Method for synchronously recovering iron and rare earth from iron separation tailings

Abstract

The invention provides a method for synchronously recovering iron and rare earth from iron separation tailings, which comprises the following steps of: roasting the pretreated iron separation tailings in a graphite crucible to obtain a roasted product; the roasted product is crushed and screened, oversize products and undersize products are obtained, and the oversize products are reduced iron products; the screen underflow and an acid solution are mixed for leaching, and leaching liquid containing rare earth elements and acid leaching residues are obtained; and adding a precipitant into the leachate to obtain a precipitate containing rare earth elements, and recovering fluorite concentrate from the acid leaching residues. According to the method, the traditional disjunction mode of pre-enrichment and then smelting is broken through, a roasting-acid leaching integrated process with deep beneficiation and smelting coupling is provided, the complex pre-enrichment link is abandoned, efficient and synchronous recovery of iron and rare earth is achieved, and the method has the advantages of being short in process, low in energy consumption, few in flotation reagent, high in rare earth recovery rate, high in resource comprehensive utilization rate and the like.

Description

Technical Field

[0001] This invention relates to the field of comprehensive utilization technology of mineral resources, and in particular to a method for simultaneously recovering iron and rare earth elements from iron ore tailings. Background Technology

[0002] The Bayan Obo mine is a unique, super-large polymetallic deposit in my country, containing iron, rare earth elements, niobium, and other associated minerals. The tailings produced after iron ore beneficiation remain a highly valuable secondary resource. These tailings not only contain a considerable grade of iron (Fe) but are also rich in high-value rare earth elements (REO grade as high as 5-15%), and are associated with strategic and scarce minerals such as fluorite and niobium. Failure to efficiently and comprehensively utilize these tailings would not only result in a huge waste of precious resources but also pose environmental and safety risks. Therefore, developing economically efficient integrated recovery technologies is of great strategic significance for realizing the resource utilization, high-value utilization, and sustainable development of Bayan Obo tailings. Currently, researchers have proposed several technical solutions for the comprehensive utilization of Bayan Obo iron ore tailings, which can be broadly divided into two categories based on their core technical routes.

[0003] One type is the combined separation process based on physical beneficiation. This type of process mainly relies on physical methods such as flotation, magnetic separation, and gravity separation for multi-element separation. For example, CN104096633A discloses a process for comprehensively recovering multiple concentrates such as rare earth, iron, sulfur, and niobium from rare earth tailings through multi-stage flotation, magnetic separation, and gravity separation; CN103263977A proposes a flotation-magnetic separation-flotation process to recover iron and rare earth from weakly magnetically separated tailings. The advantage of this type of method is that it avoids high-temperature processes and has relatively low energy consumption. However, its core drawback is that, due to the fine particle size and complex symbiotic relationships of minerals in the tailings, it is difficult to achieve efficient liberation and separation by relying solely on physical separation. This results in an extremely lengthy process flow, the consumption of large amounts of complex reagents, and unsatisfactory final concentrate grade and recovery rate (especially the rare earth recovery rate, which is mostly only around 70%), and minimal recovery effect on difficult-to-separate elements such as niobium.

[0004] Secondly, a combined beneficiation and smelting process involving roasting and activation has been introduced. To improve recovery efficiency, more processes have incorporated roasting as an activation or conversion method. For example, CN108580031A employs a combined process of "fluidized magnetic roasting - weak magnetic separation - strong magnetic separation - rare earth / fluorite flotation - acid washing" to synergistically recover iron, rare earth elements, fluorite, and niobium; CN104611541A extracts iron and rare earth elements through "coal-based calcium-coated roasting - magnetic separation - ammonium sulfate roasting - water leaching"; and CN115625031A utilizes "hydrogen-based mineral phase conversion - magnetic separation - reverse flotation defluorination - rare earth flotation" to separate iron and rare earth elements. These methods alter the mineral phases through roasting, thereby improving the recovery rate of iron and rare earth elements to some extent. However, they generally have more prominent problems: (1) The process is more complex, often involving multiple roasting and beneficiation operations, resulting in high equipment investment and operating costs; (2) The energy consumption is huge, and it usually requires the addition of solid reducing agents (such as coal powder and coke), which leads to complex operation and may introduce impurities; (3) They have not yet broken away from the inherent idea of ​​"mineral beneficiation pre-enrichment". The recovery of rare earth still requires obtaining rare earth concentrate through flotation and other processes before subsequent processing, which fails to achieve direct extraction, resulting in a long process and low efficiency.

