A combined beneficiation and smelting method for comprehensively recovering niobium, silicon and scandium from niobium-containing tailings

Abstract

The application discloses a beneficiation-metallurgy combined method for comprehensively recovering niobium, silicon and scandium from niobium-containing tailings, and belongs to the technical field of mineral processing and metallurgical engineering. The application first adopts one roughing and two cleaning flotation to obtain a niobium rough concentrate, and then adopts hydrogen-based mineral phase conversion roasting to activate the niobium rough concentrate, and the roasting product is subjected to acid leaching, so that niobium and scandium enter into a leaching solution, and silicon is enriched in a leaching residue; the leaching solution is subjected to solvent extraction, precipitation and calcination to obtain a niobium oxide product; the leaching residue is subjected to alkali leaching and pH adjustment to obtain white carbon black; and the raffinate is subjected to selective extraction to obtain a scandium oxide concentrate. The preparation method has the advantages of low-temperature hydrogen-based roasting, low carbon and low energy consumption, and can realize the step-by-step recovery of three valuable products, realizes the high-value utilization of all components of the niobium tailings, and has the advantages of mild process conditions, a niobium recovery rate of greater than or equal to 70% in the whole process, a scandium recovery rate of greater than or equal to 60% in the whole process, a white carbon black SiO2 content of greater than or equal to 95%, and good industrial application prospect.

Description

Technical Field

[0001] This invention belongs to the field of mineral processing and metallurgical engineering technology, and specifically relates to a combined beneficiation and smelting method for the comprehensive recovery of niobium, silicon and scandium from niobium-containing tailings. Background Technology

[0002] Niobium is an important rare metal widely used in cutting-edge fields such as the steel industry, aerospace, nuclear industry, superconducting materials, and high-strength alloys. my country has the second largest niobium reserves in the world, with the Bayan Obo mine accounting for over 97% of the country's niobium resources, with proven reserves of 6.6 million tons and basic industrial reserves of 1.57 million tons. However, the Bayan Obo tailings contain a variety of niobium minerals, mainly distributed in rutile, columbite (manganese) ore, calcite, and a small amount of pyrochlore. These niobium minerals are closely associated with iron, rare earth elements, fluorite, etc., with extremely fine grain size and significant differences in physicochemical properties. It is extremely difficult to obtain high-grade niobium concentrate using conventional physical beneficiation methods. Currently, production can only obtain low-grade niobium concentrate with an Nb₂O₅ grade of 5%-8%, far from meeting the requirements of downstream smelting industries. Statistics show that my country's annual niobium consumption exceeds 16,000 tons, while domestic production of niobium concentrate is only about 50 tons per year, resulting in a very high dependence on imported niobium resources. At the same time, these tailings also contain valuable associated components, with scandium content reaching 300ppm-450ppm and silicon content reaching about 55%, which have extremely high comprehensive utilization value.

[0003] Niobium-bearing tailings typically contain various valuable components such as scandium and silicon, making their comprehensive recovery and utilization significant and economically valuable. Some research reports have been published regarding the development and utilization of these resources. Chinese patent CN117285073A (if rejected or under examination, the wording should be changed to "Chinese patent application") discloses a method for separating niobium oxide and scandium oxide from Bayan Obo tailings. This method employs a hydrometallurgical process involving atmospheric pressure acid leaching, sodium sulfate precipitation of rare earth elements, extraction enrichment, and oxalic acid reaction to obtain rare earth sulfate double salts, niobium oxide, and scandium oxide products, respectively. However, this process uses a high-temperature sulfuric acid system (150℃-200℃) for atmospheric pressure acid leaching, resulting in high acid consumption, severe equipment corrosion, and failure to recover the considerable silicon resources in the tailings. Chinese patent CN113897485A discloses a method for enriching scandium from niobium-titanium ore. Using solid carbon as a reducing agent and silicon slag as additives, a carbothermic reduction reaction is carried out at a high temperature of 1250℃-1400℃ to obtain a niobium-silicon-phosphorus-iron multi-element alloy and scandium-rich slag. This method involves high reaction temperatures and high energy consumption. Using carbon as a reducing agent generates a large amount of CO2 emissions, which contradicts the current trend of green and low-carbon development. Furthermore, this process mainly focuses on the enrichment of niobium and scandium, without addressing the high-value utilization of silicon. Chinese patent CN106636614A discloses a method for leaching niobium, scandium, and rare earth elements from tailings. This involves adding calcium hydroxide and sodium chloride to the tailings for roasting (750℃-900℃), followed by hydrochloric acid leaching to extract rare earth elements and scandium. The leaching residue is then further leached with concentrated sulfuric acid at high temperature to extract niobium. Although this method achieves the leaching of niobium, scandium, and rare earth elements, it has the following shortcomings: First, the roasting temperature is still relatively high (750℃-900℃), and the addition of sodium chloride during roasting may produce harmful gases such as chlorine, resulting in a large environmental burden; Second, the two-step leaching process is relatively complex; Third, the silicon resources in the leaching residue are not recycled, resulting in secondary resource waste.

