leaching system, and potassium chloride and lead chloride are also precipitated simultaneously, resulting in a decrease in the leaching rate. In addition, although heating can accelerate the reaction rate, it may promote the volatilization of NH3, leading to an increase in the pH of the solution and the formation of precipitation, so the temperature should not be too high. The inventor of the present invention has found that when the leaching time is 30 min, the leaching rate of lead is more than 94%, but after the leaching time exceeds 90 min, as the time prolongs, the increase amplitude of the leaching rate of lead is not large; when the leaching time exceeds 120 min, as the time prolongs, the leaching rate of lead decreases, which is because continuous heating causes the evaporation of water in the leaching system, resulting in the simultaneous crystallization and precipitation of sodium chloride and lead chloride, causing a decrease in the leaching rate of lead.

[0053] In some embodiments of the present invention, the above-mentioned filtrate 1 is cooled to 15-35 °C and then subjected to two cooling precipitations. The first precipitation time is 80-100 min, and the second precipitation time is 110-130 min. The precipitation rate of lead chloride obtained by cooling and precipitating the filtrate is 99.5%. Continuing to extend the time, the precipitation rate of lead chloride no longer increases.

[0054] Preferably, after the first precipitation, water is added for dilution, and the volume ratio of the mother liquor after precipitation to water is 1:(1.4-1.8).

[0055] In some embodiments of the present invention, the pH value of the above-mentioned filtrate 2 is specifically adjusted by concentrated ammonia water, and after adjusting the pH value, stirring is continued for 20-40 min. Filtration and separation are carried out to generate a filtrate 3 containing [Zn(NH3)4] 2+ , Ca 2+ , K + , NH4 + , Cl - ions, and a filter residue 3 containing Al(OH)3↓, Fe(OH)3↓, Mg(OH)2↓.

[0056] In some embodiments of the present invention, the molar amount ratio of K2S in the saturated K2S solution added to the above-mentioned filtrate 3 to the molar amount of Zn 2+ is 1:(0.9-1.1). After adding the saturated K2S solution, stirring is continued for 20-40 min. First, by adjusting the pH value of the solution, the [Zn(NH3)4] 2+ complex is destroyed to release , and then K2S is added to react to generate a product containing Ca 2+ , K+ , NH4 + , Cl - filtrate 4 containing NH4 and Cl ions, and residue 4 containing ZnS precipitate.

[0057] In some embodiments of the present invention, after adding saturated potassium carbonate solution to the above-mentioned residue 2, stir for 10 - 30 min; the molar ratio of saturated potassium carbonate to lead chloride in residue 2 is 1:(0.9 - 1.1). Lead chloride reacts with potassium carbonate to form a precipitate, obtaining PbCO3↓.

[0058] In some embodiments of the present invention, the above-mentioned residue 4 is calcined at 800 - 1000 °C for 1 - 2 h. In the temperature range of 800 - 1000 °C, ZnS can react fully with oxygen, making the reaction proceed more completely in the direction of forming ZnO, and a higher ZnO formation rate and better product quality can be obtained. At the same time, this temperature range can ensure that the reaction proceeds under reasonable kinetic conditions, with a moderate reaction rate, neither being too slow due to too low temperature nor causing problems such as excessive energy consumption, serious equipment loss, and possible generation of other complex side reactions due to too high temperature.

[0059] In some embodiments of the present invention, the above-mentioned residue 5 is calcined at 660 - 720 °C for 2 - 3 h. Lead carbonate decomposes under heating conditions to form lead oxide and carbon dioxide, and the reaction equation is The reason for not directly calcining PbCl2 solid but first converting it to PbCO3 solid and then calcining is that calcining PbCl2 will generate toxic gas Cl2↑ or HCl vapor, which will corrode the equipment and pollute the environment. In the temperature range of 660 - 720 °C, lead carbonate can decompose more fully and stably, with a higher conversion rate of forming lead oxide. At the same time, it can avoid further reaction of lead oxide or other side reactions due to too high temperature, and also reduce energy consumption, lower costs, and equipment requirements, etc. The purity of the obtained PbO is 99.85%, meeting the requirements of the HG / T2325 - 2004 industrial first-class product index standard.

[0060] In some embodiments of the present invention, the above-mentioned filtrate 5 and filtrate 4 are mixed, CO2 is introduced, and the pH value is adjusted to 7 - 8 to obtain filtrate 6 and residue 6. Residue 6 is CaCO3. After drying, filtrate 6 obtains a (KCl + NH4Cl) mixed crystal, which can be directly sold as a potassium-nitrogen compound fertilizer for crops.

[0061] Preferably, the above-mentioned three-stage cyclic vacuum concentration - normal temperature cooling crystallization is specifically through a three-stage vacuum concentration process of a preheating section, a concentration section, and a flash evaporation section, and then a circulation mechanism is adopted in which some materials that do not reach the concentration are returned to the previous section through a circulation pump for further treatment. Finally, normal temperature cooling crystallization is carried out, and the final crystallization slurry is separated by centrifugation or filtration. The mother liquor can be returned to the concentration section for recycling.

[0062] Example 1

[0063] S1. Ball-mill 100 g of vanadium-titanium magnetite sintering dust removal ash raw materials with an average particle size of 0.13 mm to 0.17 mm to obtain raw material powder with an average particle size of 300 mesh.

[0064] S2. Perform three-stage magnetic separation on the raw material powder in a high-voltage pulsed magnetic field: first, perform magnetic separation at a magnetic field of 0.3 T to obtain gangue minerals and strongly magnetic minerals, then perform magnetic separation at a magnetic field of 0.9 T to obtain medium magnetic minerals, and finally perform magnetic separation at a magnetic field of 1.5 T to obtain weakly magnetic minerals. The remaining materials other than magnetic minerals are dry tailings.

[0065] S3. Mix the dry tailings, NH4Cl, and KCl, add water to obtain a mixed solution. The molar concentration of KCl in the mixed solution is 4 mol / L, the molar concentration of NH4Cl is 2 mol / L, the solid-liquid ratio is 1:1.5, the pH value is 4, stir and react at 70 °C for 60 min, filter and separate to obtain filtrate 1 and filter residue 1;

[0066] Cool filtrate 1 to 15 °C and then perform two-stage cooling precipitation. The first precipitation time is 80 min. After precipitation, add water to the mother liquor, and the volume ratio of the mother liquor to water is 1:1.4. The second precipitation time is 110 min. After secondary precipitation, filter and separate to obtain filtrate 2 and filter residue. Combine the filter residues precipitated twice to obtain filter residue 2;

[0067] Adjust the pH value of filtrate 2 to 9 with concentrated ammonia water, continue to stir for 20 min, filter and separate to obtain filtrate 3 and filter residue 3;

[0068] Adjust the pH of filtrate 3 to 2, then add a saturated K2S solution. The molar amount ratio of K2S to Zn 2+ is 1:0.9. After stirring for 20 min, filter and separate to obtain filtrate 4 and filter residue 4;

[0069] Add a saturated potassium carbonate solution to filter residue 2. The molar amount ratio of the saturated potassium carbonate to the molar amount of lead chloride in filter residue 2 is 1:0.9. After stirring for 10 min, filter and separate to obtain filtrate 5 and filter residue 5;

[0070] Roast filter residue 4 at 800 °C for 1 h to obtain roasted product 1;

[0071] Roast filter residue 5 at 660 °C for 2 h to obtain roasted product 2;

[0072] Mix filtrate 5 and filtrate 4, introduce CO2, adjust the pH value to 7 - 8, then filter and separate to obtain filter residue 6 and filtrate 6;

[0073] Perform three-stage cyclic vacuum concentration - normal temperature cooling crystallization on filtrate 6, filter, dry, and pulverize to obtain a potassium-nitrogen mixture.

[0074] Example 2

[0075] S1. Ball-mill 100 g of vanadium-titanium magnetite sintering dust removal ash raw materials with an average particle size of 0.13 mm to 0.17 mm to obtain raw material powder with an average particle size of 400 mesh.