[0005] In summary, the fundamental deficiency of both simple physical beneficiation and combined beneficiation and smelting processes involving roasting lies in the severe disconnect between the "beneficiation" and "smelting" stages. For rare earth recovery, almost without exception, a complex pre-enrichment process (such as multi-stage flotation or roasting-weak magnetic separation) is required to obtain rare earth concentrate, which is then used as smelting raw material. This pre-enrichment process is the key bottleneck causing the entire process to be lengthy, energy and reagent-intensive, and resulting in low metal recovery rates (especially low rare earth recovery and difficulty in recovering niobium). It increases losses in intermediate stages, raises costs, and significantly reduces the economic viability and feasibility of many technologies.

[0006] Therefore, a new approach and process are urgently needed to break the traditional model of "obtaining concentrate before smelting." Developing a short-process technology that deeply couples beneficiation and smelting, greatly simplifies the process, aims to simultaneously and efficiently recover iron elements and bypass the pre-enrichment step to directly extract rare earth elements from tailings through acid leaching. This has become an inevitable direction and urgent need to overcome existing technological bottlenecks and significantly improve the resource value and economic efficiency of iron ore tailings. Summary of the Invention

[0007] In view of the shortcomings of existing technologies for recovering valuable elements from iron ore tailings, such as lengthy process flow, high reagent consumption, and disconnect between beneficiation and metallurgy, this invention aims to provide a new combined process with a shorter process flow, lower reagent consumption, and deep coupling between beneficiation and metallurgy.

[0008] To achieve this objective, the present invention adopts the following technical solution:

[0009] This invention provides a method for simultaneously recovering iron and rare earth elements from iron ore tailings, comprising the following steps:

[0010] (1) The pretreated iron tailings are placed in a graphite crucible and roasted using the reducing atmosphere provided by the graphite crucible at high temperature to obtain roasted products.

[0011] (2) The roasted product is crushed and sieved to obtain the oversize and undersize products, wherein the oversize product is a reduced iron product;

[0012] (3) The sieved material is mixed with an acid solution for leaching to obtain a leaching solution containing rare earth elements and an acid leaching residue;

[0013] (4) Add a precipitant to the leachate to obtain a precipitate containing rare earth elements, and recover fluorite concentrate from the acid leaching residue.

[0014] This invention utilizes a core "roasting-acid leaching" process. First, iron tailings are roasted at high temperatures to reduce the iron in the iron ore to metallic iron. The difference in toughness between metallic iron and its oxides allows for the sieving and recovery of the metallic iron. Simultaneously, the high-temperature roasting effectively activates rare earth minerals, enabling them to be directly and efficiently leached with acid without subsequent pre-enrichment. This method eliminates the complex multi-stage flotation-magnetic separation pre-enrichment stage of traditional processes, achieving efficient iron recovery and direct acid leaching extraction of rare earths, thereby significantly improving the economy and efficiency of the entire resource recovery process.

[0015] It is worth noting that the present invention uses a graphite crucible for calcination, which is not only a simple choice of container, but also the key to achieving efficient reduction. The graphite crucible itself can act as a reducing agent in the reaction at high temperatures: firstly, it can consume oxygen in the environment and generate strong reducing gases such as carbon monoxide, providing a stable, endogenous reducing atmosphere for the system, greatly promoting the conversion of iron oxides to metallic iron; secondly, it can effectively isolate external air, creating and maintaining a necessary oxygen-free or low-oxygen environment for the reduction reaction. This is fundamentally different from the traditional process that uses ordinary ceramic crucibles and requires the addition of a large amount of solid reducing agents (such as coal powder or coke), simplifying the process and reducing the risk of impurity introduction. As a preferred technical solution of the present invention, the pretreatment in step (1) includes a refining treatment, which includes any one or a combination of at least two of grinding, ball milling, crushing, tapping, and impact.

[0016] As a preferred technical solution of the present invention, the proportion of powder with a particle size of no more than 75 μm in the iron tailings after the refinement treatment is more than 90 wt%, such as 90 wt%, 91 wt%, 92 wt%, 93 wt%, 94 wt%, 95 wt%, 96 wt%, 97 wt%, 98 wt%, 99 wt%, 100 wt%, etc., but not limited to the listed values, other unlisted values ​​within this range are also applicable.

[0017] As a preferred technical solution of the present invention, the iron tailings include the Bayan Obo iron tailings, whose main useful minerals include hematite, magnetite, pyrite, bastnaesite, monazite and fluorite, and whose main gangue minerals are dolomite, quartz and apatite.