[0004] In summary, existing technologies for recovering valuable components from niobium-bearing tailings generally suffer from problems such as high roasting temperatures, high energy consumption, high carbon emissions, complex processes, and failure to achieve simultaneous recovery of silicon resources. In particular, a green and efficient method has yet to emerge that can achieve cascaded synergistic recovery of niobium, scandium, and silicon in the same main process. Summary of the Invention

[0005] This invention addresses the problems of high roasting temperature, high energy consumption, high carbon emissions, complex processes, and lack of simultaneous recovery of niobium, scandium, and silicon resources in existing technologies. It proposes a combined beneficiation and metallurgical method for the comprehensive recovery of niobium, silicon, and scandium from niobium-containing tailings. The method first involves flotation and tailings removal from the niobium-containing tailings to obtain a rough niobium concentrate. Then, a hydrogen-based mineral phase transformation roasting process is used to activate the niobium concentrate, disrupting the stable crystal structure of niobium minerals and silica-containing gangue, significantly improving its chemical reactivity. The roasting product is then leached using an acid leaching process, allowing niobium, silicon, and scandium to enter different phases. The leachate is then subjected to solvent extraction, precipitation, and calcination to obtain niobium oxide. The leaching residue is then subjected to alkaline leaching and pH adjustment to prepare silica. The raffinate is selectively extracted to obtain scandium oxide concentrate, achieving full utilization of the three components (niobium, silicon, and scandium). This provides a systematic technical solution for the clean, efficient, and comprehensive utilization of niobium-containing tailings.

[0006] To achieve the above objectives, the technical solution adopted by the present invention is as follows:

[0007] A combined beneficiation and metallurgical method for the comprehensive recovery of niobium, silicon, and scandium from niobium-bearing tailings includes the following steps:

[0008] Step 1: Flotation pretreatment and tail removal;

[0009] Using niobium-containing tailings as raw material, a flotation process of one rougher and two cleaners is employed to remove the tailings and low-density or non-magnetic gangue, thus obtaining niobium rough concentrate.

[0010] Step 2: Hydrogen-based mineral phase transformation roasting;

[0011] The niobium crude concentrate obtained in step 1 was placed in a hydrogen-based mineral phase transformation roasting device and roasted in a mixed atmosphere of hydrogen and nitrogen. After roasting, the hydrogen gas was turned off and nitrogen gas was continuously introduced to cool to room temperature.

[0012] Step 3: Mixed acid leaching;

[0013] The roasted ore obtained in step 2 was added to a mixed acid solution for leaching. After leaching, the mixture was filtered and separated to obtain a leaching solution containing niobium and scandium, as well as a silica-rich leaching residue.

[0014] Step 4: Extraction and enrichment of niobium and preparation of silica;

[0015] The niobium-containing leachate obtained in step 3 was enriched and separated by solvent extraction. After extraction, the organic phase loaded with metal ions was washed with dilute acid, back-extracted, and the pH was adjusted to obtain niobium hydroxide precipitate. After filtration, washing, and calcination, industrial niobium oxide product was obtained. The silica-rich leachate residue obtained in step 3 was subjected to alkaline leaching. After leaching and filtration, a sodium silicate solution was obtained. The pH was adjusted to precipitate white carbon black (hydrated silica). After filtration and washing, white carbon black product was obtained.

[0016] Step 5: Selective extraction and removal of scandium;

[0017] The raffinate from the niobium extraction in step 4 was used as the raw material for scandium extraction. A synergistic extraction system was employed for selective scandium extraction. The organic phase loaded with metal ions was back-extracted, and oxalic acid was added to the back-extract to precipitate scandium oxalate. After filtration, washing, and calcination, scandium oxide concentrate was obtained.

[0018] Further, the niobium-containing tailings raw material described in step 1 has a particle size of -0.074 mm, the pulp concentration of the flotation roughing is 25%, the pulp pH is 7-8, the inhibitor is one or more of water glass, sodium hexametaphosphate, oxalic acid, tannic acid and ethylenediaminetetraacetic acid, the collector is a hydroxamic acid reagent, one or more of benzohydroxyxamic acid, octylhydroxyxamic acid, salicylic acid and pentylhydroxyxamic acid, the activator is lead nitrate, the frother is No. 2 oil (pine oil), the Nb2O5 grade in the niobium tailings is 0.2%-0.5%, after pretreatment, the Nb2O5 grade in the niobium concentrate is increased to 2%-5%, and the niobium recovery rate is ≥75%.

[0019] Furthermore, in step 2, the hydrogen-based mineral phase transformation roasting is carried out at a roasting temperature of 500℃-700℃, a roasting time of 30min-120min, a hydrogen volume ratio of 10%-40%, and a total gas flow rate of 200mL / min-1000mL / min; the roasting temperature is preferably 550℃-650℃, the roasting time is preferably 45min-90min, and the hydrogen volume ratio is preferably 20%-30%.

[0020] Further, in the mixed acid leaching described in step 3, the liquid-to-solid ratio of the mixed acid to the roasted ore is 2:1-5:1. The mixed acid consists of hydrochloric acid and hydrofluoric acid, with a hydrochloric acid concentration of 2 mol / L-6 mol / L and a hydrofluoric acid concentration of 0.5 mol / L-2 mol / L. The leaching temperature is 60℃-95℃, and the leaching time is 2h-6h. Preferably, the hydrochloric acid concentration is 3 mol / L-5 mol / L, the hydrofluoric acid concentration is 0.8 mol / L-1.5 mol / L, the leaching temperature is 75℃-85℃, and the leaching time is 3h-5h. The niobium leaching rate is ≥90%, and the scandium leaching rate is ≥85%.

[0021] Furthermore, in step 4, the solvent extraction method uses tributyl phosphate (TBP) or methyl isobutyl ketone (MIBK) as the extractant, and sulfonated kerosene as the diluent. The extraction conditions are: extractant concentration 10%-30%, O / A ratio 1:1-1:3, extraction time 5-15 min. The dilute acid is 0.5 mol / L-2 mol / L sulfuric acid, the back-extraction agent is 2 mol / L-6 mol / L ammonium fluoride or ammonium chloride solution, the back-extraction solution is adjusted to pH 8-10 with ammonia, the calcination temperature is 800℃-1000℃, the calcination time is 1-3 h, the overall niobium recovery rate is ≥70%, and the Nb₂O₅ grade is ≥20%.

