Beneficiation method for pre-enriching niobium ore and recovering pyrochlore from calcined digestion heavy ore

Abstract

The invention discloses a beneficiation method for pre-enriching niobium ore and recovering pyrochlore from calcined digestion heavy beneficiation, in a pretreatment stage, through ore grinding, magnetic separation and acid leaching, mineral structure rearrangement is induced on the basis of calcination, active sites of minerals such as pyrochlore, phlogopite and the like are remarkably exposed, and the selectivity and interface reactivity of subsequent flotation separation are improved. A direct and reverse flotation combined process is adopted in the flotation stage, reverse flotation is conducted firstly to remove phlogopite, then direct flotation is conducted to enrich pyrochlore, or combined enrichment and subsequent separation of pyrochlore and barite are achieved. Aiming at the problem of insufficient selectivity of a traditional agent, a molecular simulation technology is introduced, a mineral surface adsorption model is constructed, a modified collecting agent with high affinity with pyrochlore is screened, and the separation performance of the modified collecting agent is verified in actual flotation. According to the technology, under the synergistic effect of multi-scale mineral dissociation regulation and molecular interface engineering, the comprehensive recovery rate of niobium reaches 75% or above, and the concentrate grade is stabilized to be 51.5% or above.

Description

Technical Field

[0001] The present invention belongs to the technical field of mineral processing. More specifically, it relates to a beneficiation method for pre-enriching niobium ore and recovering pyrochlore from calcined digestion re-concentration, and particularly relates to a flotation method for recovering pyrochlore from ores with carbonate and silicate as the main gangue minerals. Background Art

[0002] Niobium is a rare and refractory metal with unique properties such as high boiling point, high melting point, low vapor pressure, excellent ductility, strong corrosion resistance, and good superconducting properties. Therefore, niobium is widely used in fields such as aerospace, biomedicine, optoelectronic materials, catalysts, electronic machinery, and military industry.

[0003] Niobium minerals are mainly distributed in carbonate-type and granitic pegmatite-type deposits. Although there are many types, the main 10 types as industrial raw materials include columbite, tantalite series minerals, fergusonite, euxenite, niobium euxenite, niobium rutile, and pyrochlore, etc. Among them, carbonate-type pyrochlore has become the main source of niobium metal due to its high niobium content, accounting for 95% of the global niobium supply. Therefore, the development of carbonate-type pyrochlore is crucial for the stability of the global niobium supply.

[0004] The existing beneficiation methods for niobium ore usually involve crushing and grinding the ore and then separating it by flotation. However, due to the high brittleness of niobium-containing minerals, they are easily broken, while carbonate gangue minerals have low hardness and are prone to slime formation during the grinding process, which has an adverse impact on the subsequent flotation operation. Therefore, a desliming process is usually required after grinding to reduce the interference of fine slime on the flotation operation, but this process will also result in the loss of some valuable metals. For example, in the "grinding-desliming-flotation" method for a certain pyrochlore-type niobium ore, the niobium loss rate in the single desliming process is as high as 14%.

[0005] Compared with the present invention, directly grinding and flotation without calcination will result in a low total recovery rate of niobium. The research object of the patent document (application number 201910721915.4) is weathered carbonate-type niobium polymetallic ore. First, spiral tailing discarding is used to obtain rough concentrate. For the rough concentrate, grinding and flotation are carried out to obtain phosphorus concentrate. The tailings after phosphorus flotation are subjected to weak magnetic separation to obtain magnetite concentrate. The tailings after weak magnetic separation are subjected to strong magnetic separation to obtain strong magnetic products. The tailings after strong magnetic separation are separated by a shaking table to obtain pyrochlore concentrate, with the Nb2O5 content reaching 50.18% and the recovery rate reaching 82.07%, realizing the recovery of barium-strontium pyrochlore from weathered carbonate-type niobium polymetallic ore. A method for recovering calcium-rich pyrochlore from weathered high-slime carbonate-type niobium polymetallic ore (application number 201910722543.7) in a patent document uses stirring and scrubbing to remove the fine slime adhering to the mineral surface, then uses spiral tailing discarding and desliming to obtain rough concentrate. For the rough concentrate, grinding and flotation are carried out to obtain phosphorus concentrate. The tailings after phosphorus flotation are subjected to calcium-rich pyrochlore flotation to obtain niobium concentrate, with the Nb2O5 grade of 27.93% and the total recovery rate of 79.43%. The raw materials used in the above two patents are both weathered ores, and the raw ore is first deslimed, then ground and then floated. The desliming process will cause a relatively high loss of niobium, reducing the total recovery rate of niobium, so it is difficult to achieve the best indicators for both grade and recovery rate at the same time.

[0006] Compared with the present invention, after calcination, digestion, gravity separation and then carbonitriding calcination, the maximum enrichment of niobium cannot be achieved, and the niobium grade is relatively low. A patent document (application number 202111348210.6) proposed a pre-enrichment and beneficiation method for niobium ore. The niobium ore is calcined at 800°C to 1100°C and then digested in water, and then gravity separation is carried out. The heavy product is the pre-enriched concentrate, and the light product is the pre-enriched tailings. Finally, the pre-enriched concentrate is subjected to carbonitriding calcination. After the pre-treatment of calcination, digestion and gravity separation in this invention, the tailing discarding rate is about 40%, and the niobium recovery rate is 94.7%. It is mentioned in this patent that carbonitriding roasting of niobium ore and combined control of temperature can unexpectedly improve the recovery rate and grade of niobium ore, but no specific data on grade improvement are provided. In the previous pre-treatment stage, niobium is only enriched 1.37 times, and the carbonitriding calcination process has the problems of high energy consumption and it is finally difficult to obtain high-grade pyrochlore niobium concentrate.

[0007] The present invention discloses a method for the pre - enrichment of niobium ore and the efficient recovery of pyrochlore applicable to the calcination - digestion - re - beneficiation system, which carries out systematic innovation around the three core issues of "ore - phase regulation - interface evolution - intelligent flotation". The technological breakthroughs are mainly reflected in three aspects: First, in terms of the collaborative regulation of the process chain, the present invention realizes the selective dissociation and interface differential regulation of pyrochlore and gangue minerals by introducing calcination to induce the crystal - structure reconstruction of calcite and magnetite, and the transformation of the surface physical and chemical properties of pyrochlore, phlogopite, barite and ilmenite, combined with magnetic separation and multi - stage acid leaching purification, thereby establishing an ideal physical - phase and surface - energy basis for subsequent flotation separation. Second, for the main minerals pyrochlore, phlogopite and barite after pretreatment, in terms of constructing the flotation - separation path, aiming at the interface characteristics of multi - component minerals such as pyrochlore, phlogopite and barite, the present invention proposes two flotation strategies of "first reverse then positive" and "positive - flotation combined enrichment + subsequent differential separation", breaking through the bottleneck of insufficient separation selectivity of the traditional single - flotation mode, and significantly improving the directional recognition and recovery efficiency of the flotation system. Finally, in terms of the intelligent screening of the reagent system and the analysis of the interface - action mechanism, the present invention introduces first - principles and molecular - dynamics simulation methods to reconstruct the surface model of the treated minerals at the atomic scale, accurately reveals the adsorption configuration and energy change of the collector on different crystal planes, constructs a "coupling mechanism of adsorption energy - structure response - functional - group orientation", realizes the functional optimization design of the reagent - molecule structure, and promotes the evolution of the flotation system towards high selectivity and low reagent consumption. This method realizes the efficient utilization of complex niobium - ore resources under multi - stage pretreatment and interface regulation, has both theoretical innovation and industrial adaptability, and has broad promotion value. Summary of the Invention

[0008] The present invention aims to propose a beneficiation method for pre - enriching niobium ore and recovering pyrochlore from calcination - digestion - re - beneficiation to solve the following technical problems: First, by grinding, magnetic separation and acid - leaching treatment of the calcination - digestion - re - beneficiation concentrate, calcite, iron - containing minerals and phosphorus - containing minerals can be effectively removed, while the slime content of pyrochlore minerals can be reduced, and the selectivity of subsequent flotation can be improved. Second, the calcination and acid - leaching treatment can activate the mineral surface, making its interaction with reagents in the subsequent flotation process more adaptable, thereby realizing the efficient enrichment and recovery of pyrochlore - type niobium ore. Compared with the BV and BVFR processes, the present invention can significantly improve the recovery rate of niobium while maintaining the grade of niobium concentrate, making the final grade of niobium concentrate stable above 51.5% and the recovery rate increased to above 75%.

[0009] A beneficiation method for pre - enriching niobium ore and recovering pyrochlore from calcination - digestion - re - beneficiation according to the present invention comprises the following steps:

[0010] Step 1: Pretreatment

[0011] Taking the calcined digestion and re-concentration ore as the treatment object, the treatment object is pre-treated to obtain the slag after acid leaching; the pre-treatment includes grinding, magnetic separation, and acid leaching; the particle size of the original ore is less than 30 mm, and the particle size of the calcined digestion and re-concentration ore is less than 3 mm;

[0012] Step 2: Flotation operation

[0013] The slag after acid leaching is subjected to the combination of positive flotation and reverse flotation to finally obtain flotation concentrate and flotation tailings.

