A flotation process combining pre-disposal of coarse particles with fractional recovery of fine particles

Abstract

The present invention relates to a flotation process combining coarse particle pre-discarding with fine particle fractionation and recovery. The process comprises the following steps: grinding, classification, coarse particle pre-discarding, and fine particle fractionation and recovery. The raw ore is slurried and ground to a pulp fineness of 50-75% of -200 mesh; the grinding pulp is graded according to particle sizes of 0.15 mm and 1-1.2 mm; the coarse particle size pulp is floated in a microbubble and water fluidization composite interference bed to promptly remove gangue minerals from the process; during fine particle fractionation and recovery, roughing is performed to obtain coarse concentrate and rougher tailings with different floatable properties; the coarse concentrate is selected, and the rougher tailings are scavenged and / or roughed and / or selected by flotation columns to obtain target concentrate and final tailings. The process combines coarse particle pre-discarding for optimizing flotation effect with targeted fine particle fractionation and recovery. The process can significantly reduce feed particle size requirements and flotation process costs, reduce grinding energy consumption by more than 30%, and achieve efficient flotation separation of target minerals and maximize overall recovery.

Description

Technical Field

[0001] The present invention belongs to the technical field of mine flotation, and in particular relates to a flotation process combining pre-discarding of coarse particles with quality-separated recovery of fine particles. Background Art

[0002] my country has abundant mineral resources. However, as high-quality sulfide and oxide ores are depleted, the mineral resources are becoming increasingly poor, fine, and mixed, and low-grade minerals are becoming the main focus of development and utilization. Flotation is the most commonly used method for recovering low-grade minerals. However, the existing flotation process of "grinding-classification-flotation" still has the following problems:

[0003] (1) The requirements for feed particle size are high and the cost is high: the feed particle size usually requires that the proportion of particles below 200 mesh is about 60% to 90%, and the upper limit of feed particle size is usually 0.1 to 0.15 mm. For particles with a particle size above 0.15 mm and relatively coarse particles, it is usually necessary to return to the ball mill for fine grinding to reach the feed particle size for flotation. Ores with a particle size of about +0.15 mm to 1 mm, especially brittle minerals, are easily over-grinded and muddied when regrinding, resulting in an unnecessary increase in the amount of fine mud. At the same time, the grade of coarse-grained ores is usually lower than that of fine-grained ores. Performing all grinding operations unnecessarily increases the energy consumption and cost of grinding, affecting the operation efficiency.

[0004] (2) The flotation effect of coarse-grained minerals with a particle size of more than 0.15 mm is poor: the dissociation degree of coarse-grained ore monomers is poor, and the exposure of lead-zinc sulfide minerals, tungsten minerals, cassiterite, etc. on the ore surface is relatively low, and the effect of the collector is poor, resulting in insufficient hydrophobicity on the ore particle surface and a low probability of adhesion to bubbles. In addition, the interference bed distribution, bubble dispersion and bubble size of traditional flotation equipment are poor, and the coarse particles adhere to the bubbles and then fall off, resulting in poor recovery effect; therefore, coarse-grained minerals often sink into the tailings during the flotation process, resulting in the loss of valuable minerals and the inability to achieve efficient recovery.

[0005] (3) Poor overall recovery rate: Due to the influence of surface physical and chemical properties, the proportion of floatable minerals in the coarse and fine particle sizes in the overall minerals is different, and the required flotation conditions are different. Traditional flotation technology often adopts a process of multiple selection and scavenging, sequential return of middlings, and flotation under the same reagent system and process flow. However, this process fails to carry out targeted flotation and can only sacrifice recovery rate to meet the concentrate grade requirements, resulting in a low flotation recovery rate and the inability to maximize the overall recovery rate. Summary of the Invention

[0006] The present invention aims to solve at least one of the above-mentioned technical problems to a certain extent. The present invention provides a flotation process that combines coarse particle pre-discarding with fine particle fractionation and recovery. The process flow of "grinding-classification-coarse particle pre-discarding-fine particle fractionation and recovery" is adopted. The combination of coarse particle pre-discarding for optimizing flotation effect and targeted fine particle fractionation and recovery can significantly reduce the particle size requirement of the feed, the energy consumption of grinding and the cost of the flotation process, thereby achieving efficient flotation separation of the target mineral and maximizing the overall recovery rate.

[0007] The technical solution adopted by the present invention to solve its technical problem is:

[0008] A flotation process that combines pre-discarding of coarse particles with recovery of fine particles by quality separation, the process comprising:

[0009] Grinding: After the raw ore is slurried, it is ground to a pulp fineness of -200 mesh accounting for 50-75%;

[0010] Classification: The grinding pulp is classified according to the particle size of 0.15mm and 1-1.2mm to obtain the fine-grained pulp of -0.15mm particle size and the coarse-grained pulp of +0.15mm-(1-1.2)mm particle size. The pulp of +1-1.2mm particle size is returned to the grinding process for re-grinding.

[0011] Coarse particle pre-disposal: After adding reagents to the coarse particle pulp and adjusting the pulp, it is floated in the micro-bubble and water fluidization composite interference bed of the fluidized coarse particle flotation machine to obtain concentrate A1 and final tailings. Concentrate A1 is returned to the grinding process for regrinding;

[0012] Fine particle fractionation and recovery: After the fine particle size slurry is slurried with reagents, it enters the flotation column for roughing to obtain coarse concentrate B1 and roughing tailings C1; after the coarse concentrate B1 is slurried, it undergoes at least one cleaning to obtain concentrate A2; after the roughing tailings C1 is slurried, it undergoes at least one scavenging and / or flotation column roughing and / or cleaning to obtain concentrate A3 and / or final tailings.

[0013] The above-mentioned flotation process further controls the pulp concentration to be 30-50% in the pulp adjustment of the grinding process, the coarse-grained pulp adjustment of the coarse-grained pre-discarding process, and the coarse-grained pulp adjustment of the fine-grained fractionation and recovery process, so as to further control the mineral recovery rate, concentrate quality, pulp adjustment dosage and grinding energy consumption.

[0014] Furthermore, the fine-grained pulp and the +1-1.2 mm particle size pulp after classification are respectively concentrated to further control the fine-grained fractionation recovery slurry concentration and the grinding pulp concentration.

[0015] Furthermore, after the concentrate A1 is concentrated and de-doped, the underflow is returned to the grinding process as return sand to further control the pulp concentration.

[0016] In the above flotation process, the raw ore includes sulfide ore and oxide ore, and the particle size of the coarse particle size pulp is determined according to the positive correlation between the floatability of the raw ore, the jet flow rate of the fluidized coarse particle flotation machine, and the pressure and the upper limit of the classification particle size:

[0017] Sulfide ore has better floatability, which can increase the upper limit of the classification particle size; oxide ore has poorer floatability, so the upper limit of the classification particle size should be lowered;

[0018] The larger the number of water-air jet guns in the fluidized coarse particle flotation machine, the larger the jet volume. The higher the water pressure and air pressure, the higher the upper limit of the classification particle size. Otherwise, the upper limit of the classification particle size needs to be lowered. This further improves the flotation effect and overall recovery rate of coarse particle minerals.

[0019] Furthermore, in the coarse particle size slurry of +0.15mm-(1~1.2)mm particle size, the -0.8mm particle size accounts for 60~90%, which is used to take into account both the flotation effect and the grinding energy consumption. If the particle size is smaller than this range, the proportion of coarse particles is too high and the flotation effect cannot be good. If the particle size is higher than this range, the ball mill will grind too finely, resulting in an unnecessary increase in grinding energy consumption.

[0020] The above flotation process further comprises: a fluidized coarse particle flotation machine comprising a flotation column, a connected water-gas mixer and a plurality of water-gas jet guns, the water-gas jet guns being provided with a jet zone extending into the interior of the flotation column for vertically spraying a mixed fluid of microbubbles and water-gas upward; the size of the microbubbles in the mixed fluid of microbubbles and water-gas is 50-200 μm; the dispersion of the microbubbles in the mixed fluid of water-gas is more than 80%, so that the degree of bubble dispersion and the size of the bubbles can achieve the purpose of microbubble flotation; the angle α between the jet zone and the radial direction of the circumference of the flotation column is 0-45°, which has a better flow field distribution; if the angle exceeds 45°, the flow field in the central area of the column will be weak, affecting the flotation effect;

[0021] The coarse-grained slurry is sprayed in the flotation column, and the microbubbles and water fluidization composite interference bed formed by the jet zone and the stable flow field are used for flotation, which can optimize the flotation effect, obtain the concentrate A1 at the top of the flotation column, and obtain the final tailings at the bottom of the flotation column.

[0022] Furthermore, the water-air jet gun can adjust the angle between the jet zone and the radial direction of the flotation column, thereby simplifying the optimization exploration of flotation conditions by adjusting the flotation of coarse-particle ores of different types and properties.

[0023] Furthermore, a damping structure is provided inside the flotation column above the jet zone, which is used to interfere with the bed layer and form a suitable flow field and turbulence by adjusting the number, distribution form and position of the damping structure: increasing the number and dense distribution of damping rings can reduce the turbulence of the flow field in the column, reduce the turbulent resistance in the process of coarse particles adhering to bubbles and floating, and promote the floating of coarse particles; avoid the position of the damping structure being too low, which will interfere with the formation of the bed layer too early and be detrimental to the stability of the bed layer; avoid the position being too high, which will have too weak an impact on the turbulence.

[0024] In the above flotation process, further, the reagents added for slurry adjustment include one or more of a pH regulator, an activator, a collector, a frother, and an inhibitor, so as to optimize the flotation effect through a targeted reagent system.

[0025] Furthermore, the pH regulator includes sodium carbonate, which is used to further optimize the floatability of the mineral at a suitable pH.

[0026] Furthermore, the activator includes copper sulfate, which is used to make the mineral surface easy to adsorb the collector, thereby further optimizing the mineral flotation.

[0027] Furthermore, the collector includes one or more of butyl xanthate, complex, and ethyl xanthate; and is used to change the hydrophobicity of the mineral surface, making it easier for mineral particles to adhere to bubbles, thereby further optimizing mineral flotation.

[0028] Furthermore, the foaming agent includes pine oil, which is used to effectively enrich useful minerals at the interface between air and water, thereby further optimizing mineral flotation.

[0029] Furthermore, the depressant includes one or more of lime, zinc sulfate, sodium sulfite, and modified water glass, which are used to selectively adsorb on the surface of the mineral to depress the mineral and further optimize the mineral flotation.

[0030] Furthermore, the complex is formed by coordinating lead nitrate with an organic ligand, and the organic ligand includes one or more of octylhydroxamic acid, benzohydroxamic acid, salicylic hydroxamic acid, cupralin, sodium oleate, styrenephosphonic acid, and sodium dodecyl sulfate, which is used to further change the hydrophobicity of the mineral surface, so that the floating mineral particles adhere to the bubbles to optimize the flotation effect.

