Method for efficient separation of iron and manganese components in iron-manganese ore

Abstract

This invention relates to a method for the efficient separation of iron and manganese components in iron-manganese ore. The method involves mixing iron-manganese ore with intermediate products or vanadium products from the vanadium industry to obtain a mixture; roasting the mixture at high temperature to obtain roasted clinker; leaching the roasted clinker and then filtering it to obtain a manganese- and vanadium-containing leachate and an iron-containing leaching residue; adding a vanadium-precipitating additive to the obtained manganese- and vanadium-containing leachate and adjusting the pH value to induce precipitation; filtering after vanadium precipitation to obtain intermediate products or vanadium products from the vanadium industry and a manganese-containing supernatant; precipitating or drying the obtained manganese-containing supernatant to obtain a manganese product and manganese extraction waste liquid. This invention utilizes intermediate products or vanadium products from the vanadium industry to achieve mineral phase reconstruction of manganese minerals in iron-manganese ore. During this process, the intermediate products or vanadium products can be recycled, achieving efficient separation of iron and manganese in iron-manganese ore and efficient recycling of intermediate products or vanadium products from the vanadium industry.

Description

Technical Field

[0001] This invention relates to a method for the efficient separation of iron and manganese components in iron-manganese ore, specifically a method for the efficient separation of iron and manganese components in iron-manganese ore from vanadium industry intermediates or vanadium products, while simultaneously enabling the recycling of these intermediates or products. This method belongs to the field of iron-manganese ore resource utilization and vanadium chemical technology and resource recycling technology. Background Technology

[0002] Manganese is a major strategic metal in short supply to sustain the national economy. The vast majority of manganese ore in China is iron-manganese ore with a low Mn / Fe mass ratio (<3), and its output can no longer meet the rapid development of the metallurgical industry. The dependence on foreign sources for manganese ore is as high as 90%. At present, the main separation processes for iron and manganese are mechanical separation, manganese-rich slag separation, reduction roasting magnetic separation, and leaching separation. Although these processes have been industrialized, they still face unavoidable problems.

[0003] Mechanical separation methods mainly include gravity separation, flotation, and magnetic separation. Since the physicochemical properties of iron and manganese in ore are similar, the ore characteristics do not change significantly during the separation process, making it difficult for a single mechanical separation technology to effectively separate valuable minerals from gangue minerals.

[0004] The manganese-rich slag separation method is a relatively mature separation process. It utilizes the difference in solubility between manganese slag and molten iron to allow manganese to enter the slag phase, resulting in manganese-rich slag, while the iron remains in a liquid state, thus achieving the separation of iron and manganese. However, this method requires coke as a reducing agent and smelting at high temperatures, which is characterized by high energy consumption and high pollution.

[0005] Reduction roasting magnetic separation refers to roasting the ore with reducing solids or gases before magnetic separation, causing iron to be enriched in the strongly magnetic phase while manganese remains in the weakly magnetic phase, further increasing the magnetic difference between manganese and iron. However, commonly used carbothermic reduction equipment includes vertical shaft furnaces and rotary kilns, which are complex processes, have difficulties in mass and heat transfer, and have high production and maintenance costs.

[0006] Leaching separation methods include chemical leaching and bioleaching. In leaching processes, iron and manganese are commonly leached together, making it difficult to remove iron from manganese. Furthermore, the amount of chemical waste and wastewater generated during the process increases, causing significant environmental harm.

[0007] From the perspectives of cost, environmental protection, and recycling efficiency, the above-mentioned iron and manganese separation processes all have shortcomings. Therefore, there is an urgent need to find a more suitable process for iron and manganese ore separation. Summary of the Invention

[0008] (a) Technical problems to be solved

[0009] To address the aforementioned problems in the prior art, this invention proposes a method for the efficient separation of iron and manganese components in iron-manganese ore. This method utilizes vanadium to directionally reconstruct and transform sparingly soluble manganese minerals in iron-manganese ore into acid-soluble manganese vanadate, as well as to regulate the phase composition of iron minerals. The method then achieves efficient separation of iron and manganese components and recycling of the roasting medium through acid leaching.

