Method for utilizing iron, rare earth and niobium elements from easily dissolvable stone type niobium concentrate

Abstract

The invention relates to a method for utilizing iron, rare earth and niobium elements from easily dissolvable stone type niobium concentrate, belongs to the field of mineral processing and hydrometallurgy, and solves at least one of the problems of difficulty in cooperative utilization of recovery of niobium, iron and rare earth elements, complicated process flow, high energy consumption and cost, harsh reaction conditions, high environmental pollution risk and the like in the prior art. The method comprises the following steps that niobium concentrate and a reducing agent are evenly mixed, after pelletizing, reduction roasting is conducted, and roasted ore is obtained; the roasted ore is smashed and then subjected to magnetic separation, and iron ore concentrate and tailings are obtained; acidizing and roasting to obtain re-roasted ore; the re-roasted ore is subjected to water leaching in boiling water; water leaching liquid and water leaching residues are obtained; adjusting the pH value of the water immersion liquid to 1.5-2.0, and adding an oxalic acid precipitator to obtain a rare earth compound; and mixing the water leaching residues with an acid solution for acid leaching. The method realizes efficient synergistic recovery of niobium, iron and rare earth, reduces energy consumption and production cost, does not need harsh conditions, and is environment-friendly.

Description

Technical Field

[0001] This invention relates to the fields of mineral processing and hydrometallurgical technology, and in particular to a method for utilizing iron, rare earth elements and niobium elements in easily calcified niobium concentrate. Background Technology

[0002] Niobium (Nb) is an indispensable strategic metal for the steel, aerospace, superconducting, and information industries. Although my country has a large total niobium resource, it is mainly found in polymetallic symbiotic deposits in areas such as Bayan Obo. Among them, easily calcified niobium minerals, due to their complex composition, similar floatability, and fine grain size, can only yield low-grade niobium concentrate using conventional beneficiation techniques. This has led to my country's long-term predicament of "having resources but difficulty in utilization," resulting in a persistently high dependence on imports.

[0003] Current technologies for the separation and purification of easily calcite-type niobium concentrate are still in the laboratory research stage and have not yet been industrialized. Mainstream technical routes have significant limitations: while hydrofluoric acid decomposition-solvent extraction can separate niobium, the system is highly toxic and corrosive, posing significant environmental risks; alkali fusion-water leaching methods are energy-intensive, cause significant equipment wear, and lack comprehensive recovery capabilities for associated valuable elements. Furthermore, existing processes do not adequately address the separation and recovery of iron and rare earth elements in niobium concentrate. A few iron removal technologies rely on high-cost methods such as high-temperature reduction and protective atmospheres, making it difficult to achieve efficient and selective iron separation. Regarding rare earth recovery, traditional methods are mainly designed for simple, single-element rare earth ores. When directly applied to easily calcite, the close association and similar chemical properties of rare earth elements with niobium, titanium, thorium, and other elements in the mineral lead to the complete leaching of all elements, making subsequent separation difficult and resulting in radioactive pollution and environmental costs.

[0004] Existing technologies struggle to achieve economical, efficient, and synergistic comprehensive recycling of niobium, iron, and rare earth elements when dealing with the complex symbiotic structure of easily calcite. Summary of the Invention

[0005] In view of the above analysis, the present invention aims to provide a method for utilizing iron, rare earth and niobium elements in easily calcified niobium concentrate, in order to solve at least one of the problems in the prior art, such as the difficulty in synergistic utilization of niobium, iron and rare earth elements in the recovery of low-grade easily calcified niobium concentrate, complex process flow, high energy consumption and cost, harsh reaction conditions and high environmental pollution risk.

[0006] The objective of this invention is achieved through the following technical solution:

[0007] This invention provides a method for utilizing iron, rare earth elements, and niobium in easily cleavable niobium concentrate, comprising the following steps: (1) Niobium concentrate is mixed with reducing agent, pelletized, and then reduced and roasted at 450~600℃ to obtain roasted ore; (2) The roasted ore is crushed and then subjected to magnetic separation to obtain iron concentrate and tailings; (3) The tailings are mixed with sulfuric acid and leaching aid and then acidified and roasted to obtain re-roasted ore. The acidification and roasting temperature is 300~400℃. (4) The re-roasted ore is leached in boiling water, and after solid-liquid separation, a water leaching solution and a water leaching residue are obtained; (5) The pH value of the aqueous extract is adjusted to 1.5~2.0 using a regulator, oxalic acid precipitant is added, and rare earth compounds are obtained after solid-liquid separation. The regulator is at least one of magnesium oxide or magnesium carbonate. The water-leached residue is mixed with an acid solution for acid leaching, and after solid-liquid separation, a niobium-containing acid leaching solution and acid leaching residue are obtained.

[0008] Further, in step (5), the amount of oxalic acid precipitant used is 1.5 to 3.5 times the mass content of rare earth oxides in the niobium concentrate; and / or, The rare earth oxide content in the rare earth compound obtained in step (5) is greater than 45 wt.%.

[0009] Further, in step (5), the acid solution is a mixture of hydrofluoric acid and sulfuric acid, wherein the mass concentration of hydrofluoric acid in the mixture is 5-40 wt.%, and the mass concentration of sulfuric acid in the mixture is 20-40 wt.%; or, The acid solution is selected from at least one of oxalic acid, tartaric acid, and hydroxysuccinic acid, with a mass concentration of 20-80 wt.%.