[0031] Furthermore, the lead nitrate and the organic ligand are in a molar ratio of (1-2):1. Within this range, the collector has the strongest collecting ability, and below or above this range, the collecting ability decreases.

[0032] Furthermore, the preparation method of the complex includes: preparing an organic ligand into an organic ligand solution with a concentration of 2-3%, preparing lead nitrate into a lead nitrate solution with a concentration of 5-10%, and slowly adding the lead nitrate solution to the organic ligand solution at a temperature of 35-65°C and stirring for 10-30 minutes to prepare the complex collector. The complex has the advantages of simple preparation and uniform composition, and is used to obtain a lead-organic ligand complex with optimal capture ability.

[0033] Furthermore, the modified water glass is a composite metal silicate formed by the reaction of metal ions and water glass, wherein the metal ions include Al 3+ 、Cu 2+ 、Fe 3+ , Pb 2+ One of them is used to further optimize the selectivity for minerals by introducing active metal ions into water glass.

[0034] Furthermore, the mass ratio of the metal ions to the water glass is (1-2):4. Within this range, the selectivity of the modified water glass is optimal. If the ratio is too low, the selectivity is poor, and if the ratio is too high, the inhibition ability will be significantly weakened.

[0035] Furthermore, the preparation method of the modified water glass includes: preparing metal ions into a metal ion solution with a concentration of 5 to 10%, using a water glass stock solution with a concentration of 100%, slowly adding the metal ion solution to the water glass stock solution and stirring for 10 to 20 minutes to prepare the modified water glass. It has the advantages of simple preparation and uniform composition, and is used to obtain modified water glass with the best selectivity.

[0036] The above flotation process is further characterized in that the raw ore is lead-zinc sulfide ore, and the grinding pulp fineness is: -200 mesh accounts for 50-70%;

[0037] The reagents added during the slurry preparation of the coarse particles include copper sulfate, xanthate, and pine oil, and the pH value of the slurry is 6 to 8;

[0038] The reagents added during the roughing and slurry adjustment of fine-grained slurry include lime, zinc sulfate, sodium sulfite, dithiothreitol, and pine oil. The pH value of the slurry is 6-8. It is not necessary to add acid or alkali to adjust the pH within the range of 6-8. It is the natural pH value of the ore in water.

[0039] The reagents added during the selection and slurrying of the coarse concentrate B1 include zinc sulfate, and the lead concentrate obtained after two rounds of selection is used as concentrate A2;

[0040] The rougher tailings C1 are slurried by adding reagents including ethyl thiocarb and pine oil, and after two scavenging processes, the reagents including lime, copper sulfate, ethyl xanthate and pine oil are added to slurry, and after roughing by flotation columns, the rough concentrate B2 and rougher tailings C2 are obtained;

[0041] The coarse concentrate B2 is slurried by adding reagents including lime and modified water glass, and is subjected to two rounds of cleaning to obtain zinc concentrate as concentrate A3; the rougher tailings C2 are slurried by adding reagents including ethyl xanthate and pine oil, and are subjected to two rounds of scavenging to obtain final tailings;

[0042] The closed-circuit processes for both selection and sweeping use sequential returns;

[0043] The above method can achieve the pre-disposal of coarse-grained gangue minerals, significantly reduce the grinding energy consumption and flotation process costs, and realize the efficient flotation separation of lead and zinc through the preferential flotation of fine-grained lead-zinc ore. The recovery rates of lead concentrate and zinc concentrate of sulfide lead-zinc ore are both increased by more than 1.5 percentage points, and the grinding energy consumption is reduced by more than 24%.

[0044] Furthermore, the raw ore is lead-zinc sulfide ore, and the grinding slurry is classified according to the particle sizes of 0.15 mm and 1.2 mm to obtain fine-particle slurry with a particle size of -0.15 mm and coarse-particle slurry with a particle size of +0.15 mm-1.2 mm, in which the particle size of -0.8 mm accounts for 60-90%, and the pulp with a particle size of +1.2 mm is returned to the grinding process for regrinding; this is used to further optimize the classification flotation and grinding effects of the lead-zinc sulfide ore.

[0045] Furthermore, the raw ore is a lead-zinc sulfide ore, and the slurry adjustment in the grinding process, the coarse-grained slurry adjustment in the coarse-grained pre-discarding process, and the coarse-grained slurry adjustment in the fine-grained fractionation and recovery process control the slurry concentration to be 30-40%; this is used to further optimize the grinding, coarse-grained pre-discarding, and fine-grained fractionation and recovery effects of the lead-zinc sulfide ore.

[0046] Furthermore, the raw ore is lead-zinc sulfide ore, and the amounts of reagents added relative to the raw ore during slurry preparation after the coarse particles are pre-discarded are: 50-200 g / t of copper sulfate, 100-300 g / t of xanthate, and 30-60 g / t of pine oil; this is used to further optimize the effect of pre-discarding the coarse particles of the lead-zinc sulfide ore.

[0047] Furthermore, the raw ore is lead-zinc sulfide ore, and the amounts of reagents added relative to the raw ore during roughing and slurrying of fine-grained ore pulp are: lime 500-2000 g / t, zinc sulfate 500-1000 g / t, sodium sulfite 200-500 g / t, ethyl dithiocarbamide 50-300 g / t, and pine oil 30-60 g / t; the order of adding the three inhibitors, lime, zinc sulfate, and sodium sulfite, is: first add lime and slurry for 3-5 minutes, and then add zinc sulfate and sodium sulfite simultaneously and slurry for 3-5 minutes; this is used to further optimize the roughing effect of fine-grained ore pulp of lead-zinc sulfide ore.

[0048] Furthermore, the raw ore is a lead-zinc sulfide ore, and the amount of reagent added relative to the raw ore during the selection and slurry adjustment of the coarse concentrate B1 is: zinc sulfate is 100-3000g / t, which is used to further optimize the lead flotation effect of fine particle fractionation recovery.

[0049] Furthermore, the raw ore is a sulfide lead-zinc ore, and the amounts of reagents added relative to the raw ore during scavenging and slurry adjustment of the roughing tailings C1 are: 20-50 g / t of ethyl disulfide and 10-45 g / t of pine oil; the amounts of reagents added relative to the raw ore during roughing and slurry adjustment of the flotation column are: 1000-2500 g / t of lime, 100-200 g / t of copper sulfate, 50-200 g / t of ethyl xanthate, and 20-50 g / t of pine oil; this is used to further optimize the zinc scavenging and roughing effects of fine particle fractionation recovery.

[0050] Furthermore, the raw ore is lead-zinc sulfide ore, and the amounts of reagents added relative to the raw ore during the first beneficiation and slurrying of the coarse concentrate B2 are: 200-800 g / t of lime, 50-200 g / t of water glass, and 20-60 g / t of metal ions; the amounts of reagents added relative to the raw ore during the second beneficiation and slurrying of the coarse concentrate B2 are: 25-100 g / t of water glass, and 10-30 g / t of metal ions; after modified water glass is prepared from water glass and metal ions, the modified water glass is added for slurrying; this is used to further optimize the zinc beneficiation effect of fine particle fractionation and recovery.

[0051] Furthermore, the raw ore is a lead-zinc sulfide ore, and the amount of reagent added relative to the raw ore during the C2 scavenging and slurry adjustment of the roughing tailings is: ethyl xanthate is 10-40g / t, and pine oil is 5-20g / t, which is used to further optimize the zinc flotation effect of fine particle fractionation recovery.

[0052] The above flotation process is further characterized in that the raw ore is tungsten ore, and the grinding pulp fineness is: -200 mesh accounts for 60-75%;

[0053] The reagents added during the coarse-grained pre-discarding and fine-grained slurry roughing include sodium carbonate, complex, and pine oil. The pH value of the slurry is 9-10. The floatability of tungsten ore is optimal within the alkaline range of 9-10.

[0054] The reagents added during the selection and slurrying of the coarse concentrate B1 include modified water glass, and the concentrate A2 is obtained after two rounds of selection;

[0055] The roughing tailings C1 are added with reagents including complexes to prepare the slurry, and the final tailings are obtained after three scavenging processes;

[0056] The closed-circuit processes for both selection and sweeping use sequential returns;

[0057] The above method realizes the pre-disposal of coarse-grained gangue minerals, significantly reduces the energy consumption of grinding and the cost of the flotation process, and improves the efficiency of the flotation operation. It has a good sorting effect on scheelite, wolframite, and mixed scheelite and wolframite ores, and can increase the tungsten recovery rate of tungsten concentrate by more than 3 percentage points and reduce the energy consumption of grinding by more than 20%.

[0058] Furthermore, the raw ore is tungsten ore, and the slurry adjustment in the grinding process and the coarse-grained slurry adjustment in the coarse-grained pre-discarding process control the slurry concentration to be 30-40%, and the slurry adjustment in the fine-grained fractionation and recovery process controls the slurry concentration to be 40-50%, so as to further optimize the grinding, coarse-grained pre-discarding and fine-grained fractionation and recovery effects of tungsten ore.

[0059] Furthermore, the raw ore is tungsten ore, and the amounts of reagents added relative to the raw ore during slurry preparation of coarse particles in advance are: 200-500 g / t of lead nitrate, 200-500 g / t of organic ligand, 20-50 g / t of pine oil, and 200-600 g / t of sodium carbonate; a lead-organic ligand complex is prepared from lead nitrate and the organic ligand, and the order of addition is sodium carbonate, lead-organic ligand complex, and pine oil; and the method is used to further optimize the roughing effect of the coarse particles in advance of tungsten ore.

[0060] Furthermore, the raw ore is tungsten ore, and the amounts of reagents added relative to the raw ore during rough selection and slurry adjustment of fine-grained slurry are: 300-600 g / t of lead nitrate, 200-600 g / t of organic ligand, 20-60 g / t of pine oil, and 300-800 g / t of sodium carbonate; a lead-organic ligand complex is prepared from lead nitrate and the organic ligand, and the order of addition is sodium carbonate, lead-organic ligand complex, and pine oil; this is used to further optimize the rough selection effect of tungsten for fine-grained fractionation and recovery.

[0061] Furthermore, the raw ore is tungsten ore, and the amount of reagent added relative to the raw ore during the first selection and slurrying of the coarse concentrate B1 is: 100-300 g / t of water glass, and 30-200 g / t of metal ions; the amount of reagent added relative to the raw ore during the second selection and slurrying of the coarse concentrate B1 is: 50-200 g / t of water glass, and 30-100 g / t of metal ions; this is used to further optimize the tungsten selection effect of fine-grained fractionation and recovery.

[0062] Furthermore, the raw ore is tungsten ore, and the amount of reagent added relative to the raw ore during the scavenging and slurry adjustment of the rougher tailings C1 is: 50-200 g / t of lead nitrate, and 30-100 g / t of organic ligand; a lead-organic ligand complex is prepared from lead nitrate and the organic ligand, and then added to the slurry; this is used to further optimize the tungsten scavenging effect of fine-particle fractionation recovery.