[0014] Preferably, the calcined digestion and re-concentration ore refers to the re-concentration concentrate obtained after the original ore is treated by the processes of calcination, digestion, and re-concentration. During the calcination process, calcite in the ore undergoes a decomposition reaction to generate calcium oxide and release carbon dioxide; in the subsequent digestion process, the calcium oxide is further converted into slaked lime. The re-concentration step effectively separates and removes the slaked lime, thereby removing the calcium-containing gangue minerals in the original ore and improving the occurrence grade of niobium. After the above multi-stage treatment, the structures of calcite and magnetite are broken, calcite is basically removed, and part of the magnetite is oxidized and transformed into hematite; at the same time, the surface structures of pyrochlore, barite, phlogopite, and ilmenite also undergo rearrangement, creating favorable conditions for subsequent deep treatment processes such as grinding, magnetic separation, acid leaching, and flotation.

[0015] Preferably, for the calcined digestion and re-concentration treatment, the calcination temperature is 800 - 1050 °C, the calcination time is 5 - 120 min, the calcination atmosphere is an oxidizing atmosphere, the liquid-solid ratio of the starting solution for digestion is 1:1 - 6:1 L / kg; the temperature during the digestion process is 40 °C - 95 °C, the digestion time is 5 min - 60 min. There is a large density difference and particle size difference between the pyrochlore mineral and the slaked lime. The digestion ore is re-concentrated using a shaking table, and the angle of the shaking table is 3 - 20°, obtaining the re-concentration concentrate and the re-concentration tailings of slaked lime.

[0016] Preferably, the pre-treatment process sequentially includes the steps of grinding, magnetic separation, and acid leaching. Before the grinding treatment, the pyrochlore mineral still maintains a relatively large particle size of -3 mm, effectively avoiding its slime phenomenon. Based on the calcination treatment, during the grinding process, it is easier to achieve the efficient dissociation of the target mineral pyrochlore compared to the uncalcined ore, and at the same time, iron-containing minerals such as magnetite, hematite transformed by calcination, and ilmenite are released.

[0017] To further remove the above-mentioned iron impurities, the grinding product needs to be subjected to magnetic separation to obtain magnetic concentrate and magnetic tailings. Subsequently, the magnetic tailings are subjected to acid leaching treatment to preferentially dissolve residual impurity minerals such as slaked lime and apatite. This acid leaching process further exposes the active sites of minerals such as pyrochlore, barite, and phlogopite on the basis of the calcination-induced rearrangement of the mineral surface structure, thereby providing more ideal interfacial chemical conditions and selectivity basis for subsequent flotation separation.

[0018] Preferably, the grinding fineness is that the proportion of -0.074mm is 40.0 - 85.0%.

[0019] The magnetic separation intensity is 0.2 - 1.6T, preferably 0.2 - 0.6T.

[0020] In the acid leaching process, hydrochloric acid or sulfuric acid is used for acid leaching. The concentration of the acid is 0.5 - 10mol / L, the ratio of solid mass to liquid volume is 1:1 - 10:1, and the leaching time is 5 - 60min.

[0021] Preferably, after pretreatment, the tailing rejection rate is greater than 55% (tailing rejection rate R t = , : mass of tailings, : mass of raw ore), the niobium recovery rate is greater than 90% (total niobium recovery rate , : mass of the final niobium concentrate, : niobium grade in the niobium concentrate, : mass of raw ore, : niobium grade in the raw ore), and the niobium grade in the slag after acid leaching is greater than 1.75%.

[0022] Preferably, after pretreatment, the main minerals of the calcined digestion re - beneficiation include pyrochlore, barite, phlogopite, etc. The tailing rejection rate is greater than 55%, the niobium recovery rate is greater than 90%, and the niobium grade in the slag after acid leaching is greater than 1.75%.

[0023] Preferably, the pyrochlore exposes Nb 5+ , Ca 2+ sites, generating a metal oxide - like surface (such as Ti - OH / Nb - OH), enhancing the adsorption capacity; the surface of phlogopite exposes unsaturated bond sites such as Al - OH and Mg - OH, becoming highly active adsorption centers; the surface exposes Ba²⁺ sites, which are suitable for forming collector complex precipitation, resulting in a generally enhanced affinity of the above - mentioned minerals for the collector. In this system, it is difficult to effectively separate pyrochlore by traditional single positive flotation or reverse flotation methods. Therefore, a differential flotation strategy is proposed: First, preferentially use the reverse flotation process to selectively remove phlogopite, and then recover pyrochlore by positive flotation; Second, the combined enrichment of pyrochlore and barite can be achieved by positive flotation, and then the two can be separated by the selective inhibition of barite by an inhibitor, so as to achieve the efficient extraction of pyrochlore and the effective removal of impurity minerals.

[0024] It is found in the research of the present invention that by innovatively calcining and acid-leaching pyrochlore-type niobium ore, the pyrochlore ore phase remains unchanged. The calcination and acid-leaching treatment not only make the minerals easier to dissociate, change the surface chemical composition of the minerals, but also significantly enhance the adsorption effect of flotation reagents (such as collectors, inhibitors and activators) on the mineral surface by increasing the active sites on the mineral surface. At the same time, when the tailing rejection rate is relatively high, by innovatively flotation of the pre-concentrated concentrate, it is possible to reduce the dosage of reagents and save costs while significantly improving the recovery rate and grade of niobium in the pyrochlore-type niobium ore. The research also finds that by further combining the forward and reverse flotation based on the pre-concentration process of the present invention, the reverse flotation removes silicon-containing or barium-containing minerals and inhibits the useful mineral pyrochlore, while the forward flotation inhibits gangue minerals, and finally floats out pyrochlore, ultimately significantly improving the recovery rate and grade of niobium in the pyrochlore-type niobium ore.

[0025] Preferably, according to the exposure of more active sites in pyrochlore, phlogopite and barite after calcination and acid leaching, the combined forward and reverse flotation process further optimizes and screens the types of reagents. In the reverse flotation process, the inhibitors used can be selected from at least one of carboxymethyl cellulose, starch, gellan gum, dextrin, water glass, polyacrylamide, tannic acid, sodium lignosulfonate, sodium hexametaphosphate or sodium fluorosilicate, preferably two of them; the collectors used can be at least one of oleic acid, sodium oleate, oleic acid soap, alkylamine, oleoyl alanine or oleic acid amide, preferably two of them.

[0026] In the reverse flotation, hydrochloric acid and sodium hydroxide are used to adjust the pH, and the pH is controlled at 2-11, preferably 7-11, and further preferably 10-11.

[0027] In the roughing process of the reverse flotation, the dosage of the inhibitor is 250 g / t to 2000 g / t, preferably 400-800 g / t, the dosage of the collector is 250 g / t to 2000 g / t, preferably 750-1250 g / t, and the dosage of the collector in the first scavenging is 50-500 g / t, preferably 300-500 g / t.

[0028] In the present invention, the reverse flotation can achieve floating phlogopite and inhibiting pyrochlore and barite.

[0029] Through the theoretical calculation and exploration of the present invention and the actual application in the factory, in the reverse flotation, the inhibitor is preferably starch, and the collector is preferably a mixed collector composed of ether amine and oleoyl alanine in a mass ratio of 1:3.

[0030] In the positive flotation, the niobium regulator is at least one of oxalic acid, acetic acid, hydrofluoric acid, fluorosilicic acid, sodium fluorosilicate, and sodium silicate; the activator is at least one of lead nitrate, calcium chloride, and sodium chloride, preferably two of them; the collector is at least one of sodium oleate, benzohydroxamic acid, salicylhydroxamic acid, xanthate, dithiophosphate, oleic acid, palmitic acid, alkylamine, tallow diamine, modified sodium oleate, modified benzohydroxamic acid, modified salicylhydroxamic acid, modified xanthate, modified dithiophosphate, modified oleic acid, modified palmitic acid, modified alkylamine, and modified tallow diamine, preferably two of them. In actual application, in the positive flotation, when the adsorption energy between the collector and the target mineral is less than or equal to -140 kJ / mol, it can be regarded as an available collector. The modified group is selected from at least one of ureido, hydroxamic, phosphonic, and carboxyl groups.

[0031] The pH in the positive flotation is adjusted with hydrochloric acid and sodium hydroxide, and the pH is controlled at 2 - 11, preferably 6 - 8.

[0032] In the rough selection process of pyrochlore by positive flotation, the dosage of the regulator is 250 g / t - 2000 g / t, preferably 400 - 600 g / t; the dosage of the activator is 500 g / t - 2000 g / t, preferably 800 - 1200 g / t; the dosage of the collector is 250 g / t - 2000 g / t, preferably 750 - 1250 g / t; the dosage of the collector in the first scavenging is 250 g / t - 2000 g / t, preferably 750 - 1250 g / t.

[0033] Preferably, when the conventional flotation reagent system cannot achieve the dual improvement of niobium grade and recovery rate, the optimization of the collector will be regulated based on the DFT (density functional theory) simulation results of the collector with the surfaces of pyrochlore, phlogopite, and barite minerals.

