A flotation separation method for lithium mica ore

Abstract

The present invention relates to the field of mineral processing and flotation technology, and in particular to a flotation separation method for lithium mica ore. The process method comprises: (1) ore preparation; (2) pulp flotation and (3) concentrate treatment; wherein, in the pulp flotation process, a capture agent composed of alkyl hydroxamic acid and nanoparticles is added, and a lithium mica concentrate with high grade and recovery rate can be harvested through one flotation, wherein the grade is not less than 4.52% and the recovery rate is not less than 85.3%, which is comparable to the traditional "one roughing, multiple refining, multiple sweeping" process flow, and realizes the short-process flotation recovery of lithium mica ore.

Description

Technical Field

[0001] The invention relates to the technical field of ore dressing and flotation, and in particular to a flotation separation method for lepidolite ore. Background Art

[0002] Lepidolite is an important lithium resource. Its main component is lithium aluminum silicate, which is widely used in battery materials, ceramics, glass, and chemical industries. With the surge in global demand for electric vehicles and energy storage devices, the development and utilization of lithium mineral resources have become increasingly important. As an important mineral of lithium resources, the effective extraction and separation of lepidolite has become a research hotspot. However, the flotation process of lepidolite faces many challenges, especially the identification and processing of its embedded particle size, mineral composition and their interactions during the flotation process.

[0003] First, lepidolite usually exists in the ore as a fine embedded particle size, with a diameter of no more than tens of microns. This fine particle size makes it difficult to effectively separate and recover it in the traditional mineral processing process. Conventional flotation methods significantly reduce the flotation speed and efficiency of lepidolite with too small particle size, resulting in a decrease in the recovery rate of lepidolite. Therefore, it is particularly important to develop an efficient flotation process based on the unique particle size characteristics of lepidolite.

[0004] Secondly, the mineral composition of lithium mica is often complex, containing a variety of other minerals, such as potassium feldspar, dolomite and calcite. These minerals may interfere with the selective flotation of lithium mica during the flotation process, affecting the final recovery rate and purity of lithium mica. Traditional flotation processes usually require multiple roughing, concentrating and scavenging to achieve the expected head ore quality and lithium recovery rate, increasing production costs and time. Patent technology document CN201810567144.3 discloses a lithium mica flotation method, comprising the following steps: 1) ball milling and desludging; 2) first roughing; 3) second roughing; 4) scavenging; 5) first concentrating; 6) second concentrating; 7) third concentrating. The lithium mica flotation method of the invention uses hexadecyltrimethylammonium bromide and dodecylamine as collectors, and adopts a process flow of two roughing, one scavenging and three concentrating. The Li2O grade in the obtained lithium mica concentrate is 3.9%-4.3%, and the recovery rate of Li2O is 85%-95%. However, the entire flotation process of this process is long, has many dosing points, requires large investment and is complex to operate. Summary of the invention

[0005] In view of this, the purpose of the present invention is to propose a flotation separation method for lithium mica ore to solve the problem that the lithium mica has a fine embedded particle size and a complex mineral composition, and traditional mineral processing methods often require multiple roughing and fine separations to improve the recovery rate and purity of the lithium mica.

[0006] Based on the above purpose, the present invention provides a flotation separation method of lepidolite ore, comprising the following steps:

[0007] (1) Ore preparation: crushing, grinding, and adding water to obtain slurry;

[0008] (2) Slurry flotation: add a regulator to the slurry to adjust the pH to 8-9, then add a collector and a frother, and flotate to obtain flotation concentrate and flotation tailings;

[0009] (3) Concentrate processing: Dehydrate, dry and purify the lithium mica concentrate to obtain high-purity lithium mica.

[0010] Preferably, in step (1), crushing to -2 mm and grinding to -74 μm account for 68%-73%.

[0011] Preferably, the concentration of the slurry in step (1) is 32-35wt%.

[0012] Furthermore, the collector in step (2) is a mixture of an alkyl hydroxamic acid collector and a nanoparticle collector in a weight ratio of 10:1.2-1.6.

[0013] Preferably, the molecular formula of the alkyl hydroxamic acid collector is RCONHOH, where R=C6-8.

[0014] More preferably, the molecular formula of the alkyl hydroxamic acid collector is RCONHOH, where R=C7.

[0015] Furthermore, the preparation method of the nanoparticle collector is as follows:

[0016] S1: Add nano-silica to deionized water, then add a toluene solution containing hexadecyltrimethoxysilane at a concentration of 1wt%-3wt%, stir at room temperature for 2-3min, then add triethylamine, continue stirring at room temperature for 2-3h, wash with water, wash with alcohol, and freeze-dry in vacuum to obtain partially hydrophobic nano-silica;

[0017] S2: adding part of the hydrophobic nano-silica to a mixed solution of deionized water and ethanol, stirring at room temperature for 2-3 minutes, then adding ureapropyltrimethoxysilane, heating to 50-55°C, stirring for 5-6 hours, washing with water, washing with alcohol, and vacuum freeze-drying to obtain a nanoparticle collector;

[0018] Preferably, in step S1, the weight ratio of nano-silicon dioxide, deionized water, toluene solution containing hexadecyltrimethoxysilane and triethylamine is 0.6-1.2:0.3-1:20-40:0.4-0.8;

[0019] Preferably, in step S2, the weight ratio of partially hydrophobic nano-silica, deionized water, ethanol and ureapropyltrimethoxysilane is 0.6-1.2:20-40:5-15:0.1-0.3.