[0063] The flotation process is further characterized in that the raw ore is cassiterite, and the grinding pulp fineness is: -200 mesh accounts for 60-75%;

[0064] The reagents added during the roughing and slurry preparation of coarse-grained pre-discarding and fine-grained slurry include sodium carbonate, complex, and pine oil. The pH value of the slurry is 9-11. The floatability of cassiterite is best in the alkaline range of 9-11.

[0065] The reagents added during the selection and slurrying of the coarse concentrate B1 include modified water glass, and after three rounds of selection, a high-grade concentrate is obtained as concentrate A2;

[0066] The rougher tailings C1 are slurried with reagents including complexes, and after rough separation by flotation columns, a rough concentrate B2 and rougher tailings C2 are obtained; the rougher concentrate B2 is slurried with reagents including modified water glass, and after two rounds of separation, a medium-grade concentrate is obtained as concentrate A3; the rougher tailings C2 are scavenged twice to obtain the final tailings;

[0067] The closed-circuit processes for both selection and sweeping use sequential returns;

[0068] The above method realizes the pre-disposal of coarse-grained gangue minerals, significantly reduces the grinding energy consumption and flotation process costs, and realizes differentiated flotation of cassiterite with different floatability through the fractional recovery of fine-grained cassiterite, which can increase the overall cassiterite recovery rate by more than 5 percentage points and reduce the grinding energy consumption by more than 20%.

[0069] Furthermore, the raw ore is cassiterite, and the grinding slurry is classified according to the particle sizes of 0.15 mm and 1 mm to obtain fine-particle slurry with a particle size of -0.15 mm and coarse-particle slurry with a particle size of +0.15 mm-1 mm, in which the particle size of -0.8 mm accounts for 70 to 90%, and the pulp with a particle size of +1 mm is returned to the grinding process for re-grinding; this is used to further optimize the classification flotation and grinding effects of cassiterite.

[0070] Furthermore, the raw ore is cassiterite, and the slurry adjustment in the grinding process, the coarse-grained slurry adjustment in the coarse-grained pre-discarding process, and the coarse-grained slurry adjustment in the fine-grained fractionation and recovery process control the slurry concentration to be 30-40%, so as to further optimize the grinding, coarse-grained pre-discarding and fine-grained fractionation and recovery effects of cassiterite.

[0071] Furthermore, the raw ore is cassiterite, and the amounts of reagents added relative to the raw ore during slurry preparation after the coarse particles are discarded in advance are: 300-600 g / t of lead nitrate, 200-500 g / t of octylhydroxamic acid, 20-60 g / t of pine oil, and 300-800 g / t of sodium carbonate; this is used to further optimize the roughing effect of the pre-discarding of the coarse cassiterite particles.

[0072] Furthermore, the raw ore is cassiterite, and the amounts of reagents added relative to the raw ore during rough selection and slurry adjustment of fine-grained slurry are: 300-600 g / t of lead nitrate, 200-500 g / t of octylhydroxamic acid, 20-60 g / t of pine oil, and 400-800 g / t of sodium carbonate. A lead-octylhydroxamic acid complex is prepared from lead nitrate and octylhydroxamic acid, and the order of addition is sodium carbonate, lead-octylhydroxamic acid complex, and pine oil; this is used to further optimize the rough selection effect of cassiterite for fine-grained fractionation and recovery.

[0073] Furthermore, the original ore is cassiterite, and the coarse concentrate B1 is easy-to-float cassiterite. The amount of reagent added relative to the original ore during the first selection and slurrying is: 100-300 g / t of water glass, and 20-200 g / t of metal ions; the amount of reagent added relative to the original ore during the second selection and slurrying of the coarse concentrate B1 is: 50-200 g / t of water glass, and 20-100 g / t of metal ions; the amount of reagent added relative to the original ore during the third selection and slurrying of the coarse concentrate B1 is: 0-100 g / t of water glass, and 0-50 g / t of metal ions; modified water glass is prepared from water glass and metal ions and then added to the slurry; this is used to further optimize the selection effect of easy-to-float cassiterite for fine-grained fractionation and recovery.

[0074] Furthermore, the raw ore is cassiterite, and the roughing tailings C1 are difficult-to-float cassiterite. The amounts of reagents added relative to the raw ore during roughing slurry adjustment are: 100-300 g / t of lead nitrate, and 50-200 g / t of octylhydroxamic acid. A lead-octylhydroxamic acid complex is prepared from lead nitrate and octylhydroxamic acid and then added to the slurry; this is used to further optimize the roughing effect of tin for fine-grained fractionation and recovery.

[0075] Furthermore, the raw ore is cassiterite, and the amount of reagent added relative to the raw ore during the first selection and slurrying of the coarse concentrate B2 is: 50-200 g / t of water glass, and 20-100 g / t of metal ions; the amount of reagent added relative to the raw ore during the second selection and slurrying of the coarse concentrate B2 is: 20-100 g / t of water glass, and 10-50 g / t of metal ions; this is used to further optimize the tin selection effect of fine particle fractionation recovery.

[0076] Compared with the prior art, the present invention has the following beneficial effects:

[0077] (1) The traditional "grinding-classification-flotation" process is optimized to the process of "grinding-classification-coarse particle pre-discarding-fine particle classification and recovery". By pre-discarding the coarse particles of the coarse particle size of +0.15mm-(1~1.2)mm, the gangue minerals are removed from the process in time, which significantly reduces the load of the ball mill and the processing capacity of the flotation process, avoids over-grinding and mudification of brittle ores, and can solve the problem of high particle size requirements and high costs of mineral flotation feed, improve the efficiency of flotation operations, and reduce the energy consumption of grinding by more than 30%.

[0078] (2) The coarse particles are discarded in advance by vertically spraying a fully mixed microbubble and water-gas mixed fluid, and the dispersion of the microbubble in the water-gas mixed fluid reaches more than 80%, forming a stable microbubble and water fluidized composite interference bed, which strengthens the flotation effect of the water-gas fluid on the coarse-grained ore pulp, and can solve the problem of poor flotation effect of coarse-grained minerals with a particle size of more than 0.15mm. It is combined with the reagent system to improve the hydrophobicity of the ore particle surface and further optimize the flotation effect.

[0079] (3) The coarse particle pre-discarding can be flexibly adjusted by the number of water-air jet guns, the angle between the jet area and the radial circumference of the flotation column, and the free combination of the damping structure to adjust the water flow field and turbulence. It can flexibly adapt to different ore properties, significantly shorten the exploration and development of optimized flotation conditions, and further optimize the coarse particle pre-discarding indicators and the wear and consumption reduction effects.

[0080] (4) Fine particle fractionation and recovery: Through roughing, coarse concentrate B2 and rougher tailings C2 with different floatable properties can be obtained. The coarse concentrate B2 can be finely selected, and the rougher tailings C2 can be scavenged and / or finely selected by flotation columns to obtain the target concentrate and final tailings. Targeted separation is carried out to take into account the flotation recovery rate and concentrate grade requirements. This can achieve efficient flotation separation of target minerals and solve the problem of poor overall recovery rate. Compared with the existing flotation process, the recovery rate of sulfide lead-zinc ore can be increased by more than 3 percentage points, the tungsten recovery rate of tungsten concentrate by more than 3 percentage points, and the overall recovery rate of cassiterite by more than 5 percentage points. The grade of the target mineral in the tailings is extremely low. The differentiated flotation can be further optimized with the appropriate reagent system to maximize the overall recovery rate. BRIEF DESCRIPTION OF THE DRAWINGS

[0081] The above and / or additional aspects and advantages of the present invention will become apparent and readily understood from the following description of the embodiments with reference to the accompanying drawings, in which:

[0082] Figure 1 It is a schematic diagram of the process flow of the present invention;

[0083] Figure 2 Schematic diagram of the process flow of Example 1 and Example 2 of the present invention;

[0084] Figure 3 Schematic diagram of the process flow of Example 3 and Example 4 of the present invention;

[0085] Figure 4 Schematic diagram of the process flow of Examples 5 and 6 of the present invention;

[0086] Figure 5 1 is a schematic structural diagram of a fluidized coarse particle flotation machine according to an embodiment of the present invention;

[0087] Figure 62. It is a schematic structural diagram of a water-air mixer of a fluidized coarse particle flotation machine according to an embodiment of the present invention;

[0088] Figure 7 2. It is a schematic structural diagram of a water-air jet gun of a fluidized coarse particle flotation machine according to an embodiment of the present invention;

[0089] Figure 8 Schematic diagram of the combined use of the water-air jet gun of the fluidized coarse particle flotation machine according to an embodiment of the present invention.

[0090] Figure 9 Schematic diagram of the damping ring of the fluidized coarse particle flotation machine according to an embodiment of the present invention.

[0091] Markings in the figure: flotation column 1, foam overflow trough 2, concentrate discharge port 3, feeding pipe 4, feeding distributor 5, discharge port 6, water-gas distribution pipe 7, water-gas jet gun 8, vertical pipe 801, horizontal square pipe 802, fine grid 803, angle adjuster 804, rotary joint 805; water-gas mixer 9, shell 91, water-gas mixing pipe 911, contraction pipe 912, throat 913, diffuser 914, pressure water jet pipe 915, high-pressure gas jet pipe 916, water-gas shear disperser 92, tip 921, damping ring 10, pressure water pump 11, water supply tank 12, high-pressure gas tank 13. DETAILED DESCRIPTION

[0092] The following describes embodiments of the present invention in detail. Examples of the embodiments are illustrated in the accompanying drawings, wherein identical or similar reference numerals throughout denote identical or similar elements or elements having identical or similar functions. The embodiments described below with reference to the accompanying drawings are illustrative and intended to explain the present invention, and are not to be construed as limiting the present invention. In the description of the present invention, it should be understood that terms such as "center," "longitudinal," "transverse," "length," "width," "thickness," "upper," "lower," "axial," "radial," "vertical," "horizontal," "inner," and "outer" indicate positions or locations based on the positions or locations shown in the accompanying drawings. These terms are intended solely to facilitate the description of the present invention and simplify the description, and are not intended to indicate or imply that the devices or elements referred to must have a specific orientation, be constructed, or operate in a specific orientation. Therefore, they are not to be construed as limiting the present invention. In the description of the present invention, "plurality" and "several" mean two or more, unless otherwise specifically defined. In the present invention, unless otherwise specified or defined, terms such as "mounted," "connected," "connected," and "fixed" should be interpreted broadly, meaning, for example, fixedly connected, removably connected, or integrally connected; directly connected, indirectly connected through an intermediary, or internally connected between two elements. For those skilled in the art, the specific meanings of the above terms in the present invention can be understood according to specific circumstances.

[0093] Example 1:

[0094] like Figure 2 As shown in FIG. 1 , a preferred embodiment of the flotation process of the present invention combining pre-discarding of coarse particles with recovery of fine particles by quality separation is shown. The process includes:

[0095] S1. Grinding: After the raw ore is crushed, water is added to the slurry to a slurry concentration of 35%. The raw ore is a lead-zinc sulfide ore from Hunan Province with a lead grade of 1.73% and a zinc grade of 2.22%. The ore is then fed into a ball mill for a first grinding step to a slurry fineness of -200 mesh, accounting for 70%, to obtain the ball mill discharge slurry.