[0034] Molecular simulation uses the first principles to optimize the crystal structures of three minerals, pyrochlore, phlogopite, and barite, constructs a thermodynamically stable crystal plane model, and simulates the surface exposure characteristics under flotation conditions. The model is based on the mineral structure after calcination and acid leaching treatment, reflecting the reconstruction behavior of the crystal surface active sites: more Nb and Ca active sites are exposed on pyrochlore, Al and Mg sites appear on phlogopite due to dehydroxylation, and the surface exposure of Ba ions is enhanced on barite.

[0035] On this basis, the selective adsorption ability of the collector for pyrochlore is improved by introducing functional groups (such as ureido, hydroxamic, phosphonic, and carboxyl groups) into the collector molecule. The calculation of the adsorption energy and the configuration simulation of various functional groups on the surfaces of different minerals reveal the matching relationship between the crystal surface structure and the interaction of the reagent molecules, providing a theoretical basis for achieving high - selectivity separation.

[0036] Preferably, the adsorption energy calculation is through the adsorption energy formula: ∆E = Ecomplex -E mineral +E agent Calculate the adsorption energy between the mineral surface and the reagent molecule, or determine whether there is an adsorption behavior between the mineral surface and the reagent molecule; wherein, ∆E represents the adsorption energy between the mineral surface and the reagent molecule, and E complex represents the total energy of the adsorption model of the mineral surface and the reagent molecule, and E mineral represents the energy of the mineral surface model, and E agent represents the energy of the reagent molecule model; the units of ∆E, E complex , E mineral and E agent are in eV;

[0037] The selection basis of the collector is to select the reagent with the adsorption energy less than or equal to -140 kJ / mol between the mineral surface as the available flotation collector. In actual application, the best flotation collector is explored through experiments among the available flotation collectors.

[0038] In actual application, during positive flotation, the present invention reduces the adsorption energy between the collector and the target mineral surface to less than or equal to -140 kJ / mol through compounding, modification and other methods.

[0039] Preferably, according to the simulation calculation results, the modified collector is selected to conduct flotation experiments on the actual ore sample. The flotation experiments include positive flotation and reverse flotation; the reagent regimes of positive flotation and reverse flotation are carried out according to the corresponding regimes described above in the present invention.

[0040] Preferably, the grade of niobium in the finally obtained pyrochlore niobium concentrate is greater than 51.5%, and the recovery rate is greater than 75%.

[0041] Preferably, the Nb2O5 in the original pyrochlore niobium ore is greater than or equal to the minimum industrial grade, that is, the grade of Nb2O5 is greater than or equal to 0.016%.

[0042] For the present invention, the grades of the components in the particularly applicable original pyrochlore niobium ore are as follows:

[0043] The grade of Nb2O5 is less than or equal to 1.71%, preferably 0.3% - 1.71%, more preferably 0.4 - 1.71%, the grade of SiO2 is 15.23% - 30.06%, the grade of BaO is 0.53% - 6.32%, the grade of Fe is 9.44% - 14.76%, the grade of CaO is 12.23% - 25.80%, the grade of S is 0.21% - 1.49%, and the grade of P is 0.21% - 1.32%.

[0044] When the present invention processes the raw ore with the grade of Nb2O5 being 0.4%, its effect is far better than the prior art, specifically manifested in its enrichment effect and recovery rate.

[0045] Compared with the prior art, the beneficial effects of the present invention are as follows:

[0046] Synergistic optimization of mineral liberation degree and process efficiency. The present invention induces the weakening of the mineral crystal interface through calcination digestion and gravity separation treatment, significantly improving the liberation degree of pyrochlore and gangue minerals. At the same time, it effectively inhibits the over-grinding and slime-forming phenomenon of minerals. Due to the phase reconstruction and surface energy regulation of the calcined mineral system, it shows higher mechanical liberation efficiency during the grinding process, greatly reducing the grinding energy consumption and time cost, and providing high-purity pre-concentrated raw materials for subsequent separation.

[0047] Intelligent modeling enabling of mineral phase reconstruction and surface characteristics. The calcination-acid leaching coupling process triggers the directional evolution of the mineral phase and surface physicochemical properties of minerals: the selective exposure of active sites on the surface of pyrochlore, and the phase change or dissolution and stripping of gangue minerals (such as calcite and phlogopite). The surface model of the calcined mineral constructed based on the first-principles accurately reveals the spatial distribution and electronic state evolution law of Nb / Ca active sites, and combines molecular dynamics simulation to quantify the adsorption thermodynamic advantages of the collector-mineral interface, breaking through the traditional empirical drug selection mode, and realizing the intelligent screening of flotation reagents and the in-depth analysis of the surface reaction mechanism.

[0048] Differentiated modeling driving the innovation of flotation process. The comparative study of the mineral surface models before and after calcination-acid leaching shows that the calcination-acid leaching treatment significantly enhances the chemical reaction activity of the pyrochlore surface, while the gangue minerals show differentiated interface characteristics due to phase change. This differentiation provides a theoretical support for the positive-reverse flotation combined process: in the reverse flotation stage, silicon / barium minerals are selectively removed through surface hydrophobicity regulation; in the positive flotation stage, the efficient directional separation of niobium minerals is achieved by the strong specific adsorption of the activated sites on the pyrochlore surface and the reagent. This strategy breaks through the technical bottleneck of single flotation, and achieves the synergistic improvement of flotation selectivity and recovery efficiency under low reagent intervention, providing a solution for the high-value utilization of complex niobium ore resources. Description of the Drawings

[0049] Figure 1 It is a flow schematic diagram of the combined beneficiation method of pre-concentration - reverse flotation - positive flotation for primary pyrochlore niobium ore.

[0050] Figure 2 It is a flow schematic diagram of the combined beneficiation method of pre-concentration - positive flotation - reverse flotation for primary pyrochlore niobium ore.

[0051] Figure 3 It is the modeling of the collector on the mineral surface in the molecular simulation calculation of Example 7. Detailed Embodiments

[0052] The present invention will be further described below in conjunction with embodiments, but the embodiments do not limit the present invention in any form. Without departing from the spirit and essence of the present invention, simple modifications or substitutions made to the methods, steps or conditions of the present invention all fall within the scope of the present invention; if not specifically specified, the technical means used in the embodiments are conventional means well-known to those skilled in the art. The verification examples of the present invention include Examples 8 and 9, and Examples 8 and 9 are used to verify whether the conclusion of the adsorption strength between the collector and the calcined acid-leached ore in Example 7 conforms to the process route designed by the present invention.

[0053] In the examples and comparative examples, for the process involving gravity separation, a shaking table is used for heavy separation, and the angle of the shaking table is about 3 - 6°.

[0054] Example 1

[0055] Step (1): The content of Nb2O5 in a certain pyrochlore ore is 1.2%, and the main associated minerals are calcite, dolomite, feldspar, etc. The content of CaO is 19.4%, and the total loss on ignition is 17.3%. First, the ore is crushed to less than 30 mm, and then pre-calcined in a calciner at a pre-calcination temperature of 900 °C for 60 minutes, and the calcination environment is an oxidizing atmosphere (air). The loss on ignition after calcination accounts for 82.7% of the total loss on ignition of the ore. The pre-calcined product is digested in water with a liquid-solid ratio of 3:1 L / kg, a digestion water temperature of 90 °C, and a digestion time of 10 minutes. Finally, gravity separation is performed on the digested product to obtain the heavy product as the pre-enriched concentrate and the light product as the pre-enriched tailings.

[0056] Step (2): Weak magnetic operation. The pre-enriched concentrate (with a particle size of less than 2 mm) is ground to a proportion of -0.074 mm accounting for 72 wt%, and then enters a weak magnetic operation with a magnetic field intensity of 0.2 T to obtain a weak magnetic concentrate and a weak magnetic tailings.

[0057] Step (3): Strong magnetic operation. The obtained weak magnetic concentrate enters a strong magnetic operation with a magnetic field intensity of 0.6 T to obtain a strong magnetic concentrate and a strong magnetic tailings, and the finally obtained strong magnetic concentrate enters the next operation.

[0058] Step (4): Leaching operation. The obtained strong magnetic concentrate is added with 2 mol / L hydrochloric acid according to the ratio of solid mass to liquid volume of 3:1, stirred for 1 h, and then filtered to obtain a high-grade niobium concentrate.

[0059] Step (5): Desilication reverse flotation operation. The leached niobium concentrate is placed in a flotation cell with a pulp concentration of 27%. Sodium hydroxide and hydrochloric acid are used to adjust the pH to about 10.5. 500 g / t of inhibitor starch is added. After stirring for 3 minutes, 1000 g / t of combined collector ether amine / oiloyl alanine is added for roughing operation (the mass ratio of ether amine to oiloyl alanine in the collector is 1:3), obtaining a first roughing concentrate and a first roughing tailing; 250 g / t of inhibitor starch is added to the first roughing tailing. After stirring for 3 minutes, 500 g / t of combined collector ether amine / oiloyl alanine is added for scavenging operation (the mass ratio of ether amine to oiloyl alanine in the collector is 1:3), obtaining a first scavenging concentrate and a first scavenging tailing; the first roughing concentrate is stirred for 3 minutes and then subjected to cleaning operation, obtaining a first cleaning concentrate and a first cleaning tailing.

