Method for recovering iron concentrate from laterite nickel ore smelting slag

Abstract

The application provides a method for recovering iron concentrate from laterite nickel ore smelting slag, which comprises the following steps: subjecting the laterite nickel ore smelting slag to water immersion treatment to obtain water immersion residue; subjecting the water immersion residue to alkali leaching treatment to obtain leaching residue; uniformly mixing a biomass reducing agent and the leaching residue, and performing reduction roasting to obtain roasted material; and grinding and magnetically separating the roasted material to obtain iron concentrate. The method for recovering iron concentrate from laterite nickel ore smelting slag provided by the application effectively reduces the entry of aluminum, chromium and silicate impurities into the magnetite phase, successfully recovers high-quality iron concentrate products, and significantly improves the purity and magnetic properties thereof; meanwhile, the non-magnetic tailings obtained after magnetic separation can be further used for the production of building materials, cement additives or other industrial products, thereby realizing comprehensive resource utilization of the residue.

Description

Technical Field

[0001] This invention relates to the field of hydrometallurgical technology, and in particular to a method for recovering iron concentrate from laterite nickel ore smelting slag. Background Technology

[0002] Laterite nickel ore smelting mainly involves two processes: hydrometallurgy and pyrometallurgy, each producing different waste slag characteristics. Hydrometallurgical processes are conducted at temperatures ranging from room temperature to 250℃, producing acidic leaching slag with high sulfur content, fine particle size, and an iron content typically greater than 40%. Pyrometallurgy (RKEF process) is carried out at high temperatures of 1500-1600℃, producing high-temperature molten slag. After water quenching, the slag exhibits complex phases, requiring further enrichment of iron content. Existing main technologies for recovering iron concentrate include reduction roasting-magnetic separation, magnetizing roasting-weak magnetic separation, reduction smelting, dedicated recovery processes for pyrometallurgical tailings, and co-processing technologies.

[0003] Reduction roasting-magnetic separation is currently the most widely used technical route. Its core principle is to reduce weakly magnetic iron oxides to strongly magnetic iron minerals, which are then enriched by magnetic separation. A typical process involves smelting slag being treated with carbon and additives, followed by reduction roasting, grinding, and finally magnetic separation to obtain iron concentrate. Specific technical variations include microwave reduction roasting, traditional pulverized coal reduction, and biochar reduction. Microwave reduction roasting mixes wet leaching slag with calcium-containing additives and performs microwave reduction roasting at 900-1200℃, followed by grinding and magnetic separation to obtain high-grade iron concentrate. Traditional pulverized coal reduction uses pulverized coal as a reducing agent, roasting pellets at 600-1000℃ for 30-180 minutes, followed by magnetic separation to obtain nickel-iron concentrate. Biochar reduction uses biochar instead of coal or coke, preparing the reducing agent through nitrogen carbonization and CO2 activation to achieve efficient recovery of iron concentrate.

[0004] The magnetized roasting-weak magnetic separation process addresses the issue of high sulfur content in wet leaching residue by converting iron oxides into magnetite through magnetized roasting. Existing technologies can yield iron concentrate with an iron grade of 64%, sulfur content of 0.16%, and an iron recovery rate of 94%. For example, by using dodecylamine polyoxyethylene ether dispersant for magnetic separation, the iron grade of the concentrate reaches 69.26%, the sulfur content is reduced to 0.18%, and the recovery rate is 90.83%. A gas-based reduction process can yield a concentrate with a TFe content of 68.77%, sulfur content of 0.12%, and a recovery rate of 95.15%.

[0005] Compared with magnetization roasting, reduction smelting can produce higher-grade products. When smelting for 30 minutes at a carbon ratio of 0.85, basicity of 0.9, and 1550℃, the resulting alloy has an iron grade of 95.14% and a chromium grade of 2.92%, with an iron recovery rate of 94.41%. The tailings can be directly used in building materials, realizing the resource utilization of all components.

[0006] The pyrometallurgical tailings recovery process is designed for the characteristics of high-temperature molten slag (1400-1600℃) from the RKEF process. It employs centrifugal cooling and air classification technology, utilizing the residual heat of the molten slag to separate gangue fibers and iron-containing particles through centrifugal crushing. Then, dry strong magnetic pre-selection is carried out with a magnetic field strength of 10000-16000Gs to pre-discard the tailings. Finally, high-grade iron concentrate with a grade of ≥65% is obtained through suspended conical gravity separation and high-frequency fine screening.

[0007] The co-processing technology combines laterite nickel ore leaching slag with copper smelting slag to prepare iron concentrate. By optimizing the ratio and adding calcium-containing slag-forming agent, strong mixing and grinding are used to improve the fineness. Then, the concentrate is pelletized, reduced, roasted, water-quenched, and weakly magnetically separated to obtain high-grade iron concentrate with an iron content of 85-93%.

[0008] However, the existing processes generally suffer from the following shortcomings: environmental issues are one of the main drawbacks, with the use of fossil reducing agents such as coal or coke leading to high carbon emissions, and tailings stockpiling or landfill causing environmental pollution; resource waste is prominent, with water quenching wasting the waste heat of high-temperature slag, which accounts for more than 35% of the total energy consumption; in terms of product quality, the iron concentrate has a high sulfur content (0.12-0.18%), resulting in low recovery rate and low grade in direct magnetic separation; process defects include poor grindability of metallic iron, difficulty in individual dissociation, and complex phase composition of water-quenched slag, with uncontrollable mineral crystallization state; in terms of economy, the amount of reducing agent used is large, the cost is high, and the adaptability to raw materials is poor.

[0009] Therefore, how to provide a method for recovering laterite nickel ore smelting slag that can produce iron concentrate products with higher purity, stable magnetic properties, and that meet higher industrial standards is a technical problem that urgently needs to be solved. Summary of the Invention

[0010] To address the aforementioned technical problems, this invention provides a method for recovering iron concentrate from laterite nickel ore smelting slag. The method includes sequential water leaching, alkali leaching, reduction roasting, grinding, and magnetic separation. This effectively reduces the entry of aluminum, chromium, and silicate impurities into the magnetite phase, thereby improving product quality and stability. It successfully recovers high-quality iron concentrate (magnetite phase Fe3O4), significantly enhancing its purity and magnetic properties. Simultaneously, the non-magnetic tailings obtained after magnetic separation can be further used in the production of building materials, cement additives, or other industrial products, achieving comprehensive resource utilization of the residue. This integrated technical approach significantly improves the overall resource recovery efficiency of laterite nickel ore smelting slag, reduces the environmental burden of residue disposal, and provides a sustainable and economically feasible technical approach for the recycling of laterite nickel ore resources.

[0011] To achieve this objective, the present invention adopts the following technical solution:

[0012] In a first aspect, the present invention provides a method for recovering iron concentrate from laterite nickel ore smelting slag, the method comprising the following steps:

[0013] S1. The laterite nickel ore smelting slag is subjected to water leaching treatment to obtain water leaching residue;

[0014] S2. The water-leached residue from step S1 is subjected to alkaline leaching treatment to obtain leaching residue material;

[0015] S3. Mix the biomass reducing agent and the leaching residue from step S2 evenly, and then perform reduction roasting to obtain roasted material;

[0016] S4. Grind and magnetically separate the roasted material described in step S3 to obtain iron concentrate.