[0010] (II) Technical Solution

[0011] To achieve the above objectives, the main technical solutions adopted by the present invention include:

[0012] This invention proposes a method for the efficient separation of iron and manganese components in iron-manganese ore, which includes the following steps:

[0013] S1. Roasting: Iron-manganese ore is mixed with intermediate products or vanadium products from the vanadium industry to obtain a mixture. The mixing ratio of iron-manganese ore and intermediate products or vanadium products from the vanadium industry is calculated as n(MnO2) / n(V2O5), where MnO2 is manganese in the iron-manganese ore (calculated as MnO2) and V2O5 is vanadium products or vanadium oxides in intermediate products or vanadium products from the vanadium industry (calculated as V2O5). The ratio of n(MnO2) / n(V2O5) is 0.5 to 3:1. The mixture is roasted at high temperature to obtain roasted clinker.

[0014] S2, Leaching: The roasted clinker obtained in step S1 is leached and then filtered to obtain a manganese- and vanadium-containing leachate and an iron-containing leaching residue. The iron-containing leaching residue is the separated iron product. The iron-containing leaching residue has a very high iron content, reaching the standard of iron concentrate (TFe>60%), and can be used directly as a raw material for blast furnace ironmaking or as a raw material for preparing ferroalloys.

[0015] S3. Vanadium recovery: Add vanadium precipitation additive to the manganese- and vanadium-containing leachate obtained in step S2, adjust the pH value to precipitate, and filter after vanadium precipitation to obtain intermediate products or vanadium products in the vanadium industry and manganese-containing supernatant.

[0016] S4. Manganese recovery: The manganese-containing supernatant obtained in step S3 is precipitated or dried and crystallized, and then filtered to obtain manganese products and manganese extraction waste liquid.

[0017] Furthermore, the method also includes:

[0018] S5. Utilization of iron-containing tailings: The iron-containing leaching residue obtained in step S2 can be directly used as the feed material of the blast furnace or as a raw material for the preparation of ferroalloys.

[0019] S6. Utilization of intermediate products or vanadium products in the vanadium industry: The intermediate products or vanadium products obtained in step S3 can be output as products, or recycled back to the roasting system as additives for the next round of roasting; the recovered intermediate products or vanadium products can be directly output as products, or recycled again as roasting additives.

[0020] S7. Manganese extraction waste liquid recycling: After impurity removal, the manganese extraction waste liquid obtained in step S4 is recycled back to the leaching system as the leaching medium in step S2. The impurities removed from the vanadium extraction waste liquid are ammonia nitrogen.

[0021] In the method described above, preferably, in step S1, the iron-manganese ore is a manganese ore with a Mn / Fe mass ratio of <3, and may be at least one of oxide minerals, hydrogenated minerals, carbonate minerals, sulfide minerals, silicate minerals, and borate minerals; its particle size is 200-300 mesh.

[0022] The vanadium industry intermediates or vanadium products include at least one of vanadates, vanadium oxides, and vanadium slag. Vanadates include polyvanadates, ammonium metavanadate, pyrovanadate, and orthovanadate, specifically ammonium polyvanadate, ammonium metavanadate, sodium metavanadate, sodium pyrovanadate, sodium orthovanadate, calcium metavanadate, calcium pyrovanadate, calcium orthovanadate, manganese metavanadate, manganese pyrovanadate, and manganese orthovanadate. Vanadium oxides include at least one of vanadium trioxide, vanadium dioxide, and vanadium pentoxide. Vanadium slag is an intermediate product obtained from the smelting process of vanadium-titanium magnetite, with a vanadium content of 1-15% (V₂O₅).

[0023] In the actual roasting process, vanadium directionally reconstructs and transforms the insoluble manganese minerals in the iron-manganese ore into acid-soluble manganese vanadate, and regulates the phase composition of the iron minerals. Then, leaching with an acid solution achieves efficient separation of the iron and manganese components and allows for the recycling of the roasting medium. Preferably, in step S1, the high-temperature roasting temperature is 700–1000℃, the isothermal time is 10–300 min, and the heating rate is 2–10℃ / min.

[0024] In the method described above, preferably, in step S2, leaching is carried out at a deionized water to calcined clinker liquid-solid ratio of 2 to 20:1 mL / g, and the pH value of the leaching system is controlled to be 2.0 to 3.0 by adding acid solution during the leaching process. The leaching temperature is 30 to 100°C, and the leaching time is 30 to 180 min.