[0060] Step (6): Niobium flotation operation. The above-mentioned first scavenging concentrate and first cleaning concentrate are combined and then subjected to direct flotation. Sodium hydroxide and hydrochloric acid are used to adjust the pH to about 7.0. 500 g / t of regulator sodium fluorosilicate is added. After stirring for 3 minutes, 1000 g / t of activator lead nitrate is added. After stirring for 3 minutes, 1000 g / t of combined collector sodium oleate / salicylohydroxamic acid is added for a first roughing operation (the mass ratio of sodium oleate to salicylohydroxamic acid in the collector is 1:1), obtaining a roughing concentrate and a roughing tailing; 250 g / t of regulator sodium fluorosilicate is added to the roughing tailing. After stirring for 3 minutes, 1000 g / t of activator lead nitrate is added. After stirring for 3 minutes, 1000 g / t of combined collector sodium oleate / salicylohydroxamic acid is added for a first scavenging operation (the mass ratio of sodium oleate to salicylohydroxamic acid in the collector is 1:1), obtaining a scavenging concentrate and a scavenging tailing; the roughing concentrate is stirred for 3 minutes and then subjected to a first cleaning operation, obtaining a first cleaning concentrate and a first cleaning middling. The scavenging concentrate and the first cleaning middling are combined and returned to the roughing operation; the first cleaning concentrate is stirred for 3 minutes and then subjected to a second cleaning operation, obtaining a second cleaning concentrate and a second cleaning middling. The second cleaning middling is returned to the first cleaning operation. Similarly, the third cleaning middling is returned to the second cleaning operation, and the fourth cleaning middling is returned to the third cleaning operation. The obtained fourth cleaning concentrate is the final pyrochlore concentrate.

[0061] The tailing rejection rate in the pre-concentration stage is 56.6%, the grade of Nb2O5 is 2.65%, and the recovery rate is 91.6%; in the combined direct and reverse flotation stage, the grade of Nb2O5 is 52.63% and the recovery rate is 85.47%. Using traditional flotation-based beneficiation methods to treat this pyrochlore ore, the niobium recovery rate is only about 50%. The beneficiation method based on pre-concentration - flotation can increase the Nb2O5 recovery rate by about 35% while maintaining the Nb2O5 grade of the traditional flotation-based beneficiation method.

[0062] Example 2

[0063] Step (1): The Nb2O5 content of a certain pyrochlore ore is 1.2%. The main associated minerals are calcite, dolomite, feldspar, etc. The CaO content is 19.4%, and the total loss on ignition is 17.3%. First, the ore is crushed to less than 30 mm, and then pre-calcined in a calciner at a pre-calcination temperature of 900 °C for 60 minutes. The calcination environment is an oxidizing atmosphere (air). The loss on ignition after calcination accounts for 82.7% of the total loss on ignition of the ore. The pre-calcined product is digested in water with a liquid-solid ratio of 3:1 L / kg, a digestion water temperature of 90 °C, and a digestion time of 10 minutes. Finally, gravity separation is performed on the digested product to obtain the heavy product as the pre-concentrated concentrate and the light product as the pre-concentrated tailings.

[0064] Step (2): Low-intensity magnetic separation operation. The pre-concentrated concentrate (with a particle size of less than 2 mm) is ground to a proportion of -0.074 mm accounting for 72 wt% and enters the low-intensity magnetic separation operation with a magnetic field strength of 0.2 T to obtain low-intensity magnetic concentrate and low-intensity magnetic tailings.

[0065] Step (3): High-intensity magnetic separation operation. The obtained low-intensity magnetic concentrate enters the high-intensity magnetic separation operation with a magnetic field strength of 0.6 T to obtain high-intensity magnetic concentrate and high-intensity magnetic tailings. The finally obtained high-intensity magnetic concentrate enters the next operation.

[0066] Step (4): Leaching operation. The obtained high-intensity magnetic concentrate is added with 2 mol / L hydrochloric acid according to the ratio of solid mass to liquid volume of 3:1 and stirred for 1 h, and then filtered to obtain high-grade niobium concentrate.

[0067] Step (5): Desilication reverse flotation operation. The leached niobium concentrate is placed in a flotation cell with a pulp concentration of 27%. Sodium hydroxide and hydrochloric acid are used to adjust the pH to about 10.5. 500 g / t of inhibitor starch is added, and after stirring for 3 minutes, a combined collector ether amine / oiloyl alanine 1000 g / t is added for roughing operation (the mass ratio of ether amine to oiloyl alanine in the collector is 1:3) to obtain a first roughing concentrate and a first roughing tailings; 250 g / t of inhibitor starch is added to the first roughing tailings, and after stirring for 3 minutes, a combined collector ether amine / oiloyl alanine 500 g / t is added for scavenging operation (the mass ratio of ether amine to oiloyl alanine in the collector is 1:3) to obtain a first scavenging concentrate and a first scavenging tailings; the first roughing concentrate is stirred for 3 minutes and then subjected to cleaning operation to obtain a first cleaning concentrate and a first cleaning tailings.

[0068] Step (6): Niobium flotation operation. The above-mentioned first scavenging concentrate and first cleaning concentrate are combined and then enter the positive flotation. Sodium hydroxide and hydrochloric acid are used to adjust the pH to about 5.0. 500 g / t of the modifier fluorosilicic acid is added. After stirring for 3 minutes, 500 g / t of the combined collector tallow diamine / oleic acid (the mass ratio of tallow diamine to oleic acid in the combined collector is 1:1) is added for a first roughing operation to obtain a roughing concentrate and a roughing tailing; 250 g / t of the modifier sodium fluorosilicate is added to the roughing tailing. After stirring for 3 minutes, 250 g / t of the combined collector tallow diamine / oleic acid (the mass ratio of tallow diamine to oleic acid in the combined collector is 1:1) is added for a first scavenging operation to obtain a scavenging concentrate and a scavenging tailing; the roughing concentrate is stirred for 3 minutes and then undergoes a first cleaning operation to obtain a first cleaning concentrate and a first cleaning middling; the scavenging concentrate and the first cleaning middling are combined and returned to the roughing operation; the first cleaning concentrate is stirred for 3 minutes and then undergoes a second cleaning operation to obtain a second cleaning concentrate and a second cleaning middling; the second cleaning middling is returned to the first cleaning operation. Similarly, the third cleaning middling is returned to the second cleaning operation, and the fourth cleaning middling is returned to the third cleaning operation. The obtained fourth cleaning concentrate is the final pyrochlore concentrate.

[0069] The tailing rejection rate in the pre-enrichment stage is 56.6%, the Nb2O5 grade is 2.65%, and the recovery rate is 91.6%; in the combined positive and reverse flotation stage, the Nb2O5 grade is 52.56% and the recovery rate is 82.47%. For the traditional flotation-based ore dressing method to treat this pyrochlore ore, the niobium recovery rate is only about 50%. The ore dressing method based on pre-enrichment - flotation can increase the Nb2O5 recovery rate by about 32.0% while maintaining the Nb2O5 grade of the traditional flotation-based ore dressing method.

[0070] Example 3

[0071] Step (1): The Nb2O5 content of a certain pyrochlore ore is 0.4%. The main associated minerals are calcite, dolomite, apatite, etc. The CaO content is 25%, and the total ignition loss is 18.9%. First, the ore is crushed to less than 30 mm, and then pre-calcined in a calciner. The pre-calcination temperature is 900 °C, lasting for 60 minutes, and the calcination environment is an oxidizing atmosphere (air). The ignition loss rate after calcination accounts for 80.6% of the total ignition loss of the ore. The pre-calcined product is placed in water for digestion. The liquid-solid ratio is 3:1 L / kg, the digestion water temperature is 90 °C, and the digestion time is 10 minutes. Finally, gravity separation is performed on the digested product to obtain the heavy product as the pre-enrichment concentrate and the light product as the pre-enrichment tailing.

[0072] Step (2): Low-intensity magnetic separation operation. The pre-enrichment concentrate (with a particle size of less than 2 mm) is ground to a -0.074 mm content of 72 wt% and enters the low-intensity magnetic separation operation with a magnetic field intensity of 0.2 T to obtain a low-intensity magnetic separation concentrate and a low-intensity magnetic separation tailing.

[0073] Step (3): High-intensity magnetic separation operation. The obtained low-intensity magnetic concentrate enters the high-intensity magnetic separation operation with a magnetic field intensity of 0.6 T to obtain high-intensity magnetic concentrate and high-intensity magnetic tailings. The finally obtained high-intensity magnetic concentrate enters the next operation.

[0074] Step (4): Leaching operation. The obtained high-intensity magnetic concentrate is added with 2 mol / L hydrochloric acid according to the ratio of solid mass to liquid volume of 3:1, stirred for 1 h, and filtered to obtain high-grade niobium concentrate.