[0017] The method for recovering iron concentrate from laterite nickel ore smelting slag provided by the present invention includes sequential water leaching treatment, alkali leaching treatment, reduction roasting, grinding and magnetic separation. Water leaching pre-dissolves soluble salts and removes residual acid-soluble, water-soluble impurities, and some heavy metal ions, thus purifying the slag phase. Subsequently, alkaline leaching removes impurities such as aluminum and silicon and enriches the iron-containing phase. Enriching the iron-containing phase before reduction roasting not only improves product purity but also enhances the efficiency and stability of subsequent reduction and magnetic separation processes, laying a solid foundation for later steps. Reduction roasting is then performed using a biomass reducing agent. The high carbon activity and excellent reducing properties of the biomass reducing agent improve roasting and reduction efficiency, effectively reducing operating costs, carbon emissions, significantly increasing biomass utilization, and achieving efficient resource conversion. It also effectively promotes the efficient conversion of iron oxides into the magnetite phase. Next, grinding and magnetic separation are performed. Grinding induces mineral embrittlement, refines particles, and effectively inhibits magnetite particle agglomeration, significantly improving the degree of liberation between Fe3O4 and gangue minerals. This optimizes screening and magnetic separation effects, creating favorable conditions for the efficient recovery of high-grade magnetite. In summary, the present invention combines water leaching, alkali leaching, reduction roasting, grinding, and magnetic separation processes in a synergistic manner. This effectively removes impurities before reduction and precisely controls phase transformation during roasting, achieving efficient recovery of high-quality iron concentrate while also taking into account advantages such as low cost, high resource utilization, and environmental friendliness.

[0018] This invention features a simple and efficient process with excellent cost control and significant recovery results. It achieves the targeted conversion of leaching residue from high-pressure acid leaching (HPAL) of rich iron-rich laterite nickel ore into high-value-added magnetic materials, effectively solving the environmental problems caused by long-term stockpiling and landfilling of laterite nickel ore leaching residue. Not only does it have high resource utilization, but it is also environmentally friendly, possessing the potential for large-scale application and providing strong technical support for the sustainable development of laterite nickel ore resources. The resulting iron concentrate can be widely used in magnetic separation media, pigments, catalysts, wastewater treatment, and metallurgical raw materials. The non-magnetic tailings obtained after magnetic separation can be further used in the production of building materials, cement additives, or other industrial products, significantly improving the comprehensive resource recovery efficiency while achieving sustainable waste utilization.

[0019] As a preferred technical solution of the present invention, the median particle size D50 of the laterite nickel ore smelting slag in step S1 is 50μm-200μm, such as 50μm, 80μm, 100μm, 120μm, 150μm, 180μm or 200μm.

[0020] Preferably, based on 100wt% of the total mass of the laterite nickel ore smelting slag, the laterite nickel ore smelting slag includes 35wt%-45wt% iron, for example 35wt%, 36wt%, 37wt%, 38wt%, 39wt%, 40wt%, 41wt%, 42wt%, 43wt%, 44wt%, or 45wt%, etc.; 3wt%-6wt% aluminum, for example 3wt%, 3.5wt%, 4wt%, 4.5wt%, 5wt%, 5.5wt%, or 6wt%, etc.; and 2wt% sulfur. wt%-8wt%, for example 2wt%, 3wt%, 4wt%, 5wt%, 6wt%, 7wt% or 8wt%, etc.; silicon 2wt%-6wt%, for example 2wt%, 3wt%, 4wt%, 5wt% or 6wt%, etc.; cobalt 1wt%-3wt%, for example 1wt%, 1.5wt%, 2wt%, 2.5wt% or 3wt%, etc.; magnesium 0.1wt%-1wt%, for example 0.1wt%, 0.3wt%, 0.5wt%, 0.8wt% or 1wt%, etc.

[0021] As a preferred technical solution of the present invention, the solid-liquid ratio of the laterite nickel ore smelting slag and deionized water in step S1 is 1g:(3-5)mL, for example 1g:3mL, 1g:3.5mL, 1g:4mL, 1g:4.5mL or 1g:5mL, etc.

[0022] Preferably, the water immersion treatment temperature in step S1 is 50°C-90°C, such as 50°C, 55°C, 60°C, 65°C, 70°C, 75°C, 80°C, 85°C, or 90°C.

[0023] Preferably, the stirring speed of the water immersion treatment in step S1 is 200rpm-500rpm, such as 200rpm, 250rpm, 300rpm, 350rpm, 400rpm, 450rpm or 500rpm.

[0024] Preferably, the water immersion treatment time in step S1 is 4 min to 20 min, such as 4 min, 8 min, 10 min, 12 min, 15 min, 18 min, or 20 min.

[0025] Preferably, after the water immersion treatment in step S1 is completed, the process further includes solid-liquid separation and washing steps.

[0026] It should be noted that the present invention does not impose specific requirements or special limitations on solid-liquid separation and washing after water immersion treatment. Commonly used solid-liquid separation and washing methods in the art are applicable to the present invention, such as filtration, vacuum filtration or centrifugation. Those skilled in the art can make adaptive selections and adjustments according to actual conditions.

[0027] As a preferred technical solution of the present invention, the alkaline solution used in the alkaline leaching treatment in step S2 includes sodium hydroxide solution and / or sodium carbonate solution.

[0028] Preferably, the concentration of the alkaline solution used in the alkaline leaching treatment in step S2 is 0.5 mol / L-4 mol / L, such as 0.5 mol / L, 1 mol / L, 1.5 mol / L, 2 mol / L, 2.5 mol / L, 3 mol / L, 3.5 mol / L, or 4 mol / L.

[0029] Preferably, in step S2, the solid-liquid ratio of the water-leached residue and the alkaline solution in step S1 is 1g:(3-8)mL, for example, 1g:3mL, 1g:3.5mL, 1g:4mL, 1g:4.5mL, 1g:5mL, 1g:5.5mL, 1g:6mL, 1g:6.5mL, 1g:7mL, 1g:7.5mL, or 1g:8mL, etc.

[0030] Preferably, the temperature of the alkaline leaching treatment in step S2 is 50°C-90°C, such as 50°C, 55°C, 60°C, 65°C, 70°C, 75°C, 80°C, 85°C or 90°C.

[0031] Preferably, the stirring speed for the alkali leaching treatment in step S2 is 200 rpm to 500 rpm, such as 200 rpm, 250 rpm, 300 rpm, 350 rpm, 400 rpm, 450 rpm or 500 rpm.

[0032] Preferably, the alkaline leaching time in step S2 is 30 min to 140 min, for example, 30 min, 40 min, 50 min, 60 min, 70 min, 80 min, 90 min, 100 min, 110 min, 120 min, 130 min or 140 min.