[0076] S2. Perform three-stage magnetic separation on the raw material powder in a high-pressure pulsed magnetic field: adopt a three-stage dry magnetic separation process (rough selection: 0.5 T → medium selection: 1.1 T → fine selection: 1.6 T). After magnetic separation, iron concentrate and vanadium-titanium slag rich in potassium, lead, and zinc are obtained. The grade of the iron concentrate is TFe = 59.01%, and the recovery rate is 93.89%.

[0077] S3. Mix the dry tailings, 118 g of NH4Cl, and 412 g of KCl, add 1 L of water to obtain a mixed solution. The molar concentration of KCl in the mixed solution is 6 mol / L, the molar concentration of NH4Cl is 2.2 mol / L, the solid-liquid ratio is 1:2, the pH value is 4.5, stir and react at 75 °C for 90 min, filter and separate to obtain filtrate 1 and filter residue 1. The leaching rate of Zn is measured to be 98.9%, and the leaching rate of Pb is measured to be 97.7%.

[0078] Cool filtrate 1 to 25 °C and then perform two cooling precipitations. The first precipitation time is 90 min. After precipitation, add water to the mother liquor, and the volume ratio of the mother liquor to water is 1:1.6. The second precipitation time is 120 min. After the second precipitation, filter and separate to obtain filtrate 2 and filter residue. Combine the filter residues precipitated twice to obtain filter residue 2. The precipitation rate of Pb in filter residue 2 is measured to be 96.3%.

[0079] Adjust the pH value of filtrate 2 to 9.5 with concentrated ammonia water, continue to stir for 30 min, and filter and separate to obtain filtrate 3 and filter residue 3.

[0080] Adjust the pH of filtrate 3 to 2.5, and then add a saturated K2S solution containing 4.6 g of K2S. The molar ratio of K2S to Zn 2+ is 1:1. After stirring for 30 min, filter and separate to obtain filtrate 4 and filter residue 4. The precipitation rate of Zn is measured to be 98.6%.

[0081] Add 50 ml of a saturated potassium carbonate solution containing 3.7 g of K2CO3 to filter residue 2. The molar ratio of the saturated potassium carbonate to the lead chloride in filter residue 2 is 1:1. After stirring for 20 min, filter and separate to obtain filtrate 5 and filter residue 5.

[0082] Roast filter residue 4 at 900 °C for 1.5 h to obtain roasted product 1. Roasted product 1 is a white ZnO product. After detection, the purity is 97.2%.

[0083] The filter residue 5 was calcined at 690 °C for 2.5 h to obtain a calcined product 2. The calcined product 2 was a light yellow PbO product, and after testing, the purity was 99.85%. The ZnO product and the PbO product were directly sold as industrial products after being dried and pulverized.

[0084] The filtrate 5 and the filtrate 4 were mixed, and CO2 was introduced. When the pH value was adjusted to 7.5, filtration separation was carried out to obtain a filter residue 6 and a filtrate 6; the filter residue 6 was CaCO3. After the filtrate 6 was subjected to three-stage cyclic vacuum concentration - normal temperature cooling crystallization, filtration, drying, and pulverization, a (KCl + NH4Cl) crystal mixture with a KCl purity of 98.6% and an NH4Cl purity of 99.3% could be obtained, which was directly sold as a potassium-nitrogen compound fertilizer for crops. The remaining was demineralized water. After testing, the purity of the solid CaCO3 residue was 92%. Due to the high purity of this solid residue, it was directly sold as an industrial CaCO3 product after being dried and pulverized.

[0085] The leaching toxicity analysis of the filter residue 1 and the filter residue 3 was carried out respectively in accordance with the "Identification Standard for Leaching Toxicity of Hazardous Wastes" (GB5085.3 - 2007). The test results showed that the zinc and lead in the mixed residue of the filter residue 1 and the filter residue 3 fully met the standards, which were 6.2 mg / L and 0.46 mg / L respectively. According to the relevant regulations of the "National List of Hazardous Wastes (2025 Edition)", this filter residue no longer belongs to hazardous waste but belongs to general solid waste and can be directly stockpiled or landfilled.

[0086] In the above experiment, the total iron content in iron concentrate was determined by the stannous chloride reduction titration method. This method is applicable to the determination of iron content with a mass fraction of more than 20% in natural ore, iron concentrate, sinter and pellet. The sample is decomposed by acid or alkali fusion. The iron in the test solution is reduced by stannous chloride, and mercuric chloride is added to oxidize the excessive stannous chloride. Sodium diphenylamine sulfonate is used as the indicator, and the test solution is titrated with a potassium dichromate standard solution. Operating procedure: Weigh 0.15 - 0.50 g of the sample. Place the weighed sample in a 250 mL beaker, add 30 mL of hydrochloric acid (1 + 1), cover with a watch glass, and heat at low temperature until decomposed (if the sample is difficult to dissolve, a little ammonium fluoride can be added). Avoid boiling during heating. Remove, and wash the watch glass and the beaker wall with water. Filter through a medium-speed filter paper into a 300 mL beaker, and wipe the beaker wall with a glass rod. Wash the beaker 3 times with hot water and wash the residue about 5 times. The filtrate and washing solution are reserved as the main solution. Place the filter paper and the residue in a corundum crucible and incinerate. Burn at about 900 °C for 15 min and cool. Add 3 g of mixed flux (sodium peroxide + sodium carbonate = 2 + 1), and mix well. Melt at 800 °C for about 10 min and cool. Wash into the main solution with hot hydrochloric acid (1 + 2), and heat and evaporate the test solution to a volume of about 30 mL. Avoid boiling during heating. Heat the test solution to near boiling, and while stirring, dropwise add stannous chloride solution (60 g / L) until the yellow color of the test solution disappears, and then add 1 - 2 more drops. Cool to room temperature with running water. Immediately add 5 mL of saturated mercuric chloride solution, and mix well. Let stand for 3 min. Dilute the test solution with water to about 150 mL, add 15 mL of sulfuric acid - phosphoric acid mixed acid, add 6 drops of sodium diphenylamine sulfonate solution, and immediately titrate with a potassium dichromate standard solution until a stable purple color appears. Calculate the total iron content according to the following formula, expressed as a mass fraction:

[0087]

[0088] In the formula: c - the concentration of the potassium dichromate standard solution [c(1 / 6K2CrO7)], mol / L; V1 - the volume of the potassium dichromate standard solution required for titrating the blank, mL; V2 - the volume of the potassium dichromate standard solution required for titrating the test solution, mL; m - the mass of the weighed sample, g; 55.85 - the molar mass of iron, g / mol.

[0089] In the above experiment, the measurement of lead and zinc elements was determined by atomic absorption spectrometry. Specifically, a Zolix AA - 5000 atomic absorption spectrophotometer was used for the determination. This method has the advantages of wide application range and convenient operation.

[0090] The instrument conditions are as follows:

[0091]

[0092] First, prepare a series of standard solutions of lead and zinc elements according to the following table. After the instrument is stable, inject samples successively, and draw a standard curve based on the concentration and absorbance.

[0093]

[0094] Pb element - 216.999

[0095] Linear correlation coefficient: 0.999494

[0096] First - order curve: y = 235.3796x + 3.8303

[0097] Zn element - 202.548

[0098] Linear correlation coefficient: 0.999967

[0099] First - order curve: y = 6354.6431x - 36.8441

[0100] To ensure the rigor and precision of the experiment, for the treatment of the sample, it is crushed using a self - sealing bag hammer. After being completely crushed, it is screened through a 60 - mesh sieve, and only the undersize is retained for subsequent pretreatment steps.

