A method for combined treatment of laterite nickel ore smelting slag and phosphating slag

Abstract

The application discloses a combined treatment method for laterite nickel ore smelting slag and phosphating slag, and belongs to the technical field of hydrometallurgy. The combined treatment method comprises the following steps: uniformly mixing the laterite nickel ore smelting slag, the phosphating slag and a sodium alkaline salt to obtain a mixture; performing sodium reduction roasting on the mixture to obtain a roasted material; performing water leaching on the roasted material to obtain a water leaching solution and a water leaching residue through filtration; and performing magnetic separation on the water leaching residue to obtain an iron concentrate. The laterite nickel ore smelting slag and the phosphating slag are mixed and then subjected to sodium reduction roasting, so that iron exists in the form of Fe3O4, manganese and zinc exist in the form of oxides in the slag, and aluminum oxide is converted into sodium aluminate which is easy to dissolve in water by reacting with the alkali in the roasting process, and then the aluminum in the roasted material can be efficiently separated through subsequent water leaching; meanwhile, the reaction of phosphating zinc and phosphating manganese with the alkali also promotes the transformation of iron oxide from non-magnetic to magnetic, and then the iron can be efficiently recovered through subsequent magnetic separation.

Description

Technical Field

[0001] This invention relates to the field of hydrometallurgical technology, and in particular to a method for the combined treatment of laterite nickel ore smelting slag and phosphating slag. Background Technology

[0002] With the rapid development of the new energy industry, nickel, as a supporting metal in this sector, has seen extensive research into its mineral resources. Laterite nickel ore, as a major nickel mineral resource, is attracting increasing attention due to its hydrometallurgical process. The hydrometallurgical process of laterite nickel ore produces a large amount of smelting slag, of which iron oxide is a major component. CN113279048B discloses a method for preparing high-purity ferric phosphate from iron-containing slag. This patent uses iron-containing slag and high-concentration phosphoric acid as raw materials to prepare high-purity ferric phosphate. Although this patent can convert laterite nickel ore smelting slag into high-purity ferric phosphate, it requires the use of a large amount of high-concentration phosphoric acid, and the subsequent recovery and treatment of the phosphoric acid is extremely cumbersome.

[0003] The phosphating process in the electroplating industry generates a large amount of phosphating slag. The phosphating slag produced during zinc-manganese phosphating mainly consists of zinc phosphide and manganese phosphide. Chinese patent CN102593450B discloses a method for preparing multi-element doped lithium iron phosphate using phosphating waste slag as the main raw material, achieving high-value utilization of phosphating waste slag into lithium iron phosphate. However, this method requires acid washing to remove elements in the phosphating waste slag that adversely affect the electrochemical performance of lithium iron phosphate, resulting in a significant loss of iron.

[0004] If laterite nickel ore smelting slag and phosphating slag can be co-treated, not only can iron and phosphorus be efficiently recovered from these slags and phosphating slags, but the process can also be simplified and the use of high-concentration phosphoric acid can be avoided. Therefore, it is necessary to provide a method for the co-treatment of laterite nickel ore smelting slag and phosphating slag, and to prepare sodium iron phosphate cathode materials based on this method, so as to improve the added value of solid waste. Summary of the Invention

[0005] The purpose of this invention is to overcome the above-mentioned technical deficiencies and propose a combined treatment method for laterite nickel ore smelting slag and phosphating slag, thereby solving the technical problems of cumbersome steps and large iron loss in the separate treatment of laterite nickel ore smelting slag and phosphating slag in the prior art.

[0006] This invention provides a method for the combined treatment of laterite nickel ore smelting slag and phosphating slag, comprising the following steps:

[0007] Raw material mixing: The laterite nickel ore smelting slag, phosphating slag and sodium alkaline salt are mixed evenly to obtain a mixture;

[0008] Sodium reduction roasting: The mixture is subjected to sodium reduction roasting to obtain roasted material;

[0009] Water leaching: The roasted material is leached with water, and the leaching solution and water leaching residue are obtained by filtration;

[0010] Magnetic separation: The water-leached residue is subjected to magnetic separation to obtain iron concentrate; among which,

[0011] In the sodium reduction roasting step, the roasting temperature is 500-1000℃, the roasting time is 1-4h, and the roasting atmosphere is a reducing gas.