[0022] Further, in the preparation process of white carbon black described in step 4, the leaching solution is a sodium hydroxide (NaOH) solution. The silica-rich leaching residue is mixed with the sodium hydroxide (NaOH) solution at a solid-liquid ratio of 1:2 to 1:5. The NaOH solution concentration is 15%-30%, the leaching temperature is 90℃-120℃, the leaching time is 1h-3h, and the pH range is 8-9. The preferred leaching temperature is 100℃-110℃, the preferred leaching time is 1.5h-2.5h, and the preferred NaOH solution concentration is 20%-25%. The SiO2 content in the white carbon black is ≥95%.

[0023] Further, the synergistic extraction system described in step 5 is composed of di(2-ethylhexyl) phosphate (P204) or di(2,4,4-trimethylpentyl)phosphonic acid (Cyanex 272) and tributyl phosphate (TBP). The diluent is sulfonated kerosene. The concentration of P204 or Cyanex 272 is 5%-15%, the concentration of TBP is 3%-8%, the ratio (O / A) is 1:1-1:3, the extraction time is 5-15 min, and the extraction temperature is 20℃-40℃. The back-extraction agent is a 1 mol / L-3 mol / L HCl solution. The molar ratio of oxalic acid to scandium ions is (1.5~3):1. The calcination temperature is 600℃-800℃, the calcination time is 1h-2h, the overall scandium recovery rate is ≥60%, and the Sc2O3 grade in the scandium oxide enrichment is ≥5%.

[0024] The beneficial effects and innovative points of this invention are as follows:

[0025] 1. A combined beneficiation and metallurgical process of "flotation tailings disposal—hydrogen-based mineral phase transformation roasting—mixed acid leaching—niobium extraction—silicon-alkali leaching to produce silica—scandium selective extraction" is proposed, achieving the cascaded synergistic recovery of three valuable components: niobium, silicon, and scandium. In existing technologies, a process flow for the simultaneous recovery of niobium, silicon, and scandium has not been publicly disclosed. This invention, through a stepwise recovery strategy of "flotation pre-enrichment—leaching extract for niobium—raffinate extraction for scandium—alkali leaching of leaching residue to produce silicon," produces three valuable products—industrial niobium oxide, silica, and scandium oxide concentrate—from different nodes of a single process flow, achieving full utilization of metal resources in niobium-containing tailings.

[0026] 2. The introduction of hydrogen-based mineral phase transformation technology for the activation pretreatment of flotation niobium rough concentrate significantly enhances the leaching activity of niobium and scandium, while also possessing the unique advantages of being green and low-carbon. Addressing the common challenge of niobium minerals' stable chemical properties and difficulty in dissolving them through conventional acid leaching, this invention uses hydrogen as a reducing and activating medium to treat niobium minerals and silica-containing gangues at a low temperature of 500℃-700℃, thereby reducing lattice distortion and loosening the structure. Compared to traditional carbothermal reduction (1250℃-1400℃) or calcium chloride roasting (750℃-900℃) processes, the hydrogen-based roasting temperature is reduced to 500℃-700℃, and the use of carbon-based fuels is eliminated at the source, achieving near-zero carbon emissions and aligning with green development.

[0027] 3. A stepwise leaching and recovery process strategy for the three components of niobium, silicon, and scandium was proposed, effectively solving the problems of poor selectivity and low recovery rate in the separation of niobium, silicon, and scandium in traditional processes. This invention fully utilizes the differences in the dissolution behavior of the three components under different leaching systems: in the mixed acid (HCl+HF) leaching stage, niobium and scandium are selectively dissolved into the leaching solution, while silicon is enriched in the leaching residue due to the formation of insoluble silicates, achieving one-step separation of niobium, scandium, and silicon; the raffinate after niobium extraction is further used for scandium extraction and recovery, avoiding the loss of scandium during the niobium extraction process. The stepwise separation mode enables the three components to be optimally enriched, avoiding the difficulties of separation due to mixing.

[0028] 4. Achieve high-value utilization of silicon resources in tailings by preparing white carbon black products (SiO2≥95%) through alkaline leaching and pH adjustment, which can be used as reinforcing fillers in industries such as rubber and plastics. Transforming waste tailings slag into high-value-added products not only solves the environmental problems associated with stockpiling but also realizes the value-added utilization of secondary resources.

[0029] 5. The process is compact, the operating conditions are mild, and the prospects for industrial application are promising. It uses atmospheric pressure leaching (60℃-95℃) instead of high-pressure high-temperature leaching, resulting in lower equipment requirements and higher safety. Flotation tailings removal reduces the amount of wet processing, decreasing reagent consumption and costs. The entire process achieves a niobium recovery rate of ≥70% and a scandium recovery rate of ≥60%, providing an industrially feasible solution for the clean and efficient utilization of low-grade niobium-containing tailings. Attached Figure Description

[0030] Figure 1 This is a schematic diagram of the process flow of the present invention. Detailed Implementation

[0031] The technical solutions of the present invention will be clearly and completely described below with reference to the embodiments of the present invention. It should be noted that the embodiments described in this invention are only used for further explanation and illustration, and are not intended to limit the scope of their application.