[0020] Preferably, in step (2), the air flow rate of flotation is 0.8-1 m³ / min, the stirring speed is 1000-1400 rpm, and the time is 8-12 min.

[0021] Preferably, based on the weight of the slurry, the amount of the collector added in step (2) is 0.18-0.24 kg / t, and the amount of the frother added is 0.15-0.3 kg / t.

[0022] Preferably, in step (2), the adjusting agent is sodium carbonate and the foaming agent is a fatty alcohol foaming agent.

[0023] Furthermore, in step (2), the grade of the flotation concentrate is 4.52%-4.68%, and the recovery rate is 85.3%-88.4%.

[0024] Furthermore, the grade of the high-purity lepidolite in step (3) is 6.23%-6.42%.

[0025] Beneficial effects of the present invention:

[0026] The flotation separation method for lithium mica ore provided by the present invention adopts a composite of an alkyl hydroxamic acid collector and a nanoparticle collector as a novel capture agent, and lithium mica concentrate with high grade and recovery rate can be harvested through one flotation, wherein the grade is not less than 4.52% and the recovery rate is not less than 85.3%, which is comparable to the traditional "one coarse, multiple fine, and multiple sweep" process.

[0027] The nanoparticle collector used in the present invention can be adsorbed on minerals, which, on the one hand, increases the adsorption concentration of the reagent on the mineral surface, and on the other hand, increases the surface roughness of the mineral to form a larger hydrophobic surface, thereby improving the froth flotation effect. When combined with an alkyl hydroxamic acid collector with high selectivity, the flotation effect on lepidolite is synergistically improved.

[0028] The nanoparticle collector used in the present invention has excellent amphoteric properties. The alkyl chain and urea thereon respectively improve the adhesion between mineral and air bubble and mineral and nanoparticle. In addition, since the nanoparticles have high temperature resistance, the process of the present invention is applicable to a wider temperature range and has broad application prospects. DETAILED DESCRIPTION

[0029] In order to make the objectives, technical solutions and advantages of the present invention more clearly understood, the present invention is further described in detail below in conjunction with specific embodiments.

[0030] The lithium mica ore in the specific embodiment of the present invention comes from Jiangte Mining Company, which is a lithium feldspar ore for production. The mineral embedding characteristics and embedding particle sizes in the ore are different. The feldspar minerals are mostly in the form of fine plate columns with a long diameter of about 0.1mm-0.2mm. The tantalum-niobium minerals are mostly in the form of translucent brown or dark brown fine needle columns with a long diameter of less than 0.1mm. The embedding particle size is fine and some of them are not dissociated into monomers.

[0031] Preparation Example 1:

[0032] S1: 6 g of nano-silica with a particle size of 80 nm was added to 3 g of deionized water, and then 200 g of a toluene solution containing hexadecyltrimethoxysilane with a concentration of 1 wt %, and stirred at room temperature for 2 min, and then 4 g of triethylamine was added, and stirring was continued at room temperature for 2 h, washed with water, washed with alcohol, and vacuum freeze-dried to obtain partially hydrophobic nano-silica;

[0033] S2: Add 6 g of partially hydrophobic nano-silica to a mixed solution of 200 g of deionized water and 50 g of ethanol, stir at room temperature for 2 min, then add 1 g of ureapropyltrimethoxysilane, heat to 50° C., stir for 5 h, wash with water, wash with alcohol, and freeze-dry in vacuum to obtain a nanoparticle collector;

[0034] S3: mixing 10 g of a nanoparticle collector and 1.2 g of an alkyl hydroxamic acid collector to obtain a collector;

[0035] Among them, the molecular formula of the alkyl hydroxamic acid collector is RCONHOH, R=C6.

[0036] Preparation Example 2:

[0037] S1: 9 g of nano-silica with a particle size of 80 nm was added to 6 g of deionized water, and then 300 g of a toluene solution containing hexadecyltrimethoxysilane with a concentration of 2 wt% was added, and the mixture was stirred at room temperature for 2 min, and then 6 g of triethylamine was added, and the mixture was stirred at room temperature for 2.5 h, washed with water, washed with alcohol, and vacuum freeze-dried to obtain partially hydrophobic nano-silica;

[0038] S2: Add 9 g of partially hydrophobic nano-silica to a mixed solution of 300 g of deionized water and 100 g of ethanol, stir at room temperature for 3 min, then add 2 g of ureapropyltrimethoxysilane, heat to 55° C., stir for 6 h, wash with water, wash with alcohol, and freeze-dry in vacuum to obtain a nanoparticle collector;

[0039] S3: mixing 10 g of a nanoparticle collector and 1.4 g of an alkyl hydroxamic acid collector to obtain a collector;

[0040] Among them, the molecular formula of the alkyl hydroxamic acid collector is RCONHOH, R=C7.