[0012] Compared with the prior art, the beneficial effects of the present invention include:

[0013] This invention involves mixing laterite nickel ore smelting slag and phosphating slag, followed by sodium reduction roasting. By controlling the roasting temperature and time, excessive iron reduction can be effectively avoided, ensuring that iron exists in the form of Fe3O4 while elements such as manganese and zinc exist in the slag as oxides. During roasting, aluminum oxides in the slag react with alkali to transform into water-soluble sodium aluminate, which can then be efficiently separated from the roasted material through a simple water leaching process. Simultaneously, zinc phosphide and manganese phosphide react with alkali to transform into sodium phosphate and their corresponding oxides, which also promotes the transformation of iron oxides from non-magnetic to magnetic. This allows for efficient iron recovery through a simple magnetic separation process. This invention achieves efficient separation and recovery of iron and phosphorus from waste slag through the synergistic treatment of the two types of slag. Attached Figure Description

[0014] Figure 1 This is a process flow diagram of one embodiment of the combined treatment method for laterite nickel ore smelting slag and phosphating slag provided by the present invention. Detailed Implementation

[0015] To make the objectives, technical solutions, and advantages of this invention clearer, the invention will be further described in detail below with reference to the accompanying drawings and embodiments. It should be understood that the specific embodiments described herein are merely illustrative and not intended to limit the invention.

[0016] Please see Figure 1 This invention provides a method for the combined treatment of laterite nickel ore smelting slag and phosphating slag, comprising the following steps:

[0017] Step S1, Raw material mixing: Mix laterite nickel ore smelting slag, phosphating slag and sodium alkaline salt evenly to obtain a mixture;

[0018] Step S2, Sodium Reduction Roasting: The mixture is subjected to sodium reduction roasting to obtain roasted material;

[0019] Step S3, water leaching: The roasted material is leached with water, and the leaching solution and water leaching residue are obtained by filtration;

[0020] Step S4, Magnetic Separation: The water-leached residue is subjected to magnetic separation to obtain iron concentrate; in the magnetic separation step, the magnetic part is iron concentrate, the main component of which is Fe3O4, and other impurity elements are less than 0.5%; the non-magnetic part is a mixture of zinc oxide, manganese oxide, gypsum and silicon dioxide.

[0021] The mechanism of this invention is as follows: the main components of laterite nickel ore smelting slag are iron oxide, aluminum oxide, calcium sulfate and silicon dioxide, etc., and the main components of phosphating slag are zinc phosphide, manganese phosphide and silicon dioxide, etc. In the sodium reduction roasting step, the reduction pathway of iron oxides in the waste residue is Fe2O3-Fe3O4-FeO-Fe. This invention, by mixing laterite nickel ore smelting slag and phosphating slag for sodium reduction roasting, and by controlling the roasting temperature and time, effectively avoids excessive iron reduction. Iron is preferentially reduced to metals before manganese and zinc, ensuring that iron exists in the form of Fe3O4 while ensuring that elements such as manganese and zinc exist in the slag in oxide form. During roasting, aluminum oxides in the slag react with alkali to transform into water-soluble sodium aluminate. A simple water leaching process can then achieve efficient separation of aluminum from the roasted material, reducing the Al content in the water-leached slag to below 0.05%. Simultaneously, zinc phosphide and manganese phosphide react with alkali to transform into sodium phosphate and their corresponding oxides, which also promotes the transformation of iron oxides from non-magnetic to magnetic. A simple magnetic separation process can then achieve efficient iron recovery. This invention achieves efficient separation and recovery of iron and phosphorus from waste residue through the synergistic treatment of two types of slag. Some of the reactions are as follows:

[0022] MnP+10.5Fe2O3+3NaOH=MnO+7Fe3O4+Na3PO4+1.5H2O

[0023] Zn3P2+24Fe2O3+6NaOH=3ZnO+2Na3PO4+3H2O+16Fe3O4

[0024] In this embodiment, the composition of the laterite nickel ore smelting slag includes: Fe 38-40%, Al 5-7%, Si 8-10%, Ca 5-8%, and the remaining metal elements such as Ni, Co, Zn, and Mg are all below 0.1%; the composition of the phosphating slag includes: Zn 15-18%, Mn 10-12%, Fe 2-3%, Ca 7-10%, Si 6-8%, P 12-15%, and the remaining metal elements are all below 0.1%.

[0025] In this embodiment, in step S1, the mass ratio of laterite nickel ore smelting slag to phosphating slag is 1:(0.5-5), including but not limited to 1:0.5, 1:1, 1:2, 1:3, 1:4, 1:5, etc.