[0032] A combined beneficiation and smelting method for the comprehensive recovery of niobium, silicon, and scandium from niobium-bearing tailings is illustrated in the schematic diagram below. Figure 1 As shown, it includes the following steps:

[0033] Step 1: Flotation pretreatment and tail removal;

[0034] Niobium-bearing tailings with a particle size of -0.074 mm were used as raw material. A flotation process of one rougher and two cleaners was employed for tailings disposal. The pulp concentration of the rougher was 25%, and the pulp pH was 7-8. The depressants were one or more of water glass, sodium hexametaphosphate, oxalic acid, tannic acid, and ethylenediaminetetraacetic acid. The collectors were hydroxyxamic acid reagents such as benzoyl hydroxamic acid and octyl hydroxamic acid. The activator was lead nitrate, and the frother was No. 2 oil (pine oil). Low-density or non-magnetic gangue was discarded to obtain niobium rough concentrate. The Nb2O5 grade in the niobium-bearing tailings was 0.2%-0.5%. After pretreatment, the Nb2O5 grade in the niobium rough concentrate was increased to 2%-5%, and the niobium recovery rate was ≥75%.

[0035] Step 2: Hydrogen-based mineral phase transformation roasting;

[0036] The niobium crude concentrate obtained in step 1 was placed in a hydrogen-based mineral phase transformation roasting apparatus. Roasting was carried out under a mixed atmosphere of hydrogen and nitrogen at a temperature of 500℃-700℃ for 30-120 minutes, with hydrogen comprising 10%-40% of the volume and a total gas flow rate of 200 mL / min-1000 mL / min. After roasting, the hydrogen gas was shut off, and nitrogen was continuously introduced to cool to room temperature.

[0037] Step 3: Mixed acid leaching;

[0038] The roasted ore obtained in step 2 was added to a mixed acid solution for leaching. The mixed acid was a combination of hydrochloric acid (HCl) and hydrofluoric acid (HF). The leaching conditions were as follows: the liquid-to-solid ratio of the mixed acid to the roasted ore was 2:1-5:1; the leaching temperature was 60℃-95℃; the leaching time was 2h-6h; the HCl concentration was 2mol / L-6mol / L; and the HF concentration was 0.5mol / L-2mol / L. After leaching, the ore was filtered and separated to obtain a leaching solution containing niobium and scandium, as well as a silica-rich leaching residue. The niobium leaching rate was ≥90%, and the scandium leaching rate was ≥85%.

[0039] Step 4: Extraction and enrichment of niobium and preparation of silica;

[0040] The niobium-containing leachate obtained in step 3 was enriched and separated by solvent extraction. The extractant was tributyl phosphate (TBP) or methyl isobutyl ketone (MIBK), and the diluent was sulfonated kerosene. The extraction conditions were: extractant concentration in the organic phase 10%-30%, O / A ratio 1:1-1:3, and extraction time 5-15 min. The extracted organic phase loaded with metal ions was washed with dilute acid (0.5 mol / L-2 mol / L H₂SO₄) to remove co-extraction impurities, and then back-extracted with 2 mol / L-6 mol / L NH₄F or NH₄Cl solution. Ammonia was added to the back-extraction solution to adjust the pH to 8-10, precipitating niobium hydroxide. After filtration and washing, the precipitate was calcined at 800℃-1000℃ for 1-3 h to obtain an industrial niobium oxide product with Nb₂O₅ grade ≥20%, and the overall niobium recovery rate was ≥70%. The silica-rich leaching residue obtained in step 3 is subjected to alkaline leaching treatment. The leaching residue is mixed with sodium hydroxide (NaOH) solution at a solid-liquid ratio of 1:2-1:5, the NaOH solution concentration is 15%-30%, the leaching temperature is 90℃-120℃, and the leaching time is 1h-3h. After filtration, a sodium silicate solution is obtained. CO2 gas is bubbled into the sodium silicate solution or dilute sulfuric acid is added to adjust the pH to 8-9, precipitating out white carbon black (hydrated silica). After filtration and washing, the white carbon black product is obtained, with a SiO2 content ≥95%.

[0041] Step 5: Selective extraction and removal of scandium;

[0042] The raffinate from the niobium extraction in step 4 was used as the raw material for scandium extraction. Scandium was selectively extracted using a synergistic extraction system consisting of di(2-ethylhexyl) phosphate (P204) or bis(2,4,4-trimethylpentyl)phosphonic acid (Cyanex 272) and tributyl phosphate (TBP). Sulfonated kerosene was used as the diluent. The extraction conditions were: 5%-15% concentration of P204 or Cyanex 272 in the organic phase, 3%-8% concentration of TBP, an O / A ratio of 1:1-1:3, an extraction time of 5-15 min, and an extraction temperature of 20℃-40℃. The organic phase loaded with metal ions was back-extracted with 1 mol / L-3 mol / L HCl solution. The back-extract was precipitated with oxalic acid to obtain scandium oxalate. The molar ratio of oxalic acid to scandium ions was (1.5~3):1. After filtration and washing, the sample was calcined at 600℃-800℃ for 1h-2h to obtain a scandium oxide enrichment with a Sc2O3 grade ≥5%. The overall scandium recovery rate was ≥60%.