[0101] Weigh 0.5 g of the sample (accurate to 0.0001), place it in a polytetrafluoroethylene beaker, moisten it with a small amount of water, and successively add 4 mL of hydrochloric acid and 4 mL of nitric acid. Cover the beaker and heat it at 110 °C for 1 hour. Then add 4 mL of hydrofluoric acid and 2 mL of perchloric acid, heat it at 110 °C for 1 hour and at 130 °C for 1 hour. Raise the temperature of the hot plate to about 200 °C and heat for 2 hours. Take it off and let it cool. Adjust the temperature of the hot plate to 170 °C, uncover the beaker, add 2 mL of hydrofluoric acid, and continue to heat it on the hot plate, shaking the beaker from time to time until the fluidity of the solution is poor. Take it off and let it cool. Add 1 mL of aqua regia and 2 mL of hydrochloric acid to heat and redissolve the salts, make the volume up to 50 mL, shake well, filter (due to black precipitate), and wait for measurement. After comparing with the standard curve, the measurement result can be obtained.

[0102] The following is the same test method.

[0103] Example 3

[0104] S1. Ball - mill 100 g of vanadium - titanium magnetite sintering dust removal ash raw materials with an average particle size of 0.13 mm - 0.17 mm to obtain raw material powder with an average particle size of 400 mesh.

[0105] S2. Conduct two - stage magnetic separation on the raw material powder in a high - voltage pulsed magnetic field: Utilize a high - voltage pulsed magnetic field (1.1 - 1.6 T), adopt a two - stage dry magnetic separation process (rougher: 0.5 T → cleaner: 1.6 T). After magnetic separation, iron concentrate and vanadium - titanium slag rich in potassium, lead, and zinc are obtained. The grade of the iron concentrate is TFe = 58.53%, and the recovery rate is 90.36%.

[0106] S3. Mix dry tailings, 118 g of NH4Cl and 375 g of KCl, add 1 L of water to obtain a mixed solution. In the mixed solution, the molar concentration of KCl is 5.5 mol / L, the molar concentration of NH4Cl is 2.2 mol / L, the solid-liquid ratio is 1:2, the pH value is 4.5. Stir and react at 75 °C for 45 min, then filter and separate to obtain filtrate 1 and residue 1. The leaching rate of Zn is measured to be 95.6%, and the leaching rate of Pb is measured to be 96.3%.

[0107] Cool filtrate 1 to 25 °C and then perform two cooling precipitations. The first precipitation time is 90 min. After precipitation, add water to the mother liquor, and the volume ratio of the mother liquor to water is 1:1.6. The second precipitation time is 120 min. After the second precipitation, filter and separate to obtain filtrate 2 and residue. Combine the residues precipitated twice to obtain residue 2. The precipitation rate of Pb is measured to be 96.1%.

[0108] Adjust the pH value of filtrate 2 to 9.5 with concentrated ammonia water, continue to stir for 30 min, and then filter and separate to obtain filtrate 3 and residue 3.

[0109] Adjust the pH of filtrate 3 to 2.5, and then add a saturated K2S solution containing 4.6 g of K2S. The molar ratio of K2S to Zn 2+ is 1:1. After stirring for 30 min, filter and separate to obtain filtrate 4 and residue 4. The precipitation rate of Zn is measured to be 98.4%.

[0110] Add 50 ml of a saturated potassium carbonate solution containing 3.7 g of K2CO3 to residue 2. The molar ratio of the saturated potassium carbonate to lead chloride in residue 2 is 1:1. After stirring for 20 min, filter and separate to obtain filtrate 5 and residue 5.

[0111] Roast residue 4 at 900 °C for 1.5 h to obtain roasted product 1. Roasted product 1 is a white ZnO product. After detection, the purity is 97.2%.

[0112] Roast residue 5 at 690 °C for 2.5 h to obtain roasted product 2. Roasted product 2 is a pale yellow PbO product. After detection, the purity is 95.31%. The ZnO product is directly sold as an industrial product after drying and pulverization. The purity of the PbO product is insufficient and it is not suitable to be sold as an industrial product.

[0113] Mix filtrate 5 and filtrate 4, introduce CO2, adjust the pH value to 7.5, then filter and separate to obtain filter residue 6 and filtrate 6; subject filtrate 6 to three-stage cyclic vacuum concentration - normal temperature cooling crystallization, filter, dry and crush it to obtain a (KCl + NH4Cl) crystal mixture with a KCl purity of 98.6% and an NH4Cl purity of 99.3%, which is directly sold as a potassium-nitrogen compound fertilizer for crops. The remaining is demineralized water. Filter residue 6 is CaCO3. After detection, the purity of the solid CaCO3 residue is 92%. Since the purity of this solid residue is relatively high, it is directly sold as an industrial CaCO3 product after drying and crushing.

[0114] The above-mentioned demineralized water is returned to the two processes of "ammonium salt complexation" and "lead extraction by fractional crystallization" for recycling.

[0115] The leaching toxicity analysis of filter residue 1 and filter residue 3 was carried out respectively according to the "Identification Standard for Leaching Toxicity of Hazardous Wastes" (GB5085.3 - 2007). The test results show that the zinc and lead in the mixed residue of filter residue 1 and filter residue 3 are completely up to standard, which are 6.2 mg / L and 0.46 mg / L respectively. According to the relevant regulations of the "National List of Hazardous Wastes (2025 Edition)", this filter residue no longer belongs to hazardous waste but to general solid waste and can be directly stockpiled or landfilled.

[0116] Example 4

[0117] S1. Grind 100 g of vanadium-titanium magnetite sintering dust removal ash raw materials with an average particle size of 0.13 mm - 0.17 mm to obtain raw material powder with an average particle size of 400 mesh.

[0118] S2. Perform three-stage magnetic separation on the raw material powder in a high-voltage pulsed magnetic field: use a common magnetic field (0.5 - 1.0 T) and adopt a three-stage dry magnetic separation process (rougher: 0.5 → middler: 0.8 T → finisher: 1.0 T). After magnetic separation, iron concentrate and vanadium-titanium slag rich in potassium, lead and zinc are obtained. The grade of the iron concentrate is TFe = 53.26% and the recovery rate is 87.4%.

[0119] S3. Mix the dry tailings, 118 g of NH4Cl and 412 g of KCl, add 1 L of water to obtain a mixed solution. The molar concentration of KCl in the mixed solution is 6 mol / L, the molar concentration of NH4Cl is 2.2 mol / L, the solid-liquid ratio is 1:3, the pH value is 4.5, stir and react at 50 °C for 90 min, filter and separate to obtain filtrate 1 and filter residue 1. The leaching rate of Zn is measured to be 92.1% and the leaching rate of Pb is measured to be 90.4%.

[0120] The filtrate 1 was cooled to 25°C and then cooled and precipitated twice. The first precipitation time was 90 minutes. After precipitation, water was added to the mother liquor. The volume ratio of the mother liquor to water was 1:1.3. The second precipitation time was 120 minutes. After the second precipitation, the filtrate 2 and the filter residue were separated by filtration. The filter residues from the two precipitations were combined to obtain the filter residue 2. The precipitation rate of Pb in the filter residue 2 was measured to be 92.5%.

[0121] The pH value of the filtrate 2 was adjusted to 9.5 by concentrated ammonia water, stirring was continued for 30 min, and the filtrate 3 and the residue 3 were separated by filtration.

[0122] The pH of filtrate 3 was adjusted to 2.5, and then a saturated solution of K2S was added. The molar amount of K2S was equal to that of Zn 2+ The molar ratio was 1:1, and after stirring for 30 minutes, the filtrate 4 and the residue 4 were separated by filtration. The precipitation rate of Zn was measured to be 93.2%.

[0123] Add 50 ml of a saturated potassium carbonate solution containing 3.7 g of K2CO3 to the residue 2, the molar ratio of the saturated potassium carbonate to the molar ratio of the lead chloride in the residue 2 being 1:1. After stirring for 20 minutes, filter and separate the filtrate 5 and the residue 5.

[0124] The filter residue 4 was calcined at 900°C for 1.3 hours to obtain a calcined product 1. The calcined product 1 was a white ZnO product with a purity of 95.5%. The purity of the ZnO product was insufficient and could barely be sold as an industrial product.

[0125] The filter residue 5 was calcined at 690°C for 2.5 hours to obtain a calcined product 2. The calcined product 2 was a light yellow PbO product, and the purity was 95.31% after testing. The purity of the PbO product was insufficient and it was not suitable to be sold as an industrial product.