[0018] As a preferred technical solution of the present invention, the TFe grade in the iron tailings is 15-40 wt%, such as 15 wt%, 20 wt%, 25 wt%, 30 wt%, 35 wt%, 40 wt%, and the REO grade is 2-15 wt%, such as 2 wt%, 5 wt%, 7 wt%, 9 wt%, 11 wt%, 13 wt%, 15 wt%, etc., but is not limited to the listed values. Other unlisted values ​​within this range are also applicable.

[0019] As a preferred technical solution of the present invention, the roasting temperature in step (2) is 1200-1400 ℃, such as 1200 ℃, 1210 ℃, 1220 ℃, 1230 ℃, 1240 ℃, 1250 ℃, 1260 ℃, 1270 ℃, 1280 ℃, 1290 ℃, 1300 ℃, 1310 ℃, 1320 ℃, 1330 ℃, 1340 ℃, 1350 ℃, 1360 ℃, 1370 ℃, 1380 ℃, 1390 ℃, 1400 ℃, etc., but is not limited to the listed values. Other unlisted values ​​within this range are also applicable.

[0020] This invention reduces iron minerals to metallic iron through high-temperature roasting, while simultaneously activating rare earth minerals, disrupting their stable crystal structure, and transforming them into a phase that is more easily leached by acid, allowing them to be directly and efficiently leached by acid without the need for subsequent pre-enrichment.

[0021] If the reduction roasting temperature is too low, on the one hand, the iron minerals will not be fully reduced and will be difficult to completely convert into metallic iron; on the other hand, the crystallization and growth of iron particles will be restricted, which is not conducive to effective separation by sieving. Conversely, if the roasting temperature is too high, rare earth minerals may undergo phase transformation, forming a stable phase that is highly resistant to acid leaching, thereby weakening their activation effect and reducing the leaching efficiency of rare earths.

[0022] As a preferred technical solution of the present invention, the reduction calcination time is 60-180 min, such as 60 min, 70 min, 80 min, 90 min, 100 min, 110 min, 120 min, 130 min, 140 min, 150 min, 160 min, 170 min, 180 min, etc., but is not limited to the listed values. Other unlisted values ​​within this range are also applicable.

[0023] In the high-temperature reduction roasting process, the coordinated control of roasting temperature and time is crucial, and is the key to simultaneously achieving high iron recovery rate and high rare earth leaching rate. If the roasting time is too short, the iron reduction reaction will be incomplete, and the iron particles will not grow sufficiently, affecting their subsequent physical sorting; while if the roasting time is too long, it may promote the transformation of rare earth elements into acid-insoluble phases, which is also not conducive to the leaching of rare earths.

[0024] As a preferred embodiment of the present invention, the calcination is carried out in a graphite crucible.

[0025] As a preferred technical solution of the present invention, the crushing method in step (2) includes impact crushing, grinding crushing, crusher crushing, and extrusion equipment crushing.

[0026] It is worth noting that the metal iron can be screened out in step (2) based on the difference between the toughness of metal iron and the brittleness of oxides. During the crushing process, metal iron is not easy to be over-crushed and is mostly enriched in the coarse particles. Then, after screening, the reduced iron concentrate can be obtained.

[0027] As a preferred technical solution of the present invention, the proportion of powder with a particle size of no more than 75 μm in the sieve in step (3) is 95 wt% or more, such as 95 wt%, 96 wt%, 97 wt%, 98 wt%, 99 wt%, 100 wt%, etc., but not limited to the listed values. Other unlisted values ​​within this range are also applicable.

[0028] As a preferred technical solution of the present invention, the acid solution in step (3) is any one or a combination of at least two of hydrochloric acid, sulfuric acid or nitric acid.

[0029] As a preferred embodiment of the present invention, the concentration of the acid solution is 1-4 mol / L, such as 1.0 mol / L, 1.4 mol / L, 1.8 mol / L, 2.2 mol / L, 2.6 mol / L, 3.0 mol / L, 3.4 mol / L, 3.8 mol / L, 4.0 mol / L, etc., but is not limited to the listed values. Other unlisted values ​​within this range are also applicable.

[0030] As a preferred technical solution of the present invention, the acid leaching temperature in step (3) is 40-90 ℃, such as 40 ℃, 45 ℃, 50 ℃, 55 ℃, 60 ℃, 65 ℃, 70 ℃, 75 ℃, 80 ℃, 85 ℃, 90 ℃, etc., but is not limited to the listed values. Other unlisted values ​​within this range are also applicable.