[0075] Step (5): Desilication reverse flotation operation. The leached niobium concentrate is placed in a flotation cell with a pulp concentration of 27%. Sodium hydroxide and hydrochloric acid are used to adjust the pH to about 10.5. 500 g / t of inhibitor starch is added, and after stirring for 3 minutes, 1000 g / t of combined collector etheramine / oiloyl alanine (the mass ratio of etheramine to oiloyl alanine in the collector is 1:3) is added for roughing operation to obtain roughing concentrate and roughing tailings; 250 g / t of inhibitor starch is added to the roughing tailings, and after stirring for 3 minutes, 500 g / t of combined collector etheramine / oiloyl alanine (the mass ratio of etheramine to oiloyl alanine in the collector is 1:3) is added for scavenging operation to obtain scavenging concentrate and scavenging tailings; the roughing concentrate is stirred for 3 minutes and then subjected to cleaning operation to obtain cleaning concentrate and cleaning tailings.

[0076] Step (6): Niobium flotation operation. The above-mentioned scavenging concentrate and cleaning concentrate are combined and enter the positive flotation. Sodium hydroxide and hydrochloric acid are used to adjust the pH to about 7.0. 500 g / t of regulator sodium fluorosilicate is added, and after stirring for 3 minutes, 1000 g / t of activator lead nitrate is added. After stirring for 3 minutes, 1000 g / t of collector sodium oleate is added for roughing operation to obtain roughing concentrate and roughing tailings; 250 g / t of regulator sodium fluorosilicate is added to the roughing tailings, and after stirring for 3 minutes, 1000 g / t of activator lead nitrate is added. After stirring for 3 minutes, 1000 g / t of collector sodium oleate is added for scavenging operation to obtain scavenging concentrate and scavenging tailings; the roughing concentrate is stirred for 3 minutes and then subjected to first cleaning operation to obtain first cleaning concentrate and first cleaning middlings. The scavenging concentrate and the first cleaning middlings are combined and returned to the roughing operation; the first cleaning concentrate is stirred for 3 minutes and then subjected to second cleaning operation to obtain second cleaning concentrate and second cleaning middlings. The second cleaning middlings are returned to the first cleaning operation. Similarly, the third cleaning middlings are returned to the second cleaning operation, and the fourth cleaning middlings are returned to the third cleaning operation. The obtained fourth cleaning concentrate is the final pyrochlore concentrate.

[0077] The tailing rejection rate in the pre-concentration stage is 60.4%, the grade of Nb2O5 is 1.64%, and the recovery rate is 92.3%. In the combined stage of reverse and forward flotation, the grade of Nb2O5 is 52.63% and the recovery rate is 76.3%. When treating this pyrochlore ore by traditional flotation-based beneficiation methods, the niobium recovery rate is only about 50%. The beneficiation method based on pre-concentration - flotation can increase the Nb2O5 recovery rate by about 26.3% while maintaining the Nb2O5 grade of the traditional flotation-based beneficiation methods.

[0078] Example 4

[0079] Step (1): The content of Nb2O5 in a certain pyrochlore ore is 1.2%. The main associated minerals are calcite, dolomite, feldspar, etc. The content of CaO is 19.4% and the total loss on ignition is 17.3%. First, the ore is crushed to less than 30 mm, and then pre-calcined in a calciner. The pre-calcination temperature is 900 °C, lasting for 60 minutes, and the calcination environment is an oxidizing atmosphere (air). The loss on ignition after calcination accounts for 82.7% of the total loss on ignition of the ore. The pre-calcined product is digested in water with a liquid-solid ratio of 3:1 L / kg, a digestion water temperature of 90 °C, and a digestion time of 10 minutes. Finally, the digested product is subjected to gravity separation to obtain the pre-concentration concentrate as the heavy product and the pre-concentration tailings as the light product.

[0080] Step (2): Low-intensity magnetic separation operation. The pre-concentration concentrate (with a particle size of less than 2 mm) is ground to a proportion of -0.074 mm accounting for 72%, and enters the low-intensity magnetic separation operation with a magnetic field intensity of 0.2 T to obtain the low-intensity magnetic concentrate and the low-intensity magnetic tailings.

[0081] Step (3): High-intensity magnetic separation operation. The obtained low-intensity magnetic concentrate enters the high-intensity magnetic separation operation with a magnetic field intensity of 0.6 T to obtain the high-intensity magnetic concentrate and the high-intensity magnetic tailings. The finally obtained high-intensity magnetic concentrate enters the next operation.

[0082] Step (4): Leaching operation. 2 mol / L hydrochloric acid is added to the obtained high-intensity magnetic concentrate according to a solid mass to liquid volume ratio of 3:1, and stirred for 1 h, and then filtered to obtain a high-grade niobium concentrate.

[0083] Step (5): Floating niobium operation. The leached niobium concentrate is placed in a flotation cell with a pulp concentration of 27%. Sodium hydroxide and hydrochloric acid are used to adjust the pH to about 10.5. 500 g / t of sodium fluorosilicate as a regulator is added. After stirring for 3 minutes, 1000 g / t of a combined collector, sodium oleate / salicylhydroxamic acid (the mass ratio of sodium oleate to salicylhydroxamic acid in the combined collector is 1:1), is added for a roughing operation to obtain a rougher concentrate and a rougher tailing; 250 g / t of sodium fluorosilicate as a regulator is added to the rougher tailing. After stirring for 3 minutes, 500 g / t of the combined collector, sodium oleate / salicylhydroxamic acid (the mass ratio of sodium oleate to salicylhydroxamic acid in the combined collector is 1:1), is added for a scavenging operation to obtain a scavenger concentrate and a scavenger tailing; after stirring the rougher concentrate for 3 minutes, a cleaning operation is carried out to obtain a first cleaner concentrate and a first cleaning middling; the scavenger concentrate and the first cleaning middling are combined and returned to the roughing operation; after stirring the first cleaner concentrate for 3 minutes, a second cleaning operation is carried out to obtain a second cleaner concentrate and a second cleaning middling, and the second cleaning middling is returned to the first cleaning operation. Similarly, the third cleaning middling is returned to the second cleaning operation, and the fourth cleaning middling is returned to the third cleaning operation. The obtained fourth cleaner concentrate is the fourth cleaner concentrate.

[0084] Step (6): Barium removal reverse flotation operation. The above-mentioned fourth cleaner concentrate is adjusted to a pH of about 7.0 with sodium hydroxide and hydrochloric acid. 12 mg / L of an inhibitor, gellan gum / tannic acid (the mass ratio of gellan gum to tannic acid in the inhibitor is 1:1), is added for a reverse flotation operation to obtain the final pyrochlore concentrate and a reverse flotation tailing.

[0085] The tailing rejection rate in the pre-concentration stage is 56.6%, the grade of Nb2O5 is 2.65%, and the recovery rate is 91.6%; in the combined stage of forward and reverse flotation, the grade of Nb2O5 is 55.38% and the recovery rate is 83.47%. When treating this pyrochlore ore by traditional flotation-based beneficiation methods, the niobium recovery rate is only about 50%. The beneficiation method based on pre-concentration - flotation can increase the Nb2O5 recovery rate by about 33.0% while maintaining the Nb2O5 grade of the traditional flotation-based beneficiation methods.

[0086] Example 5

[0087] Step (1): The Nb2O5 content of a pyrochlore ore is 1.2%. The main associated minerals are calcite, dolomite, feldspar, etc. The CaO content is 19.4%, and the total loss on ignition is 17.3%. First, the ore is crushed to less than 30 mm, and then pre-calcined in a calciner at a pre-calcination temperature of 900 °C for 60 minutes. The calcination environment is an oxidizing atmosphere (air). The loss on ignition after calcination accounts for 82.7% of the total loss on ignition of the ore. The pre-calcined product is digested in water with a liquid-solid ratio of 3:1 L / kg, a digestion water temperature of 90 °C, and a digestion time of 10 minutes. Finally, the digested product is subjected to gravity separation to obtain the heavy product as the pre-enriched concentrate and the light product as the pre-enriched tailings.

[0088] Step (2): Low-intensity magnetic separation operation. The pre-enriched concentrate (with a particle size of less than 2 mm) is ground to a proportion of -0.074 mm accounting for 72%, and then enters the low-intensity magnetic separation operation with a magnetic field intensity of 0.2 T to obtain the low-intensity magnetic concentrate and the low-intensity magnetic tailings.

[0089] Step (3): High-intensity magnetic separation operation. The obtained low-intensity magnetic concentrate enters the high-intensity magnetic separation operation with a magnetic field intensity of 0.6 T to obtain the high-intensity magnetic concentrate and the high-intensity magnetic tailings. The finally obtained high-intensity magnetic concentrate enters the next operation.

[0090] Step (4): Leaching operation. The obtained high-intensity magnetic concentrate is added with 2 mol / L hydrochloric acid according to the ratio of solid mass to liquid volume of 3:1 and stirred for 1 h, and then filtered to obtain the high-grade niobium concentrate.