[0043] Example 1

[0044] A combined beneficiation and smelting method for the comprehensive recovery of niobium, silicon, and scandium from niobium-bearing tailings, using Bayan Obo niobium-bearing fluorite flotation tailings as raw material, with an Nb₂O₅ grade of 0.36%, a Sc content of 0.038%, and a SiO₂ content of 54.20%, includes the following steps:

[0045] (1) Flotation pretreatment and tailings disposal: Niobium-bearing tailings with a particle size of -0.074 mm were used as raw materials, and a flotation process of one rougher and two cleaners was adopted for tailings disposal. The pulp concentration of the rougher was 25%, the pulp pH was 7.5, the depressant was water glass, the dosage of the rougher was 1000 g / t, and the dosage of the cleaner was 500 g / t; the activator was lead nitrate, the dosage of the rougher was 500 g / t, and the dosage of the cleaner was 250 g / t; the collector was benzohydroxyxamic acid, the dosage of the rougher was 1500 g / t, and the dosage of the cleaner was 750 g / t; the frother was No. 2 oil, the dosage of the rougher was 100 g / t, and the dosage of the cleaner was 50 g / t; no reagents were added in the second cleaner stage. The tailings were discarded after flotation to obtain niobium concentrate. The Nb2O5 grade in the niobium crude concentrate is 3.52%, and the niobium recovery rate is 76.80%.

[0046] (2) Hydrogen-based mineral phase transformation roasting: The niobium crude concentrate obtained in step 1 was placed in a roasting tube, nitrogen was introduced to purge the air, and then a mixture of hydrogen and nitrogen was introduced with a total gas flow rate of 600 mL / min and a hydrogen content of 25%. The roasting was carried out at 600℃ for 60 min, and then the hydrogen was turned off and nitrogen was continuously introduced to cool to room temperature.

[0047] (3) Mixed acid leaching: Roasted ore was taken and a mixed acid solution was added at a liquid-to-solid ratio of 4:1. The mixed acid contained 4 mol / L HCl and 1.2 mol / L HF. The leaching was carried out by stirring at 80℃ for 4 h, followed by filtration to obtain a leaching solution containing niobium and scandium, as well as a silica-rich leaching residue. The niobium leaching rate was 93.20%, and the scandium leaching rate was 87.60%.

[0048] (4) Niobium extraction, enrichment and preparation of silica: The pH of the leachate from step (3) was adjusted to 2.5, and 20% TBP-sulfonated kerosene was used as the organic phase with a ratio of O / A = 1:2 for 10 min. The organic phase loaded with metal ions was washed with 1 mol / L H2SO4 and then back-extracted with 4 mol / L NH4F solution. The back-extracted solution was adjusted to pH 9.0 with ammonia water to generate niobium hydroxide precipitate. After filtration and washing, it was calcined at 900℃ for 2 h to obtain industrial niobium oxide product with a Nb2O5 grade of 26.30%. The niobium recovery rate of the whole process was 72.50%. The silica-rich leachate obtained in step (3) was mixed with 20% NaOH solution at a solid-liquid ratio of 1:3 and leached at 105℃ for 2 h. The sodium silicate solution was obtained by filtration. CO2 gas was introduced into the sodium silicate solution to adjust the pH to 8.5, and silica precipitated. After filtration and washing, silica product was obtained. Tests showed that the SiO2 content in the white carbon black was 96.50%.

[0049] (5) Selective extraction and removal of scandium: The raffinate after niobium extraction in step 4 was used as raw material, and scandium was extracted using a synergistic extraction system of 10% P2O4-5% TBP-sulfonated kerosene, with a ratio of O / A = 1:2, an extraction time of 8 min, and an extraction temperature of 30℃. The organic phase loaded with metal ions was back-extracted with 2 mol / L HCl, and 14 ml of 10% oxalic acid was added to the back-extract to precipitate scandium oxalate. After filtration and washing, the precipitate was calcined at 700℃ for 1.5 h to obtain a scandium oxide enrichment with a Sc2O3 grade of 6.90%. The overall scandium recovery rate was 65.30%.

[0050] Example 2

[0051] A combined beneficiation and smelting method for the comprehensive recovery of niobium, silicon, and scandium from niobium-containing tailings, using niobium tailings as raw material with an Nb₂O₅ grade of 0.28%, a Sc content of 0.031%, and a SiO₂ content of 56.30%, includes the following steps:

[0052] (1) Flotation pretreatment tailings removal: Niobium-containing tailings with a particle size of -0.074mm are used as raw materials and a flotation process of one rougher and two cleaners is adopted for tailings removal. The pulp concentration for roughing flotation was 25%, and the pulp pH was 7.5. The depressant was a combination of water glass and sodium hexametaphosphate (mass ratio 2:1), with a dosage of 900 g / t for roughing, 450 g / t for primary cleaning, and 150 g / t for secondary cleaning. The activator was lead nitrate, with a dosage of 400 g / t for roughing, 200 g / t for primary cleaning, and 100 g / t for secondary cleaning. The collector was benzohydroxyxamic acid, with a dosage of 1400 g / t for roughing, 700 g / t for primary cleaning, and 300 g / t for secondary cleaning. The frother was No. 2 oil, with a dosage of 100 g / t for roughing, 50 g / t for primary cleaning, and 25 g / t for secondary cleaning. Niobium crude concentrate was obtained after flotation. The Nb2O5 grade in the niobium crude concentrate is 3.16%, and the niobium recovery rate is 78.2%.

[0053] (2) Hydrogen-based mineral phase transformation roasting: Niobium crude concentrate was placed in a roasting tube, nitrogen was introduced to purge the air, and then a mixture of hydrogen and nitrogen was introduced with a total gas flow rate of 500 mL / min and a hydrogen content of 22%. The roasting was carried out at 580℃ for 80 min, and then the hydrogen was turned off and nitrogen was continuously introduced to cool to room temperature.

[0054] (3) Mixed acid leaching: Roasted ore was taken and a mixed acid solution was added at a liquid-to-solid ratio of 3.5:1. The mixed acid contained 3.5 mol / L HCl and 0.9 mol / L HF. Leaching was carried out by stirring at 78℃ for 4.5 h, followed by filtration to obtain a leaching solution containing niobium and scandium, as well as a silica-rich leaching residue. The niobium leaching rate was 91.5%, and the scandium leaching rate was 86.8%.