[0010] Furthermore, in step (5), the acid leaching temperature is 50~90℃ and the leaching time is 120~240min.

[0011] Further, in step (4), the solid-liquid mass ratio of the water immersion is 1:(2.0~8.0); and / or, The immersion time in water is 60–180 minutes.

[0012] Further, in step (1), the mass ratio of the niobium concentrate to the reducing agent is 10:(5.0~3.0); and / or, The holding time for the reduction roasting is 60~180min.

[0013] Furthermore, the magnetic field strength of the magnetic separation in step (2) is 120~300mT.

[0014] Further, in step (3), the leaching aid is at least one of ammonium sulfate or ammonium bisulfate; and / or, The sulfuric acid has a mass concentration of 60~98 wt.%; and / or, The holding time for the acidification roasting is 60~180min.

[0015] Furthermore, the total iron (TFe) content in the iron concentrate is greater than 60 wt.%, the rare earth recovery rate is greater than 85%, and the niobium leaching rate (calculated as Nb2O5) is greater than 95%.

[0016] Furthermore, the niobium concentrate is a low-grade calcite-type niobium concentrate, wherein the calcite content is 1.0~35 wt.%, the hematite content is 30~70 wt.%, the Nb2O5 content is 1.0~10.0 wt.%, the total iron (TFe) content is 26~50 wt.%, the rare earth oxide (REO) content is 1~10 wt.%, and the Ti content is 1.0~10 wt.

[0017] Compared with the prior art, the present invention can achieve at least one of the following beneficial effects: This invention provides a method for the synergistic recovery of niobium, iron, and rare earth elements (REEs) from low-grade easily calcified niobium concentrate. Through a series of interconnected steps—"medium-temperature selective reduction—magnetic separation for iron removal—mild acidification transformation—boiling water selective niobium leaching—pH-controlled preferential precipitation of rare earth elements—acid leaching for deep niobium extraction"—effective dissociation and targeted recovery of complex symbiotic minerals are achieved under relatively mild conditions. The technical advantages of this invention include: (1) It has achieved efficient and synergistic recovery of niobium, iron and rare earth elements. Efficient iron separation: Using medium-temperature selective reduction at 450–600℃, iron oxides such as hematite are directionally converted into strongly magnetic Fe3O4, while maintaining the structural stability of niobium and rare earth minerals; then, through magnetic separation, efficient physical separation of iron is achieved, directly obtaining qualified iron concentrate, and significantly reducing acid consumption and impurity interference in subsequent processes.

[0018] Efficient Separation and Recovery of Niobium and Rare Earth Elements: This invention achieves the directional separation of niobium and rare earth elements in a solid-liquid two-phase process through an "acidification roasting-boiling water leaching" step. The core mechanism lies in utilizing the difference in hydrolysis behavior of their salts in high-temperature aqueous solutions: niobium sulfate generated after acidification roasting rapidly hydrolyzes in near-neutral boiling water, generating insoluble metaniobic acid precipitate which enters the water leaching residue; while rare earth sulfates dissolve stably under the same conditions and remain in the water leaching solution, thus achieving physical phase separation of the two elements.

[0019] The two phases are efficiently purified and recovered through a "pH adjustment-oxalic acid precipitation" process. Magnesium oxide and / or magnesium carbonate are selected as pH adjusters because Mg²⁺... + Both sulfates and oxalates of rare earth elements have high solubility. This adjustment system can achieve precise and mild pH control of the aqueous extract without introducing interfering solid impurities, ensuring the chemical cleanliness of the separation system. Under suitable acidic conditions, rare earth ions (RE³) can... +It can form rare earth oxalate precipitate with oxalate ions with extremely low solubility, while a small amount of niobium is retained in the solution in the form of stable oxalate complex anions. At the same time, through the dual mechanism of "precise pH control" and "oxalate complexation stabilization", the hydrolysis and precipitation of impurities such as iron and aluminum are effectively inhibited, thereby achieving highly selective precipitation separation and recovery of rare earth.

[0020] For niobium-rich water-leached residue, subsequent acid leaching is used to efficiently dissolve the niobium, ensuring the total niobium leaching rate.

[0021] (2) The process of multi-element recovery has been integrated and systematized, overcoming the limitations of traditional technologies that make it difficult to recover niobium, iron, and rare earth elements simultaneously and that the processes are fragmented. This invention innovatively integrates the traditionally cumbersome and independent processes of recovering iron, niobium, and rare earth elements separately into a continuous main line of "prioritizing iron extraction - co-leaching - sequential separation". The multi-element dissociation and purification are completed in a single material flow, which significantly simplifies equipment configuration and operation, improves production intensification and system operating efficiency, and provides an efficient and simple solution for the industrial comprehensive utilization of low-grade easily calcified complex symbiotic mineral resources.

[0022] (3) Significant reduction in energy consumption and production costs has been achieved. This invention, through a systematic low-temperature process design (reduction 450~600℃, acidification 300~400℃), effectively avoids the high energy consumption problems associated with traditional processes that rely on high-temperature (>1000℃) reduction, high-temperature roasting, or alkali fusion (>800℃). Simultaneously, the entire process eliminates the need for high-pressure equipment or high-energy-consuming auxiliary means such as ultrafine grinding, further simplifying operation and reducing equipment investment and operating costs. The simultaneous recovery of iron and rare earth elements as high-value-added byproducts achieves synergistic optimization of energy saving, cost reduction, and resource value enhancement.