[0096] S2. Classification:

[0097] S201: The ball mill discharge slurry is classified into 0.15mm particle size using a cyclone to obtain a fine-particle slurry of -0.15mm particle size and a slurry of +0.15mm particle size. The -0.15mm particle size is concentrated in a thickener and then enters the fine-particle fractionation and recovery process;

[0098] S202: The ore pulp with a particle size of +0.15 mm enters the second-stage cyclone and is classified into a particle size of 1.2 mm to obtain a coarse particle size of +0.15 mm-1.2 mm and a particle size of +1.2 mm. The particle size of -0.8 mm accounts for 75% of the coarse particle size of +0.15 mm-1.2 mm. The ore pulp with a particle size of +1.2 mm is returned to the ball mill in step S1 as return sand for further grinding.

[0099] S3. Pre-discarding of coarse particles:

[0100] S301: 10 water-gas jet guns connected to the water-gas mixer are configured, the angle between the jet zone and the radial direction of the flotation column is 31.5°, and the number of damping rings is 2. The water-gas mixer sprays microbubbles and water-gas mixed fluid vertically upward inside the flotation column through the water-gas jet guns through the jet zone. After the flow field stabilizes, a microbubble and water fluidized composite interference bed is formed;

[0101] S302: The coarse-grained slurry is fully slurried by adding reagents. The amounts of reagents added during slurrying relative to the original ore are: 100 g / t copper sulfate, 200 g / t butyl xanthate, and 50 g / t pine oil. After slurrying, the slurry has a pH of 6-8 and a concentration of 35%. The slurry is then floated in a fluidized coarse-grained flotation machine to obtain concentrate A1 and tailings. The concentrate A1 is concentrated and de-refined in a thickener, and the underflow is returned to the ball mill in step S1 as return sand for regrinding.

[0102] S4, fine particle separation and recovery:

[0103] S401: Fine-grained slurry is fully slurried by adding reagents. The amount of reagents added relative to the original ore during slurry adjustment is: lime 1000g / t, zinc sulfate 600g / t, sodium sulfite 300g / t, dithiothreitol 100g / t, and pine oil 40g / t. After slurry adjustment, the pH value of the slurry is 6-8 and the slurry concentration is 35%. After slurry adjustment, the slurry enters the flotation column for roughing, obtaining a rough concentrate B1 and a roughing tailing C1.

[0104] S402: The coarse concentrate B1 is fully slurried with a reagent. The amount of reagent added during slurrying relative to the original ore is 200g / t of zinc sulfate. The slurry after slurrying is subjected to two concentrations to obtain lead concentrate as concentrate A2. The tailings obtained from the first concentration are returned to step S401, and the tailings obtained from the second concentration are returned to the first concentration.

[0105] S403: Add reagents to the roughing tailings C1 for full slurrying. After slurrying, the pulp is scoured twice by a flotation machine. The amount of reagent added in the first scouring relative to the original ore is: 30 g / t of ethyl disulfide and 20 g / t of pine oil. The concentrate obtained by the first scouring is returned to step S401. The amount of reagent added in the second scouring relative to the original ore is: 10 g / t of pine oil. The concentrate obtained by the second scouring is returned to the first scouring. Add reagents to the tailings obtained by the second scouring for full slurrying. The amount of reagent added relative to the original ore is: 1500 g / t of lime, 120 g / t of copper sulfate, 100 g / t of ethyl xanthate, and 40 g / t of pine oil. Lime is added first during slurrying, and then zinc sulfate and sodium sulfite are added simultaneously. After slurrying, the pulp is subjected to roughing by a flotation column to obtain a rough concentrate B2 and a roughing tailing C2.

[0106] S404: preparing aluminum ion-modified water glass from water glass and aluminum sulfate; adding a reagent including aluminum ion-modified water glass to the coarse concentrate B2 to fully slurry, with the added amount of the reagent relative to the original ore being: 400 g / t of lime, 100 g / t of water glass, and 50 g / t of aluminum sulfate. After slurrying, the slurry undergoes a first concentration, and the tailings obtained from the first concentration are returned to the flotation column for roughing in step S403; adding an aluminum ion-modified water glass reagent to the concentrate obtained from the first concentration to fully slurry, with the added amount of the reagent relative to the original ore being: 50 g / t of water glass, and 25 g / t of aluminum sulfate. After slurrying, the slurry undergoes a second concentration to obtain a zinc concentrate as concentrate A3, and the tailings obtained from the second concentration are returned to the first concentration;

[0107] S405: Add reagents to the roughing tailings C2 for slurry adjustment. The amount of reagents added relative to the original ore is: 30g / t of ethyl xanthate and 10g / t of pine oil. After slurry adjustment, the slurry is scoured twice by a flotation machine. The concentrate obtained from the first scavenging is returned to the flotation column for roughing in step S403, and the concentrate obtained from the second scavenging is returned to the first scavenging. The tailings obtained from the second scavenging are combined with the tailings from step S302 as the final tailings.

[0108] The yield, Pb grade, Zn grade, Pb recovery rate, and Zn recovery rate of the lead concentrate, zinc concentrate, and final tailings of Example 1 were measured and calculated. The results are shown in Table 1 below:

[0109] Table 1 Flotation test results of a lead-zinc sulfide ore in Hunan

[0110]

[0111] As can be seen from Table 1 above, the present invention can achieve the separate recovery of fine and coarse fractions of lead-zinc sulfide ore, and can increase the recovery rates of lead and zinc by more than 2 percentage points compared with the existing flotation process. The grade of the target mineral contained in the tailings is extremely low, and the energy consumption of grinding is reduced by about 28%.

[0112] Example 2:

[0113] like Figure 2 As shown in FIG. 1 , a preferred embodiment of the flotation process of the present invention combining pre-discarding of coarse particles with recovery of fine particles by quality separation is shown. The process includes:

[0114] S1. Grinding: After the raw ore is crushed, water is added to adjust the slurry concentration to 35%. The raw ore is a lead-zinc sulfide ore in Jiangxi Province with a lead grade of 1.85% and a zinc grade of 2.15%. The ore is then fed into a ball mill for a first grinding step until the slurry fineness reaches -200 mesh, accounting for 65%, to obtain the ball mill discharge slurry.

[0115] S2. Classification:

[0116] S201: The ball mill discharge slurry is classified into 0.15mm particle size using a cyclone to obtain a fine-particle slurry of -0.15mm particle size and a slurry of +0.15mm particle size. The -0.15mm particle size is concentrated in a thickener and then enters the fine-particle fractionation and recovery process;

[0117] S202: The slurry with a particle size of +0.15 mm enters the second-stage cyclone and is classified into a particle size of 1.2 mm to obtain a coarse particle size of +0.15 mm-1.2 mm and a particle size of +1.2 mm. The particle size of -0.8 mm accounts for 80% of the coarse particle size of +0.15 mm-1.2 mm. The slurry with a particle size of +1.2 mm is returned to the ball mill in step S1 as return sand for further grinding.

[0118] S3. Pre-discarding of coarse particles:

[0119] S301: 8 water-gas jet guns connected to the water-gas mixer are configured, the angle between the jet zone and the radial direction of the flotation column is 26.5°, and the number of damping rings is 2. The water-gas mixer sprays microbubbles and water-gas mixed fluid vertically upward inside the flotation column through the water-gas jet guns through the jet zone. After the flow field stabilizes, a microbubble and water fluidized composite interference bed is formed;

[0120] S302: The coarse-grained slurry is fully slurried by adding reagents. The amounts of reagents added during slurrying relative to the original ore are: 150 g / t copper sulfate, 200 g / t butyl xanthate, and 60 g / t pine oil. After slurrying, the slurry has a pH of 6-8 and a concentration of 35%. The slurry is then floated in a fluidized coarse-grained flotation machine to obtain concentrate A1 and tailings. The concentrate A1 is concentrated and de-refined in a thickener, and the underflow is returned to the ball mill in step S1 as return sand for regrinding.

[0121] S4, fine particle separation and recovery:

[0122] S401: Fine-grained slurry is fully slurried by adding reagents. The amount of reagents added relative to the original ore during slurry adjustment is: lime 1500g / t, zinc sulfate 800g / t, sodium sulfite 400g / t, dithiothreitol 120g / t, and pine oil 40g / t. After slurry adjustment, the slurry pH is 6-8 and the slurry concentration is 40%. After slurry adjustment, the slurry enters the flotation column for roughing to obtain a rough concentrate B1 and a roughing tailing C1.

[0123] S402: The coarse concentrate B1 is fully slurried with a reagent. The amount of reagent added during slurrying relative to the original ore is 300g / t of zinc sulfate. The slurry after slurrying is subjected to two concentrations to obtain lead concentrate as concentrate A2. The tailings obtained from the first concentration are returned to step S401, and the tailings obtained from the second concentration are returned to the first concentration.

[0124] S403: The roughing tailings C1 are fully slurried by adding reagents. After slurrying, the pulp is scoured twice by a flotation machine. The amount of reagent added in the first scouring relative to the original ore is: 30 g / t of ethyl disulfide and 30 g / t of pine oil. The concentrate obtained by the first scouring is returned to step S401. The amount of reagent added in the second scouring relative to the original ore is: 15 g / t of pine oil. The concentrate obtained by the second scouring is returned to the first scouring; the tailings obtained by the second scouring are fully slurried by adding reagents. The amount of reagent added relative to the original ore is: 2000 g / t of lime, 150 g / t of copper sulfate, 80 g / t of ethyl xanthate, and 50 g / t of pine oil. During slurrying, lime is first added, and then zinc sulfate and sodium sulfite are added simultaneously. After slurrying, the pulp is subjected to roughing by a flotation column to obtain a rough concentrate B2 and a roughing tailings C2.

[0125] S404: preparing aluminum ion-modified water glass from water glass and aluminum sulfate; adding a reagent including aluminum ion-modified water glass to the coarse concentrate B2 to fully slurry, with the added amount of reagent relative to the original ore being: 600 g / t of lime, 150 g / t of water glass, and 75 g / t of aluminum sulfate. After slurrying, the slurry undergoes a first concentration, and the tailings obtained from the first concentration are returned to the flotation column for roughing in step S403; adding aluminum ion-modified water glass to the concentrate obtained from the first concentration to fully slurry, with the added amount of reagent relative to the original ore being: 100 g / t of water glass, and 50 g / t of aluminum sulfate. After slurrying, the slurry undergoes a second concentration to obtain zinc concentrate as concentrate A3, and the tailings obtained from the second concentration are returned to the first concentration;

[0126] S405: Add reagents to the roughing tailings C2 for slurry adjustment. The amount of reagents added relative to the original ore is: 30g / t of ethyl xanthate and 20g / t of pine oil. After slurry adjustment, the slurry is scoured twice by a flotation machine. The concentrate obtained from the first scavenging is returned to the flotation column for roughing in step S403, and the concentrate obtained from the second scavenging is returned to the first scavenging. The tailings obtained from the second scavenging are combined with the tailings from step S302 as the final tailings.