[0025] The acid solution is at least one of sulfuric acid, hydrochloric acid, or nitric acid; leaching can be carried out in a water bath or heating mantle, and the iron leaching residue is repeatedly rinsed with water at a pH of 2.0 to 3.0 and a temperature of 20 to 100°C until the solution is transparent.

[0026] In the method described above, preferably, in step S3, the vanadium precipitation additive is an ammonium salt, which is at least one of ammonium sulfate, ammonium carbonate, or ammonium bicarbonate; the amount of the vanadium precipitation additive added is such that the molar ratio of ammonium in the ammonium salt to vanadium in the leachate is 1.0 to 3.0; the pH value of the system is maintained at 2 to 3 by adding acid during the vanadium precipitation process; the acid is at least one of sulfuric acid, hydrochloric acid, or nitric acid; the vanadium precipitation temperature is 60 to 100°C, and the vanadium precipitation time is 20 to 180 minutes.

[0027] Preferably, in step S4, the manganese precipitation is achieved through either carbonization precipitation or direct precipitation. The carbonization precipitation method involves adding at least one of carbon dioxide gas, sodium carbonate solution, and ammonium carbonate solution to the manganese-containing supernatant. The carbon dioxide is either pure carbon dioxide gas, industrial carbon dioxide, or carbon dioxide-containing flue gas released during the roasting of vanadium slag mixed with calcium and manganese salts. During carbonization, an alkaline solution is added to maintain the pH of the carbonization system at 6-8. The alkaline solution is at least one of sodium hydroxide solution, ammonia, sodium carbonate, and ammonium carbonate solution. The carbonization temperature is 20-100°C, and the carbonization time is 10-120 minutes. The manganese salt obtained from the carbonization precipitation is manganese carbonate. The direct precipitation method involves adding an alkaline substance to the manganese-containing supernatant to maintain the pH of the precipitation system at 7-10. The alkaline substance is at least one of sodium hydroxide solution and ammonia. The direct precipitation temperature is 20-100°C, and the precipitation time is 10-120 minutes. The manganese salt obtained from the direct precipitation is manganese tetroxide.

[0028] In step S5, the drying and crystallization method involves placing the manganese-containing supernatant in at least one of a rotary evaporator, a vacuum drying oven, an oven, or a centrifuge. The drying temperature is 20–110°C, and the drying time is 12–48 h. The resulting manganese salt is obtained through drying and crystallization.

[0029] (III) Beneficial Effects

[0030] The method for enhancing vanadium-chromium separation by synergistic roasting of vanadium slag using calcium and manganese salts provided by this invention has the following advantages:

[0031] (1) The vanadium yield of the method of the present invention is higher than that of the existing iron and manganese separation process in iron and manganese ore, which reflects the synergistic effect of vanadium and manganese. This is because, starting from the low temperature roasting stage (600℃), manganese and vanadium combine to form a stable manganese pyrovanadate phase. The phase does not change even when the roasting temperature is further increased. The manganese pyrovanadate phase is easily soluble in acid solution, which can realize the efficient extraction of manganese and vanadium.

[0032] (2) The method of the present invention does not release toxic gases during the roasting process;

[0033] (3) The separation efficiency of iron and manganese in iron-manganese ore by the method of the present invention is far higher than that of existing iron-manganese ore separation methods;

[0034] (4) The method of the present invention enriches vanadium in the slag and manganese in the leachate by adding vanadium precipitation additive.

[0035] (5) The method of the present invention enriches the iron in the manganese iron ore into a phase in the iron tailings, and its content reaches that of iron concentrate (TFe>60%), which can be directly used for blast furnace smelting of various ferroalloys.

[0036] (6) The addition of intermediate products or vanadium products in the vanadium industry reduces the high roasting reaction temperature during the separation of iron and manganese in traditional iron-manganese ores, further saving costs and improving the recovery rate of valuable components of iron and manganese in iron-manganese ores.

[0037] (7) Due to the addition of intermediate products or vanadium products in the vanadium industry, the problem of iron and manganese being leached together during the leaching of iron-manganese ore is avoided, and the subsequent process of removing iron from manganese is reduced.

[0038] (8) Vanadium industry intermediates or vanadium products used as additives can be recycled. Detailed Implementation

[0039] To better explain and facilitate understanding of the present invention, the present invention will be described in detail below through specific embodiments.