[0023] (4) No stringent conditions required, making it more industrially feasible. This invention uses hydroxy organic acids instead of hydrofluoric acid, and does not require high-pressure or supercritical equipment. All reactions are carried out at normal pressure, medium and low temperatures, and in conventional acid and alkaline media, which greatly reduces equipment requirements and operational safety risks, and enhances the potential for industrial application.

[0024] (5) Improved environmental friendliness. The present invention preferably uses a chemical system with more controllable environmental risks (such as oxalic acid) to replace highly toxic reagents such as hydrofluoric acid, thereby reducing the risk of high-risk substances from the source; by prioritizing iron extraction and multi-element synergistic recovery, the acid consumption, waste residue amount and resource comprehensive utilization efficiency are simultaneously optimized; and radioactive elements are directionally enriched in a specific slag phase, which is convenient for centralized disposal and reduces the environmental pollution load.

[0025] In this invention, the above-described technical solutions can be combined with each other to achieve more preferred combinations. Other features and advantages of this invention will be set forth in the following description, and some advantages may become apparent from the description or be learned by practicing the invention. The objects and other advantages of this invention can be realized and obtained from what is particularly pointed out in the description and drawings. Attached Figure Description

[0026] The accompanying drawings are for illustrative purposes only and are not intended to limit the invention. Throughout the drawings, the same reference numerals denote the same parts.

[0027] Figure 1 This is a schematic flowchart illustrating a method for utilizing iron, rare earth elements, and niobium in easily cleaved niobium concentrate, as provided in an embodiment of the present invention. Detailed Implementation

[0028] Preferred embodiments of the present invention will now be described in detail with reference to the accompanying drawings, which form part of this application and are used together with the embodiments of the present invention to illustrate the principles of the present invention, but are not intended to limit the scope of the present invention.

[0029] This invention provides a method for utilizing iron, rare earth elements, and niobium in easily cleavable niobium concentrate, comprising the following steps: (1) Niobium concentrate is mixed with reducing agent, pelletized, and then reduced and roasted at 450~600℃ to obtain roasted ore; (2) The roasted ore is crushed and then subjected to magnetic separation to obtain iron concentrate and tailings; (3) The tailings are mixed with sulfuric acid and leaching aid and then acidified and roasted to obtain re-roasted ore. The acidification and roasting temperature is 300~400℃. (4) The re-roasted ore is leached in boiling water, and after solid-liquid separation, a water leaching solution and a water leaching residue are obtained; (5) The pH value of the aqueous extract is adjusted to 1.5~2.0 using a regulator, oxalic acid precipitant is added, and rare earth compounds are obtained after solid-liquid separation. The regulator is at least one of magnesium oxide or magnesium carbonate. The water-leached residue is mixed with an acid solution for acid leaching, and after solid-liquid separation, a niobium-containing acid leaching solution and acid leaching residue are obtained.

[0030] Compared with existing technologies, this invention controls the temperature conditions of reduction roasting, matching the thermodynamic and kinetic window for the conversion of iron oxides to Fe3O4. While ensuring a high conversion rate, it minimizes side reactions, achieving a balance between high efficiency and energy saving. If the temperature is too low or the time is too short, the reduction reaction cannot be completed, resulting in low iron recovery. If the temperature is too high or the time is too long, it not only increases energy consumption but may also cause Fe3O4 decomposition or over-reduction, and may lead to unfavorable phase transformations in niobium minerals.

[0031] Specifically, in step (1), the mass ratio of niobium concentrate to reducing agent is 10:(5.0~3.0). The core purpose of controlling the ratio of niobium concentrate to reducing agent in this invention is to ensure that iron oxides are fully reduced to strongly magnetic Fe3O4 while preventing excessive carbon. Excessive carbon is not only uneconomical, but may also cause Fe3O4 to be further reduced to weakly magnetic FeO, affecting the efficiency of subsequent magnetic separation; at the same time, excessive carbon will cause the local reducing atmosphere to be too strong, which may destroy the crystal structure of niobium minerals and is not conducive to the subsequent recovery of niobium.

[0032] For example, in step (1), the reducing agent is a carbonaceous reducing agent.

[0033] For example, the reducing agent is at least one of carbon powder, coke powder, or coal powder. Preferably, it is anthracite with a fixed carbon content ≥76 wt.%, volatile matter ≤7.8 wt.%, ash content ≤15 wt.%, moisture content ≤1 wt.%, and particle size ≤74 μm.

[0034] For example, in step (1), the mass ratio of niobium concentrate to reducing agent is 10:(5.0~3.0), such as 10:5, 10:4.5, 10:4, 10:3.5, 10:3; preferably 10:4.5 to 10:3.5, which achieves an optimized balance between reduction efficiency and selectivity.

[0035] For example, in step (1), the particle size of the pellets obtained after pelletizing is 10~15mm. Controlling the particle size of the pellets is beneficial to ensure sufficient reduction while taking into account a high metal conversion rate and efficiency; if the particle size is too small, it will easily lead to severe pulverization during roasting, increasing the dust collection burden and metal loss; if the particle size is too large, it will be difficult for heat and gas to be transferred to the interior of the pellets, resulting in uneven reduction.

[0036] Specifically, in step (1), the reduction calcination temperature is 450~600℃; for example, the reduction calcination temperature is 450℃, 470℃, 500℃, 520℃, 550℃, 570℃, or 600℃.