[0126] Filtrate 5 and filtrate 4 are mixed, CO2 is introduced, and when the pH value is adjusted to 7.5, the residue 6 and filtrate 6 are obtained by filtration and separation; filtrate 6 is subjected to three-stage circulation vacuum concentration-normal temperature cooling crystallization, filtration, drying, and crushing to obtain a (KCl+NH4Cl) crystal mixture with a KCl purity of 98.6% and an NH4Cl purity of 99.3%, which is directly sold as a potassium-nitrogen compound fertilizer for crops. The rest is desalted water. The residue 6 is CaCO3, and after testing, the purity of the CaCO3 solid residue is 92%. Due to the high purity of the solid residue, it is directly sold as a CaCO3 industrial product after drying and crushing.

[0127] The leaching toxicity analysis of Filter Residue 1 and Filter Residue 3 was carried out in accordance with the "Identification Standard for Leaching Toxicity of Hazardous Wastes" (GB5085.3 - 2007). The test results showed that the zinc in Filter Residue 1 did not meet the standard at all, and the lead basically met the standard, being 21.9 mg / L and 384.2 mg / L respectively. The zinc and lead in Filter Residue 3 completely met the standard. According to the relevant regulations of the "National List of Hazardous Wastes (2025 Edition)", Filter Residue 1 belongs to hazardous waste, while Filter Residue 3 does not belong to hazardous waste but is ordinary solid waste, which can be directly stockpiled or landfilled, but should be closely observed.

[0128] Example 5

[0129] S1. 100 g of vanadium-titanium magnetite sintering dust removal ash raw materials with an average particle size of 0.13 mm - 0.17 mm were ball-milled to obtain raw material powder with an average particle size of 400 mesh.

[0130] S2. The raw material powder was subjected to three-stage magnetic separation in a high-pressure pulsed magnetic field: Using a high-pressure pulsed magnetic field (1.1 - 1.6 T), a three-stage dry magnetic separation process was adopted (rougher separation: 0.5 T → medium separation: 1.1 T → fine separation: 1.6 T). After magnetic separation, iron concentrate and vanadium-titanium slag rich in potassium, lead, and zinc were obtained. The grade of the iron concentrate was TFe = 59.01%, and the recovery rate was 93.89%.

[0131] S3. The dry tailings, 96.3 g of NH4Cl, and 412 g of KCl were mixed, and 1 L of water was added to obtain a mixed solution. The molar concentration of KCl in the mixed solution was 6 mol / L, the molar concentration of NH4Cl was 1.8 mol / L, the solid-liquid ratio was 1:2, the pH value was 3.5, and it was stirred and reacted at 90 °C for 90 min, followed by filtration and separation to obtain Filtrate 1 and Filter Residue 1. The leaching rate of Zn was measured to be 90.3%, and the leaching rate of Pb was measured to be 88.7%.

[0132] After cooling Filtrate 1 to 25 °C, it was subjected to two-stage cooling precipitation. The first precipitation time was 60 min. After precipitation, water was added to the mother liquor, and the volume ratio of the mother liquor to water was 1:1.6. The second precipitation time was 90 min. After the second precipitation, filtration and separation were carried out to obtain Filtrate 2 and Filter Residue. The filter residues precipitated in the two stages were combined to obtain Filter Residue 2. The precipitation rate of Pb in Filter Residue 2 was measured to be 92.1%.

[0133] The pH value of Filtrate 2 was adjusted to 9.5 with concentrated ammonia water, and it was continuously stirred for 30 min, followed by filtration and separation to obtain Filtrate 3 and Filter Residue 3.

[0134] The pH of Filtrate 3 was adjusted to 4.5, and then a saturated K2S solution containing 4.6 g of K2S was added. The molar ratio of K2S to Zn 2+ was 1:1. After stirring for 30 min, filtration and separation were carried out to obtain Filtrate 4 and Filter Residue 4. The precipitation rate of Zn was measured to be 93.5%.

[0135] Add 50 ml of saturated potassium carbonate solution containing 3.7 g of K2CO3 to the filter residue 2. The molar ratio of saturated potassium carbonate to lead chloride in the filter residue 2 is 1:1. After stirring for 20 min, filter to separate the filtrate 5 and the filter residue 5.

[0136] Calcine the filter residue 4 at 700 °C for 1.5 h to obtain the calcined product 1. The calcined product 1 is a white ZnO product. After testing, the purity is 95.7%.

[0137] Calcine the filter residue 5 at 600 °C for 2.5 h to obtain the calcined product 2. The calcined product 2 is a pale yellow PbO product. After testing, the purity is 95.56%. The ZnO product and the PbO product are directly sold as industrial products after drying and pulverizing.

[0138] Mix the filtrate 5 and the filtrate 4, introduce CO2, and adjust the pH value to 7.5, then filter to separate, obtaining the filter residue 6 and the filtrate 6; perform three-stage cyclic vacuum concentration - normal temperature cooling crystallization on the filtrate 6, filter, dry, and pulverize, then a (KCl + NH4Cl) crystal mixture with a KCl purity of 98.6% and an NH4Cl purity of 99.3% can be obtained, which is directly sold as a potassium-nitrogen compound fertilizer for crops. The remaining is demineralized water. The filter residue 6 is calcium carbonate. After testing, the purity of the CaCO3 solid residue is 92%. Since the purity of this solid residue is relatively high, it is directly sold as an industrial product of CaCO3 after drying and pulverizing.

[0139] The leaching toxicity analysis of the filter residue 1 and the filter residue 3 was carried out respectively according to the "Identification Standard for Leaching Toxicity of Hazardous Wastes" (GB5085.3 - 2007). The test results show that the zinc in the filter residue 1 fully meets the standard, being 25.72 mg / L, while the lead completely fails to meet the standard, reaching 433.8 mg / L. The zinc and lead in the filter residue 3 fully meet the standard. According to the relevant regulations of the "National List of Hazardous Wastes (2025 Edition)", this filter residue 1 belongs to hazardous waste, while the filter residue 3 belongs to general solid waste and can be directly stockpiled or landfilled.

[0140] Example 6

[0141] S1. Grind 100 g of vanadium-titanium magnetite sintering dust removal ash raw materials with an average particle size of 0.13 mm - 0.17 mm to obtain raw material powder with an average particle size of 400 mesh.

[0142] S2. Perform three-stage magnetic separation on the raw material powder in a high-voltage pulsed magnetic field: adopt a three-stage dry magnetic separation process (rougher: 0.5 T → middler: 1.1 T → finisher: 1.6 T). After magnetic separation, iron concentrate and vanadium-titanium slag rich in potassium, lead, and zinc are obtained. The grade of the iron concentrate is TFe = 59.01%, and the recovery rate is 93.89%.

[0143] S3. Mix dry tailings, 118 g of NH4Cl, and 372.5 g of KCl, add 1 L of water to obtain a mixed solution. The molar concentration of KCl in the mixed solution is 5 mol / L, the molar concentration of NH4Cl is 2.2 mol / L, the solid-liquid ratio is 1:2, and the pH value is 2.5. Stir and react at 75 °C for 90 min, then filter and separate to obtain filtrate 1 and residue 1. The leaching rate of Zn is measured to be 92.1%, and the leaching rate of Pb is measured to be 89.3%.

[0144] Cool filtrate 1 to 25 °C and then perform two cooling precipitations. The first precipitation time is 90 min. After precipitation, add water to the mother liquor, and the volume ratio of the mother liquor to water is 1:1.6. The second precipitation time is 120 min. After the second precipitation, filter and separate to obtain filtrate 2 and residue. Combine the residues precipitated twice to obtain residue 2. The precipitation rate of Pb in residue 2 is measured to be 96.3%.

[0145] Adjust the pH value of filtrate 2 to 7.5 with concentrated ammonia water, continue to stir for 30 min, and then filter and separate to obtain filtrate 3 and residue 3.