[0040] The reaction mechanism of this invention is:

[0041] The method provided by this invention uses iron-manganese ore as raw material and vanadium industry intermediates or vanadium products as roasting media. It leverages vanadium's excellent selective binding and directional reduction capabilities for manganese to achieve phase reconstruction of manganese minerals in iron-manganese ore into acid-soluble vanadates without altering the iron mineral phase. Furthermore, acid leaching enables efficient separation of iron and manganese and the recycling of the vanadium roasting media, effectively avoiding the drawbacks of existing iron-manganese ore separation methods.

[0042] This invention innovatively uses intermediate products or vanadium products from the vanadium industry as additives, mixed with iron-manganese ore and roasted. Vanadium and manganese combine at a low temperature (600℃) to form manganese pyrovanadate, which is easily soluble in acid. Even with further increases in roasting temperature, the phase remains manganese pyrovanadate, exhibiting relative stability, and allowing for the separation and extraction of iron and manganese during subsequent leaching. In this process, manganese and vanadium work synergistically and are indispensable. The manganese in the iron-manganese ore is defined as MnO2, and the vanadium in the intermediate products or vanadium products is defined as V2O5, with a molar ratio of n(MnO2) / n(V2O5) of 0.5–3:1, as shown in equation (2). If n(MnO2) / n(V2O5) is less than 0.5:1, the reaction between vanadium and the elements in the iron-manganese ore will be insufficient, resulting in the formation of Fe4Mn3(VO4)6, as shown in equation (1). If n(MnO2) / n(V2O5) is greater than 3:1, it will result in too little vanadium, which is insufficient to react completely with the manganese in the iron-manganese ore to form manganese pyrovanadate. The remaining manganese will combine with iron to form FeMnO3, as shown in equation (3).

[0043] 8Fe2O3+6Mn2O3+12V2O5=4Fe4Mn3(VO4)6+3O2↑(1)

[0044] 2MnO2 + V2O5 = Mn2V2O7 + O2↑ (2)

[0045] 2Fe2O3+4MnO2=4FeMnO3+O2↑(3).

[0046] To better explain and facilitate understanding of the present invention, the present invention will be described in detail below through specific embodiments.

[0047] Example 1

[0048] A method for efficiently separating iron and manganese components in iron-manganese ore by mixing intermediate products or vanadium products from the vanadium industry with iron-manganese ore, while simultaneously recycling the intermediate products or vanadium products, includes the following steps:

[0049] (1) Roasting: The iron-manganese ore is ground and sieved to obtain an iron-manganese ore particle size of 200-250 mesh. Vanadium pentoxide is added and mixed, followed by roasting. The vanadium pentoxide particle size is required to be 200-250 mesh so that the vanadium in the vanadium pentoxide can further combine with the iron-manganese ore. The manganese content in the iron-manganese ore is 26.97% (MnO2), and the vanadium content in the vanadium pentoxide is 56.02% (V2O5). The iron-manganese ore and vanadium pentoxide are mixed at a molar ratio of n(MnO2) / n(V2O5) of 2.5. The mixed material is roasted at 850℃ for 120 min to obtain roasted clinker at a heating rate of 2° / min. The roasted clinker is then crushed to a particle size of 250-300 mesh.

[0050] (2) Leaching: The leaching process was carried out in a water bath. Deionized water was added to the leaching container at a liquid-to-solid ratio of 5 mL / g to calcined clinker. The calcined clinker was added when the temperature was heated to 70°C. The pH of the leaching system was controlled to be around 2.2 by adding 10% sulfuric acid by volume. The leaching temperature was 70°C. After leaching for 120 min, the solid and liquid were separated by filtration to obtain iron-containing leaching residue and manganese- and vanadium-containing leaching solution. The leaching residue was repeatedly rinsed with hot water at a pH of 2.2 and a temperature of 70°C. After leaching, the content of manganese and vanadium in the leaching solution was measured, and the vanadium leaching rate was found to be 89.26%. The vanadium leaching rate was calculated using the formula: Vanadium leaching rate = (mass of vanadium in the leaching solution) / mass of vanadium in the calcined clinker. The manganese leaching rate was 85.97%, while the iron leaching rate in the solution was only 0.00492%, meaning that the leachate contained both manganese and vanadium. Iron was concentrated in the leaching residue, thus achieving efficient separation of iron and manganese.