[0037] Specifically, in step (1), the holding time for reduction calcination is 60~180min. For example, the reduction calcination time is 60min, 90min, 120min, 150min, or 180min.

[0038] Preferably, in step (1), the reduction calcination temperature is 470~570℃ and the time is 60~120min. Such synergistic conditions achieve efficient and selective reduction to Fe3O4 while taking into account the optimal balance between reaction thoroughness, energy consumption economy and phase stability.

[0039] Specifically, in step (1), the reduction roasting is carried out in a muffle furnace. Exemplarily, the muffle furnace is protected by a reducing atmosphere.

[0040] In step (1), the roasted ore obtained by reduction roasting is cooled by rapid cooling, preferably by water quenching.

[0041] Specifically, the magnetic field strength of the magnetic separation in step (2) is 120~300mT. For example, the magnetic field strength of the magnetic separation is 120mT, 140mT, 160mT, 180mT, 200mT, 220mT, 240mT, 260mT, 280mT, or 300mT; preferably 140~240mT.

[0042] For example, in step (2), the roasted ore is crushed to a particle size of -74μm and a content of 80~95wt.%, and then magnetic separation is performed.

[0043] Specifically, in step (2), the total iron (TFe) content in the iron concentrate is >60 wt.%.

[0044] Specifically, in step (3), the mass ratio of tailings, sulfuric acid and leaching aid is 1:(2.0~3.5):(0.2~1.5).

[0045] For example, the mass ratio of tailings, sulfuric acid and leaching aid is 1:(2.0, 2.2, 2.5, 2.6, 2.7, 2.8, 2.9, 3.0, 3.2 or 3.5):(0.2, 0.5, 0.6, 0.7, 0.8, 0.9, 1.0, 1.1, 1.3 or 1.5).

[0046] Preferably, the mass ratio of tailings, sulfuric acid and leaching aid is 1:(2.5~3.0):(0.5~1.1).

[0047] It should be noted that a specific ratio of sulfuric acid ensures sufficient acidolysis and transformation of the tailings, while the appropriate leaching aid selectively destroys the structure of impurities and inhibits their dissolution. This synergistic formulation achieves a high niobium conversion rate while reducing ineffective acid consumption and impurity interference, making the resulting re-roasted ore more suitable for subsequent boiling water leaching systems.

[0048] Specifically, the holding time for the acidification roasting is 60~180 min.

[0049] For example, the acidification calcination temperature is 300℃, 320℃, 330℃, 340℃, 350℃, 360℃, 380℃, or 400℃.

[0050] For example, the holding time for acidification roasting is 60 min, 80 min, 100 min, 120 min, 130 min, 140 min, 150 min, 160 min, or 180 min.

[0051] Preferably, the acidification roasting temperature is 320~360℃, and the holding time is 120~160min.

[0052] It should be noted that the acid roasting temperature and time window are the key process balance points for achieving efficient and selective transformation. If the temperature is too low or the time is too short, the sulfuric acid decomposition will be insufficient, the niobium mineral transformation will be incomplete, and the subsequent niobium leaching rate will be significantly reduced. If the temperature is too high or the time is too long, it will easily promote the formation of insoluble polymer oxides of niobium and aggravate the dissolution and activation of impurities such as silicon and aluminum, while causing unnecessary increases in energy consumption and equipment load.

[0053] Specifically, in step (3), the acid roasting process involves directly heating to the target acid roasting temperature at a rate of 5-10℃ / min; or heating in stages, specifically including: heating to 200-250℃ at a rate of 5-10℃ / min and holding for 20-40min, and then continuing to heat to the target acid roasting temperature. The staged heating procedure, through a strategy of gentle dehydration and preliminary activation followed by deep reaction, effectively prevents material splashing, improves sulfuric acid utilization and reaction uniformity, thereby ensuring the safety, efficiency, and product quality stability of the acid roasting process.

[0054] Specifically, in step (3), the acid roasting is carried out in a muffle furnace. Exemplarily, the muffle furnace is protected by a reducing atmosphere.

[0055] In step (3), the re-roasted ore obtained after acid roasting is cooled by rapid cooling, preferably by water quenching.

[0056] Specifically, in step (3), the mass concentration of the sulfuric acid is 60 wt.% to 98 wt.%, preferably 78 wt.% to 92 wt.%. For example, the mass concentration of the sulfuric acid is 60 wt.%, 70 wt.%, 80 wt.%, 85 wt.%, 88 wt.%, 90 wt.%, 95 wt.%, and 98 wt.%.

[0057] Specifically, in step (3), the immersion aid is at least one of ammonium sulfate or ammonium bisulfate; preferably ammonium sulfate.

[0058] Specifically, in step (4), the solid-liquid mass ratio of the water immersion is 1:(2.0~8.0).

[0059] For example, the solid-liquid mass ratio of the water immersion is 1:2, 1:3, 1:4, 1:5, 1:6, 1:7, or 1:8; preferably 1:(4.0~6.0).