[0127] The yield, Pb grade, Zn grade, Pb recovery rate, and Zn recovery rate of the lead concentrate, zinc concentrate, and final tailings of Example 2 were measured and calculated. The results are shown in Table 2 below:

[0128] Table 2 Flotation test results of a lead-zinc sulfide ore in Jiangxi

[0129]

[0130] As can be seen from Table 2 above, the present invention can achieve separate recovery of fine and coarse fractions of lead-zinc sulfide ore, and can increase the recovery rates of lead and zinc by more than 1.5 percentage points compared with the existing flotation process. The grade of the target mineral contained in the tailings is extremely low, and the energy consumption of grinding is reduced by about 24%.

[0131] Example 3:

[0132] like Figure 3 As shown in FIG. 1 , a preferred embodiment of the flotation process of the present invention combining pre-discarding of coarse particles with recovery of fine particles by quality separation is shown. The process includes:

[0133] S1. Grinding: After the raw ore is crushed, water is added to the slurry to adjust the slurry concentration to 35%. The raw ore is a mixed black and white tungsten ore from Hunan: the WO3 grade is 0.47%. The ore is then fed into a ball mill for a first grinding step to a slurry fineness of -200 mesh accounting for 70%, obtaining the ball mill discharge slurry.

[0134] S2. Classification:

[0135] S201: The ball mill discharge slurry is classified into 0.15mm particle size using a cyclone to obtain a fine-particle slurry of -0.15mm particle size and a slurry of +0.15mm particle size. The -0.15mm particle size is concentrated in a thickener and then enters the fine-particle fractionation and recovery process;

[0136] S202: The ore pulp with a particle size of +0.15 mm enters the second-stage cyclone and is classified into a particle size of 1.2 mm to obtain a coarse particle size of +0.15 mm-1.2 mm and a particle size of +1.2 mm. The particle size of -0.8 mm accounts for 75% of the coarse particle size of +0.15 mm-1.2 mm. The ore pulp with a particle size of +1.2 mm is returned to the ball mill in step S1 as return sand for further grinding.

[0137] S3. Pre-discarding of coarse particles:

[0138] S301: 12 water-gas jet guns connected to the water-gas mixer are configured, the angle between the jet zone and the radial direction of the flotation column is 27 degrees, and the number of damping rings is 3. The water-gas mixer sprays microbubbles and water-gas mixed fluid vertically upward inside the flotation column through the water-gas jet guns through the jet zone. After the flow field stabilizes, a microbubble and water fluidized composite interference bed is formed;

[0139] S302: The coarse-grained slurry is fully slurried by adding reagents. The amounts of reagents added during slurrying relative to the original ore are: 300 g / t of lead nitrate, 250 g / t of benzohydroxamic acid, 20 g / t of pine oil, and 360 g / t of sodium carbonate. A lead-benzohydroxamic acid complex is prepared from lead nitrate and benzohydroxamic acid, and the order of addition is sodium carbonate, lead-benzohydroxamic acid complex, and pine oil. After slurrying, the pH value of the slurry is 9.5 and the slurry concentration is 35%. The slurry is floated in a fluidized coarse-grained flotation machine to obtain concentrate A1 and tailings. The concentrate A1 is concentrated and the reagents are removed in a thickener. The underflow is returned to the ball mill in step S1 as return sand for regrinding.

[0140] S4, fine particle separation and recovery:

[0141] S401: Fine-grained ore pulp is fully slurried by adding reagents: 400 g / t of lead nitrate, 300 g / t of benzohydroxamic acid, 30 g / t of pine oil, and 400 g / t of sodium carbonate; a lead-benzohydroxamic acid complex is prepared from lead nitrate and benzohydroxamic acid, and the order of addition is sodium carbonate, lead-benzohydroxamic acid complex, and pine oil; after slurrying, the pH value of the ore pulp is 9.5 and the ore pulp concentration is 40%. After slurrying, the ore pulp enters the flotation column for roughing, obtaining a rough concentrate B1 and a roughing tailing C1;

[0142] S402: Aluminum ion-modified water glass is prepared from water glass and aluminum sulfate; aluminum ion-modified water glass reagent is added to the coarse concentrate B1 and fully slurried. After slurrying, the slurry is subjected to two concentrations to obtain concentrate A2. The amount of reagent added to the original ore during the first concentration is: 120g / t water glass, 60g / t aluminum sulfate, and the tailings obtained from the first concentration are returned to step S401; the amount of reagent added to the original ore during the second concentration is: 60g / t water glass, 30g / t aluminum sulfate; the tailings obtained from the second concentration are returned to the first concentration;

[0143] S403: Add reagents to the roughing tailings C1 for full slurry adjustment. The amount of reagents added relative to the original ore is: 100g / t of lead nitrate and 50g / t of benzohydroxamic acid; a lead-benzohydroxamic acid complex is prepared from lead nitrate and benzohydroxamic acid and then added to the slurry; After slurry adjustment, the slurry is scavenged three times by a flotation machine. The concentrate obtained from the first scavenging is returned to step S401, the concentrate obtained from the second scavenging is returned to the first scavenging, and the concentrate obtained from the third scavenging is returned to the second scavenging. The tailings obtained from the third scavenging are combined with the tailings from step S302 to form the final tailings. The yield, WO3 grade, and WO3 recovery rate of the concentrate A2 and the final tailings of Example 3 are measured and calculated, and the results are shown in Table 3 below:

[0144] Table 3 Flotation test results of a black and white tungsten mixed ore in Hunan

[0145]

[0146] As can be seen from Table 3 above, the present invention can achieve the separate recovery of fine and coarse tungsten ore, and can increase the recovery rate of WO3 by more than 2 percentage points compared with the existing flotation process. The grade of the target mineral contained in the tailings is extremely low, and the grinding energy consumption is reduced by about 30%.

[0147] Example 4:

[0148] like Figure 3 As shown in FIG. 1 , a preferred embodiment of the flotation process of the present invention combining pre-discarding of coarse particles with fractionation and recovery of fine particles is shown. The process includes:

[0149] S1. Grinding: After the raw ore is crushed, water is added to the slurry to adjust the slurry concentration to 35%. The raw ore is a Hunan scheelite ore with a WO3 grade of 0.22%. The ore is then fed into a ball mill for a first grinding step to a slurry fineness of -200 mesh, accounting for 65%, to obtain the ball mill discharge slurry;

[0150] S2. Classification:

[0151] S201: The ball mill discharge slurry is classified into 0.15mm particle size using a cyclone to obtain a fine-particle slurry of -0.15mm particle size and a slurry of +0.15mm particle size. The -0.15mm particle size is concentrated in a thickener and then enters the fine-particle fractionation and recovery process;

[0152] S202: The ore pulp with a particle size of +0.15 mm enters the second-stage cyclone and is classified into a particle size of 1.2 mm to obtain a coarse particle size of +0.15 mm-1.2 mm and a particle size of +1.2 mm. The particle size of -0.8 mm accounts for 70% of the coarse particle size of +0.15 mm-1.2 mm. The ore pulp with a particle size of +1.2 mm is returned to the ball mill in step S1 as return sand for further grinding.

[0153] S3. Pre-discarding of coarse particles:

[0154] S301: 10 water-gas jet guns connected to the water-gas mixer are configured, the angle between the jet zone and the radial direction of the flotation column is 21 degrees, and the number of damping rings is 2. The water-gas mixer sprays microbubbles and water-gas mixed fluid vertically upward inside the flotation column through the water-gas jet guns through the jet zone. After the flow field stabilizes, a microbubble and water fluidized composite interference bed is formed;

[0155] S302: The coarse-grained slurry is fully slurried by adding reagents. The amounts of reagents added during slurrying relative to the original ore are: 350 g / t of lead nitrate, 300 g / t of salicylic acid, 20 g / t of pine oil, and 430 g / t of sodium carbonate. A lead-salicylic acid complex is prepared from lead nitrate and salicylic acid, and the order of addition is sodium carbonate, lead-salicylic acid complex, and pine oil. After slurrying, the pH value of the slurry is 9.2 and the slurry concentration is 35%. The slurry is floated in a fluidized coarse-grained flotation machine to obtain concentrate A1 and tailings. The concentrate A1 is concentrated and the reagents are removed in a thickener. The underflow is returned to the ball mill in step S1 as return sand for regrinding.

[0156] S4, fine particle separation and recovery:

[0157] S401: Fine-grained ore pulp is fully slurried by adding reagents: 450 g / t of lead nitrate, 400 g / t of salicylic acid, 30 g / t of pine oil, and 400 g / t of sodium carbonate; a lead-salicylic acid complex is prepared from lead nitrate and salicylic acid, and the order of addition is sodium carbonate, lead-salicylic acid complex, and pine oil; after slurrying, the pH value of the ore pulp is 9.2, and the ore pulp concentration is 45%. After slurrying, the ore pulp enters a flotation column for roughing, obtaining a rough concentrate B1 and a roughing tailing C1;

[0158] S402: Aluminum ion-modified water glass is prepared from water glass and aluminum sulfate; aluminum ion-modified water glass reagent is added to the coarse concentrate B1 and fully slurried. After slurrying, the slurry is subjected to two concentrations to obtain concentrate A2. The amount of reagent added to the original ore during the first concentration is: 150g / t water glass, 75g / t aluminum sulfate, and the tailings obtained from the first concentration are returned to step S401; the amount of reagent added to the original ore during the second concentration is: 80g / t water glass, 40g / t aluminum sulfate; the tailings obtained from the second concentration are returned to the first concentration;

[0159] S403: Add reagents to the roughing tailings C1 to fully adjust the slurry. The amount of reagents added relative to the original ore is: 80g / t of lead nitrate and 80g / t of benzohydroxamic acid; a lead-benzohydroxamic acid complex is prepared from lead nitrate and benzohydroxamic acid and then added to the slurry; After slurry adjustment, the slurry is scavenged three times by a flotation machine. The concentrate obtained from the first scavenging is returned to step S401, the concentrate obtained from the second scavenging is returned to the first scavenging, and the concentrate obtained from the third scavenging is returned to the second scavenging. The tailings obtained from the third scavenging are combined with the tailings from step S302 to form the final tailings. The yield, WO3 grade, and WO3 recovery rate of the concentrate A2 and the final tailings of Example 4 are measured and calculated, and the results are shown in Table 4 below:

[0160] Table 4 Flotation test results of a black and white tungsten mixed ore in Hunan

[0161]

[0162] As can be seen from Table 4 above, the present invention can achieve the separate recovery of fine and coarse tungsten ore, and can increase the recovery rate of WO3 by more than 3 percentage points compared with the existing flotation process. The grade of the target mineral contained in the tailings is extremely low, and the grinding energy consumption is reduced by about 20%.