[0031] As a preferred technical solution of the present invention, the liquid-solid ratio of the acid solution to the sieved material is (6-12):1 mL / g, such as 6:1 mL / g, 6.5:1 mL / g, 7:1 mL / g, 7.5:1 mL / g, 8:1 mL / g, 8.5:1 mL / g, 9:1 mL / g, 9.5:1 mL / g, 10:1 mL / g, 10.5:1 mL / g, 11:1 mL / g, 11.5:1 mL / g, 12:1 mL / g, etc., but is not limited to the listed values. Other unlisted values ​​within this range are also applicable.

[0032] As a preferred technical solution of the present invention, the acid leaching time is 30-120 min, such as 30 min, 40 min, 50 min, 60 min, 70 min, 80 min, 90 min, 100 min, 110 min, 120 min, etc., but is not limited to the listed values. Other unlisted values ​​within this range are also applicable.

[0033] By controlling the leaching conditions, rare earth elements can be selectively dissolved into the solution, while most gangue minerals (such as fluorite and quartz) remain in the residue. Insufficient leaching time leads to incomplete reaction between the acid and rare earth minerals, resulting in a decrease in the rare earth leaching rate; while excessive leaching time may cause the dissolution of a large number of impurity elements, which not only increases the burden of subsequent purification but also leads to higher operating costs, thus negatively impacting the overall rare earth recovery efficiency.

[0034] As a preferred technical solution of the present invention, the precipitant in step (4) includes any one or a combination of at least two of oxalic acid, carbonate, hydroxide or fluoride.

[0035] Preferably, the carbonate includes any one or a combination of at least two of ammonium bicarbonate, sodium carbonate, or ammonium carbonate.

[0036] Preferably, the hydroxide includes ammonia.

[0037] As a preferred embodiment of the present invention, the concentration of the precipitant is 0.5-3 mol / L, such as 0.5 mol / L, 0.7 mol / L, 0.9 mol / L, 1.1 mol / L, 1.3 mol / L, 1.5 mol / L, 1.7 mol / L, 1.9 mol / L, 2.1 mol / L, 2.3 mol / L, 2.5 mol / L, 2.7 mol / L, 2.9 mol / L, 3.0 mol / L, etc., but is not limited to the listed values; other unlisted values ​​within this range are also applicable.

[0038] Preferably, the recovery of fluorite concentrate includes the use of flotation.

[0039] As a preferred technical solution of the present invention, the precipitation temperature is 40-80 ℃, such as 40 ℃, 45 ℃, 50 ℃, 55 ℃, 60 ℃, 65 ℃, 70 ℃, 75 ℃, 80 ℃, etc., but is not limited to the listed values. Other unlisted values ​​within this range are also applicable.

[0040] In one specific embodiment of the present invention, the method for simultaneously recovering iron and rare earth elements from iron ore tailings includes the following steps:

[0041] (1) Take the powder with a particle size of no more than 75 μm in the pretreated iron tailings, accounting for more than 90 wt%, and reduce roast at a roasting temperature of 1200-1400 ℃ for 60-180 min to obtain the roasted product.

[0042] (2) The obtained roasted product is crushed and sieved to obtain the oversize and undersize. The undersize has a particle size of no more than 75 μm and the proportion of powder is more than 95 wt%. The oversize is a reduced iron product.

[0043] (3) Mix the sieve material with an acid solution with a concentration of 1-4 mol / L, wherein the liquid-solid ratio of the acid solution to the sieve material is (6-12):1 mL / g, and stir continuously for 30-120 min at an acid leaching temperature of 40-90 ℃ to obtain a leaching solution containing rare earth elements and an acid leaching residue.

[0044] (4) Add a precipitant to the leachate with a concentration of 0.5-3 mol / L, react at a temperature of 40-80 ℃ for 30-90 min to precipitate rare earth elements, and calcine at 800-1000 ℃ for 50-70 min to obtain mixed rare earth oxides, and recover fluorite concentrate from the acid leaching residue.

[0045] In this invention, key process parameters such as roasting temperature and time, and liquid-solid ratio and time in acid leaching must be systematically optimized and coordinated to efficiently reduce iron minerals to easily sortable metallic iron and simultaneously fully activate rare earth minerals, so that they can be directly and efficiently leached by acid without pre-enrichment, thereby achieving the process goal of simultaneous recovery of iron and rare earths.