[0051] (3) Vanadium precipitation: The vanadium precipitation process was carried out in a water bath, and ammonium sulfate was used as the vanadium precipitation additive. The amount of vanadium precipitation additive added was based on a molar ratio of ammonium in the ammonium salt to vanadium in the leachate of 1.5. During the vanadium precipitation process, the pH of the system was maintained at 2 by adding a 10% sulfuric acid solution. After precipitation at 95℃ for 60 min, vanadium precipitated, and the solid and liquid were separated by filtration to obtain ammonium polyvanadate precipitate and manganese-containing supernatant. After vanadium precipitation, the vanadium content in the manganese-containing supernatant was measured, and the vanadium precipitation rate was found to be 99.28%. The vanadium precipitation rate was calculated according to the following formula: Vanadium precipitation rate = (mass of vanadium in the leachate - mass of vanadium in the manganese-containing supernatant) / mass of vanadium in the leachate.

[0052] (4) Obtaining vanadium product: The ammonium polyvanadate precipitate is dried at 120℃ for 12h before being heated to 550℃ for 120min. The vanadium pentoxide obtained is measured to have a purity of 99.23% by chemical analysis. It can be output as vanadium pentoxide product or returned to the roasting system in step (1) to participate in the next round of roasting process.

[0053] (5) Utilization of iron-containing tailings: Iron-containing leaching slag is iron concentrate (TFe>60%), which is directly fed into the blast furnace for smelting. The smelting atmosphere and temperature are controlled to obtain high-quality ferroalloys.

[0054] (6) Manganese recovery: Carbon dioxide is introduced into the manganese-containing supernatant obtained in step (3) for carbonization, with a gas flow rate of 0.1 L / min; during the carbonization process, the pH value of the carbonization system is maintained at 6.8 by adding 5 g / L sodium hydroxide solution; the carbonization temperature is 40 °C; the carbonization time is 60 min; then the solid and liquid are separated by filtration to obtain manganese carbonate precipitate and vanadium extraction waste liquid.

[0055] Alternatively, electrolysis of the manganese-containing supernatant obtained in step (3) can yield an electrolytic manganese product with a manganese content >99.7%.

[0056] (7) Impurity removal: The main impurity to be removed from the vanadium extraction waste liquid obtained in step (5) is ammonia nitrogen. The ammonia nitrogen is removed by chemical precipitation and then recycled back to the leaching system as the leaching medium in step (2) for reuse.

[0057] Example 2

[0058] A method for efficiently separating iron and manganese components in iron-manganese ore by mixing intermediate products or vanadium products from the vanadium industry with iron-manganese ore, while simultaneously recycling the intermediate products or vanadium products, includes the following steps:

[0059] (1) Roasting: The iron-manganese ore is ground and sieved to obtain an iron-manganese ore particle size of 200-250 mesh. Ammonium polyvanadate is added and mixed, followed by roasting. The particle size of the ammonium polyvanadate is required to be 200-250 mesh so that the vanadium in the ammonium polyvanadate can further combine with the iron-manganese ore. The manganese content in the iron-manganese ore is 26.97% (MnO2), and the vanadium content in the ammonium polyvanadate is 50.50% (V2O5). The iron-manganese ore and ammonium polyvanadate are mixed at a molar ratio of n(MnO2) / n(V2O5) of 2.25. The mixed material is roasted at 850℃ for 180 min to obtain roasted clinker at a heating rate of 5° / min. The roasted clinker is crushed to a particle size of 250-300 mesh.

[0060] (2) Leaching: The leaching was carried out in a water bath. Deionized water was added to the leaching container at a liquid-to-solid ratio of 10 mL / g to the calcined clinker. The calcined clinker was added when the temperature was heated to 60°C. The pH of the leaching system was controlled to 2.0 by adding 20% ​​sulfuric acid. The leaching temperature was 60°C. After leaching for 120 min, the solid and liquid were separated by filtration to obtain iron-containing leaching residue and manganese- and vanadium-containing leaching solution. The leaching residue was repeatedly rinsed with hot water at 60°C and pH 2.0. After leaching, the contents of iron, manganese, and vanadium in the manganese- and vanadium-containing leaching solution were measured. The leaching rates were found to be 0.00732% for iron, 84.73% for manganese, and 86.02% for vanadium, thus achieving efficient separation of iron and manganese.