[0163] Example 5:

[0164] like Figure 4 As shown in FIG. 1 , a preferred embodiment of the flotation process of the present invention combining pre-discarding of coarse particles with recovery of fine particles by quality separation is shown. The process includes:

[0165] S1. Grinding: After the raw ore is crushed, water is added to the slurry to adjust the slurry concentration to 35%. The raw ore is a cassiterite from Hunan: the SnO2 grade is 0.40%, and then enters the ball mill for a stage of grinding until the slurry fineness is: -200 mesh accounts for 70%, and the ball mill discharge slurry is obtained;

[0166] S2. Classification:

[0167] S201: The ball mill discharge slurry is classified into 0.15mm particle size using a cyclone to obtain a fine-particle slurry of -0.15mm particle size and a slurry of +0.15mm particle size. The -0.15mm particle size is concentrated in a thickener and then enters the fine-particle fractionation and recovery process;

[0168] S202: The ore pulp with a particle size of +0.15 mm enters the second-stage cyclone and is classified into a particle size of 1 mm to obtain a coarse particle size of +0.15 mm-1 mm and a particle size of +1 mm. The particle size of -0.8 mm accounts for 80% of the coarse particle size of +0.15 mm-1 mm. The ore pulp with a particle size of +1 mm is returned to the ball mill in step S1 as return sand for further grinding.

[0169] S3. Pre-discarding of coarse particles:

[0170] S301: 14 water-gas jet guns connected to the water-gas mixer are configured, the angle between the jet zone and the radial direction of the flotation column is 36.5°, and the number of damping rings is 4. The water-gas mixer sprays microbubbles and water-gas mixed fluid vertically upward inside the flotation column through the water-gas jet guns through the jet zone. After the flow field stabilizes, a microbubble and water fluidized composite interference bed is formed;

[0171] S302: The coarse-grained slurry is fully slurried by adding reagents. The amounts of reagents added during slurrying relative to the original ore are: 400 g / t of lead nitrate, 300 g / t of octylhydroxamic acid, 30 g / t of pine oil, and 530 g / t of sodium carbonate. A lead-octylhydroxamic acid complex is prepared from the lead nitrate and octylhydroxamic acid. The order of addition is sodium carbonate, lead-octylhydroxamic acid complex, and pine oil. After slurrying, the pH value of the slurry is 10 and the slurry concentration is 35%. The slurry is floated in a fluidized coarse-grained flotation machine to obtain concentrate A1 and tailings. The concentrate A1 is concentrated and the reagents are removed in a thickener. The underflow is returned to the ball mill in step S1 as return sand for regrinding.

[0172] S4, fine particle separation and recovery:

[0173] S401: Fine-grained ore pulp is fully slurried by adding reagents. The amounts of reagents added relative to the original ore during slurrying are: 450 g / t of lead nitrate, 350 g / t of octylhydroxamic acid, 40 g / t of pine oil, and 650 g / t of sodium carbonate. A lead-octylhydroxamic acid complex is prepared from lead nitrate and octylhydroxamic acid. The order of addition is sodium carbonate, lead-octylhydroxamic acid complex, and pine oil. After slurrying, the pH value of the ore pulp is 10 and the ore pulp concentration is 35%. After slurrying, the ore pulp enters a flotation column for roughing to obtain a rough concentrate B1 and a roughing tailing C1.

[0174] S402: preparing aluminum ion-modified water glass from water glass and aluminum sulfate; the coarse concentrate B1 is easy-to-float cassiterite, to which aluminum ion-modified water glass is added for full slurrying, and the slurry after slurrying is subjected to three concentrations: the amount of reagent added relative to the original ore during the first concentration and slurrying is: 150 g / t of water glass, 60 g / t of aluminum sulfate, and the tailings obtained from the first concentration are returned to step S401; the concentrate obtained from the first concentration is further concentrated by adding aluminum ion-modified water glass for full slurrying, and the slurry after slurrying is subjected to a second concentration, and the amount of reagent added relative to the original ore during slurrying is: 60 g / t of water glass, 30 g / t of aluminum sulfate, and the tailings obtained from the second concentration are returned to the first concentration; the concentrate obtained from the second concentration is subjected to a third concentration, and the tailings obtained from the third concentration are returned to the second concentration, and the high-grade concentrate is obtained from the third concentration as concentrate A2;

[0175] S403: The rougher tailings C1 is difficult-to-float cassiterite. Reagents are added to fully prepare the slurry. The amounts of reagents added relative to the original ore are: 200g / t of lead nitrate and 100g / t of octylhydroxamic acid. A lead-octylhydroxamic acid complex is prepared from lead nitrate and octylhydroxamic acid and then added to the slurry. After slurry preparation, the slurry is subjected to rough separation by flotation columns to obtain a rough concentrate B2 and rougher tailings C2.

[0176] S404: preparing modified aluminum ion-modified water glass from water glass and aluminum sulfate; adding a reagent including aluminum ion-modified water glass to the coarse concentrate B2 and fully slurrying the coarse concentrate B2, wherein the amount of the reagent added relative to the original ore is: water glass is 80g / t, and aluminum sulfate is 40g / t. After slurrying, the slurry undergoes a first concentration; the tailings obtained from the first concentration are returned to the flotation column for roughing in step S403; the concentrate obtained from the first concentration is added with an aluminum ion-modified water glass reagent to fully slurrying the coarse concentrate B2, wherein the amount of the reagent added relative to the original ore is: water glass is 40g / t, and aluminum sulfate is 20g / t. After slurrying, the slurry undergoes a second concentration, and the tailings obtained from the second concentration are returned to the first concentration. The second concentration obtains a medium-grade concentrate as concentrate A3;

[0177] S405: The roughing tailings C2 are scavenged twice by the flotation machine. The concentrate obtained from the first scavenging is returned to the flotation column for roughing in step S403. The concentrate obtained from the second scavenging is returned to the first scavenging. The tailings obtained from the second scavenging are combined with the tailings from step S302 as the final tailings.

[0178] The yield, SnO2 grade, and SnO2 recovery of the high-grade concentrate, medium-grade concentrate, and final tailings of Example 5 were measured and calculated. The results are shown in Table 5 below:

[0179] Table 5 Cassiterite flotation test results in Hunan

[0180]

[0181] As can be seen from Table 5 above, the present invention can achieve the separate recovery of fine and coarse fractions of cassiterite, and can increase the recovery rate of SnO2 by more than 2 percentage points compared with the existing flotation process, and the overall recovery rate of cassiterite is increased by more than 5 percentage points. The grade of the target mineral contained in the tailings is extremely low, and the energy consumption of grinding is reduced by about 28%.

[0182] Example 6:

[0183] like Figure 4 As shown in FIG. 1 , a preferred embodiment of the flotation process of the present invention combining pre-discarding of coarse particles with fractionation and recovery of fine particles is shown. The process includes:

[0184] S1. Grinding: After the raw ore is crushed, water is added to the slurry to adjust the slurry concentration to 35%. The raw ore is a cassiterite from Yunnan with a SnO2 grade of 0.31%. The ore is then fed into a ball mill for a first grinding step to a slurry fineness of -200 mesh, accounting for 65%, to obtain the ball mill discharge slurry.

[0185] S2. Classification:

[0186] S201: The ball mill discharge slurry is classified into 0.15mm particle size using a cyclone to obtain a fine-particle slurry of -0.15mm particle size and a slurry of +0.15mm particle size. The -0.15mm particle size is concentrated in a thickener and then enters the fine-particle fractionation and recovery process;

[0187] S202: The ore pulp with a particle size of +0.15 mm enters the second-stage cyclone and is classified into a particle size of 1.2 mm to obtain a coarse particle size of +0.15 mm-1.2 mm and a particle size of +1.2 mm. The particle size of -0.8 mm accounts for 70% of the coarse particle size of +0.15 mm-1.2 mm. The ore pulp with a particle size of +1.2 mm is returned to the ball mill in step S1 as return sand for further grinding.

[0188] S3. Pre-discarding of coarse particles:

[0189] S301: 10 water-gas jet guns connected to the water-gas mixer are configured, the angle between the jet zone and the radial direction of the flotation column is 41.5°, and the number of damping rings is 3. The water-gas mixer sprays microbubbles and water-gas mixed fluid vertically upward inside the flotation column through each water-gas jet gun through the jet zone. After the flow field stabilizes, a microbubble and water fluidized composite interference bed is formed;

[0190] S302: The coarse-grained slurry is fully slurried by adding reagents. The amounts of reagents added during slurrying relative to the original ore are: 350 g / t of lead nitrate, 300 g / t of octylhydroxamic acid, 30 g / t of pine oil, and 600 g / t of sodium carbonate. A lead-octylhydroxamic acid complex is prepared from the lead nitrate and octylhydroxamic acid. The order of addition is sodium carbonate, lead-octylhydroxamic acid complex, and pine oil. After slurrying, the pH value of the slurry is 9.0 and the slurry concentration is 35%. The slurry is floated in a fluidized coarse-grained flotation machine to obtain concentrate A1 and tailings. The concentrate A1 is concentrated and the reagents are removed in a thickener. The underflow is returned to the ball mill in step S1 as return sand for regrinding.

[0191] S4, fine particle separation and recovery:

[0192] S401: Fine-grained ore pulp is fully slurried by adding reagents. The amounts of reagents added relative to the original ore during slurrying are: 400 g / t of lead nitrate, 300 g / t of octylhydroxamic acid, 30 g / t of pine oil, and 550 g / t of sodium carbonate. A lead-octylhydroxamic acid complex is prepared from lead nitrate and octylhydroxamic acid. The order of addition is sodium carbonate, lead-octylhydroxamic acid complex, and pine oil. After slurrying, the pH value of the ore pulp is 9.0 and the ore pulp concentration is 40%. After slurrying, the ore pulp enters a flotation column for roughing to obtain a rough concentrate B1 and a roughing tailing C1.

[0193] S402: preparing aluminum ion-modified water glass from water glass and aluminum sulfate; the coarse concentrate B1 is easy-to-float cassiterite, to which aluminum ion-modified water glass is added for full slurrying, and the slurry after slurrying is subjected to three concentrations: the amount of reagent added relative to the original ore during the first concentration and slurrying is: 120 g / t of water glass, 60 g / t of aluminum sulfate, and the tailings obtained from the first concentration are returned to step S401; the concentrate obtained from the first concentration is further subjected to a second concentration, and the amount of reagent added relative to the original ore during slurrying is: 80 g / t of water glass, 40 g / t of aluminum sulfate, and the tailings obtained from the second concentration are returned to the first concentration; the concentrate obtained from the second concentration is subjected to a third concentration, and the tailings obtained from the third concentration are returned to the second concentration, and the high-grade concentrate is obtained from the third concentration as concentrate A2;

[0194] S403: The rougher tailings C1 is difficult-to-float cassiterite. Reagents are added to fully prepare the slurry. The amount of reagents added relative to the original ore is: 150g / t of lead nitrate and 100g / t of octylhydroxamic acid. A lead-octylhydroxamic acid complex is prepared from lead nitrate and octylhydroxamic acid and then added to the slurry. After slurry preparation, the slurry is roughly separated by a flotation column to obtain a rough concentrate B2 and rougher tailings C2.