[0146] Adjust the pH of filtrate 3 to 5.5, and then add a saturated K2S solution containing 4.6 g of K2S. The molar ratio of K2S to Zn 2+ is 1:1. After stirring for 30 min, filter and separate to obtain filtrate 4 and residue 4. The precipitation rate of Zn is measured to be 92.3%.

[0147] Add 50 ml of a saturated potassium carbonate solution containing 3.7 g of K2CO3 to residue 2. The molar ratio of the saturated potassium carbonate to lead chloride in residue 2 is 1:1. After stirring for 20 min, filter and separate to obtain filtrate 5 and residue 5.

[0148] Calcine residue 4 at 900 °C for 1.5 h to obtain calcined product 1. Calcined product 1 is a white ZnO product, and after testing, the purity is 95.4%.

[0149] Calcine residue 5 at 690 °C for 2.2 h to obtain calcined product 2. Calcined product 2 is a light yellow PbO product, and after testing, the purity is 96.32%. The ZnO product and the PbO product can be barely sold as industrial products after drying and pulverization.

[0150] Mix filtrate 5 and filtrate 4, introduce CO2, adjust the pH value to 7.5, then filter and separate to obtain filter residue 6 and filtrate 6; subject filtrate 6 to three-stage cyclic vacuum concentration - normal temperature cooling crystallization, filter, dry, and pulverize to obtain a (KCl + NH4Cl) crystal mixture with a KCl purity of 98.6% and an NH4Cl purity of 99.3%, which can be directly sold as a potassium-nitrogen compound fertilizer for crops. The remaining is demineralized water. Filter residue 6 is CaCO3, and after testing, the purity of the solid CaCO3 residue is 92%. Since the purity of this solid residue is relatively high, it can be directly sold as an industrial CaCO3 product after drying and pulverizing.

[0151] The leaching toxicity analysis of filter residue 1 and filter residue 3 was carried out respectively in accordance with the "Identification Standard for Leaching Toxicity of Hazardous Wastes" (GB5085.3 - 2007). The test results show that zinc in filter residue 1 fully meets the standard, being 26.9 mg / L, while lead completely fails to meet the standard, being 426.3 mg / L. Zinc and lead in filter residue 3 fully meet the standard. According to the relevant regulations of the "National List of Hazardous Wastes (2025 Edition)", filter residue 1 belongs to hazardous waste, and filter residue 3 does not belong to hazardous waste but belongs to general solid waste and can be directly stockpiled or landfilled.

[0152] Example 7

[0153] S1. Ball-mill 100 g of vanadium-titanium magnetite sintering dust removal ash raw materials with an average particle size of 0.13 mm - 0.17 mm to obtain raw material powder with an average particle size of 400 mesh.

[0154] S2. Conduct three-stage magnetic separation of the raw material powder in a high-voltage pulsed magnetic field: adopt a three-stage dry magnetic separation process (rougher: 0.5 T → middler: 1.1 T → finisher: 1.6 T). After magnetic separation, iron concentrate and vanadium-titanium slag rich in potassium, lead, and zinc are obtained. The grade of the iron concentrate is TFe = 59.01%, and the recovery rate is 93.89%.

[0155] S3. Mix the dry tailings, 118 g of NH4Cl, and 412 g of KCl, add 1 L of water to obtain a mixed solution. The molar concentration of KCl in the mixed solution is 6 mol / L, the molar concentration of NH4Cl is 2.2 mol / L, the solid-liquid ratio is 1:2, the pH value is 4.5, stir and react at 75 °C for 90 min, then filter and separate to obtain filtrate 1 and filter residue 1. The leaching rate of Zn is measured to be 98.9%, and the leaching rate of Pb is measured to be 97.7%.

[0156] Cool filtrate 1 to 25 °C and then perform two-stage cooling precipitation. The first precipitation time is 90 min. After precipitation, add water to the mother liquor, and the volume ratio of the mother liquor to water is 1:1.6. The second precipitation time is 120 min. After the second precipitation, filter and separate to obtain filtrate 2 and filter residue. Combine the filter residues precipitated twice to obtain filter residue 2. The precipitation rate of Pb in filter residue 2 is measured to be 96.3%.

[0157] Adjust the pH value of filtrate 2 to 9.5 with concentrated ammonia water, continue stirring for 30 min, and filter to separate filtrate 3 and residue 3.

[0158] Adjust the pH of filtrate 3 to 6.0, and then add a saturated K2S solution containing 4.6 g of K2S. The molar ratio of K2S to Zn is 1:1. After stirring for 30 min, filter to separate filtrate 4 and residue 4. The precipitation rate of Zn is measured to be 58.6%. 2+

[0159] Add 50 ml of a saturated potassium carbonate solution containing 3.7 g of K2CO3 to residue 2. The molar ratio of saturated potassium carbonate to lead chloride in residue 2 is 1:1. After stirring for 20 min, filter to separate filtrate 5 and residue 5.

[0160] Calcine residue 4 at 650 °C for 1.5 h to obtain calcined product 1. The calcined product 1 is a white ZnO product, and after detection, the purity is 84.2%.

[0161] Calcine residue 5 at 550 °C for 2.5 h to obtain calcined product 2. The calcined product 2 is a pale yellow PbO product, and after detection, the purity is 83.27%. The ZnO product and the PbO product cannot be sold as industrial products after being dried and pulverized.

[0162] Mix filtrate 5 and filtrate 4, introduce CO2, and when the pH value is adjusted to 7.5, filter to separate to obtain residue 6 and filtrate 6; subject filtrate 6 to three-stage cyclic vacuum concentration - normal temperature cooling crystallization, filter, dry, and pulverize to obtain a (KCl + NH4Cl) crystal mixture with a KCl purity of 98.6% and an NH4Cl purity of 99.3%, which can be directly sold as a potassium-nitrogen compound fertilizer for crops. The remainder is demineralized water. Residue 6 is CaCO3, and after detection, the purity of the CaCO3 solid residue is 92%. Since the purity of this solid residue is relatively high, it can be directly sold as an industrial CaCO3 product after being dried and pulverized.

[0163] The leaching toxicity analysis of residue 1 and residue 3 was carried out respectively in accordance with the "Identification Standard for Leaching Toxicity of Hazardous Wastes" (GB5085.3 - 2007). The test results show that the zinc in residue 1 does not meet the standard, being 117.6 mg / L, and the lead completely meets the standard, being 2.78 mg / L. The zinc and lead in the solid residue (No. 3) completely meet the standard. According to the relevant regulations of the "National List of Hazardous Wastes (2025 Edition)", residue 1 belongs to hazardous waste, and residue 3 does not belong to hazardous waste but belongs to general solid waste and can be directly stockpiled or landfilled.

[0164] Example 8

[0165] ​S1. Ball mill 100 g of vanadium-titanium magnetite sintering dust removal ash raw materials with an average particle size of 0.13 mm to 0.17 mm to obtain raw material powder with an average particle size of 400 mesh.

[0166] S2. Perform three-stage magnetic separation on the raw material powder in a high-voltage pulsed magnetic field: adopt a three-stage dry magnetic separation process (rougher separation: 0.5 T → medium separation: 1.1 T → cleaner separation: 1.6 T). After magnetic separation, iron concentrate and vanadium-titanium slag rich in potassium, lead, and zinc are obtained. The grade of the iron concentrate is TFe = 59.01%, and the recovery rate is 93.89%.

[0167] S3. Mix the dry tailings, 118 g of NH4Cl, and 412 g of KCl, add 1 L of water to obtain a mixed solution. The molar concentration of KCl in the mixed solution is 6 mol / L, the molar concentration of NH4Cl is 2.2 mol / L, the solid-liquid ratio is 1:2, the pH value is 4.5, stir and react at 75 °C for 90 min, filter and separate to obtain filtrate 1 and filter residue 1. The leaching rate of Zn is measured to be 98.9%, and the leaching rate of Pb is measured to be 97.7%.

[0168] Cool filtrate 1 to 25 °C and then perform two-stage cooling precipitation. The first precipitation time is 90 min. After precipitation, add water to the mother liquor, and the volume ratio of the mother liquor to water is 1:1.6. The second precipitation time is 120 min. After the second precipitation, filter and separate to obtain filtrate 2 and filter residue. Combine the filter residues precipitated twice to obtain filter residue 2. The precipitation rate of Pb in filter residue 2 is measured to be 96.3%.