[0060] It should be noted that in the boiling water leaching step, by controlling the solid-liquid mass ratio within a suitable range and conducting the process under boiling water conditions, the hydrolysis-dissolution equilibrium was systematically regulated, thereby enhancing the phase separation selectivity of niobium and rare earth elements. This synergistic condition allows niobyl sulfate to rapidly and completely hydrolyze and precipitate into metaniobic acid in boiling water and a suitable weakly acidic environment, entering the slag phase. Simultaneously, it maintains the stable retention of rare earth sulfates in the liquid phase by inhibiting the hydrolysis reaction of rare earth ions. An excessively high solid-liquid ratio would limit mass transfer and dissolution, leading to incomplete niobium hydrolysis and impairing the separation effect; an excessively low solid-liquid ratio would dilute the reaction system, weaken the driving force of niobium hydrolysis, and reduce the rare earth concentration, affecting economic efficiency. Therefore, the optimized solid-liquid ratio, combined with the boiling water medium, achieves the key objective of promoting efficient and directional separation of "niobium into the slag and rare earth into the liquid" from both thermodynamic and kinetic perspectives.

[0061] For example, "water immersion in boiling water" refers to a leaching process carried out under normal pressure, where the leaching medium water reaches a local boiling state (gas-liquid equilibrium). For instance, the boiling state corresponds to a water temperature of 95~100℃.

[0062] Specifically, in step (4), the immersion time in water is 60~180min.

[0063] For example, the immersion time in water is 60 min, 80 min, 100 min, 120 min, 140 min, 160 min, or 180 min; preferably 80 to 160 min.

[0064] In the boiling water leaching step, controlling the leaching time within a suitable range and coordinating it with boiling water conditions ensures the directional separation of niobium into the slag and rare earth elements into the liquid. This allows niobium sulfate to rapidly and completely hydrolyze and precipitate into metaniobic acid in boiling water and a suitable weakly acidic environment, entering the slag phase. Simultaneously, the hydrolysis reaction of rare earth ions is suppressed to maintain the stable retention of rare earth sulfates in the liquid phase. Too short a time will lead to incomplete niobium hydrolysis and precipitation or insufficient dissolution of rare earth elements, reducing separation efficiency; too long a time may cause the fine niobic acid precipitate to colloidalize, increasing the difficulty of solid-liquid separation and reducing process energy efficiency.

[0065] Specifically, in step (4), the water immersion is a stirring immersion with a stirring speed of 200~1200 rpm.

[0066] For example, the stirring speed of the water immersion is 200 rpm, 400 rpm, 600 rpm, 800 rpm, 1000 min, or 1200 rpm; preferably 400~800 rpm.

[0067] In the boiling water leaching step, by controlling the stirring speed within a suitable range, a uniform and stable reaction environment is created for the efficient directional separation of "niobium into slag and rare earth into liquid". Appropriate stirring intensity can promote mass transfer, prevent solid-phase sedimentation, ensure sufficient hydrolysis and precipitation of niobium sulfate and sufficient dissolution and diffusion of rare earth ions, and avoid uneven reaction and decreased separation efficiency due to excessively low stirring speed, or fine granulation, colloidalization, and phase interface damage caused by excessively high stirring speed.

[0068] Specifically, in step (5), the amount of oxalic acid precipitant used is 1.5 to 3.5 times the mass content of rare earth oxides (REO) in the niobium concentrate.

[0069] For example, the amount of oxalic acid precipitant used is 1.5 times, 1.6 times, 1.8 times, 2.0 times, 2.4 times, 2.6 times, 2.8 times, 3.0 times, 3.2 times, 3.4 times, or 3.5 times the mass content of rare earth oxides (REO) in niobium concentrate; preferably 2.4 to 3.2 times.

[0070] By controlling the relationship between the amount of oxalic acid precipitant and the total amount of rare earth oxides, combined with pH adjustment and the use of magnesium oxide / magnesium carbonate as regulators, efficient precipitation of rare earths can be achieved. Insufficient oxalic acid precipitant dosage will lead to incomplete precipitation and decreased recovery rate; excessive oxalic acid precipitant dosage not only wastes reagents but may also promote the formation of stable soluble oxalic acid complexes from residual iron, aluminum, and other impurities in the system, increasing the complexity of subsequent precipitation and washing processes, potentially affecting the purity of the final product, and also correspondingly increasing the difficulty and overall cost of wastewater treatment.

[0071] Specifically, oxalic acid precipitant is added, and the mixture is stirred at 400-800 rpm for 30-120 minutes at room temperature. Then, solid-liquid separation is performed to obtain rare earth compound precipitate and wastewater.

[0072] In some embodiments, the rare earth oxide (REO) content in the rare earth compound obtained in step (5) is greater than 45 wt.%. The present invention can directly obtain rare earth compound intermediate products with REO content > 45 wt.% under mild conditions, laying a high-quality raw material foundation for subsequent further purification or deep processing, and improving the resource value and economy of the overall process.

[0073] Specifically, in the acid leaching of step (5), the acid solution is a mixture of hydrofluoric acid and sulfuric acid, wherein the mass concentration of hydrofluoric acid in the mixture is 5~40 wt.% and the mass concentration of sulfuric acid in the mixture is 20~40 wt.%.

[0074] Preferably, in step (5), the acid solution is selected from at least one of oxalic acid, tartaric acid, and hydroxysuccinic acid with a mass concentration of 20-80 wt.%.

[0075] In some preferred embodiments, the acid solution is selected as an oxalic acid solution with a mass concentration of 30-50 wt.%.