[0091] Step (5): Niobium flotation operation. The leached niobium concentrate is placed in a flotation cell with a pulp concentration of 27%. The pH is adjusted to about 5.0 using sodium hydroxide and hydrochloric acid. 500 g / t of the regulator fluosilicic acid is added, and after stirring for 3 minutes, 500 g / t of the combined collector tallow diamine / oleic acid is added for a roughing operation to obtain the roughing concentrate and the roughing tailings; 250 g / t of the regulator fluosilicic acid is added to the roughing tailings, and after stirring for 3 minutes, 500 g / t of the combined collector tallow diamine / oleic acid is added for a scavenging operation (the mass ratio of tallow diamine to oleic acid in the combined collector is 1:1) to obtain the scavenging concentrate and the scavenging tailings; the roughing concentrate is stirred for 3 minutes and then subjected to a cleaning operation to obtain the first cleaning concentrate and the first cleaning middlings. The scavenging concentrate and the first cleaning middlings are combined and returned to the roughing operation; the first cleaning concentrate is stirred for 3 minutes and then subjected to a second cleaning operation to obtain the second cleaning concentrate and the second cleaning middlings. The second cleaning middlings are returned to the first cleaning operation. Similarly, the third cleaning middlings are returned to the second cleaning operation, and the fourth cleaning middlings are returned to the third cleaning operation. The obtained fourth cleaning concentrate is the four-time cleaning concentrate.

[0092] Step (6): Barium removal reverse flotation operation. The pH of the above four times of concentrated concentrates was adjusted to about 7.0 using sodium hydroxide and hydrochloric acid. 12 mg / L of inhibitor gellan gum / tannic acid was added for reverse flotation operation (the mass ratio of gellan gum to tannic acid in the inhibitor was 1:1), obtaining the final pyrochlore concentrate and reverse flotation tailings.

[0093] The tailing rejection rate in the pre-concentration stage was 56.6%, the grade of Nb2O5 was 2.65%, and the recovery rate was 91.6%; in the combined stage of positive and reverse flotation, the grade of Nb2O5 was 52.36% and the recovery rate was 82.64%. When treating this pyrochlore ore by the traditional flotation-based beneficiation method, the niobium recovery rate was only about 50%. The beneficiation method based on pre-concentration - flotation could increase the Nb2O5 recovery rate by about 32.0% while maintaining the Nb2O5 grade of the traditional flotation-based beneficiation method.

[0094] Example 6

[0095] Step (1): The content of Nb2O5 in a certain pyrochlore ore was 1.2%. The main associated minerals were calcite, dolomite, feldspar, etc. The content of CaO was 19.4%, and the total loss on ignition was 17.3%. First, the ore was crushed to less than 30 mm, and then pre-calcined in a calciner at a pre-calcination temperature of 900 °C for 60 minutes. The calcination environment was an oxidative atmosphere (air). The loss on ignition after calcination accounted for 82.7% of the total loss on ignition of the ore. The pre-calcined product was digested in water with a liquid-solid ratio of 3:1 L / kg, a digestion water temperature of 90 °C, and a digestion time of 10 minutes. Finally, gravity separation was performed on the digested product, and the obtained heavy product was the pre-concentration concentrate and the light product was the pre-concentration tailings.

[0096] Step (2): Low-intensity magnetic separation operation. The pre-concentration concentrate (with a particle size of less than 2 mm) was ground to a proportion of -0.074 mm accounting for 72%, and then entered the low-intensity magnetic separation operation with a magnetic field intensity of 0.2 T, obtaining the low-intensity magnetic separation concentrate and the low-intensity magnetic separation tailings.

[0097] Step (3): High-intensity magnetic separation operation. The obtained low-intensity magnetic separation concentrate entered the high-intensity magnetic separation operation with a magnetic field intensity of 0.6 T, obtaining the high-intensity magnetic separation concentrate and the high-intensity magnetic separation tailings. The finally obtained high-intensity magnetic separation concentrate entered the next operation.

[0098] Step (4): Leaching operation. 2 mol / L hydrochloric acid was added to the obtained high-intensity magnetic separation concentrate according to the ratio of solid mass to liquid volume of 3:1, and then stirred for 1 h. After filtration, a high-grade niobium concentrate was obtained.

[0099] Step (5): Niobium flotation operation. The leached niobium concentrate is placed in a flotation cell with a pulp concentration of 27%. Sodium hydroxide and hydrochloric acid are used to adjust the pH to about 10.5. 750 g / t of sodium fluorosilicate as a modifier is added. After stirring for 3 minutes, 750 g / t of sodium oleate as a collector is added for a roughing operation to obtain a rougher concentrate and a rougher tailing; 250 g / t of sodium fluorosilicate as a modifier is added to the rougher tailing. After stirring for 3 minutes, 250 g / t of sodium oleate as a collector is added for a scavenging operation to obtain a scavenger concentrate and a scavenger tailing; the rougher concentrate is stirred for 3 minutes and then a cleaning operation is carried out to obtain a first cleaner concentrate and a first cleaning middling. The scavenger concentrate and the first cleaning middling are combined and returned to the roughing operation; the first cleaner concentrate is stirred for 3 minutes and then a second cleaning operation is carried out to obtain a second cleaner concentrate and a second cleaning middling. The second cleaning middling is returned to the first cleaning operation. Similarly, the third cleaning middling is returned to the second cleaning operation, and the fourth cleaning middling is returned to the third cleaning operation. The obtained fourth cleaner concentrate is the fourth cleaner concentrate.

[0100] Step (6): Barium removal reverse flotation operation. The above-mentioned fourth cleaner concentrate is adjusted to a pH of about 7.0 with sodium hydroxide and hydrochloric acid. 12 mg / L of gellan gum / tannic acid as an inhibitor is added for a reverse flotation operation (the mass ratio of gellan gum to tannic acid in the inhibitor is 1:1) to obtain the final pyrochlore concentrate and a reverse flotation tailing.

[0101] The tailing rejection rate in the pre-concentration stage is 56.6%, the grade of Nb2O5 is 2.65%, and the recovery rate is 91.6%; in the combined forward and reverse flotation stage, the grade of Nb2O5 is 51.52% and the recovery rate is 75.66%. When using traditional flotation-based ore dressing methods to process this pyrochlore ore, the niobium recovery rate is only about 50%. The ore dressing method based on pre-concentration - flotation can increase the Nb2O5 recovery rate by about 25.0% while maintaining the Nb2O5 grade of the traditional flotation-based ore dressing method.

[0102] Example 7

[0103] If the expected niobium grade and recovery rate cannot be achieved during actual operation, simulation calculations are used to screen selective collectors; the implementation process is as follows:

[0104] Step (1): Construct the mineral surface model and optimize the reagent structure. Use Materials Studio software to establish the crystal structure models of pyrochlore, phlogopite, and barite after calcination and acid leaching. Select the crystal planes that are most easily exposed during the flotation process for each mineral. For pyrochlore, it is (1 1 1); for phlogopite, it is (0 0 1); for barite, it is (0 0 1). Remove the surface hydroxyl oxygen when constructing all crystal planes, set a vacuum layer of 40 Å, and fix the bottom 4 layers of atoms. Take sodium oleate as an example here, and use the CASTEP module to optimize the collector sodium oleate. To enhance selectivity, introduce a urea group (–NH–CO–NH–) structure into the sodium oleate molecular chain to endow it with the ability of directional complexation and hydrogen bonding. Arrange the optimized collector on the surfaces of the three minerals at an interval of 1.5 Å, and use the first-principles method to calculate the adsorption energy, set the cutoff energy to 400 eV, and the convergence criterion is 1×10 -5 eV / atom.

[0105] The results show that: on the surface of pyrochlore after calcination and acid leaching, the adsorption energy of sodium oleate containing urea group is -165 kJ / mol, much lower than that of single sodium oleate (-125 kJ / mol), showing a significant synergistic effect; on the surface of phlogopite, the adsorption energy is -89 kJ / mol; on the surface of barite, the adsorption energy is only -75 kJ / mol. Sodium oleate after introducing the urea group has the strongest adsorption ability on the surface of pyrochlore after calcination and acid leaching, and the difference in adsorption energy with phlogopite and barite exceeds 70 kJ / mol, showing good selectivity.

[0106] Step (2): The present invention proposes the following collector screening principles: the adsorption energy needs to be lower than -140 kJ / mol; non-specific adsorption on the surfaces of phlogopite and barite should be avoided to ensure the high-selectivity recovery of pyrochlore. The results of this simulation calculation provide a theoretical basis for the selection of reagents in the subsequent flotation process and the separation of pyrochlore and associated minerals.

[0107] Example 8

[0108] Step (1): The Nb2O5 content of a certain pyrochlore ore is 0.4%, and the main associated minerals are calcite, dolomite, apatite, etc. The CaO content is 25%, and the total ignition loss is 18.9%. First, crush the ore to less than 30 mm, and then perform pre-calcination in a calcination furnace. The pre-calcination temperature is 900 °C, lasting for 60 minutes, and the calcination environment is an oxidizing atmosphere (air). The ignition loss rate after calcination accounts for 80.6% of the total ignition loss of the ore. Place the pre-calcined product in water for digestion, with a liquid-solid ratio of 3:1 L / kg, a digestion water temperature of 90 °C, and a digestion time of 10 minutes. Finally, perform gravity separation on the digested product. The obtained heavy product is the pre-enriched concentrate, and the light product is the pre-enriched tailings.