[0169] Adjust the pH value of filtrate 2 to 9.5 with concentrated ammonia water, continue to stir for 30 min, and filter and separate to obtain filtrate 3 and filter residue 3.

[0170] Adjust the pH of filtrate 3 to 2.5, and then add a saturated K2S solution containing 4.6 g of K2S. The molar ratio of K2S to the molar amount of Zn 2+ is 1:1. After stirring for 30 min, filter and separate to obtain filtrate 4 and filter residue 4. The precipitation rate of Zn is measured to be 98.6%.

[0171] Add 50 ml of saturated potassium carbonate solution containing 3.7 g of K2CO3 to filter residue 2. The molar ratio of saturated potassium carbonate to the molar amount of lead chloride in filter residue 2 is 1:1. After stirring for 20 min, filter and separate to obtain filtrate 5 and filter residue 5.

[0172] Roast filter residue 4 at 900 °C for 1.5 h to obtain roasted product 1. Roasted product 1 is a white ZnO product. After detection, the purity is 97.2%.

[0173] The filter residue 5 was calcined at 690 °C for 2.5 h to obtain a calcined product 2. The calcined product 2 was a light yellow PbO product, and after testing, its purity was 99.85%. The ZnO product and the PbO product were directly sold as industrial products after being dried and pulverized.

[0174] The filtrate 5 and the filtrate 4 were mixed, and CO2 was introduced. When the pH value was adjusted to 9, filtration separation was carried out to obtain a filter residue 6 and a filtrate 6; the filtrate 6 was subjected to two-stage cyclic vacuum concentration - normal temperature cooling crystallization, filtration, drying, and pulverization, and then a (KCl + NH4Cl) crystal mixture with a KCl purity of 96.2% and an NH4Cl purity of 96.9% could be obtained, which was directly sold as a potassium-nitrogen compound fertilizer for crops. The remaining was demineralized water. The filter residue 6 was CaCO3, and after testing, the purity of the CaCO3 solid residue was 86%. Since the purity of this solid residue was not high, it was not suitable to be sold as an industrial product of CaCO3 after being dried and pulverized.

[0175] The leaching toxicity analysis of the filter residue 1 and the filter residue 3 was carried out respectively according to the "Identification Standard for Leaching Toxicity of Hazardous Wastes" (GB5085.3 - 2007). The test results showed that the zinc and lead in the mixed residue of the filter residue 1 and the filter residue 3 fully met the standards, which were 6.2 mg / L and 0.46 mg / L respectively. According to the relevant regulations of the "National List of Hazardous Wastes (2025 Edition)", this filter residue no longer belongs to hazardous waste but belongs to general solid waste and can be directly stockpiled or landfilled.

[0176] Example 9

[0177] S1. 100 g of vanadium-titanium magnetite sintering dust removal ash raw materials with an average particle size of 0.13 mm - 0.17 mm were ball-milled to obtain raw material powder with an average particle size of 400 mesh.

[0178] S2. The raw material powder was subjected to three-stage magnetic separation in a high-voltage pulsed magnetic field: a three-stage dry magnetic separation process (rough selection: 0.5 T → medium selection: 1.1 T → fine selection: 1.6 T) was adopted. After magnetic separation, iron concentrate and vanadium-titanium slag rich in potassium, lead, and zinc were obtained. The grade of the iron concentrate was TFe = 59.01%, and the recovery rate was 93.89%.

[0179] S3. The dry tailings, 59 g of NH4Cl, and 206 g of KCl were mixed, and 1 L of water was added to obtain a mixed solution. The molar concentration of KCl in the mixed solution was 3 mol / L, the molar concentration of NH4Cl was 1.1 mol / L, the solid-liquid ratio was 1:2, the pH value was 4.5, and the mixture was stirred and reacted at 45 °C for 90 min, followed by filtration separation to obtain a filtrate 1 and a filter residue 1. The leaching rate of Zn was measured to be 44.2%, and the leaching rate of Pb was measured to be 37.5%.

[0180] After cooling the filtrate 1 to 25°C, it was subjected to two cooling and precipitation steps. The first precipitation time was 90 min. After precipitation, water was added to the mother liquor, and the volume ratio of the mother liquor to water was 1:1.6. The second precipitation time was 120 min. After the second precipitation, filtration was carried out to separate the filtrate 2 and the filter residue. The filter residues obtained from the two precipitations were combined to obtain the filter residue 2. The precipitation rate of Pb in the filter residue 2 was measured to be 92.6%.

[0181] The pH value of the filtrate 2 was adjusted to 9.5 with concentrated ammonia water, and stirring was continued for 30 min. Filtration was carried out to separate the filtrate 3 and the filter residue 3.

[0182] The pH of the filtrate 3 was adjusted to 2.5, and then a saturated K2S solution containing 4.6 g of K2S was added. The molar ratio of K2S to Zn 2+ was 1:1. After stirring for 30 min, filtration was carried out to separate the filtrate 4 and the filter residue 4. The precipitation rate of Zn was measured to be 96.5%.

[0183] 50 ml of a saturated potassium carbonate solution containing 3.7 g of K2CO3 was added to the filter residue 2. The molar ratio of the saturated potassium carbonate to lead chloride in the filter residue 2 was 1:1. After stirring for 20 min, filtration was carried out to separate the filtrate 5 and the filter residue 5.

[0184] The filter residue 4 was calcined at 900°C for 1.5 h to obtain the calcined product 1. The calcined product 1 was a white ZnO product, and after detection, the purity was 97.2%.

[0185] The filter residue 5 was calcined at 690°C for 2.5 h to obtain the calcined product 2. The calcined product 2 was a light yellow PbO product, and after detection, the purity was 99.85%. The ZnO product and the PbO product were directly sold as industrial products after drying and pulverization.

[0186] The filtrate 5 and the filtrate 4 were mixed, and CO2 was introduced. When the pH value was adjusted to 7.5, filtration was carried out to obtain the filter residue 6 and the filtrate 6; the filtrate 6 was subjected to three-stage cyclic vacuum concentration - normal temperature cooling crystallization, filtration, drying, and pulverization, and then a (KCl + NH4Cl) crystal mixture with a KCl purity of 87.6% and an NH4Cl purity of 89.1% could be obtained, which was directly sold as a potassium-nitrogen compound fertilizer for crops. The remaining was demineralized water. The filter residue 6 was CaCO3, and after detection, the purity of the CaCO3 solid residue was 92%. Since the purity of this solid residue was relatively high, it was directly sold as an industrial CaCO3 product after drying and pulverization.

[0187] The leaching toxicity analysis was carried out on Filter Residue 1 and Filter Residue 3 respectively in accordance with the "Identification Standard for Leaching Toxicity of Hazardous Wastes" (GB5085.3 - 2007). The test results showed that zinc and lead in Filter Residue 1 did not meet the standards, being 132.8 mg / L and 1238.6 mg / L respectively, belonging to hazardous wastes. Zinc and lead in Filter Residue 3 fully met the standards, being 6.2 mg / L and 0.46 mg / L respectively. According to the relevant regulations of the "National List of Hazardous Wastes (2025 Edition)", this filter residue no longer belongs to hazardous waste but ordinary solid waste and can be directly stockpiled or landfilled.

[0188] Example 10

[0189] S1. The vanadium-titanium magnetite sintering dust removal ash raw material with an average particle size of 0.13 mm - 0.17 mm was ball-milled to obtain raw material powder with an average particle size of 450 mesh.

[0190] S2. The raw material powder was subjected to three-stage magnetic separation in a high-voltage pulsed magnetic field: first, gangue minerals and strongly magnetic minerals were magnetically separated under a 0.7 T magnetic field, then medium magnetic minerals were magnetically separated under a 1.3 T magnetic field, and finally weakly magnetic minerals were magnetically separated under a 1.7 T magnetic field. The remaining material except magnetic iron minerals was dry tailings.