[0046] Compared with the prior art, the present invention has at least the following beneficial effects:

[0047] (1) The process is greatly simplified, achieving a truly “short process”: This invention eliminates the complex multi-stage flotation-magnetic separation pre-enrichment stage in the traditional process through the core path of “roasting-acid leaching”. The roasting reduction recovery of iron and the wet acid leaching extraction of rare earth are seamlessly connected, making the process flow very simple;

[0048] (2) Energy consumption and flotation reagent consumption are significantly reduced: The large amount of inhibitors, collectors and modifiers required for flotation are avoided, and the reagent cost is greatly reduced. At the same time, the shortened process also reduces the corresponding equipment operating energy consumption. Secondly, by utilizing the graphite crucible itself to provide a reducing atmosphere, the addition of additional solid reducing agents is reduced or even avoided, further reducing the raw material cost and the burden of subsequent impurity removal.

[0049] (3) High rare earth recovery rate and direct extraction: The reduction roasting process effectively activates rare earth minerals, destroys their stable crystal structure, and allows them to be directly and efficiently leached by acid without subsequent pre-enrichment. It is expected to increase the rare earth recovery rate to a level far higher than that of traditional flotation processes (>85%).

[0050] (4) High degree of comprehensive utilization of resources: This invention not only efficiently recovers iron and rare earth, but also uses acid leaching tailings as raw material for recovering fluorite, realizing the synergistic recovery of three major valuable elements: iron, rare earth and fluorite, which meets the requirements of green circular economy.

[0051] (5) Laying the foundation for solving the problem of niobium recovery: After the acid leaching residue recovers fluorite, valuable elements such as niobium may be further enriched, creating favorable conditions for subsequent economical and effective recovery of niobium. Attached Figure Description

[0052] Figure 1 This is a flowchart illustrating the specific process flow of the method for simultaneously recovering iron and rare earth elements from iron ore tailings according to the present invention. Detailed Implementation

[0053] To better explain the present invention, it will be described in detail below with reference to the accompanying drawings and specific embodiments. It should be understood that the specific embodiments described herein are merely illustrative and not intended to limit the scope of the invention.

[0054] Unless otherwise specified, the experimental methods described in the following examples are conventional methods; unless otherwise specified, the reagents and materials are all available from commercial sources.

[0055] The following method for simultaneously recovering iron and rare earth elements from iron ore tailings is illustrated in the flowchart below. Figure 1 .

[0056] Example 1

[0057] This embodiment provides a method for simultaneously recovering iron and rare earth elements from iron ore tailings. The iron ore tailings are rare earth-containing iron ore tailings produced after iron ore beneficiation at Baotou Steel Beneficiation Plant. Their main components and contents are: iron (TFe) grade of 17.13% and rare earth (RE) grade of 7.46%. The main useful minerals are hematite, magnetite, pyrite, bastnaesite, monazite, and fluorite, while the main gangue minerals are dolomite, quartz, and apatite.

[0058] The method includes the following steps:

[0059] (1) The powder with a particle size of no more than 75 μm in the pretreated iron tailings, accounting for 92.88 wt%, was placed in a graphite crucible and reduced-roasted at a roasting temperature of 1300 ℃ for 120 min to obtain the roasted product.

[0060] (2) The obtained roasted product is coarsely crushed by a jaw crusher and then finely crushed by an extrusion device. The crushed material is screened to obtain the oversize and undersize. The undersize has a particle size of no more than 75 μm and the proportion of powder is more than 95 wt%. The oversize is reduced iron product.

[0061] (3) Mix the sieve material with a hydrochloric acid solution with a concentration of 2 mol / L, wherein the liquid-solid ratio of the acid solution to the sieve material is 10:1 mL / g, and stir continuously for 60 min at an acid leaching temperature of 70 ℃ to obtain a leaching solution containing rare earth elements and an acid leaching residue.

[0062] (4) Add oxalic acid to the leachate at a concentration of 1 mol / L, react at 50 ℃ for 30 min to precipitate rare earth elements, and calcine at 900 ℃ for 60 min to obtain mixed rare earth oxides. Recover fluorite concentrate from the acid leaching residue. The fluorite grade in the acid leaching residue is 55%, which can be further recovered by flotation.

[0063] Example 2

[0064] This embodiment provides a method for simultaneously recovering iron and rare earth elements from iron ore tailings. The iron ore tailings are rare earth-containing iron ore tailings produced after iron ore beneficiation at Baotou Steel Beneficiation Plant. Their main components and contents are: iron (TFe) grade 27.15%, rare earth (RE) grade 2.37%. The main useful minerals are hematite, magnetite, pyrite, bastnaesite, monazite, and fluorite, while the main gangue minerals are dolomite, quartz, and apatite.