[0033] Preferably, after the alkaline leaching treatment in step S2 is completed, the process further includes solid-liquid separation, washing, and drying steps to obtain leaching residue.

[0034] It should be noted that the present invention does not impose specific requirements or special limitations on the solid-liquid separation, washing and drying after the alkaline leaching treatment. Commonly used solid-liquid separation, washing and drying methods in the art are applicable to the present invention. For example, they can be filtration, vacuum filtration or centrifugation, as long as a completely dry leaching residue can be obtained. Those skilled in the art can make adaptive selections and adjustments according to actual conditions.

[0035] As a preferred technical solution of the present invention, after the alkaline leaching treatment in step S2, an acid leaching treatment is further included to obtain leaching residue.

[0036] The present invention performs acid leaching after alkaline leaching, which can further remove residual impurity phases and enrich iron oxides in the solid residue on the basis of alkaline leaching, resulting in iron concentrate products with higher purity.

[0037] Preferably, the acid solution used in the acid leaching treatment includes hydrochloric acid solution.

[0038] Preferably, the concentration of the acid solution used in the acid leaching treatment is 0.5 mol / L-1 mol / L, such as 0.5 mol / L, 0.6 mol / L, 0.7 mol / L, 0.8 mol / L, 0.9 mol / L, or 1 mol / L.

[0039] Preferably, the solid-liquid ratio of the alkali leaching residue obtained after the alkali leaching treatment to the acid solution is 1g:(3-8)mL, for example, 1g:3mL, 1g:3.5mL, 1g:4mL, 1g:4.5mL, 1g:5mL, 1g:5.5mL, 1g:6mL, 1g:6.5mL, 1g:7mL, 1g:7.5mL, or 1g:8mL, etc.

[0040] In this invention, the alkaline leaching residue obtained after alkaline leaching treatment is the alkaline leaching residue obtained after solid-liquid separation and washing following alkaline leaching treatment.

[0041] Preferably, the acid leaching temperature is 50°C-90°C, such as 50°C, 55°C, 60°C, 65°C, 70°C, 75°C, 80°C, 85°C, or 90°C.

[0042] Preferably, the stirring speed for the acid leaching treatment is 200 rpm to 500 rpm, such as 200 rpm, 250 rpm, 300 rpm, 350 rpm, 400 rpm, 450 rpm, or 500 rpm.

[0043] Preferably, the acid leaching time is 30 min to 140 min, such as 30 min, 40 min, 50 min, 60 min, 70 min, 80 min, 90 min, 100 min, 110 min, 120 min, 130 min or 140 min.

[0044] Preferably, after the acid leaching treatment is completed, the process further includes solid-liquid separation, washing, and drying steps to obtain leaching residue.

[0045] It should be noted that the present invention does not impose specific requirements or special limitations on the solid-liquid separation, washing and drying after the acid leaching treatment. Commonly used solid-liquid separation and washing methods in the art are applicable to the present invention, such as filtration, vacuum filtration or centrifugation, as long as a completely dry leaching residue can be obtained. Those skilled in the art can make adaptive selections and adjustments according to actual conditions.

[0046] As a preferred technical solution of the present invention, in step S3, the amount of biomass reducing agent added is 2wt%-20wt% of the mass of the leaching residue in step S2, for example, 2wt%, 3wt%, 4wt%, 5wt%, 6wt%, 7wt%, 8wt%, 9wt%, 10wt%, 11wt%, 12wt%, 13wt%, 14wt%, 15wt%, 16wt%, 17wt%, 18wt%, 19wt%, or 20wt%, etc.

[0047] This invention regulates the amount of biomass reducing agent added to 2wt%-20wt% of the mass of the leaching residue in step S2, and, combined with a reducing roasting atmosphere, can effectively transform the iron mineral morphology, further improving the iron recovery rate. If the amount of biomass reducing agent added is too small, the reduction reaction will be incomplete, and iron oxides (such as hematite and goethite) cannot be fully converted into the magnetic iron phase (magnetite phase or metallic iron), significantly reducing the magnetic separation recovery rate; at the same time, the reaction kinetics are limited, the required reduction time is prolonged, energy consumption increases, and the residual non-magnetic iron phase causes resource waste; if the amount of biomass reducing agent added is too large, it will lead to over-reduction, generating a large amount of metallic iron instead of the target product magnetite phase. The metallic iron particles are easily wrapped and adhered to the gangue, deteriorating the magnetic separation effect and reducing the iron concentrate grade; in addition, incomplete combustion of excessive reducing agent can easily produce carbon deposits, polluting iron products, and causing waste of biomass resources, increased production costs, and increased flue gas treatment load.

[0048] Furthermore, this invention mixes the biomass reducing agent and the leaching residue from step S2 and then performs reduction roasting. By combining the amount of biomass reducing agent added and the reduction roasting atmosphere, iron oxide can be efficiently converted into magnetite phase Fe3O4 without the need to add other materials to assist in reduction roasting, thereby avoiding impurity contamination, simplifying the process, reducing production costs, and reducing subsequent separation and purification steps.

[0049] Preferably, the preparation method of the biomass reducing agent includes: carbonizing the biomass material under an inert atmosphere, followed by activation under a carbon dioxide atmosphere.

[0050] Preferably, the biomass material includes at least one of coconut shells, wheat straw, corn stalks, bamboo, rice straw, peanut shells, or tree branches.

[0051] Preferably, the inert atmosphere includes a nitrogen atmosphere and / or an argon atmosphere.

[0052] Preferably, the carbonization holding temperature is 350°C-600°C, such as 350°C, 400°C, 450°C, 500°C, 550°C or 600°C.

[0053] Preferably, the carbonization holding time is 80min-160min, such as 80min, 100min, 120min, 140min or 160min.

[0054] Preferably, the activation holding temperature is 350°C-600°C, such as 350°C, 400°C, 450°C, 500°C, 550°C or 600°C.

[0055] Preferably, the activation holding time is 40-70 minutes, such as 40 minutes, 50 minutes, 60 minutes or 70 minutes.

[0056] As a preferred technical solution of the present invention, the mixing in step S3 includes a first ball milling.

[0057] Preferably, the first ball milling time is 5 min to 120 min, such as 5 min, 10 min, 20 min, 30 min, 40 min, 50 min, 60 min, 70 min, 80 min, 90 min, 100 min, 110 min or 120 min.

[0058] Preferably, the rotational speed of the first ball mill is 200rpm-500rpm, such as 200rpm, 250rpm, 300rpm, 350rpm, 400rpm, 450rpm or 500rpm.

[0059] As a preferred technical solution of the present invention, the gas used in the reduction roasting in step S3 includes an inert gas.

[0060] Preferably, the flow rate of the inert gas is 200 mL / min to 400 mL / min, such as 200 mL / min, 250 mL / min, 300 mL / min, 350 mL / min or 400 mL / min.

[0061] Preferably, the inert gas includes nitrogen and / or argon.