[0055] (4) Niobium extraction, enrichment, and preparation of silica: The pH of the leachate from step (3) was adjusted to 2.2, and 15% TBP-sulfonated kerosene was used as the organic phase with a ratio of O / A = 1:2.5 for 12 min. The loaded organic phase was washed with 1.5 mol / L H2SO4 and then back-extracted with 3.5 mol / L NH4F solution. The back-extract was adjusted to pH 8.5 with ammonia water to generate niobium hydroxide precipitate. After filtration and washing, it was calcined at 850℃ for 2.5 h to obtain industrial niobium oxide product with a Nb2O5 grade of 25.20%. The overall niobium recovery rate was 71.20%. The silica-rich leachate obtained in step (3) was mixed with 22% NaOH solution at a solid-liquid ratio of 1:2.5 and leached at 100℃ for 2.5 h. The sodium silicate solution was obtained by filtration. Dilute sulfuric acid was added to a sodium silicate solution to adjust the pH to 8.5, causing silica to precipitate. After filtration and washing, the silica product was obtained. Testing revealed that the silica contained 95.9% SiO2.

[0056] (5) Selective extraction and removal of scandium: The raffinate after niobium extraction in step 4 was used as raw material, and scandium was extracted using a synergistic extraction system of 12% P2O4-4% TBP-sulfonated kerosene. The O / A ratio was 1:2, the extraction time was 10 min, and the extraction temperature was 25℃. The loaded organic phase was back-extracted with 2.5 mol / L HCl. 11 ml of 10% oxalic acid was added to the back-extract to precipitate scandium oxalate. After filtration and washing, the precipitate was calcined at 680℃ for 2 h to obtain a scandium oxide enrichment with a Sc2O3 grade of 6.30%. The overall scandium recovery rate was 64.50%.

[0057] Example 3

[0058] A combined beneficiation and smelting method for the comprehensive recovery of niobium, silicon, and scandium from niobium-bearing tailings, using niobium tailings as raw material, with an Nb₂O₅ grade of 0.41%, a Sc content of 0.045%, and a SiO₂ content of 52.70%. Specifically, it includes the following steps:

[0059] (1) Flotation pretreatment and tailings disposal: Niobium-bearing tailings with a particle size of -0.074mm were used as raw materials and a flotation process of one rougher and two cleaners was adopted for tailings disposal. The pulp concentration of the rougher was 25%, the pulp pH was 7.5, the inhibitor was tannic acid, the dosage of the rougher was 1100g / t, the dosage of the first cleaner was 600g / t, and the dosage of the second cleaner was 250g / t; the activator was lead nitrate, the dosage of the rougher was 320g / t, the dosage of the first cleaner was 160g / t, and the dosage of the second cleaner was 80g / t; the collector was octyl hydroxamic acid, the dosage of the rougher was 1700g / t, the dosage of the first cleaner was 850g / t, and the dosage of the second cleaner was 300g / t; the frother was No. 2 oil, the dosage of the rougher was 100g / t, the dosage of the first cleaner was 50g / t, and the dosage of the second cleaner was 25g / t. After flotation, niobium crude concentrate was obtained, with an Nb2O5 grade of 4.03% and a niobium recovery rate of 74.90%.

[0060] (2) Hydrogen-based mineral phase transformation roasting: Niobium crude concentrate was placed in a roasting tube, nitrogen was introduced to purge the air, and then a mixture of hydrogen and nitrogen was introduced with a total gas flow rate of 700 mL / min and a hydrogen content of 28%. The roasting was carried out at 630℃ for 50 min, and then the hydrogen was turned off and nitrogen was continuously introduced to cool to room temperature.

[0061] (3) Mixed acid leaching: Roasted ore was taken and a mixed acid solution was added at a liquid-to-solid ratio of 4.5:1. The mixed acid contained 4.5 mol / L HCl and 1.4 mol / L HF. Leaching was carried out by stirring at 88℃ for 3.5 h, followed by filtration to obtain a leaching solution containing niobium and scandium, as well as a silica-rich leaching residue. The niobium leaching rate was found to be 95.10%, and the scandium leaching rate was 88.20%.

[0062] (4) Niobium extraction, enrichment, and preparation of silica: The pH of the leachate was adjusted to 2.8, and 25% MIBK-sulfonated kerosene was used as the organic phase with a ratio of O / A = 1:2 for 8 min. The organic phase loaded with metal ions was washed with 1.8 mol / L H2SO4 and then back-extracted with 5 mol / L NH4Cl solution. The back-extract was adjusted to pH 9.5 with ammonia water to generate niobium hydroxide precipitate. After filtration and washing, it was calcined at 950℃ for 1.5 h to obtain industrial niobium oxide product with a Nb2O5 grade of 27.50%. The niobium recovery rate of the whole process was 73.10%. The silica-rich leachate obtained in step 3 was mixed with 25% NaOH solution at a solid-liquid ratio of 1:3.5 and leached at 110℃ for 1.5 h. The sodium silicate solution was obtained by filtration. CO2 gas was introduced into the sodium silicate solution to adjust the pH to 8.0, and silica precipitated. After filtration and washing, silica product was obtained. Tests showed that the SiO2 content in the white carbon black was 96.8%.