[0026] In this embodiment, in step S1, the sodium alkaline salt is one or more of sodium hydroxide or sodium carbonate; the mass ratio of laterite nickel ore smelting slag to sodium alkaline salt is 1:(1-5), including but not limited to 1:1, 1:2, 1:3, 1:4, 1:5, etc.

[0027] In this embodiment, in step S2, the calcination temperature is 500-1000℃, including but not limited to 500℃, 600℃, 700℃, 800℃, 900℃, 1000℃, etc.; the calcination time is 1-4h, including but not limited to 1h, 1.5h, 2h, 2.5h, 3h, 4h, etc.; the calcination atmosphere is a reducing gas, such as one or a mixture of several of H2, CH4, CO, etc. The present invention does not limit the calcination atmosphere, and those skilled in the art can select it according to the actual situation; the calcination process is carried out in a muffle furnace.

[0028] In this embodiment, in step S3, room temperature leaching is used, and the solid-liquid ratio is 100-300 g / L, including but not limited to 100 g / L, 150 g / L, 200 g / L, 250 g / L, 300 g / L, etc., and the water immersion time is 0.5-3 h, including but not limited to 0.5 h, 1 h, 1.5 h, 2 h, 2.5 h, 3 h, etc.

[0029] In this embodiment, in step S4, the magnetic separation is completed by a magnetic separator with a magnetic field strength of 10000-15000GS, including but not limited to 10000GS, 11000GS, 12000GS, 13000GS, 14000GS, 15000GS, etc. The feeding speed of the magnetic separator is 0.5-1t / h, including but not limited to 0.5t / h, 0.6t / h, 0.7t / h, 0.8t / h, 0.9t / h, 1t / h, etc.

[0030] In this embodiment, after step S3, step S5, aluminum and silicon removal, and evaporation crystallization are further included: the aqueous leaching solution is subjected to aluminum removal, silicon removal, and evaporation crystallization sequentially to obtain sodium phosphate. By sequentially subjecting the aqueous leaching solution to precipitation for aluminum removal, silicon removal, and evaporation crystallization, efficient phosphorus recovery can be achieved. After the aluminum and silicon removal processes of this invention, the concentrations of aluminum and silicon ions in the silicon-removed solution are both below 1 ppm, and the purity of sodium phosphate obtained by evaporation crystallization of the silicon-removed solution is above 99.5%.

[0031] The aluminum removal process includes: adjusting the pH of the aqueous leaching solution to 5-8, including but not limited to 5, 6, 7, 8, etc., and filtering to obtain Al(OH)3 and the aluminum-removed solution; furthermore, in the aluminum removal process, the acid used to adjust the pH is phosphoric acid. This invention does not limit the concentration of phosphoric acid, and those skilled in the art can select it according to actual conditions; the silicon removal process includes: adjusting the pH of the aluminum-removed solution to 1-2, including but not limited to 1, 1.2, 1.4, 1.6, 1.8, 2, etc., and carrying out a heating reaction, followed by filtration to obtain silica slag and the aluminum-removed solution. The desiliconized liquid; furthermore, during the desiliconization process, the acid used to adjust the pH is phosphoric acid. This invention does not limit the concentration of phosphoric acid; those skilled in the art can select it according to actual conditions. The heating reaction temperature is 50-100℃, including but not limited to 50℃, 60℃, 70℃, 80℃, 90℃, 100℃, etc., and the heating reaction time is 0.5-3h, including but not limited to 0.5h, 1h, 1.5h, 2h, 2.5h, 3h, etc. The evaporation and crystallization process includes: evaporating and crystallizing the desiliconized liquid to obtain sodium phosphate. Further, during the evaporation and crystallization process, the desiliconized liquid is evaporated at 70-90℃, including but not limited to 70℃, 75℃, 80℃, 85℃, 90℃, etc., until the relative density is 1.2-1.3 g / cm³. 3 Evaporation ends at a certain time, and the solution is cooled to 15-25℃ to precipitate sodium phosphate crystals.

[0032] In this embodiment, after step S4, step S6, the preparation of sodium iron phosphate cathode material, is also included: the iron concentrate obtained in step S4 and the sodium phosphate obtained in step S5 are mixed, ball-milled, and then calcined at high temperature to obtain sodium iron phosphate cathode material. The inventors unexpectedly discovered during the disposal process that the recovered iron and phosphorus could be directly used to regenerate and prepare sodium iron phosphate, a higher value-added sodium-ion cathode material, thus improving the resource utilization rate of laterite nickel ore smelting slag and phosphating slag.