[0062] Preferably, the heating rate of the reduction calcination in step S3 is 3°C / min-6°C / min, such as 3°C / min, 3.5°C / min, 4°C / min, 4.5°C / min, 5°C / min, 5.5°C / min or 6°C / min.

[0063] Preferably, the holding temperature for reduction calcination in step S3 is 500°C-900°C, such as 500°C, 550°C, 600°C, 650°C, 700°C, 750°C, 800°C, 850°C, or 900°C.

[0064] Preferably, the holding time for reduction calcination in step S3 is 30 min to 180 min, such as 30 min, 40 min, 50 min, 60 min, 70 min, 80 min, 90 min, 100 min, 110 min, 120 min, 130 min, 140 min, 150 min, 160 min, 170 min, or 180 min.

[0065] As a preferred technical solution of the present invention, the grinding in step S4 includes a second ball milling.

[0066] Preferably, the second ball milling time is 5 min to 15 min, such as 5 min, 8 min, 10 min, 12 min or 15 min.

[0067] Preferably, the temperature of the second ball mill is 20°C-30°C, such as 20°C, 22°C, 25°C, 28°C or 30°C.

[0068] Preferably, the rotational speed of the second ball mill is 400rpm-800rpm, such as 400rpm, 450rpm, 500rpm, 550rpm, 600rpm, 650rpm, 700rpm, 750rpm or 800rpm.

[0069] Preferably, the ball-to-material ratio of the second ball mill is (10-20):1, such as 10:1, 12:1, 15:1, 18:1 or 20:1.

[0070] Furthermore, by adjusting the parameters of the second ball mill in the grinding process, this invention can achieve the effects of material embrittlement, particle size refinement, and agglomeration reduction without the need to add other materials, such as dispersants, to assist in grinding. This avoids the introduction of impurities and simplifies subsequent washing and purification steps, reducing production costs. At the same time, it eliminates the interference of dispersants on the surface properties of iron minerals, which is beneficial to maintaining the original activity of the iron phase and improving the magnetic separation effect.

[0071] Preferably, the magnetically separated powder is obtained by sieving after grinding, and the particle size of the magnetically separated powder is >150μm.

[0072] In this invention, the particle size of the magnetically separated powder is >150μm, which means that the powder after grinding is sieved to obtain powder with a particle size >150μm.

[0073] In this invention, controlling the particle size of the magnetically separated powder to >150μm can significantly reduce the specific surface area of ​​the powder, reduce surface oxidation and exposure of active sites, thereby inhibiting the re-oxidation of the iron phase during storage and transportation, resulting in higher product stability. Simultaneously, a larger particle size helps reduce dust dispersion, improve the working environment, and reduce material loss. If the particle size is small (≤150μm), the specific surface area of ​​the powder increases sharply, the surface energy rises, and it easily leads to particle agglomeration and caking, reducing magnetic separation efficiency. Furthermore, fine powder is more prone to surface oxidation to generate Fe2O3 or FeO(OH), reducing the magnetic iron content and increasing the risk of dust explosion, posing a threat to safe production. The pelletizing properties in subsequent pelletizing or briquetting processes will also deteriorate due to excessively fine particle size.

[0074] As a preferred technical solution of the present invention, the magnetic separation in step S4 includes: mixing deionized water and the magnetic separation powder evenly to obtain a magnetic separation slurry, and performing magnetic separation on the magnetic separation slurry.

[0075] Preferably, the magnetic field strength for magnetic separation in step S4 is 500Gs-3000Gs, such as 500Gs, 600Gs, 700Gs, 800Gs, 900Gs, 1000Gs, 1200Gs, 1500Gs, 1800Gs, 2000Gs, 2200Gs, 2500Gs, 2800Gs, or 3000Gs.

[0076] Preferably, the solid-liquid ratio of the magnetically separated powder to the deionized water is 1g:(10-20)mL, for example, 1g:10mL, 1g:12mL, 1g:15mL, 1g:18mL or 1g:20mL, etc.

[0077] It should be noted that the present invention does not impose specific requirements or special limitations on the mixing of deionized water and the magnetic separation powder. As long as the magnetic separation powder can be uniformly dispersed, it is acceptable. Those skilled in the art can make adaptive selections and adjustments according to actual conditions. For example, it can be stirred for 10 minutes at 30°C and 400rpm-500rpm.

[0078] Preferably, after the magnetic separation in step S4, the obtained magnetic material is subjected to solid-liquid separation and drying to obtain iron concentrate.

[0079] It should be noted that the present invention does not impose specific requirements or special limitations on the solid-liquid separation and drying of magnetic materials. Commonly used solid-liquid separation and drying methods in the art are applicable to the present invention, such as filtration, vacuum filtration or centrifugation, as long as the powder can be completely dried. Those skilled in the art can make adaptive selections and adjustments according to actual conditions.

[0080] Compared with the prior art, the present invention has at least the following beneficial effects:

[0081] The method for recovering iron concentrate from laterite nickel ore smelting slag provided by this invention includes sequential water leaching, alkaline leaching, reduction roasting, grinding, and magnetic separation. This effectively reduces the entry of aluminum, chromium, and silicate impurities into the magnetite phase, thereby improving product quality and stability. It successfully recovers high-quality iron concentrate (magnetite phase Fe3O4), significantly enhancing its purity and magnetic properties. Simultaneously, the non-magnetic tailings obtained after magnetic separation can be further used in the production of building materials, cement additives, or other industrial products, achieving comprehensive resource utilization of the residue. This integrated technical approach significantly improves the overall resource recovery efficiency of laterite nickel ore smelting slag, reduces the environmental burden of residue disposal, and provides a sustainable and economically feasible technical approach for the recycling of laterite nickel ore resources. Attached Figure Description

[0082] Figure 1 This is a flowchart of the method for recovering iron concentrate from laterite nickel ore smelting slag provided by the present invention. Detailed Implementation

[0083] The technical solution of the present invention will be further described below with reference to the accompanying drawings and specific embodiments. However, the following examples are merely simplified examples of the present invention and do not represent or limit the scope of protection of the present invention. The scope of protection of the present invention is determined by the claims.

[0084] The scope of this invention can be defined by lower and upper limits. The selected lower and upper limits define the boundaries of a specific range. The range defined in this way can be defined by the inclusion or exclusion of endpoints. Any endpoint can be independently selected for inclusion or exclusion, and all lower and upper limits can be arbitrarily combined to form new ranges. That is, any lower limit can be combined with any upper limit to form an effective range. For example, if the ranges of 60~120 and 80~110 are listed for specific parameters, it should be understood that the ranges of 60~110 and 80~120 also fall within the scope of this invention. In addition, if the minimum range values ​​1 and 2 are listed, and the maximum range values ​​3, 4 and 5 are also listed, then all ranges of 1~3, 1~4, 1~5, 2~3, 2~4 and 2~5 fall within the scope of this invention. In this invention, the numerical range "a~b" represents a shortened representation of any combination of real numbers between a and b, where a and b are both real numbers. For example, the numerical range "0~5" means that all real numbers between 0 and 5 have been fully listed in this document, and "0~5" is only a shortened representation of this set of numerical combinations. When a parameter is expressed as an integer ≥2, it is equivalent to listing positive integers that meet the requirements, such as 2, 3, 4, 5, 6, 7, 8, 9, 10, etc. When a parameter is expressed as an integer selected from "2~10", it is equivalent to listing any integer among 2, 3, 4, 5, 6, 7, 8, 9, and 10.