[0109] Step (2): Low-intensity magnetic separation operation. The pre-concentrated concentrate (with a particle size of less than 2 mm) is ground to a proportion of -0.074 mm accounting for 72%, and enters the low-intensity magnetic separation operation with a magnetic field intensity of 0.2 T to obtain low-intensity magnetic concentrate and low-intensity magnetic tailings.

[0110] Step (3): High-intensity magnetic separation operation. The obtained low-intensity magnetic concentrate enters the high-intensity magnetic separation operation with a magnetic field intensity of 0.6 T to obtain high-intensity magnetic concentrate and high-intensity magnetic tailings. The finally obtained high-intensity magnetic concentrate enters the next operation.

[0111] Step (4): Leaching operation. The obtained high-intensity magnetic concentrate is added with 2 mol / L hydrochloric acid according to the ratio of solid mass to liquid volume of 3:1 and stirred for 1 h, and then filtered to obtain high-grade niobium concentrate.

[0112] Step (5): Desilication reverse flotation operation. The leached niobium concentrate is placed in a flotation cell with a pulp concentration of 27%, and the pH is adjusted to about 10.5 using sodium hydroxide and hydrochloric acid. 500 g / t of inhibitor starch is added, and after stirring for 3 minutes, 1000 g / t of combined collector etheramine / oiloyl alanine (the mass ratio of etheramine to oiloyl alanine in the collector is 1:3) is added for roughing operation to obtain roughing concentrate and roughing tailings; 250 g / t of inhibitor starch is added to the roughing tailings, and after stirring for 3 minutes, 500 g / t of combined collector etheramine / oiloyl alanine (the mass ratio of etheramine to oiloyl alanine in the collector is 1:3) is added for scavenging operation to obtain scavenging concentrate and scavenging tailings; the roughing concentrate is stirred for 3 minutes and then subjected to cleaning operation to obtain cleaning concentrate and cleaning tailings.

[0113] Step (6): Niobium flotation operation. The above-mentioned scavenging concentrate and cleaning concentrate are combined and enter the positive flotation. The pH is adjusted to about 7.0 using sodium hydroxide and hydrochloric acid. 500 g / t of regulator sodium fluorosilicate is added, and after stirring for 3 minutes, 1000 g / t of activator lead nitrate is added, and after stirring for 3 minutes, 1000 g / t of collector sodium oleate modified with urea group is added for one roughing operation to obtain roughing concentrate and roughing tailings; 250 g / t of regulator sodium fluorosilicate is added to the roughing tailings, and after stirring for 3 minutes, 1000 g / t of activator lead nitrate is added, and after stirring for 3 minutes, 1000 g / t of collector sodium oleate modified with urea group is added for one scavenging operation to obtain scavenging concentrate and scavenging tailings; the roughing concentrate is stirred for 3 minutes and then subjected to one cleaning operation to obtain one cleaning concentrate and one cleaning middlings, and the scavenging concentrate and one cleaning middlings are combined and returned to the roughing operation; the one cleaning concentrate is stirred for 3 minutes and then subjected to two cleaning operations to obtain two cleaning concentrate and two cleaning middlings, and the two cleaning middlings are returned to the one cleaning operation. Similarly, the three cleaning middlings are returned to the two cleaning operations, and the four cleaning middlings are returned to the three cleaning operations. The obtained four cleaning concentrate is the final pyrochlore concentrate.

[0114] The tailing rejection rate in the pre-concentration stage is 60.4%, the grade of Nb2O5 is 1.64%, and the recovery rate is 92.3%; in the combined stage of reverse and forward flotation, the grade of Nb2O5 is 54.58%, and the recovery rate is 86.32%. When using the traditional flotation-based ore dressing method to process this pyrochlore ore, the niobium recovery rate is only about 50%. The ore dressing method based on pre-concentration - flotation can increase the Nb2O5 recovery rate by about 36.3% while maintaining the Nb2O5 grade of the traditional flotation-based ore dressing method.

[0115] Example 9

[0116] Step (1): The content of Nb2O5 in a certain pyrochlore ore is 1.2%. The main associated minerals are calcite, dolomite, feldspar, etc. The content of CaO is 19.4%, and the total ignition loss is 17.3%. First, the ore is crushed to less than 30 mm, and then pre-calcined in a calcination furnace at a pre-calcination temperature of 900 °C for 60 minutes. The calcination environment is an oxidizing atmosphere (air). The ignition loss rate after calcination accounts for 82.7% of the total ignition loss of the ore. The pre-calcined product is placed in water for digestion, with a liquid-solid ratio of 3:1 L / kg, a digestion water temperature of 90 °C, and a digestion time of 10 minutes. Finally, gravity separation is performed on the digested product to obtain the heavy product as the pre-concentration concentrate and the light product as the pre-concentration tailings.

[0117] Step (2): Low-intensity magnetic separation operation. The pre-concentration concentrate (with a particle size of less than 2 mm) is ground to a proportion of -0.074 mm accounting for 72%, and then enters the low-intensity magnetic separation operation with a magnetic field intensity of 0.2 T to obtain the low-intensity magnetic concentrate and the low-intensity magnetic tailings.

[0118] Step (3): High-intensity magnetic separation operation. The obtained low-intensity magnetic concentrate enters the high-intensity magnetic separation operation with a magnetic field intensity of 0.6 T to obtain the high-intensity magnetic concentrate and the high-intensity magnetic tailings. The finally obtained high-intensity magnetic concentrate enters the next operation.

[0119] Step (4): Leaching operation. 2 mol / L hydrochloric acid is added to the obtained high-intensity magnetic concentrate according to the ratio of solid mass to liquid volume of 3:1, and then stirred for 1 h. After filtration, high-grade niobium concentrate is obtained.

[0120] Step (5): Floating niobium operation. The leached niobium concentrate is placed in a flotation cell with a pulp concentration of 27%. Sodium hydroxide and hydrochloric acid are used to adjust the pH to about 10.5. 750 g / t of sodium fluorosilicate as a regulator is added. After stirring for 3 minutes, 750 g / t of oleic acid modified with urea as a collector is added for a roughing operation to obtain a rougher concentrate and a rougher tailing; 250 g / t of sodium fluorosilicate as a regulator is added to the rougher tailing. After stirring for 3 minutes, 250 g / t of oleic acid modified with urea as a collector is added for a scavenging operation to obtain a scavenger concentrate and a scavenger tailing; after stirring the rougher concentrate for 3 minutes, a cleaning operation is carried out to obtain a first cleaner concentrate and a first cleaning middling. The scavenger concentrate and the first cleaning middling are combined and returned to the roughing operation; after stirring the first cleaner concentrate for 3 minutes, a second cleaning operation is carried out to obtain a second cleaner concentrate and a second cleaning middling. The second cleaning middling is returned to the first cleaning operation. Similarly, the third cleaning middling is returned to the second cleaning operation, and the fourth cleaning middling is returned to the third cleaning operation. The obtained fourth cleaner concentrate is the fourth cleaner concentrate.

[0121] Step (6): Barium removal reverse flotation operation. The above-mentioned fourth cleaner concentrate is adjusted to a pH of about 7.0 using sodium hydroxide and hydrochloric acid. 12 mg / L of gellan gum / tannic acid as an inhibitor is added for a reverse flotation operation (the mass ratio of gellan gum to tannic acid in the inhibitor is 1:1) to obtain the final pyrochlore concentrate and a reverse flotation tailing.

[0122] The tailing rejection rate in the pre-concentration stage is 56.6%, the grade of Nb2O5 is 2.65%, and the recovery rate is 91.6%; in the combined stage of forward and reverse flotation, the grade of Nb2O5 is 54.32% and the recovery rate is 82.36%. When treating this pyrochlore ore by the traditional flotation-based beneficiation method, the niobium recovery rate is only about 50%. The beneficiation method based on pre-concentration - flotation can increase the Nb2O5 recovery rate by about 32.3% while maintaining the Nb2O5 grade of the traditional flotation-based beneficiation method.

[0123] Comparative Example 1

[0124] This comparative example provides the beneficiation method of the existing pyrochlore in BV factory. The original ore is the same as that in Examples 1 and 2. The beneficiation steps only include the flotation process. The scale of the flotation machine is the same as that in Examples 1 and 2. The flotation is carried out according to the conventional pyrochlore beneficiation process, mainly including desliming, reverse flotation of silicate minerals, reverse flotation of carbonate minerals, magnetic separation for iron removal, then forward flotation of pyrochlore, and finally acid leaching for phosphorus removal and calcination to obtain the pyrochlore concentrate. The results show that in the finally obtained product, the grade of Nb2O5 is about 55% and the recovery rate is about 50%.