[0033] In step S6, a regulator may be added to adjust the proportion of P, so that the molar ratio of Fe:Na:P in the mixture is 1:(0.96-1):(1-1.1). Further, the regulator may be ammonium dihydrogen phosphate, diammonium hydrogen phosphate, or sodium carbonate, etc. This invention does not limit the choice of regulator, and those skilled in the art can select it according to the actual situation.

[0034] In step S6, during the ball milling process, a carbon source may be added. Adding a carbon source enhances the conductivity of sodium iron phosphate and reduces ferric iron in the iron concentrate to ferrous iron. Further, the carbon source is one or more of starch, glucose, or sucrose, and accounts for 5-10% of the total mass of sodium phosphate, iron concentrate, and regulator, including but not limited to 5%, 6%, 7%, 8%, 9%, and 10%.

[0035] In step S6, during the high-temperature roasting process, the roasting temperature is 650-850℃, including but not limited to 650℃, 700℃, 750℃, 800℃, 850℃, etc.; the roasting time is 10-20h, including but not limited to 10h, 12h, 14h, 16h, 18h, 20h, etc.; and the roasting atmosphere is an inert atmosphere.

[0036] In this embodiment, after step S4, step S7, acid leaching, is also included: zinc and manganese in the non-magnetic part are selectively leached with inorganic acid to obtain zinc-manganese enriched solution and zinc-manganese deslag. The zinc-manganese deslag can be used for brick making, and the zinc-manganese enriched solution can be further processed using existing conventional methods.

[0037] The inorganic acid used is a common inorganic acid, such as sulfuric acid, and this invention does not limit its use. In the acid leaching process, the concentration of the inorganic acid is 0.5-2 mol / L, including but not limited to 0.5 mol / L, 1 mol / L, 1.5 mol / L, and 2 mol / L; the temperature is 50-80℃, including but not limited to 50℃, 60℃, 70℃, and 80℃; the time is 1-3 hours, including but not limited to 1 hour, 1.5 hours, 2 hours, 2.5 hours, and 3 hours; and the solid-liquid ratio is 50-200 g / L, including but not limited to 50 g / L, 100 g / L, 150 g / L, and 200 g / L. Through the above acid leaching process, the zinc and manganese residues in the residue after leaching are both below 0.1%.

[0038] To avoid redundancy, the compositions of the laterite nickel ore smelting slag and phosphating slag in the following embodiments and comparative examples of the present invention are as follows:

[0039] The composition of laterite nickel ore smelting slag is: Fe 39.59%, Al 5.82%, Si 8.64%, Ca 6.88%, with other metal elements such as Ni, Co, Zn, and Mg all below 0.1%; the main composition of phosphating slag is: Zn 16.74%, Mn 11.65%, Fe 2.13%, Ca 9.57%, Si 7.50%, P 14.25%, with other metal elements all below 0.1%.

[0040] Example 1

[0041] (1) Raw material mixing: The laterite nickel ore smelting slag, phosphate slag and sodium hydroxide are mixed evenly in a mass ratio of 1:1:3 to obtain a mixture;

[0042] (2) Sodium reduction roasting: The mixture is placed in a muffle furnace and reduced gas H2 is introduced for sodium reduction roasting. After roasting, roasted material is obtained; the roasting temperature is 800℃ and the roasting time is 2h.

[0043] (3) Water leaching: The calcined material is leached with water, the solid-liquid ratio is controlled at 200 g / L, the leaching is carried out at room temperature and stirred for 1 hour, and then filtered to obtain water leaching liquid and water leaching residue; wherein, the aluminum content in the water leaching residue is less than 0.05%;

[0044] (4) Aluminum and silicon removal, evaporation and crystallization: Dilute phosphoric acid was added to the aqueous leaching solution to adjust the pH to 7, and the precipitate was obtained by filtration to obtain aluminum hydroxide and aluminum-removed solution. Dilute phosphoric acid was then added to the aluminum-removed solution to adjust the pH to 1.5, and the temperature was controlled at 80℃. After reacting for 0.5 hours, the solution was filtered to obtain silicon-removed solution and silicic acid residue. The concentrations of aluminum and silicon ions in the silicon-removed solution were both below 1 ppm. The silicon-removed solution was then evaporated at 80±2℃ until the relative density of the solution was 1.25±0.2 g / cm³. 3 Evaporation was stopped at a certain time, and the solution was cooled to 20±2℃ to precipitate sodium phosphate crystals (Na3PO4·12H2O) with a purity of 99.87%.