[0085] In this invention, "a combination of at least two" refers to a quantity greater than or equal to 2 unless otherwise specified. For example, "any one or a combination of at least two" means that any one of the listed items can be selected, or a combination of at least two of the listed items formed in a manner that does not conflict and enables the implementation of this invention. In this invention, unless otherwise specified, the features or solutions corresponding to "and / or" cover any one of two or more related listed items, as well as any and all combinations of the related listed items. The arbitrary and all combinations include any two related listed items, any more related listed items, or a combination of all related listed items. For example, "A and / or B" means a set consisting of A, B, and combinations of A and B, where "containing A and / or B" can be understood, depending on the context of the statement, as containing A, containing B, or simultaneously containing both A and B. In this invention, "optional" means that the corresponding feature, component, step or solution is not necessary, that is, it is selected from either "with" or "without". If there are multiple "optional" limitations in a technical solution, unless otherwise specified and there is no technical conflict or mutual constraint, each "optional" limitation is independent and does not affect the others.

[0086] In this invention, technical features or solutions described using open-ended terms such as "comprising" or "including" do not exclude additional non-conflicting elements beyond the listed elements unless otherwise specified. They are considered to disclose both closed-ended features or solutions consisting solely of the listed elements and open-ended features or solutions that may include additional non-conflicting elements beyond the listed elements. For example, if A includes a1, a2, and a3, unless otherwise specified, this means that A can consist only of a1, a2, and a3, or it can include other non-conflicting elements based on a1, a2, and a3. This corresponds to the disclosure of technical solutions such as "A consists of a1, a2, and a3," "A is selected from a1, a2, and a3," and "A not only includes a1, a2, and a3, but may also include other non-conflicting elements." All embodiments and optional embodiments of this invention, unless otherwise specified and without technical conflict, can be combined to form new technical solutions, and such combinations fall within the scope of this invention. The term "embodiment" as used in this invention means that a specific feature, structure, or characteristic described in connection with an embodiment may be included in at least one embodiment or implementation of the invention. The appearance of this phrase in various locations throughout the specification does not necessarily refer to the same embodiment, nor is it an independent or alternative embodiment mutually exclusive with other embodiments. Those skilled in the art will understand, explicitly and implicitly, that the embodiments described in this invention can be combined with other embodiments that do not conflict with the technology. The ordinal numbers "first," "second," "third," and "fourth," etc., used in the expressions "first aspect," "second aspect," "third aspect," and "fourth aspect" in this invention are for descriptive purposes only and should not be construed as indicating or implying relative importance or quantity, nor should they be construed as implicitly specifying the importance or quantity of the indicated technical features. They serve only as a non-exhaustive enumeration and do not constitute a closed limitation on quantity.

[0087] In this invention, the order in which the steps are written in the methods described in each embodiment does not imply a strict execution order. The actual execution order of each step should be determined based on its function and possible internal logic. Unless otherwise specified, all steps of this invention can be executed in the order they are written, or in any order without technical conflict. For example, if the method includes steps (a) and (b), it means that the method may include steps (a) and (b) executed sequentially, or it may include steps (b) and (a) executed sequentially. If the method also includes step (c), then step (c) can be added to the method in any order without conflict, including but not limited to the execution order of steps (a), (b), and (c), steps (a), (c), and (b), steps (c), (a), and (b), etc.

[0088] Figure 1 The present invention provides a method for recovering iron concentrate from laterite nickel ore smelting slag, the method comprising the following steps:

[0089] S1. The laterite nickel ore smelting slag is subjected to water leaching treatment to obtain water leaching residue;

[0090] S2. The water-immersed residue from step S1 is subjected to alkaline leaching treatment to obtain leaching residue material;

[0091] S3. Mix the biomass reducing agent and the leaching residue from step S2 evenly, and then perform reduction roasting to obtain roasted material;

[0092] S4. Grind and magnetically separate the roasted material described in step S3 to obtain iron concentrate.

[0093] The laterite nickel ore smelting slag used in the specific embodiments of this invention is obtained by drying laterite nickel ore residue at 100°C for 480 min, followed by grinding and pulverization, and then passing it through a 200-mesh sieve, with a median particle size D50 of 75 μm. The laterite nickel ore residue is produced from laterite nickel ore through a high-pressure acid leaching (HPAL) process, with a moisture content of 30 wt%. XRD analysis determined that its main phases are hematite Fe2O3, jaundice aluminum sulfate (H3O)Al3(SO4)2(OH)6, and quartz SiO2. Elemental analysis of the laterite nickel ore residue produced by the high-pressure acid leaching process was performed using inductively coupled plasma optical emission spectrometry (ICP-OES), and the main elements and their contents are shown in Table 1.

[0094] Table 1

[0095]

[0096] Example 1

[0097] This embodiment provides a method for recovering iron concentrate from laterite nickel ore smelting slag, the method comprising the following steps:

[0098] S1. Deionized water was added to the laterite nickel ore smelting slag for water leaching treatment. The solid-liquid ratio was 1g:5mL. The mixture was heated to 60°C under stirring at 400rpm and leached for 10min. After filtration and washing, the water leaching residue was obtained.

[0099] S2. The obtained water-leaching residue was placed in a sodium hydroxide solution with a concentration of 1 mol / L and a solid-liquid ratio of 1 g: 4 mL. The solution was heated to 60°C and leached for 90 min under stirring at 400 rpm. After filtration, washing and drying, the leaching residue was obtained.

[0100] S3. The leaching residue and biomass reducing agent are ball-milled at 400 rpm for 10 min. The amount of biomass reducing agent added is 2.4 wt% of the mass of the leaching residue. The mixture after the first ball milling is placed in a crucible and placed in a tube furnace. N2 is introduced at a flow rate of 300 mL / min. The temperature is raised to 800°C at a rate of 5°C / min and held for 60 min. After calcination, the mixture is cooled to room temperature under N2 to obtain the calcined material.

[0101] The biomass reducing agent was prepared by the following method: coconut shells were crushed, rinsed with deionized water to remove surface impurities, dried at 100°C for 480 min, then carbonized at 500°C for 120 min under a nitrogen atmosphere, and naturally cooled to room temperature; the resulting product was activated at 500°C for 60 min under a carbon dioxide atmosphere, and naturally cooled to room temperature; the resulting material was crushed and ground in an agate mortar, passed through a 150-mesh sieve, and rinsed four times with deionized water to remove impurities, and then dried at 100°C for 680 min to obtain the above-mentioned biomass reducing agent;

[0102] S4. The roasted material is ball-milled for 10 minutes at 400 rpm with a ball-to-material ratio of 15:1. The material is then sieved until the particle size is >150 μm to obtain magnetic separation powder. Deionized water is added to the magnetic separation powder with a solid-liquid ratio of 1 g:20 mL. The mixture is stirred at 30°C and 400 rpm for 10 minutes to obtain magnetic separation slurry. The magnetic separation slurry is then fed into a magnetic separator with a magnetic field strength of 600 Gs for separation. The magnetic fraction is collected, filtered, and dried to obtain iron concentrate.