[0063] (5) Selective extraction and removal of scandium: The raffinate after niobium extraction in step 4 was used as raw material, and scandium was extracted using a synergistic extraction system of 8% P2O4-6% TBP-sulfonated kerosene. The O / A ratio was 1:1.5, the extraction time was 12 min, and the extraction temperature was 35℃. Back-extraction was performed with 2 mol / L HCl, and 15 ml of 10% oxalic acid was added to the back-extraction solution to precipitate scandium oxalate. After filtration and washing, the precipitate was calcined at 750℃ for 1.5 h to obtain a scandium oxide enrichment with a Sc2O3 grade of 7.20%. The scandium recovery rate of the entire process was 66.80%.

[0064] Example 4

[0065] Using niobium-bearing tailings of the same type as raw material, the Nb₂O₅ grade was 0.44%, the Sc content was 0.042%, and the SiO₂ content was 55.1%. The steps were basically the same as in Example 1, except that: the inhibitor was changed to a combination of sodium hexametaphosphate and oxalic acid (mass ratio 1:1), the roughing agent was 800 g / t, and the primary cleaning agent was 400 g / t; the collector was changed to octyl hydroxamic acid, the roughing agent was 1800 g / t, and the primary cleaning agent was 900 g / t. The resulting niobium concentrate had an Nb₂O₅ grade of 3.85% and a niobium recovery rate of 75.20%. In step 2, the roasting temperature was 650℃, the roasting time was 50 min, and the hydrogen content was 30%. In step 3, the mixed acid leaching conditions were: HCl 5 mol / L, HF 1.5 mol / L, leaching temperature 85℃, and leaching time 3.5 h. The niobium leaching rate was 94.5%, and the scandium leaching rate was 86.2%. In step 4, industrial niobium oxide with a Nb₂O₅ grade of 27.10% was obtained, with an overall niobium recovery rate of 71.8%, and white carbon black with a SiO₂ content of 95.8% was obtained. In step 5, a scandium oxide enrichment with a Sc₂O₃ grade of 6.20% was obtained, with an overall scandium recovery rate of 63.70%.

[0066] Example 5

[0067] Using niobium-bearing tailings of the same type as raw material, the Nb₂O₅ grade was 0.31%, the Sc content was 0.035%, and the SiO₂ content was 53.80%. The steps were basically the same as in Example 1, except that: the inhibitor was changed to a combination of tannic acid and ethylenediaminetetraacetic acid (mass ratio 2:1), the roughing agent was 1200 g / t, and the primary cleaning agent was 600 g / t; the collector was a combination of benzoyl hydroxamic acid and octyl hydroxamic acid (mass ratio 1:1), the roughing agent was 1600 g / t, and the primary cleaning agent was 800 g / t. The resulting niobium concentrate had an Nb₂O₅ grade of 3.28% and a niobium recovery rate of 77.50%. In step 2, the roasting temperature was 550℃, the roasting time was 90 min, and the hydrogen content was 20%. In step 3, the mixed acid leaching conditions were: HCl 3.5 mol / L, HF 1.0 mol / L, leaching temperature 75℃, and leaching time 5 h. The niobium leaching rate was 91.8%, and the scandium leaching rate was 85.3%. In step 4, industrial niobium oxide with a Nb₂O₅ grade of 24.90% was obtained, with a niobium recovery rate of 70.60% for the entire process, and white carbon black with a SiO₂ content of 96.10% was obtained. In step 5, scandium oxide enrichment with a Sc₂O₃ grade of 5.80% was obtained, with a scandium recovery rate of 62.40% for the entire process.

[0068] Table 1: Results of comprehensive recovery of niobium, silicon, and scandium from niobium-bearing tailings in Examples 1-5

[0069] Based on the embodiments of the present invention, all other embodiments obtained by those skilled in the art without creative effort are within the scope of protection of the present invention.

Claims

1. A combined beneficiation and metallurgical method for the comprehensive recovery of niobium, silicon, and scandium from niobium-bearing tailings, characterized in that, Includes the following steps: Step 1: Flotation pretreatment and tail removal; Using niobium-containing tailings as raw material, a flotation process of one rougher and two cleaners is used to remove the tailings and remove low-density or non-magnetic gangue to obtain niobium rough concentrate. Step 2: Hydrogen-based mineral phase transformation roasting; The niobium crude concentrate obtained in step 1 was placed in a hydrogen-based mineral phase transformation roasting device and roasted in a mixed atmosphere of hydrogen and nitrogen. After roasting, the hydrogen was turned off and nitrogen was continuously introduced to cool to room temperature. Step 3: Mixed acid leaching; The roasted ore obtained in step 2 was added to a mixed acid solution for leaching. After leaching, the ore was filtered and separated to obtain a leaching solution containing niobium and scandium and a silica-rich leaching residue. Step 4: Extraction and enrichment of niobium and preparation of silica; The niobium-containing leachate obtained in step 3 was enriched and separated by solvent extraction. After extraction, the organic phase loaded with metal ions was washed with dilute acid, back-extracted, and the pH was adjusted to obtain niobium hydroxide precipitate. After filtration, washing, and calcination, industrial niobium oxide product was obtained. The silica-rich leachate obtained in step 3 was subjected to alkaline leaching treatment. After leaching, it was filtered to obtain sodium silicate solution. The pH was adjusted to precipitate hydrated silica. After filtration and washing, white carbon black product was obtained. Step 5: Selective extraction and removal of scandium; The raffinate after niobium extraction in step 4 was used as the raw material for scandium extraction. Scandium was selectively extracted using a synergistic extraction system. The organic phase loaded with metal ions was back-extracted. The back-extract was precipitated with oxalic acid to obtain scandium oxalate. After filtration, washing, and calcination, scandium oxide enrichment was obtained.