[0076] It should be noted that, compared to hydrofluoric acid and nitric acid, using organic acids such as oxalic acid for niobium leaching offers a comprehensive advantage: it is green, safe, efficient, and facilitates media recycling. Organic acid systems completely avoid the use of highly toxic fluorides and fluorine pollution, ensuring high operational safety. More importantly, taking oxalic acid as an example, it can form stable, soluble complexes with niobium, achieving highly efficient and selective leaching of niobium under mild conditions. After leaching, the niobium-containing oxalic acid solution can be separated from niobium through extraction or other methods. The remaining oxalic acid mother liquor can be purified, concentrated, and returned to the leaching process for recycling, thus significantly reducing the consumption of fresh reagents and wastewater discharge.

[0077] Specifically, in step (5), the acid leaching temperature is 50~90℃ and the leaching time is 120~240min.

[0078] For example, the acid leaching temperature is 50°C, 60°C, 70°C, 75°C, 80°C, 85°C, or 90°C; preferably 70-90°C.

[0079] For example, the leaching time of the acid leaching is 120 min, 140 min, 160 min, 180 min, 200 min, 220 min, or 240 min; preferably 140 to 220 min.

[0080] By controlling the acid leaching temperature and leaching time, the aim is to synergistically optimize reaction kinetics and process economy. A specific acid leaching temperature range effectively promotes the dissolution and complexation reactions of niobium compounds such as metaniobic acid in the water-leached residue in organic acids (e.g., oxalic acid), increasing the leaching rate and final niobium leaching rate, while avoiding excessively high temperatures that could lead to organic acid decomposition or uneconomical energy consumption. A corresponding leaching time window ensures that the dissolution-complexation reaction tends to be complete, while avoiding excessively long leaching times that could trigger side reactions, thus balancing reaction efficiency, leaching selectivity, and process economy.

[0081] Specifically, in step (5), the acid leaching is a stirring leaching process with a stirring speed of 200~1200 rpm.

[0082] For example, the stirring speed of the acid leaching is 200 rpm, 400 rpm, 600 rpm, 800 rpm, 1000 min, or 1200 rpm; preferably 400~800 rpm.

[0083] During acid leaching, excessively slow stirring speed can lead to insufficient solid sedimentation and mass transfer, reducing leaching efficiency; excessively fast stirring speed may cause excessive particle mudification, increasing the difficulty of solid-liquid separation and energy consumption, and may also damage the solution stability of niobium complexes. Optimized stirring speed ensures efficient leaching while taking into account the economy and operability of the process.

[0084] Specifically, in step (5), the solid-liquid mass ratio of the acid leaching is 1:(2.0~10.0).

[0085] For example, in step (5), the solid-liquid mass ratio of the acid leaching is 1:2, 1:3, 1:4, 1:5, 1:6, 1:7, 1:8, 1:9, or 1:10; preferably 1:(3~6).

[0086] By controlling the solid-liquid mass ratio of acid leaching, sufficient liquid phase is provided to ensure the full dissolution and mass transfer of niobium compounds, avoiding incomplete leaching due to insufficient liquid volume; at the same time, excessive dilution of the leachate is prevented, thereby reducing the volume of subsequent treatments and reagent consumption.

[0087] In some embodiments, the total iron (TFe) content in the iron concentrate is greater than 60 wt.%, the rare earth recovery rate is greater than 85%, and the niobium leaching rate (calculated as Nb2O5) is greater than 95%.

[0088] For example, the niobium concentrate is a low-grade calcite-type niobium concentrate, wherein the calcite content is 1.0~35 wt.%, the hematite content is 30~70 wt.%, the Nb2O5 content is 1.0~10.0 wt.%, the total iron (TFe) content is 26~50 wt.%, the rare earth oxide (REO) content is 1~10 wt.%, and the Ti content is 1.0~10 wt.

[0089] The technical solution of the present invention will be further described in detail below with reference to specific embodiments and comparative examples.

[0090] Example 1 This embodiment provides a method for the comprehensive utilization of iron, rare earth elements, and niobium in easily calcified niobium concentrate. The specific steps are as follows: (1) Reduction roasting: Low-grade easily calcified niobium concentrate obtained from the flotation of Bayan Obo mine was used. Its main components included: easily calcified 32.15 wt.%, hematite 37.36 wt.%, Nb₂O₅ 6.43 wt.%, rare earth oxides (REO) 5.81 wt.%, Ti 2.58 wt.%, and total iron (TFe) 26.15 wt.%. The niobium concentrate was mixed with coke powder (reducing agent) at a mass ratio of 10:4, pelletized, and then subjected to reduction roasting at 550℃ for 120 min in a muffle furnace (with reducing gas protection) to obtain roasted ore. The roasted ore was then cooled by water quenching. The resulting pellets had a particle size of 10–15 mm.

[0091] (2) Magnetic separation: The roasted ore was pulverized to -74μm with a content of 90wt.% using a sample preparation machine. The roasted ore was then magnetically separated using a magnetic separator under a magnetic field strength of 160mT to obtain iron concentrate and tailings.

[0092] (3) Acidification roasting: The tailings, concentrated sulfuric acid (90 wt.%), and ammonium sulfate were mixed at a mass ratio of 1:2.5:0.5 and then acidified and roasted in a muffle furnace (with reducing gas protection) at 350°C for 120 min to obtain the re-roasted ore. The heating process was carried out at a rate of 10°C / min to reach the target acidification and roasting temperature of 350°C.