[0103] The non-magnetic components are collected, filtered, dried, and then reused.

[0104] Example 2

[0105] This embodiment provides a method for recovering iron concentrate from laterite nickel ore smelting slag, the method comprising the following steps:

[0106] S1. Deionized water was added to the laterite nickel ore smelting slag for water leaching treatment. The solid-liquid ratio was 1g:5mL. The mixture was heated to 60°C under stirring at 400rpm and leached for 10min. After filtration and washing, the water leaching residue was obtained.

[0107] S2. The obtained water-leaching residue was placed in a 1 mol / L sodium hydroxide solution with a solid-liquid ratio of 1 g: 4 mL and heated to 60°C under stirring at 400 rpm for alkali leaching for 90 min. After filtration and washing, alkali leaching residue was obtained. The obtained alkali leaching residue was placed in a 0.5 mol / L hydrochloric acid solution with a solid-liquid ratio of 1 g: 4 mL and heated to 60°C under stirring at 400 rpm for acid leaching for 120 min. After filtration and washing, it was dried at 105°C to obtain leaching residue material.

[0108] S3. The leaching residue and biomass reducing agent are ball-milled at 400 rpm for 10 min. The amount of biomass reducing agent added is 2.4 wt% of the mass of the leaching residue. The mixture after the first ball milling is placed in a crucible and placed in a tube furnace. N2 is introduced at a flow rate of 300 mL / min. The temperature is raised to 800°C at a rate of 5°C / min and held for 60 min. After calcination, the mixture is cooled to room temperature under N2 to obtain the calcined material.

[0109] The biomass reducing agent was prepared by the following method: coconut shells were crushed, rinsed with deionized water to remove surface impurities, dried at 100°C for 480 min, then carbonized at 500°C for 120 min under a nitrogen atmosphere, and naturally cooled to room temperature; the resulting product was activated at 500°C for 60 min under a carbon dioxide atmosphere, and naturally cooled to room temperature; the resulting material was crushed and ground in an agate mortar, passed through a 150-mesh sieve, and rinsed four times with deionized water to remove impurities, and then dried at 100°C for 680 min to obtain the above-mentioned biomass reducing agent;

[0110] S4. The roasted material is ball-milled for 10 minutes at 400 rpm with a ball-to-material ratio of 15:1. The material is then sieved until the particle size is >150 μm to obtain magnetic separation powder. Deionized water is added to the magnetic separation powder with a solid-liquid ratio of 1 g:20 mL. The mixture is stirred at 30°C and 400 rpm for 10 minutes to obtain magnetic separation slurry. The magnetic separation slurry is then fed into a magnetic separator with a magnetic field strength of 600 Gs for separation. The magnetic fraction is collected, filtered, and dried to obtain iron concentrate.

[0111] The non-magnetic components are collected, filtered, dried, and then reused.

[0112] Example 3

[0113] This embodiment provides a method for recovering iron concentrate from laterite nickel ore smelting slag. The difference between this method and Embodiment 2 is that in step S4, deionized water is added to the magnetic separation powder with a solid-liquid ratio of 1g:10mL. The remaining preparation methods and parameters are consistent with those in Embodiment 2.

[0114] Example 4

[0115] This embodiment provides a method for recovering iron concentrate from laterite nickel ore smelting slag. The difference between this method and Embodiment 2 is that in step S2, the stirring speed during the alkaline leaching process is 500 rpm, the stirring speed during the acid leaching process is 500 rpm, and in step S4, deionized water is added to the magnetic separation powder with a solid-liquid ratio of 1g:10mL. The remaining preparation methods and parameters are consistent with those of Embodiment 2.

[0116] Example 5

[0117] This embodiment provides a method for recovering iron concentrate from laterite nickel ore smelting slag. The difference between this method and Embodiment 2 is that in step S2, a sodium hydroxide solution with a concentration of 2 mol / L is used for alkali leaching for 120 min. The remaining preparation methods and parameters are consistent with those in Embodiment 2.

[0118] Example 6

[0119] This embodiment provides a method for recovering iron concentrate from laterite nickel ore smelting slag. The difference between this method and Embodiment 2 is that in step S2, a 2 mol / L sodium hydroxide solution is used, and the solution is heated to 70°C for alkali leaching for 120 min. The remaining preparation methods and parameters are consistent with those in Embodiment 2.

[0120] Example 7

[0121] This embodiment provides a method for recovering iron concentrate from laterite nickel ore smelting slag. The difference between this method and Embodiment 2 is that in step S2, a 1 mol / L hydrochloric acid solution is used and heated to 80°C for acid leaching. The remaining preparation methods and parameters are consistent with those in Embodiment 2.

[0122] Example 8

[0123] This embodiment provides a method for recovering iron concentrate from laterite nickel ore smelting slag. The difference between this method and Embodiment 2 is that in step S3, the amount of biomass reducing agent added is 16.7 wt% of the leaching slag material, argon gas is introduced for roasting at a flow rate of 400 mL / min, the roasting is held at a temperature of 100 min, and after roasting, it is cooled to room temperature under argon gas. The remaining preparation methods and parameters are consistent with those in Embodiment 2.

[0124] Example 9

[0125] This embodiment provides a method for recovering iron concentrate from laterite nickel ore smelting slag, the method comprising the following steps:

[0126] S1. Add deionized water to laterite nickel ore smelting slag for water leaching treatment. The solid-liquid ratio is 1g:3mL. Heat to 90°C under stirring at 200rpm and leach for 4min. After filtration and washing, the water leaching residue is obtained.

[0127] S2. The obtained water-leaching residue was placed in a sodium hydroxide solution with a concentration of 4 mol / L and a solid-liquid ratio of 1 g: 3 mL. The solution was heated to 90°C and leached for 30 min under stirring at 200 rpm. After filtration, washing and drying, the leaching residue was obtained.

[0128] S3. The leaching residue and biomass reducing agent are ball-milled at 500 rpm for 5 min. The amount of biomass reducing agent added is 2 wt% of the mass of the leaching residue. The mixture after the first ball milling is placed in a crucible and placed in a tube furnace. N2 is introduced at a flow rate of 200 mL / min. The temperature is raised to 500°C at a rate of 3°C / min and held for 180 min. After calcination, the mixture is cooled to room temperature under N2 to obtain the calcined material.