[0061] (3) Vanadium precipitation: The vanadium precipitation process is carried out in a water bath, and ammonium sulfate is used as the vanadium precipitation additive. The amount of vanadium precipitation additive added is based on a molar ratio of ammonium in the ammonium salt to vanadium in the leachate of 1.3. The pH of the system is maintained at 2 by adding a 20% sulfuric acid solution. The vanadium precipitation temperature is 100℃, and after precipitation for 40 min, vanadium precipitates. The solid and liquid are separated by filtration to obtain ammonium polyvanadate precipitate and manganese-containing supernatant. After vanadium precipitation, the vanadium content in the manganese-containing supernatant is measured, and the vanadium precipitation rate is 98.36%. The purity is measured to be 98.76% by chemical analysis. It can be exported as ammonium polyvanadate product or returned to the roasting system of step (1) to participate in the next round of roasting process.

[0062] (4) Utilization of iron-containing tailings: Iron-containing leaching slag is iron concentrate (TFe>60%), which is directly fed into the blast furnace for smelting. The smelting atmosphere and temperature are controlled to obtain high-quality ferroalloys.

[0063] (5) Manganese recovery: Add ammonia to the manganese-containing supernatant obtained in step (3) to precipitate manganese ions, maintain the pH value of the precipitation system at 8.0; the precipitation temperature is 60℃; the precipitation time is 30min; then filter to separate the solid and liquid to obtain manganese tetroxide precipitate and vanadium extraction waste liquid.

[0064] (6) Impurity removal: The main impurity to be removed from the vanadium extraction waste liquid obtained in step (5) is ammonia nitrogen. The ammonia nitrogen is removed by conventional chemical precipitation method, and the removed impurity is recycled back to the leaching system as the leaching medium in step (2).

[0065] Example 3

[0066] A method for efficiently separating iron and manganese components in iron-manganese ore by mixing intermediate products or vanadium products from the vanadium industry with iron-manganese ore, while simultaneously recycling the intermediate products or vanadium products, includes the following steps:

[0067] (1) Calcination: The iron-manganese ore is ground and sieved to obtain an iron-manganese ore particle size of 200-250 mesh. Vanadium slag is added for mixing, followed by mixed roasting. The particle size of the vanadium slag is required to be 200-250 mesh so that the vanadium in the vanadium slag can further combine with the iron-manganese ore. The manganese content in the iron-manganese ore is 26.97% (MnO2), and the vanadium content in the vanadium slag is 18.22% (V2O5). The iron-manganese ore and vanadium pentoxide are mixed at a molar ratio of n(MnO2) / n(V2O5) of 2.35. The mixed material is roasted at 850℃ for 120 min to obtain roasted clinker at a heating rate of 10° / min. The roasted clinker is crushed to a particle size of 250-300 mesh.

[0068] (2) Leaching: The leaching process was carried out in a heating jacket. Deionized water was added to the leaching container at a liquid-to-solid ratio of 3 mL / g to deionized water and calcined clinker. Calcined clinker was added at 70°C. The pH of the leaching system was controlled to be around 2.7 by adding 50% sulfuric acid. The leaching temperature was 70°C and the leaching time was 50 min. After leaching, solid-liquid separation was performed by filtration to obtain leaching residue and vanadium-containing leachate. The leaching residue was repeatedly rinsed with hot water at 70°C and pH 2.7. After leaching, the contents of iron, manganese, and vanadium in the manganese- and vanadium-containing leachate were measured. The leaching rates were found to be 85.45% for manganese, 87.45% for vanadium, and 0.00412% for iron, achieving efficient separation of iron and manganese. In addition, 7.2% of the manganese in the original vanadium slag was transferred into the manganese- and vanadium-containing leachate after composite roasting and acid leaching. This part of the manganese also combined with the vanadium in the vanadium slag to form manganese pyrovanadate, which was then leached into the solution, thus realizing the recovery and utilization of part of the manganese in the vanadium slag.