[0191] S3. Dry tailings, NH4Cl, and KCl were mixed and water was added to obtain a mixed solution. The molar concentration of KCl in the mixed solution was 8 mol / L, the molar concentration of NH4Cl was 2.4 mol / L, the solid-liquid ratio was 1:2.5, the pH value was 5, and it was stirred and reacted at 80 °C for 120 min, followed by filtration and separation to obtain Filtrate 1 and Filter Residue 1;

[0192] After cooling Filtrate 1 to 35 °C, it was subjected to two-stage cooling precipitation. The first precipitation time was 100 min. After precipitation, water was added to the mother liquor, and the volume ratio of the mother liquor to water was 1:1.8. The second precipitation time was 130 min. After the second precipitation, Filtrate 2 and filter residue were separated by filtration. The filter residues precipitated twice were combined to obtain Filter Residue 2;

[0193] The pH value of Filtrate 2 was adjusted to 10 with concentrated ammonia water, and stirring was continued for 40 min, followed by filtration and separation to obtain Filtrate 3 and Filter Residue 3; the pH of Filtrate 3 was adjusted to 3, and then a saturated K2S solution was added. The molar amount of K2S and Zn 2+The molar ratio is 1:1.1. After stirring for 40 min, filtrate 4 and residue 4 are separated by filtration; saturated potassium carbonate solution is added to residue 2, and the molar ratio of saturated potassium carbonate to lead chloride in residue 2 is 1:1.1. After stirring for 30 min, filtrate 5 and residue 5 are separated by filtration; residue 4 is calcined at 1000 °C for 2 h to obtain calcined product 1; residue 5 is calcined at 720 °C for 3 h to obtain calcined product 2; filtrate 5 and filtrate 4 are mixed, and CO2 is introduced. When the pH value is adjusted to 7-8, filtration separation is carried out to obtain residue 6 and filtrate 6; filtrate 6 is subjected to three-stage cyclic vacuum concentration - normal temperature cooling crystallization, and after filtration, drying and pulverization, a potassium-nitrogen mixture is obtained.

[0194] Experimental Example 1

[0195] The vanadium-titanium magnetite sintering dust removal ash raw material with an average particle size of 0.13 mm to 0.17 mm is subjected to dry ball milling to obtain raw material powder, and raw material powder No. 1 with a 400-mesh proportion of 30.12% and raw material powder No. 2 with a 400-mesh proportion of 38.13% are obtained respectively.

[0196] Then the raw material powder is subjected to dry high-pressure pulsed magnetic separation (refer to Figure 3 ). First, gangue minerals and strongly magnetic iron minerals are obtained by magnetic separation under a magnetic field of 0.5 T, then medium magnetic iron minerals are obtained by magnetic separation under a magnetic field of 1.1 T, and finally weakly magnetic iron minerals are obtained by magnetic separation under a magnetic field of 1.6 T. Figure 1 is the relationship diagram of the grinding fineness - iron recovery rate of the dust removal ash of raw material powder No. 1; Figure 2 is the relationship diagram of the grinding fineness - iron recovery rate of the dust removal ash of raw material powder No. 2.

[0197] Observation Figure 1-2 , and the results show that for dust removal ash No. 1, when the magnetic field strength is 150 KA / m, good concentrate indexes are obtained. At this time, the concentrate grade reaches 62.13% and the recovery rate is 50.75%. The situation of dust removal ash No. 2 is similar to that of No. 1. When the magnetic field strength is 200 KA / m, concentrate indexes with an iron grade of 58.26% and a recovery rate of 40.45% can be obtained. Therefore, the magnetic field strength of dust removal ash No. 1 is controlled at 150 KA / m, and the magnetic field strength of dust removal ash No. 2 is controlled at 200 KA / m. The grinding fineness of the two kinds of dust removal ash is investigated to find the best grinding fineness of the two kinds of dust removal ash.

[0198] The influence of grinding fineness on magnetic separation indexes was investigated for two kinds of dedusted ash under the most suitable magnetic field intensity conditions. It can be seen from the curves in the figure that the influence trends of the two kinds of dedusted ash are roughly the same. For No. 1 dedusted ash, when the grinding fineness reaches 57.65% passing through -0.038 mm, the iron grade of the concentrate that can be obtained under the optimal magnetic field intensity conditions is 65.23%, and the recovery rate is 48.72%. If the grinding fineness is continuously increased, the concentrate grade will instead decrease. This may be because when the grinding fineness is increased, some impurities adhere to the magnetic separation concentrate during magnetic separation, thus reducing the iron grade of the concentrate. The dedusted ash itself belongs to fine-grained dust, and the finer the particle size, the greater the difficulty of separation, and the more serious the phenomenon of concentrate inclusion. For No. 2 dedusted ash, in the range of 59.32% to 73.21% passing through 0.038 mm of grinding fineness, the concentrate grade increases from 63.15% to 63.67%, with only a 0.52% increase in grade, and the recovery rate decreases by 0.09%, but the required grinding power consumption increases greatly. Considering from the comprehensive economic perspective, choosing 59.32% passing through -0.038 mm is the suitable grinding fineness.

[0199] Therefore, the optimal grinding particle size in this experimental study is 400 mesh, that is, 0.038 mm.

[0200] Observe Figure 3 , through Figure 3 the experimental results of the flow chart, it can be known that for the sintering dedusted ash of vanadium-titanium magnetite in the weak magnetic - medium magnetic - strong magnetic test flow, through each optimal condition test, the indexes of the weak magnetic concentrate obtained are TFe = 65.23%, recovery rate 48.65%, the indexes of the medium magnetic concentrate obtained are TFe = 56.49%, recovery rate 38.01%, the indexes of the strong magnetic concentrate are TFe = 55.32%, recovery rate 47.23%, and the indexes of the comprehensive concentrate are TFe = 59.01%, recovery rate 93.89%. Obviously, the experimental data of the traditional wet magnetic separation iron recovery rate is usually <80%, significantly lower than 93.89% of this scheme. This shows that through the three-stage magnetic separation method of the present invention, the iron recovery rate can be maximally improved.

[0201] Experimental Example 2

[0202] (1). According to the steps of Example 2, adjust the NH4Cl concentration of S3 to 0.5 mol / L, 1 mol / L, 1.5 mol / L, 2 mol / L, and 2.5 mol / L respectively. Collect the filter residue 1 respectively and detect the lead leaching rate of the filter residue 1. The results are as Figure 4 shown.

[0203] Observe Figure 4, the present inventor found that when the concentration of NH4Cl in the NH4Cl solution is 2 mol / L at 70 - 80 °C, the lead leaching rate reaches 98%. When the concentration of NH4Cl is further increased, the increase in the lead leaching rate is not significant. Among them, magnesium chloride is partially dissolved. The NH4Cl solution does not chemically react with the solid phases of SiO2, magnesium chloride, titanium dioxide, vanadium pentoxide, C, and S.

[0204] (2) According to the steps of Example 2, adjust the concentration of KCl in S3 to 100 g / L, 200 g / L, 300 g / L, 400 g / L, 500 g / L, and 600 g / L respectively. Collect the filter residue 1 respectively and detect the lead leaching rate of the filter residue 1. The results are as Figure 5 shown.

[0205] Observe Figure 5 , when the concentration of KCl is 600 g / L, the lead leaching rate reaches over 95%. When the concentration of potassium chloride is further increased, the increase in the lead leaching rate is not significant.

[0206] (3) According to the steps of Example 2, adjust the temperature of the mixed solution to be between 50 - 100 °C respectively. Collect the filter residue 1 respectively and detect the lead leaching rate of the filter residue 1. The results are as Figure 6 shown.

[0207] Observe Figure 6 , the present inventor's research found that when the leaching temperature is 76 °C, the leaching rate reaches over 95.5%; thereafter, as the leaching temperature rises, the increase in the lead leaching rate is very small and then begins to decline. This is because due to the temperature leaching, the water evaporation in the leaching system is serious, and potassium chloride and lead chloride are also precipitated simultaneously, resulting in a decrease in the leaching rate. In addition, although heating can accelerate the reaction rate, it may promote the volatilization of NH3, leading to an increase in the pH of the solution and the formation of precipitation. Therefore, the temperature should not be too high.