[0129] The biomass reducing agent was prepared by the following method: coconut shells were crushed, rinsed with deionized water to remove surface impurities, dried at 100°C for 480 min, carbonized at 350°C for 160 min under nitrogen atmosphere, and then naturally cooled to room temperature; the resulting product was activated at 350°C for 70 min under carbon dioxide atmosphere, and then naturally cooled to room temperature; the resulting material was crushed and ground in an agate mortar, passed through a 150-mesh sieve, and rinsed four times with deionized water to remove impurities; and then dried at 100°C for 680 min to obtain the above-mentioned biomass reducing agent.

[0130] S4. The roasted material is ball-milled for 5 minutes at 800 rpm with a ball-to-material ratio of 20:1. It is then sieved until the particle size is >150μm to obtain magnetic separation powder. Deionized water is added to the magnetic separation powder with a solid-liquid ratio of 1g:15mL. The mixture is stirred at 30°C and 400 rpm for 10 minutes to obtain magnetic separation slurry. The magnetic separation slurry is then fed into a magnetic separator with a magnetic field strength of 500Gs for separation. The magnetic fraction is collected, filtered, and dried to obtain iron concentrate.

[0131] The non-magnetic components are collected, filtered, dried, and then reused.

[0132] Example 10

[0133] This embodiment provides a method for recovering iron concentrate from laterite nickel ore smelting slag, the method comprising the following steps:

[0134] S1. Add deionized water to laterite nickel ore smelting slag for water leaching treatment with a solid-liquid ratio of 1g:4mL. Heat to 50°C under stirring at 500rpm and leach for 20min. After filtration and washing, obtain water leaching residue.

[0135] S2. The obtained water-leaching residue was placed in a sodium hydroxide solution with a concentration of 0.5 mol / L and a solid-liquid ratio of 1 g: 8 mL. The solution was heated to 50°C and leached for 140 min under stirring at 500 rpm. After filtration, washing and drying, the leaching residue was obtained.

[0136] S3. The leaching residue and biomass reducing agent are ball-milled at 200 rpm for 120 min. The amount of biomass reducing agent added is 20 wt% of the mass of the leaching residue. The mixture after the first ball milling is placed in a crucible and placed in a tube furnace. N2 is introduced at a flow rate of 400 mL / min. The temperature is raised to 900°C at a rate of 6°C / min and held for 30 min. After calcination, the mixture is cooled to room temperature under N2 to obtain the calcined material.

[0137] The biomass reducing agent was prepared by the following method: coconut shells were crushed, rinsed with deionized water to remove surface impurities, dried at 100°C for 480 min, carbonized at 600°C for 8 min under nitrogen atmosphere, and then naturally cooled to room temperature; the resulting product was activated at 600°C for 40 min under carbon dioxide atmosphere, and then naturally cooled to room temperature; the resulting material was crushed and ground in an agate mortar, passed through a 150-mesh sieve, and rinsed four times with deionized water to remove impurities; then dried at 100°C for 680 min to obtain the above-mentioned biomass reducing agent.

[0138] S4. The roasted material is ball-milled for 15 minutes at 400 rpm with a ball-to-material ratio of 10:1. The material is then sieved until the particle size is >150 μm to obtain magnetic separation powder. Deionized water is added to the magnetic separation powder with a solid-liquid ratio of 1 g:15 mL. The mixture is stirred at 30°C and 400 rpm for 10 minutes to obtain magnetic separation slurry. The magnetic separation slurry is then fed into a magnetic separator with a magnetic field strength of 3000 Gs for separation. The magnetic fraction is collected, filtered, and dried to obtain iron concentrate.

[0139] The non-magnetic components are collected, filtered, dried, and then reused.

[0140] Example 11

[0141] This embodiment provides a method for recovering iron concentrate from laterite nickel ore smelting slag. The difference between this method and Embodiment 1 is that in step S3, the amount of biomass reducing agent added is 1 wt% of the mass of the leaching slag, while the remaining preparation methods and parameters are consistent with Embodiment 1.

[0142] Example 12

[0143] This embodiment provides a method for recovering iron concentrate from laterite nickel ore smelting slag. The difference between this method and Embodiment 1 is that in step S3, the amount of biomass reducing agent added is 21 wt% of the mass of the leaching slag, while the remaining preparation methods and parameters are consistent with Embodiment 1.

[0144] Example 13

[0145] This embodiment provides a method for recovering iron concentrate from laterite nickel ore smelting slag. The difference between this method and Embodiment 1 is that in step S4, the particle size is controlled to be ≤150μm after sieving to obtain magnetically separated powder. The remaining preparation methods and parameters are consistent with those in Embodiment 1.

[0146] Comparative Example 1

[0147] This comparative example provides a method for recovering iron concentrate from laterite nickel ore smelting slag. The difference between this method and Example 1 is that step S1 water leaching is omitted, and alkaline leaching is performed directly. That is, the laterite nickel ore smelting slag is placed in a sodium hydroxide solution for alkaline leaching treatment. The remaining preparation methods and parameters are consistent with those of Example 1.

[0148] Comparative Example 2

[0149] This comparative example provides a method for recovering iron concentrate from laterite nickel ore smelting slag. The difference between this method and Example 1 is that the order of step S2 (alkali leaching treatment) and step S3 (reduction roasting) is reversed. That is, the water leaching residue is dried at 105°C and then mixed with a biomass reducing agent for reduction roasting. The resulting roasted material is then placed in a sodium hydroxide solution for alkali leaching treatment. The remaining preparation methods and parameters are consistent with those of Example 1.

[0150] The contents of iron, aluminum and sulfur in the iron concentrates obtained in Examples 1-13 and Comparative Examples 1-2 were tested, and the specific results are shown in Table 2.

[0151] Table 2

[0152]

[0153] The test results show that:

[0154] (1) As can be seen from Examples 1-10, the present invention effectively reduces the entry of aluminum, chromium, and silicate impurities into the magnetite phase by sequentially performing water leaching, alkali leaching, reduction roasting, grinding, and magnetic separation, thereby improving product quality and stability, successfully recovering high-quality iron concentrate products, and significantly improving their purity and magnetic properties. Specifically, the iron content in the iron concentrate can reach 60.07wt%-61.52wt%, the aluminum content is as low as 1.05wt%-1.22wt%, and the sulfur content is as low as 0.46wt%-0.71wt%.

[0155] (2) As can be seen from Examples 1 and 11-12, the present invention can effectively transform the iron mineral morphology and further improve the iron recovery rate by further adjusting the amount of biomass reducing agent added to 2wt%-20wt% of the mass of the leaching residue in step S2 and combining it with a reducing roasting atmosphere.

[0156] (3) As can be seen from Examples 1 and 13, the present invention can optimize the magnetic separation effect by controlling the particle size of the magnetic separation powder to >150μm, and significantly reduce the specific surface area of ​​the powder, reduce surface oxidation and exposure of active sites, thereby inhibiting the re-oxidation of the iron phase during storage and transportation, resulting in higher product stability; at the same time, the larger particle size is conducive to reducing dust flying, improving the working environment, and reducing material loss.