[0065] The method includes the following steps:

[0066] (1) The powder with a particle size of no more than 75 μm in the pretreated iron tailings, accounting for 94.38 wt%, was placed in a graphite crucible and reduced-roasted at a roasting temperature of 1200 ℃ for 180 min to obtain the roasted product.

[0067] (2) The obtained roasted product is coarsely crushed by a jaw crusher and then finely crushed by an extrusion device. The crushed material is screened to obtain the oversize and undersize. The undersize has a particle size of no more than 75 μm and the proportion of powder is more than 95 wt%. The oversize is reduced iron product.

[0068] (3) Mix the sieve material with a hydrochloric acid solution with a concentration of 1 mol / L, wherein the liquid-solid ratio of the acid solution to the sieve material is 10:1 mL / g, and stir continuously for 90 min at an acid leaching temperature of 60 ℃ to obtain a leaching solution containing rare earth elements and an acid leaching residue.

[0069] (4) Add oxalic acid to the leachate. The concentration of oxalic acid used is 0.5 mol / L. React at 80 ℃ for 30 min to precipitate rare earth elements. Calcinate at 900 ℃ for 60 min to obtain mixed rare earth oxides. Recover fluorite concentrate from the acid leaching residue. The fluorite grade in the acid leaching residue is 53%, which can be further recovered by flotation.

[0070] Example 3

[0071] This embodiment provides a method for simultaneously recovering iron and rare earth elements from iron ore tailings. The iron ore tailings are rare earth-containing iron ore tailings produced after iron ore beneficiation at Baotou Steel Beneficiation Plant. Their main components and contents are: iron (TFe) grade 39.24%, rare earth (RE) grade 14.58%. The main useful minerals are hematite, magnetite, pyrite, bastnaesite, monazite, and fluorite, while the main gangue minerals are dolomite, quartz, and apatite.

[0072] The method includes the following steps:

[0073] (1) 91.05 wt% of the pretreated iron tailings with a particle size of no more than 75 μm were placed in a graphite crucible and reduced and roasted at a roasting temperature of 1300 ℃ for 60 min to obtain the roasted product.

[0074] (2) The obtained roasted product is coarsely crushed by a jaw crusher and then finely crushed by an extrusion device. The crushed material is screened to obtain the oversize and undersize. The undersize has a particle size of no more than 75 μm and the proportion of powder is more than 95 wt%. The oversize is reduced iron product.

[0075] (3) Mix the sieve material with a hydrochloric acid solution with a concentration of 3 mol / L, wherein the liquid-solid ratio of the acid solution to the sieve material is 6:1 mL / g, and stir continuously for 30 min at an acid leaching temperature of 90 ℃ to obtain a leaching solution containing rare earth elements and an acid leaching residue.

[0076] (4) Add ammonia water to the leachate. The concentration of ammonia water used is 1.5 mol / L. React at 60 ℃ for 60 min to precipitate rare earth elements. Calcinate at 900 ℃ for 60 min to obtain mixed rare earth oxides. Recover fluorite concentrate from the acid leaching residue. The fluorite grade in the acid leaching residue is 56%, which can be further recovered by flotation.

[0077] Example 4

[0078] This embodiment provides a method for simultaneously recovering iron and rare earth from iron ore tailings. The iron ore tailings are rare earth-containing iron ore tailings produced after iron ore beneficiation at Baogang Iron and Steel Plant, and their main components and contents are the same as in Example 1.

[0079] The method includes the following steps:

[0080] (1) The powder with a particle size of no more than 75 μm in the pretreated iron tailings, accounting for 93.46 wt%, was placed in a graphite crucible and reduced and roasted at a roasting temperature of 1400 ℃ for 60 min to obtain the roasted product.

[0081] (2) The obtained roasted product is coarsely crushed by a jaw crusher and then finely crushed by an extrusion device. The crushed material is screened to obtain the oversize and undersize. The undersize has a particle size of no more than 75 μm and the proportion of powder is more than 95 wt%. The oversize is reduced iron product.

[0082] (3) Mix the sieve material with a hydrochloric acid solution with a concentration of 3 mol / L, wherein the liquid-solid ratio of the acid solution to the sieve material is 12:1, and stir continuously at 40 °C for 120 min to obtain a leaching solution containing rare earth elements and an acid leaching residue.