2. The combined beneficiation and smelting method for comprehensively recovering niobium, silicon, and scandium from niobium-bearing tailings according to claim 1, characterized in that, The niobium-containing tailings raw material described in step 1 has a particle size of -0.074 mm, a pulp concentration of 25% for roughing flotation, a pulp pH of 7-8, and a depressant consisting of one or more of water glass, sodium hexametaphosphate, oxalic acid, tannic acid, and ethylenediaminetetraacetic acid. The collector is a hydroxamic acid reagent, consisting of one or more of benzoyl hydroxamic acid, octyl hydroxamic acid, salicylic acid, and pentyl hydroxamic acid. The activator is lead nitrate, and the frother is No. 2 oil. The Nb2O5 grade in the niobium tailings is 0.2%-0.5%. After pretreatment, the Nb2O5 grade in the niobium concentrate is increased to 2%-5%, and the niobium recovery rate is ≥75%.

3. The combined beneficiation and smelting method for comprehensively recovering niobium, silicon, and scandium from niobium-bearing tailings according to claim 1, characterized in that, The hydrogen-based mineral phase transformation roasting in step 2 is carried out at a roasting temperature of 500℃-700℃, a roasting time of 30min-120min, a hydrogen volume ratio of 10%-40%, and a total gas flow rate of 200mL / min-1000mL / min.

4. The combined beneficiation and smelting method for comprehensively recovering niobium, silicon, and scandium from niobium-bearing tailings according to claim 3, characterized in that, The hydrogen-based mineral phase transformation roasting in step 2 is carried out at a roasting temperature of 550℃-650℃, a roasting time of 45min-90min, and a hydrogen volume ratio of 20%-30%.

5. The combined beneficiation and smelting method for comprehensively recovering niobium, silicon, and scandium from niobium-bearing tailings according to claim 1, characterized in that, In the mixed acid leaching described in step 3, the liquid-solid ratio of the mixed acid to the roasted ore is 2:1-5:

1. The mixed acid is composed of hydrochloric acid and hydrofluoric acid, with a hydrochloric acid concentration of 2 mol / L-6 mol / L and a hydrofluoric acid concentration of 0.5 mol / L-2 mol / L. The leaching temperature is 60℃-95℃, the leaching time is 2h-6h, the niobium leaching rate is ≥90%, and the scandium leaching rate is ≥85%.

6. The combined beneficiation and smelting method for comprehensively recovering niobium, silicon, and scandium from niobium-bearing tailings according to claim 5, characterized in that, In the mixed acid leaching described in step 3, the hydrochloric acid concentration is 3 mol / L-5 mol / L, the hydrofluoric acid concentration is 0.8 mol / L-1.5 mol / L, the leaching temperature is 75℃-85℃, and the leaching time is 3h-5h.

7. The combined beneficiation and smelting method for comprehensively recovering niobium, silicon, and scandium from niobium-bearing tailings according to claim 1, characterized in that, In step 4, the solvent extraction method uses tributyl phosphate or methyl isobutyl ketone as the extractant, and sulfonated kerosene as the diluent. The extraction conditions are as follows: extractant concentration of 10%-30%, organic phase to aqueous phase ratio of 1:1-1:3, extraction time of 5 min-15 min, dilute acid of 0.5 mol / L-2 mol / L sulfuric acid, back-extraction agent of 2 mol / L-6 mol / L ammonium fluoride solution or ammonium chloride solution, ammonia water is added to the back-extraction solution to adjust the pH to 8-10, calcination temperature of 800℃-1000℃, calcination time of 1 h-3 h, niobium recovery rate of the whole process ≥70%, Nb2O5 grade ≥20%.

8. The combined beneficiation and smelting method for comprehensively recovering niobium, silicon, and scandium from niobium-bearing tailings according to claim 1, characterized in that, In the white carbon black preparation process described in step 4, the leaching solution is a sodium hydroxide solution, the silicon-rich leaching residue is mixed with the sodium hydroxide solution at a solid-liquid ratio of 1:2-1:5, the NaOH solution concentration is 15%-30%, the leaching temperature is 90℃-120℃, the leaching time is 1h-3h, the pH range is 8-9, and the SiO2 content in the white carbon black is ≥95%.

9. A combined beneficiation and smelting method for comprehensively recovering niobium, silicon, and scandium from niobium-bearing tailings according to claim 8, characterized in that, In the white carbon black preparation process described in step 4, the leaching temperature is 100℃-110℃, the leaching time is 1.5h-2.5h, and the NaOH solution concentration is 20%-25%.

10. The combined beneficiation and smelting method for comprehensively recovering niobium, silicon, and scandium from niobium-bearing tailings according to claim 1, characterized in that, The synergistic extraction system described in step 5 is composed of di(2-ethylhexyl) phosphate or di(2,4,4-trimethylpentyl)phosphonic acid and tributyl phosphate. The diluent is sulfonated kerosene. The concentration of di(2-ethylhexyl) phosphate or di(2,4,4-trimethylpentyl)phosphonic acid is 5%-15%, the concentration of tributyl phosphate is 3%-8%, the ratio of organic phase to aqueous phase is 1:1-1:3, the extraction time is 5 min-15 min, the extraction temperature is 20℃-40℃, the back-extraction agent is 1 mol / L-3 mol / L HCl solution, the molar ratio of oxalic acid to scandium ions is (1.5~3):1, the calcination temperature is 600℃-800℃, the calcination time is 1 h-2 h, the scandium recovery rate throughout the process is ≥60%, and the Sc2O3 grade in the scandium oxide enrichment is ≥5%.

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