[0157] (4) As can be seen from Example 1 and Comparative Example 1, if the water leaching treatment is omitted in this invention, the acid-soluble, water-soluble impurities and some heavy metal ions in the laterite nickel ore smelting slag will remain inside. Direct alkaline leaching treatment of the laterite nickel ore smelting slag cannot remove many impurity ions, and the alkaline leaching effect is poor. This will also affect the subsequent reduction roasting and magnetic separation steps, resulting in a significant increase in the aluminum and sulfur content in the recovered iron concentrate, a significant decrease in the iron content, and a decrease in the grade of the iron concentrate.

[0158] (5) As can be seen from Example 1 and Comparative Example 2, if the present invention first performs reduction roasting and then alkaline leaching, the gangue minerals such as aluminosilicates will undergo solid-phase reaction with iron oxides during the high-temperature reduction process, generating insoluble aluminum-iron spinel (FeAl2O4) and low-melting-point slag phase. At the same time, the residual sulfate decomposes and combines with metallic iron to form a sulfide coating layer. During subsequent alkaline leaching, these high-temperature stable phases are difficult to be effectively decomposed and removed by the alkaline solution. Aluminum and sulfur are firmly contained in the iron concentrate in the form of mechanical inclusions or chemical bonding, which leads to a significant increase in the content of aluminum and sulfur in the recovered iron concentrate, a significant decrease in the iron content, and a decrease in the grade of the iron concentrate.

[0159] In summary, the method for recovering iron concentrate from laterite nickel ore smelting slag provided by this invention includes sequential water leaching, alkaline leaching, reduction roasting, grinding, and magnetic separation. This effectively reduces the entry of aluminum, chromium, and silicate impurities into the magnetite phase, thereby improving product quality and stability. It successfully recovers high-quality iron concentrate (magnetite phase Fe3O4), significantly enhancing its purity and magnetic properties. Simultaneously, the non-magnetic tailings obtained after magnetic separation can be further used in the production of building materials, cement additives, or other industrial products, achieving comprehensive resource utilization of the residue. This integrated technical approach significantly improves the overall resource recovery efficiency of laterite nickel ore smelting slag, reduces the environmental burden of residue disposal, and provides a sustainable and economically feasible technical approach for the recycling of laterite nickel ore resources.

[0160] The applicant declares that the above description is only a specific embodiment of the present invention, but the protection scope of the present invention is not limited thereto. Those skilled in the art should understand that any changes or substitutions that can be easily conceived by those skilled in the art within the technical scope disclosed in the present invention fall within the protection and disclosure scope of the present invention.

Claims

1. A method for recovering iron concentrate from laterite nickel ore smelting slag, characterized in that, The method includes the following steps: S1. The laterite nickel ore smelting slag is subjected to water leaching treatment to obtain water leaching residue; S2. The water-immersed residue from step S1 is subjected to alkaline leaching treatment to obtain leaching residue material; S3. Mix the biomass reducing agent and the leaching residue from step S2 evenly, and then perform reduction roasting to obtain roasted material; S4. Grind and magnetically separate the roasted material described in step S3 to obtain iron concentrate.

2. The method according to claim 1, characterized in that, The median particle size D50 of the laterite nickel ore smelting slag described in step S1 is 50μm-200μm; Preferably, based on the total mass of the laterite nickel ore smelting slag as 100wt%, the laterite nickel ore smelting slag includes 35wt%-45wt% iron, 3wt%-6wt% aluminum, 2wt%-8wt% sulfur, 2wt%-6wt% silicon, 1wt%-3wt% cobalt and 0.1wt%-1wt% magnesium.

3. The method according to claim 1 or 2, characterized in that, The solid-liquid ratio of the laterite nickel ore smelting slag and deionized water in step S1 is 1g:(3-5)mL; Preferably, the water immersion treatment in step S1 is carried out at a temperature of 50°C-90°C; Preferably, the stirring speed for the water immersion treatment in step S1 is 200 rpm to 500 rpm; Preferably, the water immersion treatment time in step S1 is 4 min to 20 min.

4. The method according to any one of claims 1-3, characterized in that, The alkaline solution used in the alkaline leaching treatment in step S2 includes sodium hydroxide solution and / or sodium carbonate solution; Preferably, the concentration of the alkaline solution used in the alkaline leaching treatment in step S2 is 0.5 mol / L-4 mol / L; Preferably, in step S2, the solid-liquid ratio of the water-leached residue and the alkaline solution in step S1 is 1g:(3-8)mL; Preferably, the temperature of the alkaline leaching treatment in step S2 is 50°C-90°C; Preferably, the stirring speed for the alkali leaching treatment in step S2 is 200 rpm to 500 rpm; Preferably, the alkaline leaching treatment in step S2 takes 30-140 minutes.

5. The method according to any one of claims 1-4, characterized in that, After the alkaline leaching treatment described in step S2, an acid leaching treatment is also included to obtain leaching residue. Preferably, the acid solution used in the acid leaching treatment includes hydrochloric acid solution; Preferably, the concentration of the acid solution used in the acid leaching treatment is 0.5 mol / L to 1 mol / L; Preferably, the solid-liquid ratio of the alkali leaching residue obtained after the alkali leaching treatment to the acid solution is 1g:(3-8)mL; Preferably, the acid leaching treatment temperature is 50°C-90°C; Preferably, the stirring speed for the acid leaching treatment is 200 rpm to 500 rpm; Preferably, the acid leaching treatment time is 30 min to 140 min.

6. The method according to any one of claims 1-5, characterized in that, In step S3, the amount of biomass reducing agent added is 2wt%-20wt% of the mass of the leaching residue in step S2. Preferably, the preparation method of the biomass reducing agent includes: carbonizing the biomass material under an inert atmosphere, followed by activation under a carbon dioxide atmosphere.

7. The method according to any one of claims 1-6, characterized in that, The mixing in step S3 includes a first ball milling process; Preferably, the first ball milling time is 5 min to 120 min.

8. The method according to any one of claims 1-7, characterized in that, The gas used in the reduction roasting described in step S3 includes an inert gas; Preferably, the flow rate of the inert gas is 200 mL / min to 400 mL / min; Preferably, the heating rate of the reduction calcination in step S3 is 3°C / min-6°C / min; Preferably, the holding temperature for the reduction calcination in step S3 is 500°C-900°C; Preferably, the holding time for reduction calcination in step S3 is 30 min to 180 min.

9. The method according to any one of claims 1-8, characterized in that, The grinding process described in step S4 includes a second ball milling. Preferably, the second ball milling time is 5 min to 15 min; Preferably, the temperature of the second ball mill is 20°C-30°C; Preferably, the magnetically separated powder is obtained by sieving after grinding, and the particle size of the magnetically separated powder is >150μm.

10. The method according to claim 9, characterized in that, Step S4, the magnetic separation includes: mixing deionized water and the magnetic separation powder evenly to obtain a magnetic separation slurry, and performing magnetic separation on the magnetic separation slurry; Preferably, the magnetic field strength for magnetic separation in step S4 is 500Gs-3000Gs.

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