[0083] (4) Add ammonium bicarbonate to the leachate. The concentration of ammonium bicarbonate used is 1 mol / L. React at 40 ℃ for 90 min to precipitate rare earth elements. Calcinate at 900 ℃ for 60 min to obtain mixed rare earth oxides. Recover fluorite concentrate from the acid leaching residue. The fluorite grade in the acid leaching residue is 56%, which can be further recovered by flotation.

[0084] Example 5

[0085] This embodiment provides a method for simultaneously recovering iron and rare earth from iron ore tailings. The raw materials and methods are the same as in Embodiment 1, except that: in step (1), the roasting temperature is adjusted to 1100 ℃.

[0086] Example 6

[0087] This embodiment provides a method for simultaneously recovering iron and rare earth from iron ore tailings. The raw materials and methods are the same as in Embodiment 1, except that: in step (1), the roasting temperature is adjusted to 1500 ℃.

[0088] Example 7

[0089] This embodiment provides a method for simultaneously recovering iron and rare earth from iron ore tailings. The raw materials and methods are the same as in Embodiment 1, except that the roasting time in step (1) is adjusted to 50 min.

[0090] Example 8

[0091] This embodiment provides a method for simultaneously recovering iron and rare earth from iron ore tailings. The raw materials and methods are the same as in Embodiment 1, except that the roasting time in step (1) is adjusted to 190 min.

[0092] Example 9

[0093] This embodiment provides a method for simultaneously recovering iron and rare earth from iron tailings. The raw materials and methods are the same as in Embodiment 1, except that in step (1), the liquid-solid ratio of the acid solution and the undersize material is adjusted to 5:1 mL / g.

[0094] Example 10

[0095] This embodiment provides a method for simultaneously recovering iron and rare earth from iron tailings. The raw materials and methods are the same as in Embodiment 1, except that in step (1), the liquid-solid ratio of the acid solution and the undersize material is adjusted to 13:1 mL / g.

[0096] Comparative Example 1

[0097] This embodiment provides a method for simultaneously recovering iron and rare earth elements from iron ore tailings. The raw materials and methods are the same as in Embodiment 1, except that an alumina crucible is used for roasting.

[0098] Performance testing

[0099] The total iron content in reduced iron was determined according to the method in GB / T223.7-2002, and the fluorite content in the acid leaching residue was determined according to the EDTA titration method in GB / T5195.1-2017. Other elements were analyzed by ICP. Based on the measurement data, the iron content and recovery rate, rare earth leaching rate and recovery rate, and fluorite content in the reduced iron of Examples 1-10 and Comparative Example 1 were calculated, and the results are shown in Table 1.

[0100] Table 1

[0101]

[0102] In the table, "-" indicates that there is no data.

[0103] A comprehensive comparison of Examples 1 and 5-10 shows that the present invention achieves efficient and simultaneous recovery of iron and rare earth elements through system optimization and coordinated control of key process parameters in the "roasting-acid leaching" process. Example 1, under optimized conditions (roasting temperature 1300 ℃, time 120 min, acid leaching liquid-to-solid ratio 10:1 mL / g), achieved excellent results with an iron recovery rate of 91.05% and a rare earth recovery rate of 95.68%.

[0104] (1) Effect of roasting temperature: In Example 5, the iron minerals were not fully reduced due to the low temperature, resulting in a significant decrease in the reduced iron grade and fluorite grade; In Example 6, although the iron grade was slightly higher, the rare earth leaching rate was slightly lower, indicating that excessively high temperature may cause rare earth minerals to undergo phase transformations that are not conducive to acid leaching, which confirms the necessity of controlling the roasting temperature between 1200-1400 ℃.

[0105] (2) Effect of roasting time: The rare earth recovery rate in Examples 7 and 8 remained at a high level, indicating that within the range of 60-180 min, the time variation had little effect on rare earth activation. However, too short or too long a time may still affect the full growth of iron particles or cause phase transformation of rare earth. 60-180 min is preferred.

[0106] (3) Effect of liquid-solid ratio in acid leaching: The rare earth leaching rates of Examples 9 and 10 were slightly lower than those of Example 1, indicating that an excessively low liquid-solid ratio may lead to insufficient leaching reaction, while an excessively high ratio may introduce too many impurities or increase the burden of subsequent processing. This further verifies the rationality of controlling the liquid-solid ratio at (6-12):1 mL / g.