[0125] Comparative Example 2

[0126] The original ore and pretreatment process of this comparative example are the same as those of Example 1. The ore dressing steps only include positive flotation, and the scale of the flotation machine is the same as that of Example 1. The pretreated niobium concentrate is placed in a flotation cell with a pulp concentration of 27%. Sodium hydroxide and hydrochloric acid are used to adjust the pH to about 10.5. 750 g / t of sodium fluorosilicate as a regulator is added, and after stirring for 3 minutes, a combined collector of sodium oleate / salicylhydroxamic acid at 750 g / t with a mass ratio of 1:1 is added for a roughing operation to obtain a rougher concentrate and rougher tailings; 250 g / t of sodium fluorosilicate as a regulator is added to the rougher tailings, and after stirring for 3 minutes, a combined collector of sodium oleate / salicylhydroxamic acid at 250 g / t with a mass ratio of 1:1 is added for a scavenging operation to obtain a scavenger concentrate and scavenger tailings; the rougher concentrate is stirred for 3 minutes and then undergoes a cleaning operation to obtain a first cleaner concentrate and first cleaning middlings. The scavenger concentrate and the first cleaning middlings are combined and returned to the roughing operation; the first cleaner concentrate is stirred for 3 minutes and then undergoes a second cleaning operation to obtain a second cleaner concentrate and second cleaning middlings. The second cleaning middlings are returned to the first cleaning operation. Similarly, the third cleaning middlings are returned to the second cleaning operation, and the fourth cleaning middlings are returned to the third cleaning operation. The obtained fourth cleaner concentrate is the fourth cleaner concentrate. The grade of Nb2O5 in the finally obtained niobium concentrate product is 26%, and the recovery rate is 80%.

[0127] In summary, the present invention provides a method for dressing pyrochlore, which changes the prior art process of desliming first and then flotation. First, the raw ore is pre-enriched, that is, calcination-digestion-grinding-magnetic separation-acid leaching, greatly reducing niobium loss and reducing the treatment scale of the subsequent pyrochlore ore dressing plant. By using the combined process of positive and reverse flotation, the recovery rate of Nb2O5 is increased while maintaining the grade of Nb2O5 in the original plant. Compared with the conventional process, the recovery rate of Nb2O5 by the method of the present invention is increased by about 30%.

[0128] The above are only the preferred embodiments of the present invention and are not used to limit the present invention. For those skilled in the art, the present invention can have various modifications and changes. Any modification, equivalent replacement, improvement, etc. made within the spirit and principle of the present invention shall be included within the protection scope of the present invention.

Claims

1. A beneficiation method for pre-concentrating niobium ore and recovering pyrochlore from calcined digestion re-concentration ore, characterized in that, It includes the following steps: Step 1: Pretreatment Taking the calcined digestion re-concentrated ore as the treatment object, performing pretreatment on the treatment object to obtain the slag after acid leaching; the pretreatment includes grinding, magnetic separation, and acid leaching; the particle size of the original ore is less than 30 mm, and the particle size of the calcined digestion re-concentrated ore is less than 3 mm; Step 2: Flotation operation Combining the positive flotation and reverse flotation of the slag after acid leaching to finally obtain flotation concentrate and flotation tailings.

2. The beneficiation method for pre-enriching niobium ore and recovering pyrochlore from calcined digestion re-concentration ore according to claim 1, wherein: The calcined digestion re-concentrated ore refers to the re-concentrated concentrate obtained after the original ore is treated by calcination, digestion, and re-selection processes; during the calcination process, calcite in the ore undergoes a decomposition reaction to generate calcium oxide and release carbon dioxide; in the subsequent digestion process, calcium oxide is further converted into slaked lime; the slaked lime is separated and removed by re-selection.

3. The ore dressing method according to claim 1, characterized in that: The pretreatment process sequentially includes the steps of grinding, magnetic separation, and acid leaching; Before the grinding treatment, the particle size of the pyrochlore mineral is less than 3 mm, The grinding product is subjected to magnetic separation to obtain magnetic separation concentrate and magnetic separation tailings; then the magnetic separation tailings are subjected to acid leaching treatment.

4. The ore dressing method according to claim 3, wherein: The fineness of grinding is that the proportion of -0.074 mm is 40.0 - 85.0%; The magnetic separation intensity is 0.2 - 1.6 T, preferably 0.2 - 0.6 T; In the acid leaching process, hydrochloric acid or sulfuric acid is used for acid leaching, the concentration of the acid is 0.5 - 10 mol / L, the ratio of solid mass to liquid volume is 1:1 - 10:1, and the leaching time is 5 - 60 min.

5. The ore dressing method according to claim 4, wherein: After the calcined digestion re-concentrated ore undergoes pretreatment, the minerals include pyrochlore, barite, phlogopite, etc. The tailing rejection rate of the original ore is greater than 55%, the niobium recovery rate is greater than 90%, and the niobium grade in the slag after acid leaching is greater than 1.75%.

6. The ore dressing method according to claim 1, characterized in that: After pretreatment, niobium is exposed on the pyrochlore 5+ , calcium 2+ sites, generating a metal oxide-like surface; the surface of phlogopite exposes Al-OH and Mg-OH unsaturated bond sites, and the surface of barite exposes Ba 2+ sites.

7. The ore dressing method according to claim 6, wherein: For the combined process of positive flotation and reverse flotation, based on the exposure of active sites in pyrochlore, phlogopite, and barite after calcination and acid leaching, the types of reagents are further optimized and screened; in the reverse flotation process, the inhibitors used are selected from at least one of carboxymethyl cellulose, starch, gellan gum, dextrin, water glass, polyacrylamide, tannic acid, sodium lignosulfonate, sodium hexametaphosphate, or sodium fluorosilicate; the collectors used are selected from at least one of oleic acid, sodium oleate, oleic acid soap, alkylamine, oiloyl alanine, or oleic acid amide; In the reverse flotation, hydrochloric acid and sodium hydroxide are used to adjust the pH, and the pH is controlled at 2 - 11; In the rough selection process of reverse flotation to float phlogopite and suppress pyrochlore and barite, the dosage of the inhibitor is 250 g / t - 2000 g / t, the dosage of the collector is 250 g / t - 2000 g / t, and the dosage of the collector in the first scavenging is 50 - 500 g / t; In the positive flotation, the niobium floating regulator is selected from at least one of oxalic acid, acetic acid, hydrofluoric acid, fluorosilicic acid, sodium fluorosilicate, or sodium silicate; the activator is selected from at least one of lead nitrate, calcium chloride, or sodium chloride; the collector is selected from at least one of sodium oleate, benzohydroxamic acid, salicylhydroxamic acid, xanthate, black medicine, oleic acid, palmitic acid, alkylamine, or tallow diamine; In the positive flotation, hydrochloric acid and sodium hydroxide are used to adjust the pH, and the pH is controlled at 2 - 11; During the rough selection process of positive flotation of pyrochlore, the dosage of the regulator is 250 g / t to 2000 g / t, the dosage of the activator is 500 g / t to 2000 g / t, and the dosage of the collector in the first scavenging group is 250 g / t to 2000 g / t.

8. The ore dressing method according to claim 7, characterized in that: When the flotation reagent system cannot achieve the dual improvement of niobium grade and recovery rate, the optimization selection of the collector will be regulated based on the DFT simulation results of the collector with the surfaces of pyrochlore, phlogopite, and barite minerals; For molecular simulation, the first-principles method is used to optimize the crystal structures of three minerals, namely pyrochlore, phlogopite, and barite, construct a thermodynamically stable crystal plane model, and simulate the surface exposure characteristics under flotation conditions; the model is based on the mineral structure after calcination and acid leaching treatment, reflecting the reconstruction behavior of the active sites on the crystal surface: more Nb and Ca active sites are exposed on pyrochlore, Al and Mg sites appear on phlogopite due to dehydroxylation, and the surface exposure of Ba ions is enhanced on barite; On this basis, the selective adsorption ability of the collector to pyrochlore is improved by introducing functional groups into the collector molecule, and the functional group is selected from at least one of ureido, hydroxamic acid group, phosphonic acid group, and carboxyl group; The calculation of the adsorption energy and the configuration simulation of various functional groups on the surfaces of different minerals reveal the matching relationship between the crystal surface structure and the interaction between the reagent molecules.

9. A beneficiation method for pre-enriching niobium ore and recovering pyrochlore from calcined digestion re-concentration ore, characterized in that: The adsorption energy calculation is carried out through the adsorption energy formula: ∆E = E complex - E mineral + E agent to calculate the adsorption energy between the mineral surface and the reagent molecule, or to judge whether there is an adsorption behavior between the mineral surface and the reagent molecule; where ∆E represents the adsorption energy between the mineral surface and the reagent molecule, and E complex represents the total energy of the adsorption model of the mineral surface and the reagent molecule, E mineral represents the energy of the mineral surface model, and E agent represents the energy of the reagent molecule model; the units of ∆E, E complex , E mineral and E agent are eV; The selection criterion for the collector is to select a reagent with an adsorption energy less than or equal to -140 kJ / mol with the mineral surface as the available flotation collector.

10. A beneficiation method for pre-enriching niobium ore and recovering pyrochlore from calcined digestion re-concentration ore, characterized in that: According to the simulation calculation results, the modified collector is selected to conduct flotation experiments on the actual ore samples; the flotation experiments include positive flotation and reverse flotation.

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