[0093] (4) Boiling water extraction: The re-roasted ore was placed in boiling water and leached with stirring at a solid-liquid mass ratio of 1:4, a stirring speed of 800 rpm, and a leaching time of 120 min. After leaching, solid-liquid separation was performed to obtain water-leached solution and water-leached residue.

[0094] (5) Rare earth precipitation recovery Magnesium oxide was added to the aqueous leachate to adjust the pH to 1.8, and then oxalic acid precipitant was added at a concentration 2.8 times the mass content of rare earth oxides (REO) in the niobium concentrate. After stirring at 450 rpm for 40 minutes, solid-liquid separation was performed to obtain rare earth compound precipitates and wastewater.

[0095] The water-leached residue was mixed with a 30 wt.% oxalic acid solution at a solid-liquid mass ratio of 1:5, and acid-leached at 75°C for 120 min under stirring (800 rpm). After solid-liquid separation, niobium-containing acid leaching solution and acid leaching residue were obtained.

[0096] Example 2 This embodiment provides a method for the comprehensive utilization of iron, rare earth elements, and niobium in easily calcifiable niobium concentrate. The specific steps are as follows: (1) Reduction roasting Low-grade easily calcified niobium concentrate obtained from the flotation of Bayan Obo mine was used. Its main components included: easily calcified 8.0 wt.%, hematite 35 wt.%, Nb₂O₅ 1.21 wt.%, rare earth oxide (REO) 2.51 wt.%, Ti 0.58 wt.%, and total iron (TFe) 22.54 wt.%. The niobium concentrate was mixed evenly with coke powder (reducing agent) at a mass ratio of 10:4. After pelletizing, the mixture was placed in a muffle furnace (with reducing gas protection) and subjected to reduction roasting at 550℃ for 120 min to obtain roasted ore. The roasted ore was then cooled by water quenching. The resulting pellets had a particle size of 10–15 mm.

[0097] (2) Magnetic separation The roasted ore was pulverized to -74μm with a content of 92wt.% using a sample preparation machine. The roasted ore was then magnetically separated using a magnetic separator under a magnetic field strength of 150mT to obtain iron concentrate and tailings.

[0098] (3) Acidification roasting The tailings, concentrated sulfuric acid (92 wt.%), and ammonium sulfate were mixed at a mass ratio of 1:2.5:0.5 and then acidified and roasted in a muffle furnace (with reducing gas protection) at 350°C for 120 min to obtain re-roasted ore. The temperature was increased to 250°C at a rate of 10°C / min and held for 20 min, and then increased to the target acidification and roasting temperature of 350°C at a rate of 10°C / min.

[0099] (4) Boiling water immersion The re-roasted ore was placed in boiling water and leached with stirring at a solid-liquid mass ratio of 1:4, a stirring speed of 800 rpm, and a leaching time of 120 min. After leaching, solid-liquid separation was performed to obtain water-leached solution and water-leached residue.

[0100] (5) Rare earth precipitation recovery Magnesium carbonate was added to the aqueous leachate to adjust the pH to 1.9, and then oxalic acid precipitant was added at a concentration 3.0 times the mass content of rare earth oxides (REO) in the niobium concentrate. The mixture was stirred at 500 rpm for 50 minutes, followed by solid-liquid separation to obtain rare earth compound precipitates and wastewater.

[0101] The water-leaching residue was mixed with a 30 wt.% oxalic acid solution at a solid-liquid mass ratio of 1:3, and acid-leaching was carried out at 80°C for 120 min under stirring conditions (800 rpm). After solid-liquid separation, niobium-containing acid leaching solution and acid leaching residue were obtained.

[0102] The specific process parameters for Examples 3 to 27 are summarized in Tables 1-1 and 1-2. Examples 3 to 27 are based on Example 1, with one or more of the specified process parameters shown in Tables 1-1 and 1-2 adjusted. All other process conditions and operating parameters not listed in Tables 1-1 and 1-2 are consistent with those of Example 1.

[0103] Table 1-1 Summary of process parameters for the embodiments

[0104] Table 1-2 Summary of process parameters for the embodiments

[0105] Comparative Example 1: The difference between this comparative example and Example 1 is that in step (3), the acidification and roasting temperature is 250°C. The remaining steps and conditions are the same as in Example 1.

[0106] Comparative Example 2: The difference between this comparative example and Example 1 is that ammonium sulfate (an leaching aid) is not added during acid roasting in step (3). The remaining steps and conditions are the same as in Example 1.

[0107] Comparative Example 3: The difference between this comparative example and Example 1 is that calcium oxide was used to adjust the pH value in step (5). The remaining steps and conditions are the same as in Example 1.

[0108] Comparative Example 4: The difference between this comparative example and Example 1 is that steps (1) and (2) are omitted, that is, the niobium concentrate is not subjected to reduction roasting and magnetic separation, and the raw ore is directly subjected to acid roasting in step (3). The remaining steps and conditions are the same as in Example 1.

[0109] Comparative Example 5: The difference between this comparative example and Example 1 is that step (3) is omitted, and the tailings are directly subjected to boiling water leaching in step (4). The remaining steps and conditions are the same as in Example 1.

[0110] Comparative Example 6: The difference between this comparative example and Example 1 is that step (4) is omitted, and the niobium is extracted directly from the re-roasted ore by acid leaching in step (5). The remaining steps and conditions are the same as in Example 1.

[0111] Comparative Example 7: The difference between this comparative example and Example 1 is that in step (1), the reduction calcination temperature is 400°C. The remaining steps and conditions are the same as in Example 1.