[0107] In summary, the preferred combination of process parameters represented by Example 1 achieves synergistic and efficient recovery of three valuable elements—iron, rare earth, and fluorite—while ensuring efficient iron reduction and full activation of rare earth elements, fully demonstrating the technical advantages of the present invention: "short process and high efficiency."

[0108] A comprehensive comparison of Example 1 and Comparative Example 1 reveals that the present invention utilizes a graphite crucible to achieve integrated reduction and activation through endogenous reduction. This makes it possible to simultaneously complete the reduction of iron and the activation of rare earth elements in a single calcination step, fundamentally avoiding the need for an additional reducing agent pretreatment system and establishing the technical advantages of a short process and low energy consumption. The above description is merely a preferred embodiment of the present invention and is not intended to limit the patent scope of the present invention. Therefore, all equivalent structural changes made based on the content of this specification are included within the protection scope of the present invention.

Claims

1. A method for simultaneously recovering iron and rare earth elements from iron ore tailings, characterized in that, The method includes the following steps: (1) The pretreated iron tailings are placed in a graphite crucible and roasted using the reducing atmosphere provided by the graphite crucible at high temperature to obtain roasted products. (2) The roasted product is crushed and sieved to obtain the oversize and undersize products, wherein the oversize product is a reduced iron product; (3) The sieved material is mixed with an acid solution for leaching to obtain a leaching solution containing rare earth elements and an acid leaching residue; (4) Add a precipitant to the leachate to obtain a precipitate containing rare earth elements, and recover fluorite concentrate from the acid leaching residue.

2. The method according to claim 1, characterized in that, The preprocessing described in step (1) includes refinement processing; Preferably, the refining process includes grinding; Preferably, the proportion of powder with a particle size of no more than 75 μm in the iron tailings after the refining treatment is more than 90 wt%.

3. The method according to claim 1 or 2, characterized in that, The iron tailings have a TFe grade of 15-40 wt% and an REO grade of 2-15 wt%.

4. The method according to any one of claims 1-3, characterized in that, The roasting temperature in step (1) is 1200-1400 ℃; Preferably, the roasting time is 60-180 min; Preferably, the calcination is carried out in a graphite crucible.

5. The method according to any one of claims 1-4, characterized in that, The proportion of powder with a particle size of no more than 75 μm in the sieve material in step (2) is more than 95 wt%.

6. The method according to any one of claims 1-5, characterized in that, The acid solution mentioned in step (3) is any one or a combination of at least two of hydrochloric acid, sulfuric acid, or nitric acid; Preferably, the concentration of the acid solution is 1-4 mol / L.

7. The method according to any one of claims 1-6, characterized in that, The leaching temperature in step (3) is 40-90 ℃; Preferably, the liquid-to-solid ratio of the acid solution to the sieved material is (6-12):1 mL / g; Preferably, the leaching time is 30-120 min.

8. The method according to any one of claims 1-7, characterized in that, The precipitant mentioned in step (4) includes any one or a combination of at least two of oxalic acid, carbonate, hydroxide or fluoride; Preferably, the carbonate includes any one or a combination of at least two of ammonium bicarbonate, sodium carbonate, or ammonium carbonate; Preferably, the hydroxide includes ammonia.

9. The method according to any one of claims 1-8, characterized in that, The concentration of the precipitant is 0.5-3 mol / L; Preferably, the recovery of fluorite concentrate includes the use of flotation.

10. The method according to any one of claims 1-9, characterized in that, The method includes the following steps: (1) The powder with a particle size of no more than 75 μm in the pretreated iron tailings, accounting for more than 90 wt%, is roasted at a roasting temperature of 1200-1400 ℃ for 60-180 min to obtain the roasted product. (2) The obtained roasted product is crushed and sieved to obtain the oversize and undersize. The undersize has a particle size of no more than 75 μm and the proportion of powder is more than 95 wt%. The oversize is a reduced iron product. (3) Mix the sieve material with an acid solution with a concentration of 1-4 mol / L, wherein the liquid-solid ratio of the acid solution to the sieve material is (6-12):1 mL / g, and stir continuously for 30-120 min at an acid leaching temperature of 40-90 ℃ to obtain a leaching solution containing rare earth elements and an acid leaching residue. (4) Add a precipitant to the leachate with a concentration of 0.5-3 mol / L, react at a temperature of 40-80 ℃ for 30-90 min to precipitate rare earth elements, and calcine the precipitate containing rare earth elements at 800-1000 ℃ for 50-70 min to obtain mixed rare earth oxides, and recover fluorite concentrate from the acid leaching residue.

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