[0069] (3) Vanadium precipitation: The vanadium precipitation process is carried out in a water bath, and ammonium sulfate is used as the vanadium precipitation additive. The amount of vanadium precipitation additive added is based on a molar ratio of ammonium in the ammonium salt to vanadium in the leachate of 2.0. The pH of the system is maintained at 2.2 by adding a 50% sulfuric acid solution. After precipitation at 90℃ for 90 min, vanadium precipitates, and the solid and liquid are separated by filtration to obtain ammonium polyvanadate precipitate and manganese-containing supernatant. After vanadium precipitation, the vanadium content in the manganese-containing supernatant is measured, and the vanadium precipitation rate is 97.35%. The purity is 97.54% as determined by chemical analysis. It can be exported as ammonium polyvanadate product or returned to the roasting system of step (1) to participate in the next round of roasting process.

[0070] (4) Utilization of iron-containing tailings: Iron-containing leaching slag is iron concentrate (TFe>60%), which is directly fed into the blast furnace for smelting. The smelting atmosphere and temperature are controlled to obtain high-quality ferroalloys.

[0071] (5) Manganese recovery: Add 20% sodium hydroxide solution to the manganese-containing supernatant obtained in step (3) to precipitate manganese ions, maintain the pH value of the precipitation system at about 7.0; the precipitation temperature is 40℃; the precipitation time is 120min; filter to separate solid and liquid, and obtain manganese carbonate precipitate and vanadium extraction waste liquid.

[0072] (6) Manganese salt utilization: The manganese carbonate precipitate from step (5) is oxidized and roasted at 400℃ for 12h, during which oxygen is introduced at a flow rate of 1L / min. After obtaining crude manganese dioxide product, it is placed in a sulfuric acid solution and sodium chlorate solution with a volume fraction of 15% and reacted for 120min to obtain refined manganese dioxide with a manganese dioxide content of 91.98%. This product can be used as an industrial oxidant and catalyst.

[0073] (7) Impurity removal: The main impurity to be removed from the vanadium extraction waste liquid obtained in step (3) is ammonia nitrogen. The ammonia nitrogen is removed by chemical precipitation and then recycled back to the leaching system as the leaching medium in step (2).

[0074] Comparative Example 1

[0075] When using conventional sulfation roasting-water leaching to separate iron and manganese in low-grade iron-manganese ore (TFe: 30.35%, Mn: 14.30%), not only SO2 but also a large amount of N2 needs to be introduced during the roasting process, resulting in a significant waste of gas. During leaching, co-leaching of iron and manganese is unavoidable, which brings considerable difficulties to the subsequent removal of iron from manganese.

[0076] Comparative Example 2

[0077] Endophytic fungi were used to bioleach iron-manganese ore (Fe2O3: 67.90%, MnO2: 8.79%). This method requires a 20-day experiment to separate iron and manganese in the ore, but the manganese leaching rate is only 79%, which is much lower than the manganese leaching rate of this invention. However, different types of bacteria are required for different iron-manganese ores, and the entire leaching process is slow, which greatly limits this method.

[0078] Comparative Example 3

[0079] Iron and manganese in high-iron manganese oxide ore (TFe: 21.28%, Mn: 39.05%) were separated by adding 7wt% Na2S2O3 as a reducing agent and employing a reduction roasting and magnetic separation method. This method not only requires the reaction to be carried out at a high temperature of 1100℃, resulting in high energy consumption, but also necessitates the reaction in a vertical furnace, making the process complex and incurring high production and maintenance costs. Furthermore, the manganese recovery rate is only 85.96%, far lower than the manganese recovery rate of this invention.

[0080] The above description is merely a preferred embodiment of the present invention and is not intended to limit the present invention in any other way. Any person skilled in the art can make changes or modifications to the above-disclosed technical content to create equivalent embodiments. However, any simple modifications, equivalent changes, and modifications made to the above embodiments based on the technical essence of the present invention without departing from the scope of the present invention shall still fall within the protection scope of the present invention.