[0045] (5) Magnetic separation: The water-leached residue is subjected to magnetic separation with a magnetic field strength of 12000GS and a feed rate of 0.5t / h. The Fe3O4 content in the obtained iron concentrate is 99.73%, the impurity element content is less than 0.3%, and the Fe3O4 yield is 99.59%.

[0046] (6) Preparation of sodium iron phosphate cathode material: Sodium phosphate and iron concentrate were mixed, with ammonium dihydrogen phosphate used to adjust the phosphorus ratio so that the molar ratio of Na:Fe:P in the mixture was 1:1:1. Glucose was added at 5% of the mass of the mixture as a carbon source for high-temperature reduction calcination under an inert atmosphere at 750℃ for 12 hours to obtain NaFePO4 / C cathode material. A battery was prepared using the above NaFePO4 / C as cathode material, and its electrochemical performance was tested. The initial charge capacity at 0.1C was 142.5 mAh / g, and the initial discharge capacity was 140.3 mAh / g, indicating that the cathode material has good electrochemical performance.

[0047] (7) Acid leaching: The non-magnetic slag after magnetic separation is subjected to acid leaching process. The sulfuric acid concentration is 1 mol / L, the leaching solid-liquid ratio is 100 g / L, the leaching temperature is 70℃, and the leaching time is 1 h. The zinc and manganese content in the filter residue obtained by filtration is less than 0.1%, and the filtrate is a zinc-manganese enriched solution that can be treated using existing conventional technical means.

[0048] Examples 2-5

[0049] Examples 2-5 verify the effect of roasting temperature on the Fe3O4 content and yield in the iron concentrate obtained from the magnetic separation step during the sodium reduction roasting step. The only difference compared to Example 1 is the roasting temperature in the sodium reduction roasting step. The test results are shown in Table 1.

[0050] Table 1

[0051]

[0052] Examples 6-9

[0053] Examples 6-9 verify the effect of roasting time on the Fe3O4 content and yield in the iron concentrate obtained from the magnetic separation step during the sodium reduction roasting step. The only difference compared to Example 1 is the roasting time in the sodium reduction roasting step. The test results are shown in Table 2.

[0054] Table 2

[0055]

[0056] Comparative Example 1

[0057] Compared with Example 1, the only difference is: (1) Raw material mixing: the laterite nickel ore smelting slag, phosphating slag and sodium hydroxide are mixed evenly in a mass ratio of 1:0:3 to obtain a mixture; the remaining steps are the same as in Example 1, and will not be repeated here.

[0058] Using the method in Comparative Example 1, the Fe3O4 content in the iron concentrate obtained by the magnetic separation step was 70.33%, and the Fe3O4 yield was 68.55%.

[0059] The specific embodiments of the present invention described above do not constitute a limitation on the scope of protection of the present invention. Any other corresponding changes and modifications made in accordance with the technical concept of the present invention should be included within the scope of protection of the claims of the present invention.

Claims

1. A method for the combined treatment of laterite nickel ore smelting slag and phosphating slag, characterized in that, Includes the following steps: Raw material mixing: The laterite nickel ore smelting slag, phosphating slag and sodium alkaline salt are mixed evenly to obtain a mixture; Sodium reduction roasting: The mixture is subjected to sodium reduction roasting to obtain roasted material; Water leaching: The roasted material is leached with water, and the leaching solution and water leaching residue are obtained by filtration; Magnetic separation: The water-leached residue is subjected to magnetic separation to obtain iron concentrate; wherein, The main components of the laterite nickel ore smelting slag are iron oxide, aluminum oxide, calcium sulfate and silicon dioxide, and the main components of the phosphating slag are zinc phosphide, manganese phosphide and silicon dioxide. In the sodium reduction roasting step, the roasting temperature is 500-1000℃, the roasting time is 1-4h, and the roasting atmosphere is a reducing gas.

2. The method for combined treatment of laterite nickel ore smelting slag and phosphating slag according to claim 1, characterized in that, The composition of the laterite nickel ore smelting slag includes: Fe 38-40%, Al 5-7%, Si 8-10%, Ca 5-8%, with the remaining metal elements all below 0.1%; the composition of the phosphating slag includes: Zn 15-18%, Mn 10-12%, Fe 2-3%, Ca 7-10%, Si 6-8%, P 12-15%, with the remaining metal elements all below 0.1%.

3. The method for combined treatment of laterite nickel ore smelting slag and phosphating slag according to claim 1, characterized in that, In the raw material mixing step, the mass ratio of the laterite nickel ore smelting slag to the phosphating slag is 1:(0.5-5).