[0208] (4) According to the steps of Example 2, adjust the stirring reaction time to 30 min, 60 min, 90 min, 120 min, and 150 min respectively. Collect the filter residue 1 respectively and detect the lead leaching rate of the filter residue 1. The results are as Figure 7 shown.

[0209] Observe Figure 7 , the present inventor's research found that when the leaching time is 30 min, the lead leaching rate is over 94%. However, when the leaching time exceeds 90 min, as the time prolongs, the increase in the lead leaching rate is not significant; when the leaching time exceeds 120 min, as the time prolongs, the lead leaching rate decreases. This is because continuous heating causes the water evaporation in the leaching system, resulting in the simultaneous crystallization and precipitation of sodium chloride and lead chloride, causing a decrease in the lead leaching rate.

[0210] (5). According to the steps of Example 2, adjust the solid-liquid ratio to 1:1, 1:1.5, 1:2, 1:2.5, 1:3, and 1:4 respectively. Collect the filter residue 1 separately and detect the lead leaching rate of the filter residue 1. The results are as Figure 8 shown.

[0211] Observe Figure 8 , and the inventor's research found that as the solid-liquid ratio increases, the lead leaching rate first increases and then decreases. When the solid-liquid ratio is 1:2, the lead leaching rate is the highest at 97.6%. The essence of the influence of the solid-liquid ratio is due to the change in the amount of potassium chloride solution, which causes changes in the ammonium chloride concentration (equivalent to the amount of ammonium chloride) and the total amount of chloride ions in the leaching system (equivalent to the potassium chloride concentration). Therefore, the influence of the solid-liquid ratio on the lead leaching rate is the result of the combined effect of the amount of ammonium chloride and the potassium chloride concentration on the lead leaching rate. It should be noted that since the complexation of ammonium salts with zinc is easier than with lead, as long as the lead leaching rate meets the requirements, the zinc leaching rate will also meet the requirements.

[0212] Experimental Example 3

[0213] According to the steps of Example 2, the volume ratios of the mother liquor to water are 1:1, 1:1.1, 1:1.2, 1:1.3, 1:1.4, and 1:1.5 respectively. Collect the filter residue 2 separately and detect the lead chloride precipitation rate of the filter residue 2. The results are as Figure 9 shown.

[0214] Observe Figure 9 , and the inventor's research found that at room temperature, the lead chloride precipitation rate increases with the increase of the dilution factor. When the dilution factor is 1.6, the lead chloride precipitation rate reaches 86.1%. At the same time, the larger the dilution factor, the larger the amount of mother liquor to be treated and recycled after crystallization. Therefore, the dilution factor should not be too large.

[0215] Experimental Example 4

[0216] (1). After mixing the filter residue 1 and filter residue 3 of Example 2, leaching toxicity analysis was carried out in accordance with the "Identification Standard for Leaching Toxicity of Hazardous Wastes" (GB5085.3-2007), and the test report is shown in Table 1. The results show that zinc and lead fully meet the standards, which are 6.2 mg / L and 0.46 mg / L respectively. According to the relevant regulations of the "National List of Hazardous Wastes (2025 Edition)", this solid residue no longer belongs to hazardous waste but belongs to general solid waste and can be directly stockpiled or landfilled.

[0217] Table 1

[0218]

[0219] The specific embodiments described above further elaborate on the purpose, technical solution, and beneficial effects of the present invention. It should be understood that the above description is only the specific embodiments of the present invention and is not used to limit the protection scope of the present invention. Any modifications, equivalent replacements, improvements, etc. made within the spirit and principle of the present invention shall be included within the protection scope of the present invention.

Claims

1. A recovery process for vanadium-titanium magnetite slag, characterized in that, It includes the following steps: Ball-mill the raw material of vanadium-titanium magnetite sintering dust removal ash to obtain raw material powder; Use high-pressure pulsed magnetic field dry magnetic separation to separate gangue minerals, weakly magnetic iron minerals, medium magnetic iron minerals and strongly magnetic iron minerals from the raw material powder. The residue other than magnetic iron minerals is the dry tailings; Mix the dry tailings, NH4Cl and KCl, add water to obtain a mixed solution, stir and react, filter and separate to obtain filtrate 1 and filter residue 1; cool the filtrate 1 and perform two cooling precipitations, filter and separate to obtain filtrate 2 and filter residue, combine the filter residues precipitated twice to obtain filter residue 2; adjust the pH value of filtrate 2 to 9-10, filter and separate to obtain filtrate 3 and filter residue 3; adjust the pH value of filtrate 3 to 2-3, then add saturated K2S solution, stir, filter and separate to obtain filtrate 4 and filter residue 4; add saturated potassium carbonate solution to filter residue 2, stir, filter and separate to obtain filtrate 5 and filter residue 5; roast filter residue 4 and filter residue 5 respectively to obtain roasted product 1 and roasted product 2; mix filtrate 5 and filtrate 4, introduce CO2, adjust the pH value to 7-8, filter and separate to obtain filter residue 6 and filtrate 6; perform three-stage cyclic vacuum concentration-ambient temperature cooling crystallization on filtrate 6, filter, dry and crush to obtain a potassium-nitrogen mixture.

2. The recycling process of a vanadium-titanium magnetite slag according to claim 1, wherein The average particle size of the raw material is 0.13mm - 0.17mm.

3. The recovery process of vanadium-titanium magnetite slag according to claim 1, characterized in that, The ball milling is dry ball milling; the average particle size of the raw material powder is 300 - 450 mesh.

4. The recovery process of vanadium-titanium magnetite slag according to claim 1, characterized in that, The high-pressure pulsed magnetic field dry magnetic separation is specifically to perform three-stage magnetic separation on the raw material powder: first, magnetically separate to obtain gangue minerals and strongly magnetic minerals under a magnetic field of 0.3 - 0.7T, then magnetically separate to obtain medium magnetic minerals under a magnetic field of 0.9 - 1.3T, and finally magnetically separate to obtain weakly magnetic minerals under a magnetic field of 1.5 - 1.7T. The residue other than magnetic iron minerals is the dry tailings.

5. The recovery process of vanadium-titanium magnetite slag according to claim 1, characterized in that, The molar concentration of KCl in the mixed solution is 4 - 8mol / L, the molar concentration of NH4Cl is 2 - 2.4mol / L, the solid-liquid ratio is 1:(1.5 - 2.5), and the pH value is 4 - 5; the stirring of the mixed solution is specifically to stir and react at 70 - 80°C for 60 - 120min.

6. The recycling process of vanadium-titanium magnetite slag according to claim 1, wherein Cool the filtrate 1 to 15 - 35°C and perform two cooling precipitations. The first precipitation time is 80 - 100min, and the second precipitation time is 110 - 130min.

7. The recycling process of a vanadium-titanium magnetite slag according to claim 1, characterized in that The adjustment of the pH value of filtrate 2 is specifically carried out by concentrated ammonia water, and continue to stir for 20 - 40min after adjusting the pH value.

8. The recovery process of vanadium-titanium magnetite slag according to claim 1, wherein, The molar amount ratio of K2S in the saturated K2S solution added to the filtrate 3 to Zn 2+ is 1:(0.9 - 1.1), and after adding the saturated K2S solution, stirring is continued for 20 - 40 min.

9. The recovery process of vanadium-titanium magnetite slag according to claim 1, characterized in that, After adding saturated potassium carbonate solution to filter residue 2, stir for 10 - 30min; the molar ratio of saturated potassium carbonate to the molar amount of lead chloride in filter residue 2 is 1:(0.9 - 1.1).

10. The recovery process of a vanadium-titanium magnetite slag according to claim 1, characterized in that, Roast filter residue 4 at 800 - 1000°C for 1 - 2h; roast filter residue 5 at 660 - 720°C for 2 - 3h.

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