[0195] S404: preparing modified aluminum ion-modified water glass from water glass and aluminum sulfate; adding a reagent including aluminum ion-modified water glass to the coarse concentrate B2 to fully slurry, the amount of the reagent added relative to the original ore is: water glass is 60g / t, aluminum sulfate is 30g / t, and the slurry after slurrying is subjected to a first concentration; the tailings obtained from the first concentration are returned to the flotation column roughing of step S403; the concentrate obtained from the first concentration is added with an aluminum ion-modified water glass reagent to fully slurry, the amount of the reagent added relative to the original ore is: water glass is 30g / t, aluminum sulfate is 15g / t, and the slurry after slurrying is subjected to a second concentration, the tailings obtained from the second concentration are returned to the first concentration, and the second concentration obtains a medium-grade concentrate as concentrate A3;

[0196] S405: The roughing tailings C2 are scavenged twice by the flotation machine. The concentrate obtained from the first scavenging is returned to the flotation column for roughing in step S403. The concentrate obtained from the second scavenging is returned to the first scavenging. The tailings obtained from the second scavenging are combined with the tailings from step S302 as the final tailings.

[0197] The yield, SnO2 grade, and SnO2 recovery of the high-grade concentrate, medium-grade concentrate, and final tailings of Example 6 were measured and calculated. The results are shown in Table 6 below:

[0198] Table 6 Cassiterite flotation test results in Hunan

[0199]

[0200] As can be seen from Table 6 above, the present invention can achieve the separate recovery of fine and coarse cassiterite fractions, and can increase the recovery rate of SnO2 by more than 2 percentage points compared with the existing flotation process, and the overall recovery rate of cassiterite is increased by more than 5 percentage points. The grade of the target mineral contained in the tailings is extremely low, and the energy consumption of grinding is reduced by about 20%.

[0201] The fluidized coarse particle flotation machine used in step S3 of the above embodiments 1-6 is as follows: Figure 5-9 As shown, it includes a flotation column 1, a connected water-gas mixer 9 and a plurality of water-gas jet guns 8. The top of the flotation column 1 is provided with a foam overflow trough 2, a concentrate discharge port 3 and a feeding device, and the bottom of the flotation column 1 is provided with a discharge port 6. The water-gas jet gun 8 has a jet area extending into the interior of the flotation column 1 for vertically spraying a mixed fluid of microbubbles and water-gas upwards.

[0202] The size of microbubbles in the mixed fluid of microbubbles and water and gas is 50-200 μm, the dispersion of microbubbles in the mixed fluid of water and gas is more than 80%, and the angle α between the jet zone and the radial direction of the flotation column is 0-45°; the coarse-grained slurry is sprayed in the flotation column, and the microbubbles and water fluidization composite interference bed flotation formed by stabilizing the flow field by the jet zone can optimize the flotation effect, and the concentrate A1 is obtained at the concentrate discharge port 3, and the tailings are obtained as the final tailings at the discharge port 6.

[0203] Furthermore, the feeding device includes a feeding pipe 4 extending into the interior of the flotation column 1, and a feeding distributor 5 connected to the feeding pipe 4 and located below the foam overflow trough 2, so that the slurry can slowly descend along the feeding pipe 4 and the feeding distributor 5 in the flotation column 1 over the entire cross-section, thereby achieving uniform feeding and further improving the flotation effect.

[0204] Furthermore, the middle and upper parts of the flotation column 1 are cylindrical, and the lower part of the flotation column 1 is a cone, which is used to stabilize the microbubble and water fluidization composite interference bed through the cylinder and discharge ore through the cone; the longitudinal length ratio of the cylinder and the cone is (2.5-3.5):1. The longitudinal length ratio can further form a stable microbubble and water fluidization composite interference bed suitable for coarse particle mineral separation. If the cylindrical ratio is too low, the vertical space required for bed stability cannot be achieved. If the ratio is too high, the coarse particles will have a long floating path after adhering to bubbles. , which will cause the desorption of particles and bubbles, thereby worsening the flotation effect; the angle between the cone wall and the vertical axis is 40-50°. The angle limitation can further make the discharge of the inner tailings smoother, and the discharge speed is moderate and easy to control the discharge volume. If the angle is too large, the ore moves slowly on the cone, the discharge is not smooth and it is easy to cause blockage. If the angle is too small, the ore moves too fast on the cone and the discharge volume is difficult to control; the jet zone is close to the junction area between the cylinder and the cone, which is used to further play the role of bed stabilization of the cylinder and the discharge of the cone.

[0205] Furthermore, the water-gas mixer 9 includes a shell 91 and a water-gas shear disperser 92. The shell 91 includes a connected water-gas mixing tube 911, a contraction tube 912, a throat 913 and a diffusion tube 914 in sequence along the fluid input to output direction. The water-gas mixing tube 911 is provided with a pressure water jet tube 915 and a plurality of high-pressure gas jet tubes 916 whose ends extend to the interior of the water-gas mixing tube 911. The high-pressure airflow of the high-pressure gas jet tube 916 can intersect with the pressure water flow of the pressure water jet tube 915 to form a water-gas mixed fluid. The water-gas shear disperser 92 is located in the diffusion tube 914.

[0206] Furthermore, the water-gas mixer 9 is arranged outside the flotation column 1, the pressure water jet pipe 915 is connected to the pressure water pump 11, and the pressure water pump 11 is connected to the water pool, which is used to pressure the water in the water pool into the pressure water jet pipe 915 through the pressure water pump 11 to form a pressure water jet, and several high-pressure gas jet pipes 916 are connected to the high-pressure gas tank 13, which are used to generate high-pressure airflow through the high-pressure gas tank 13 and distribute it to several high-pressure gas jet pipes 916 to form high-pressure gas jets, thereby realizing the generation of pressure water jets and high-pressure gas jets, and the flow field and turbulence can be further adjusted by adjusting the air pressure and water pressure.

[0207] Furthermore, the pressure water jet pipe 915 is provided with a conical nozzle with a cone angle of 20-30°, and the high-pressure gas jet pipe 916 is provided with a conical nozzle with a cone angle of 20-25°. The fluid injection angle between the pressure water jet pipe 915 and the high-pressure gas jet pipe 916 is 30-55°, which is used to further enhance the injection and further improve the water-gas mixing effect.

[0208] Furthermore, the pressure injection pipe is arranged axially along the shell 91, and a plurality of high-pressure gas injection pipes 916 are symmetrically arranged in pairs outside the pressure injection pipe to further improve the water-gas mixing effect.

[0209] Furthermore, the cone angle of the contraction tube 912 is 35-40°, the cone angle of the diffusion tube 914 is 25-35°, and the longitudinal length ratio of the contraction tube 912, the throat 913 and the diffusion tube 914 is 1:1:(2~3), which is used to further improve the dispersion uniformity of bubbles in the water-gas mixed fluid and refine the bubble size.

[0210] Furthermore, the water-gas shear disperser 92 includes a plurality of tips 921 perpendicular to the flow direction of the air bubble and water-gas mixed fluid, which is used to further enhance the microbubble effect by shearing and dispersing the passing air bubbles through the tips 921 .

[0211] Furthermore, the water-gas shear disperser 92 includes a plurality of spaced-apart, polygonal star-shaped cutting plates, preferably 4 to 8 in number, with the corners of each cutting plate serving as tips 921 arranged in a staggered manner for ease of arrangement and further enhancing the microbubble effect.

[0212] Furthermore, a water-gas distributor connected to the water-gas mixer 9 and the water-gas jet guns 8 is provided, and a valve is provided between the water-gas distributor and the water-gas jet guns 8, which is used to evenly distribute microbubbles and water-gas mixed fluid to the water-gas jet guns 8 through the water-gas distributor, and control the number of connected water-gas jet guns 8 by opening and closing the valve, thereby further realizing flexible adjustment of the water flow field and turbulence.

[0213] Furthermore, the water-gas jet gun 8 can adjust the angle between the jet zone and the circumferential radial direction of the flotation column 1, thereby simplifying the optimization exploration of flotation conditions by adjusting the flotation of coarse-grained ores of different types and properties.

[0214] Furthermore, the water vapor jet gun 8 is rotatably connected to the water vapor distributor, and an angle adjuster 804 for adjusting the radial angle between the jet area and the flotation column 1 is provided between the water vapor jet gun 8 and the flotation column 1. The angle adjuster 804 drives the water vapor jet gun 8 to move relative to the flotation column 1 to further realize the rapid and flexible adjustment of the radial angle between the jet area and the flotation column 1, and facilitates the adjustment of the water flow field and turbulence.

[0215] Furthermore, the water vapor distributor includes a water vapor distribution pipe 7 arranged in an annular shape outside the flotation column 1. The water vapor distribution pipe 7 is evenly spaced with a number of openings respectively controlled by valves. The water vapor jet gun 8 includes a vertical pipe 801 and a horizontal square pipe 802 connected. The horizontal square pipe 802 is arranged on the flotation column 1 and extends vertically into the flotation column 1 and is provided with a jet area. The jet area includes a fine grid 803 horizontally arranged at the fluid outlet of the water vapor jet gun 8. The area of each single hole of the fine grid is 0.04-0.36mm. 2 The single hole includes a square hole. The microbubbles and water-gas mixed fluid along the water-gas jet gun 8 are further dispersed into tiny bubbles and water columns through the fine grid 803, which can improve the flotation effect.

[0216] Furthermore, a damping structure is provided inside the flotation column above the jet zone, and the damping structure includes at least two damping rings 10 arranged continuously or at intervals. The multiple damping rings 10 on the inner wall of the flotation column 1 can effectively adjust the turbulence of the fluid to form a flow field distribution that is more suitable for the flotation of coarse particles. The longitudinal spacing from the lowest position of the damping structure to the bottom of the flotation column 1 is ≥1 / 3 of the axial height of the column, which is used to interfere with the bed layer and form a suitable flow field and turbulence through the number, distribution form and position of the damping structure: increasing the number and dense distribution of the damping rings 10 can reduce the turbulence of the flow field in the column, reduce the turbulent resistance of the coarse particles and bubbles in the floating process of adhesion, and promote the floating of coarse particles; avoid the position of the damping structure being too low, which will interfere with the formation of the bed layer too early and be detrimental to the stability of the bed layer; avoid the position being too high, which will have too weak an effect on the turbulence.

[0217] Furthermore, the damping ring 10 is annular and has a hemispherical protrusion on the inner side, which is used to reduce turbulent resistance through the hemispherical protrusion. The damping ring 10 is detachably connected to the flotation column 1 by bolts, etc., which is used to adjust the installation position, number and distribution spacing of several damping rings 10.