4. The method for combined treatment of laterite nickel ore smelting slag and phosphating slag according to claim 1, characterized in that, In the raw material mixing step, the sodium alkaline salt is one or more of sodium hydroxide or sodium carbonate, and the mass ratio of the laterite nickel ore smelting slag to the sodium alkaline salt is 1:(1-5).

5. The method for combined treatment of laterite nickel ore smelting slag and phosphating slag according to claim 1, characterized in that, In the sodium reduction roasting step, the roasting temperature is 600-1000℃ and the roasting time is 1-3h.

6. The method for combined treatment of laterite nickel ore smelting slag and phosphating slag according to claim 1, characterized in that, In the water leaching step, leaching is carried out at room temperature, with a solid-liquid ratio of 100-300 g / L and a leaching time of 0.5-3 h.

7. The method for combined treatment of laterite nickel ore smelting slag and phosphating slag according to claim 1, characterized in that, In the magnetic separation step, the magnetic separation is completed by a magnetic separator with a magnetic field strength of 10000-15000GS and a feeding speed of 0.5-1 t / h.

8. The method for combined treatment of laterite nickel ore smelting slag and phosphating slag according to claim 1, characterized in that, After the water leaching step, the process further includes aluminum and silicon removal and evaporation crystallization steps: the water leaching solution is sequentially subjected to aluminum removal, silicon removal, and evaporation crystallization to obtain sodium phosphate; wherein, The aluminum removal process includes: adjusting the pH of the aqueous leaching solution to 5-8, and filtering to obtain Al(OH)3 and the aluminum-removed solution; The silicon removal process includes: adjusting the pH of the aluminum removal solution to 1-2, heating the solution, and then filtering to obtain silica slag and silicon removal solution; The evaporation and crystallization process includes: evaporating and crystallizing the desiliconized liquid to obtain sodium phosphate.

9. The method for combined treatment of laterite nickel ore smelting slag and phosphating slag according to claim 8, characterized in that, In the processes of aluminum and silicon removal, phosphoric acid is used to adjust the pH; in the process of silicon removal, the heating reaction temperature is 50-100℃, and the heating reaction time is 0.5-3 hours; in the process of evaporation and crystallization, the silicon-removed liquid is evaporated at 70-90℃ to a relative density of 1.2-1.3 g / cm³. 3 Evaporation ends at a certain time, and the solution is cooled to 15-25℃ to precipitate sodium phosphate crystals.

10. The method for combined treatment of laterite nickel ore smelting slag and phosphating slag according to claim 8, characterized in that, After magnetic separation, the process also includes the preparation of sodium iron phosphate cathode material: the iron concentrate obtained by magnetic separation and the sodium phosphate obtained by evaporation and crystallization are mixed, ball-milled, and then calcined at high temperature to obtain sodium iron phosphate cathode material.

11. The method for combined treatment of laterite nickel ore smelting slag and phosphating slag according to claim 10, characterized in that, In the preparation steps of sodium iron phosphate cathode material, a regulator is added to control the proportion of P so that the molar ratio of Fe:Na:P in the mixture is 1:(0.96-1):(1-1.1).

12. The method for combined treatment of laterite nickel ore smelting slag and phosphating slag according to claim 11, characterized in that, In the preparation steps of the sodium iron phosphate cathode material, a carbon source is added during the ball milling and mixing process; wherein, The carbon source is one or more of starch, glucose or sucrose, and the carbon source accounts for 5-10% of the total mass of sodium phosphate, iron concentrate and regulator.

13. The method for combined treatment of laterite nickel ore smelting slag and phosphating slag according to claim 10, characterized in that, In the preparation steps of the sodium iron phosphate cathode material, the high-temperature calcination process involves a calcination temperature of 650-850℃, a calcination time of 10-20h, and an inert atmosphere.

14. The method for combined treatment of laterite nickel ore smelting slag and phosphating slag according to claim 1, characterized in that, After magnetic separation, the process also includes an acid leaching step: using inorganic acid to selectively leach zinc and manganese in the non-magnetic part to obtain a zinc-manganese enriched solution and zinc-manganese-removed residue.

15. The method for combined treatment of laterite nickel ore smelting slag and phosphating slag according to claim 14, characterized in that, In the acid leaching step, the concentration of inorganic acid is 0.5-2 mol / L, the temperature is 50-80℃, the time is 1-3 h, and the solid-liquid ratio is 50-200 g / L.

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