[0112] Comparative Example 8: The difference between this comparative example and Example 1 is that in step (1), the reduction calcination temperature is 650°C. The remaining steps and conditions are the same as in Example 1.

[0113] Comparative Example 9: The difference between this comparative example and Example 1 is that in step (3), the acidification and roasting temperature is 450°C. The remaining steps and conditions are the same as in Example 1.

[0114] Comparative Example 10: The difference between this comparative example and Example 1 is that in step (5), the pH value is adjusted to 1.0. The remaining steps and conditions are the same as in Example 1.

[0115] Comparative Example 11: The difference between this comparative example and Example 1 is that in step (5), the pH value is adjusted to 2.5. The remaining steps and conditions are the same as in Example 1.

[0116] Comparative Example 12: The difference between this comparative example and Example 1 is that in step (4), the water immersion temperature is 70°C (boiling water was not used). The remaining steps and conditions are the same as in Example 1.

[0117] Key indicators of the products in the examples and comparative examples, including the total iron (TFe) content, niobium leaching rate, and rare earth oxide content in the iron concentrate, were tested and calculated as follows: Element content determination: The contents of TFe in iron concentrate, Nb2O5 and rare earth oxides (REO) in various intermediate products and slag were determined by X-ray fluorescence spectrometry (XRF).

[0118] For example, niobium leaching rate (%) = 100% The formula is: (Acid leaching residue mass × Nb2O5 content in acid leaching residue) / (Water leaching residue mass × Nb2O5 content in water leaching residue) × 100%; where acid leaching residue is obtained in step (5) and water leaching residue is obtained in step (4).

[0119] Table 2 Test results of the embodiments

[0120] Table 3 shows the test results of the comparative examples.

[0121] The above description is only a preferred embodiment of the present invention, but the scope of protection of the present invention is not limited thereto. Any changes or substitutions that can be easily conceived by those skilled in the art within the scope of the technology disclosed in the present invention should be included within the scope of protection of the present invention.

Claims

1. A method for utilizing iron, rare earth elements, and niobium in easily calcified niobium concentrate, characterized in that, Includes the following steps: (1) Niobium concentrate is mixed with reducing agent, pelletized, and then reduced and roasted at 450~600℃ to obtain roasted ore; (2) The roasted ore is crushed and then subjected to magnetic separation to obtain iron concentrate and tailings; (3) The tailings are mixed with sulfuric acid and leaching aid and then acidified and roasted to obtain re-roasted ore. The acidification and roasting temperature is 300~400℃. (4) The re-roasted ore is leached in boiling water, and after solid-liquid separation, a water leaching solution and a water leaching residue are obtained; (5) The pH value of the aqueous extract is adjusted to 1.5~2.0 using a regulator, oxalic acid precipitant is added, and rare earth compounds are obtained after solid-liquid separation. The regulator is at least one of magnesium oxide or magnesium carbonate. The water-leached residue is mixed with an acid solution for acid leaching, and after solid-liquid separation, a niobium-containing acid leaching solution and acid leaching residue are obtained.

2. The method according to claim 1, characterized in that, In step (5), the amount of oxalic acid precipitant used is 1.5 to 3.5 times the mass content of rare earth oxides in the niobium concentrate; and / or, The rare earth oxide content in the rare earth compound obtained in step (5) is greater than 45 wt.%.

3. The method according to claim 1, characterized in that, In step (5), the acid solution is a mixture of hydrofluoric acid and sulfuric acid, wherein the mass concentration of hydrofluoric acid in the mixture is 5-40 wt.%, and the mass concentration of sulfuric acid in the mixture is 20-40 wt.%; or, The acid solution is selected from at least one of oxalic acid, tartaric acid, and hydroxysuccinic acid, with a mass concentration of 20-80 wt.%.

4. The method according to claim 1, characterized in that, In step (5), the acid leaching temperature is 50~90℃ and the leaching time is 120~240min.

5. The method according to claim 1, characterized in that, In step (4), the solid-liquid mass ratio of the water immersion is 1:(2.0~8.0); and / or, The immersion time in water is 60–180 minutes.

6. The method according to claim 1, characterized in that, In step (1), the mass ratio of the niobium concentrate to the reducing agent is 10:(5.0~3.0); and / or, The holding time for the reduction roasting is 60~180min.

7. The method according to claim 1, characterized in that, The magnetic field strength of the magnetic separation in step (2) is 120~300mT.

8. The method according to claim 1, characterized in that, In step (3), the leaching aid is at least one of ammonium sulfate or ammonium bisulfate; and / or, The sulfuric acid has a mass concentration of 60~98 wt.%; and / or, The holding time for the acidification roasting is 60~180min.

9. The method according to claim 1, characterized in that, The total iron (TFe) content in the iron concentrate is greater than 60 wt.%, the rare earth recovery rate is greater than 85%, and the niobium leaching rate (calculated as Nb2O5) is greater than 95%.

10. The method according to claim 1, characterized in that, The niobium concentrate is a low-grade calcite-type niobium concentrate, wherein the calcite content is 1.0~35 wt.%, the hematite content is 30~70 wt.%, the Nb2O5 content is 1.0~10.0 wt.%, the total iron (TFe) content is 26~50 wt.%, the rare earth oxide (REO) content is 1~10 wt.%, and the Ti content is 1.0~10 wt.

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