[0218] The working principle of the above-mentioned fluidized coarse particle flotation machine is as follows:

[0219] The pressure water pump 11 draws water from the water supply tank 12 and pressurizes it to produce pressure water. The high-pressure gas tank 13 produces high-pressure gas. The pressure water and high-pressure gas are respectively introduced into the water-gas mixer 9. The high-pressure airflow of the high-pressure gas injection pipe 916 intersects with the pressure water flow of the pressure water injection pipe 915, so that the gas is partially dissolved in the water with the help of pressure, and is initially mixed in the water-gas mixing pipe 911 to form a water-gas mixed fluid. Then, the water-gas mixed fluid passes through the contraction tube 912, the throat 913 and the diffusion pipe 914 in turn to form a large number of bubbles. The water-gas shear disperser 92 composed of a staggered polygonal star arrangement in the diffusion pipe 914 further cuts and disperses the bubbles into microbubbles, thereby improving the bubble dispersion degree and reducing the bubble size, and finally forming a mixed fluid of a large number of tiny bubbles and water at the output end of the water-gas mixer 9.

[0220] The formed microbubble and water vapor mixed fluid is evenly distributed along the circumference through the water vapor distribution pipe 7 and enters the water vapor jet gun 8. The fine grid 803 of the water vapor jet gun 8 further disperses the microbubble and water vapor mixed fluid into tiny bubbles and water flow columns. The microbubble size in the microbubble and water vapor mixed fluid ejected from the jet area can reach 50-200 μm, and the dispersion of microbubbles in the water vapor mixed fluid can reach more than 80%. This can solve the problem that the generation of bubbles and rising water flow in the flotation column 1 is relatively simple and extensive, and the bubble dispersion and bubble size are difficult to achieve the purpose of microbubble flotation.

[0221] The jet zone extends inside the flotation column 1, so that the microbubbles form a coordinated upward trend in the water-gas mixed fluid and are ejected vertically upward, forming an upward water-gas mixed flow field in the flotation column 1. After the flow field stabilizes, a composite interference bed of microbubbles and water fluidization is formed, which can solve the problem that the existing distribution of bubbles and rising water flow in the flotation column 1 is relatively simple and extensive, which is not conducive to the uniform distribution of the interference bed.

[0222] By freely combining the number and installation angle of the water-gas jet guns 8 and the number and layout of the damping rings, the problem that the flow field and turbulence are difficult to flexibly adjust due to the fixed components of the existing air bubble and rising water flow distributor as core components can be solved, thereby realizing flexible adjustment of the water flow field and turbulence in the fluidized coarse particle flotation machine; after the slurry and flotation reagent are fully slurried in the mixing barrel, they enter the flotation column 1 through the feeding pipe 4 and the feeding distributor 5, and slowly descend along the entire cross-section of the flotation column 1. The granular ore is floated in the fluidized interference bed, and the coarse particles containing the target mineral are simultaneously lifted by the buoyancy of the bubbles and the vertical lift of the rising water flow to become foam concentrate. The obtained concentrate can be further processed and recovered, while the gangue minerals have no bubble adhesion, sink and are discharged through the discharge port 6 to become final tailings. The tailings can be directly used as the final tailings in advance to complete the flotation separation process.

[0223] The series of detailed descriptions listed above are only specific descriptions of feasible embodiments of the present invention. They are not intended to limit the scope of protection of the present invention. Any equivalent embodiments or changes that do not deviate from the technical spirit of the present invention should be included in the scope of protection of the present invention.

Claims

1. A flotation process combining pre-disposal of coarse particles with recovery of fine particles by quality separation, characterized in that: The process includes: Grinding: After the raw ore is slurried, it is ground to a pulp fineness of -200 mesh accounting for 50-75%; Classification: The grinding pulp is classified according to the particle size of 0.15mm and 1-1.2mm to obtain fine-grained pulp with a particle size of -0.15mm and coarse-grained pulp with a particle size of +0.15mm-(1-1.2)mm. The particle size of -0.8mm accounts for 60% to 90% of the coarse-grained pulp, and the pulp with a particle size of +1-1.2mm is returned to the grinding process for re-grinding; Coarse particle pre-discarding: After adding reagents to the coarse particle pulp and adjusting the pulp, the coarse particle pulp is floated in the microbubble and water fluidized composite interference bed of the fluidized coarse particle flotation machine to obtain concentrate A1 and final tailings. After the concentrate A1 is concentrated and the medicine is removed, the bottom flow is returned to the grinding process as return sand for re-grinding; the pulp adjustment in the grinding process and the coarse particle pre-discarding process are controlled to have a pulp concentration of 30% to 40%; the fluidized coarse particle flotation machine includes a flotation column (1), a connected water-gas mixer (9) and a plurality of water-gas jet guns (8), the water-gas jet gun (8) is provided with a jet zone extending into the interior of the flotation column (1) for vertically spraying a microbubble and water-gas mixed fluid upward, the angle α between the jet zone and the circumference radial direction of the flotation column (1) is 0-45°, and the water-gas jet gun can adjust the angle between the jet zone and the circumference radial direction of the flotation column; Fine particle fractionation and recovery: After the fine particle size slurry is slurried with reagents, it enters the flotation column for roughing to obtain coarse concentrate B1 and roughing tailings C1; after the coarse concentrate B1 is slurried, it undergoes at least one cleaning to obtain concentrate A2; after the roughing tailings C1 is slurried, it undergoes at least one scavenging and / or flotation column roughing and / or cleaning to obtain concentrate A3 and / or final tailings.

2. The flotation process combining pre-disposal of coarse particles with recovery of fine particles by quality separation according to claim 1 is characterized in that: The coarse-grained slurry in the fine-grained fractionation and recovery process is adjusted to control the slurry concentration to be 30-50%. After classification, the fine-grained slurry and the slurry with a particle size of +1-1.2 mm are respectively concentrated.

3. The flotation process combining pre-disposal of coarse particles with recovery of fine particles by quality separation according to claim 1 is characterized in that: The raw ore includes sulfide ore and oxide ore. The particle size of the coarse particle size pulp is determined according to the positive correlation between the floatability of the raw ore, the jet flow rate of the fluidized coarse particle flotation machine, and the pressure and the upper limit of the classification particle size.

4. The flotation process combining pre-disposal of coarse particles with recovery of fine particles by quality separation according to claim 1 is characterized in that: The size of microbubbles in the mixed fluid of microbubbles and water gas is 50-200 μm, and the dispersion of microbubbles in the mixed fluid of water gas is more than 80%; The coarse-grained ore pulp is floated in a flotation column (1) by a composite interference bed of microbubbles and water fluidization formed by spraying and stabilizing the flow field in the jet zone, thereby obtaining concentrate A1 and final tailings.

5. The flotation process combining coarse particle pre-disposal and fine particle separation and recovery according to claim 1 is characterized in that: The agents added for slurry adjustment include one or more of pH regulators, activators, collectors, foaming agents, and inhibitors; The pH regulator includes sodium carbonate; the activator includes copper sulfate; The collector includes one or more of butyl xanthate, complex, and ethyl xanthate; The foaming agent includes pine oil; The inhibitor includes one or more of lime, zinc sulfate, sodium sulfite, and modified water glass.

6. The flotation process according to claim 5, characterized in that: The complex is formed by coordinating lead nitrate and an organic ligand in a molar ratio of (1-2):

1. The organic ligand comprises one or more of octylhydroxamic acid, benzohydroxamic acid, salicylic hydroxamic acid, cupferroni, sodium oleate, styrenephosphonic acid, and sodium lauryl sulfate.

7. The flotation process combining pre-disposal of coarse particles with recovery of fine particles by quality separation according to claim 5, characterized in that: The modified water glass is a composite metal silicate formed by the reaction of metal ions and water glass in a mass ratio of (1-2):4, wherein the metal ions include Al 3+ 、Cu 2+ 、Fe 3+ , Pb 2+ One of them.

8. The flotation process according to any one of claims 1 to 7, characterized in that: The raw ore is lead-zinc sulfide ore, and the grinding slurry fineness is: -200 mesh accounts for 50-70%; The reagents added during the slurry preparation of the coarse particles include copper sulfate, xanthate, and pine oil, and the pH value of the slurry is 6 to 8; The reagents added during the roughing and slurry adjustment of fine-grained slurry include lime, zinc sulfate, sodium sulfite, ethylthiocyanate, and pine oil. The pH value of the slurry is 6-8. The reagents added during the selection and slurrying of the coarse concentrate B1 include zinc sulfate, and the lead concentrate obtained after two rounds of selection is used as concentrate A2; The rougher tailings C1 are slurried by adding reagents including ethyl thiocarb and pine oil, and after two scavenging processes, the reagents including lime, copper sulfate, ethyl xanthate and pine oil are added to slurry, and after roughing by flotation columns, the rough concentrate B2 and rougher tailings C2 are obtained; The coarse concentrate B2 is slurried by adding reagents including lime and modified water glass, and is subjected to two rounds of cleaning to obtain zinc concentrate as concentrate A3; the rougher tailings C2 are slurried by adding reagents including ethyl xanthate and pine oil, and are subjected to two rounds of scavenging to obtain final tailings; The closed-loop processes for both selection and sweeping adopt sequential return.

9. The flotation process according to any one of claims 1 to 7, characterized in that: The raw ore is tungsten ore, and the grinding slurry fineness is: -200 mesh accounts for 60-75%; The reagents added during the roughing and slurry preparation of coarse-grained pre-discarded and fine-grained slurry include sodium carbonate, complex, and pine oil, and the slurry pH value is 9-10; The reagents added during the selection and slurrying of the coarse concentrate B1 include modified water glass, and the concentrate A2 is obtained after two rounds of selection; The roughing tailings C1 are added with reagents including complexes to prepare the slurry, and the final tailings are obtained after three scavenging processes; The closed-loop processes for both selection and sweeping adopt sequential return.

10. The flotation process combining pre-discarding of coarse particles and recovery of fine particles by quality separation according to any one of claims 1 to 7, characterized in that: The raw ore is cassiterite, and the grinding slurry fineness is: -200 mesh accounts for 60-75%; The reagents added during the roughing and slurry preparation of coarse-grained pre-waste and fine-grained slurry include sodium carbonate, complex, and pine oil, and the slurry pH value is 9-11; The reagents added during the selection and slurrying of the coarse concentrate B1 include modified water glass, and after three rounds of selection, a high-grade concentrate is obtained as concentrate A2; The rougher tailings C1 are added with reagents including complexes to prepare a slurry, and are subjected to rough separation by flotation columns to obtain a rough concentrate B2 and a rougher tailings C2; The coarse concentrate B2 is slurried by adding reagents including modified water glass, and after two rounds of beneficiation, a medium-grade concentrate is obtained as concentrate A3; The rougher tailings C2 are scavenged twice to obtain the final tailings; The closed-loop processes for both selection and sweeping adopt sequential return.

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