Claims

1. A method for efficient separation of iron and manganese components in iron-manganese ore, characterized in that, It includes the following steps: S1. Roasting: A mixture is prepared by mixing iron-manganese ore with intermediate products or vanadium products from the vanadium industry. The iron-manganese ore is a manganese ore with an Mn / Fe mass ratio <3. The mixing ratio of the iron-manganese ore and the intermediate products or vanadium products from the vanadium industry is as follows: n (MnO2) / n (V2O5) calculation, n (MnO2) / n (V2O5) ratio is 0.5–3:1; the mixture is roasted at high temperature to obtain roasted clinker; In step S1, the intermediate product or vanadium product in the vanadium industry includes at least one of vanadium pentoxide, ammonium polyvanadate, and vanadium slag. When it is vanadium slag, the vanadium content in the vanadium slag is 1 to 15% as V2O5. S2. Leaching: The roasted clinker obtained in step S1 is leached and then filtered to obtain a manganese- and vanadium-containing leachate and an iron-containing leaching residue; wherein the iron-containing leaching residue has an iron content (TFe) > 60%; S3. Vanadium recovery: Add vanadium precipitation additive to the manganese- and vanadium-containing leachate obtained in step S2, adjust the pH value to precipitate, and filter after vanadium precipitation to obtain intermediate products or vanadium products in the vanadium industry and manganese-containing supernatant. S4. Manganese recovery: The manganese-containing supernatant obtained in step S3 is precipitated or dried and crystallized, and then filtered to obtain manganese products and manganese extraction waste liquid.

2. The method as described in claim 1, characterized in that, The method further includes: S5. Utilization of iron-containing tailings: The iron-containing leaching residue obtained in step S2 can be directly used as the feed material of the blast furnace or as a raw material for the preparation of ferroalloys. S6. Utilization of intermediate products or vanadium products in the vanadium industry: The intermediate products or vanadium products obtained in step S3 are output as products, or the intermediate products or vanadium products in the vanadium industry are recycled back to the roasting system as additives to participate in the next round of roasting. S7. Manganese extraction waste liquid recycling: After impurities are removed from the manganese extraction waste liquid obtained in step S4, it is recycled back to the leaching system as the leaching medium in step S2.

3. The method as described in claim 1, characterized in that, In step S1, the iron-manganese ore is at least one of oxide minerals, hydrogenated minerals, carbonate minerals, sulfide minerals, silicate minerals, and borate minerals; Its particle size is 200-300 mesh.

4. The method as described in claim 1, characterized in that, In step S1, the high-temperature calcination temperature is 700-1000℃, the holding time is 10-300 min, and the heating rate is 2-10℃ / min.

5. The method as described in claim 1, characterized in that, In step S2, the leaching process is carried out by leaching deionized water to calcined clinker liquid-solid ratio of 2 to 20:1 mL / g, and the pH value of the leaching system is controlled to be 2.0 to 3.0 by adding acid solution during the leaching process. The leaching temperature is 30 to 100°C and the leaching time is 30 to 180 min.

6. The method as described in claim 5, characterized in that, In step S2, the iron-containing leaching residue is repeatedly rinsed with water at a pH of 2.0-3.0 and a temperature of 20-100℃ until the solution becomes transparent.

7. The method as described in claim 1, characterized in that, In step S3, the vanadium precipitation additive is an ammonium salt, which is at least one of ammonium sulfate, ammonium carbonate, or ammonium bicarbonate; the amount of vanadium precipitation additive added is such that the molar ratio of ammonium in the ammonium salt to vanadium in the leachate is 1.0 to 3.0; the pH of the system is maintained at 2 to 3 by adding acid during the vanadium precipitation process; the vanadium precipitation temperature is 60 to 100°C, and the vanadium precipitation time is 20 to 180 minutes.

8. The method as described in claim 1, characterized in that, In step S4, the precipitation includes two methods: carbonation precipitation or direct precipitation. The carbonation precipitation method involves adding at least one of carbon dioxide gas, sodium carbonate solution, and ammonium carbonate solution to the manganese-containing supernatant. During the carbonation process, an alkaline solution is added to maintain the pH value of the carbonation system at 6-8. The carbonation temperature is 20-100℃, and the carbonation time is 10-120 min. The manganese salt obtained by carbonation precipitation is manganese carbonate. The direct precipitation method involves adding an alkaline substance to the manganese-containing supernatant to maintain the pH value of the precipitation system at 7-10. The direct precipitation temperature is 20-100℃, and the direct precipitation time is 10-120 min. The manganese salt obtained by direct precipitation is manganese tetroxide.

9. The method as described in claim 1, characterized in that, In step S4, the drying and crystallization method involves placing the manganese-containing supernatant in at least one of a rotary evaporator, a vacuum drying oven, an oven, or a centrifuge. The drying temperature is 20-110°C, and the drying time is 12-48 hours. The manganese salt is obtained by drying and crystallization.