[0041] Preparation Example 3:

[0042] S1: 12 g of nano-silica with a particle size of 80 nm was added to 10 g of deionized water, and then 400 g of a toluene solution containing hexadecyltrimethoxysilane with a concentration of 3 wt % was added, and the mixture was stirred at room temperature for 3 min, and then 8 g of triethylamine was added, and the mixture was stirred at room temperature for 3 h, washed with water, washed with alcohol, and vacuum freeze-dried to obtain partially hydrophobic nano-silica;

[0043] S2: Add 12g of partially hydrophobic nano-silica to a mixed solution of 400g of deionized water and 150g of ethanol, stir at room temperature for 3min, then add 3g of ureapropyltrimethoxysilane, heat to 55°C, stir for 6h, wash with water, wash with alcohol, and freeze-dry in vacuum to obtain a nanoparticle collector;

[0044] S3: mixing 10 g of a nanoparticle collector and 1.5 g of an alkyl hydroxamic acid collector to obtain a collector;

[0045] Among them, the molecular formula of the alkyl hydroxamic acid collector is RCONHOH, R=C8.

[0046] Embodiment 1:

[0047] (1) Ore preparation: The lithium mica ore was crushed to -2 mm, ground to -74 μm accounting for 68.4%, and water was added to obtain a slurry with a concentration of 32 wt%;

[0048] (2) Slurry flotation: sodium carbonate was added to the slurry to adjust the pH to 8.2, and then 0.18 kg / t of the collector prepared in Preparation Example 1 and 0.15 kg / t of fatty alcohol polyoxyethylene ether AEO-7 were added based on the weight of the slurry, and flotation was performed. The flotation air flow rate was 0.8 m³ / min, the stirring speed was 1000 rpm, and the time was 8 min to obtain flotation concentrate and flotation tailings;

[0049] (3) Concentrate processing: The lithium mica concentrate is dehydrated, dried and purified to obtain high-purity lithium mica with a grade of 6.23%.

[0050] Embodiment 2:

[0051] (1) Ore preparation: The lithium mica ore was crushed to -2 mm, ground to -74 μm accounting for 71.2%, and water was added to obtain a slurry with a concentration of 33.5 wt%;

[0052] (2) Slurry flotation: sodium carbonate was added to the slurry to adjust the pH to 8.6, and then 0.2 kg / t of the collector prepared in Preparation Example 2 and 0.2 kg / t of fatty alcohol polyoxyethylene ether AEO-9 were added according to the weight of the slurry, and flotation was performed. The flotation air flow rate was 0.8 m³ / min, the stirring speed was 1200 rpm, and the time was 10 min to obtain flotation concentrate and flotation tailings;

[0053] (3) Concentrate processing: The lithium mica concentrate is dehydrated, dried and purified to obtain high-purity lithium mica with a grade of 6.32%.

[0054] Embodiment 3:

[0055] (1) Ore preparation: The lithium mica ore was crushed to -2 mm, ground to -74 μm accounting for 72.6%, and water was added to obtain a slurry with a concentration of 35 wt%;

[0056] (2) Slurry flotation: sodium carbonate was added to the slurry to adjust the pH to 8-9, and then 0.24 kg / t of the collector prepared in Preparation Example 3 and 0.3 kg / t of fatty alcohol polyoxyethylene ether AEO-9 were added according to the weight of the slurry, and flotation was performed. The flotation air flow rate was 1 m³ / min, the stirring speed was 1400 rpm, and the time was 12 min to obtain flotation concentrate and flotation tailings;

[0057] (3) Concentrate processing: The lithium mica concentrate is dehydrated, dried and purified to obtain high-purity lithium mica with a grade of 6.42%.

[0058] Comparative Example 1:

[0059] The main difference between Comparative Example 1 and Example 2 is that the alkyl hydroxamic acid collector is used alone;

[0060] (1) Ore preparation: The lithium mica ore was crushed to -2 mm, ground to -74 μm accounting for 71.6%, and water was added to obtain a slurry with a concentration of 33.5 wt%;

[0061] (2) Slurry flotation: Sodium carbonate was added to the slurry to adjust the pH to 8.4. Then, 0.2 kg / t of alkyl hydroxamic acid collector and 0.2 kg / t of fatty alcohol polyoxyethylene ether AEO-9 were added according to the weight of the slurry. The flotation was carried out with an air flow rate of 0.8 m³ / min, a stirring speed of 1200 rpm, and a time of 10 min to obtain flotation concentrate and flotation tailings.

[0062] Among them, the molecular formula of the alkyl hydroxamic acid collector is RCONHOH, R=C7.

[0063] Comparative Example 2:

[0064] The difference between Comparative Example 2 and Example 2 is that: the nanoparticle collector prepared in Preparation Example 2 is used alone;

[0065] (1) Ore preparation: The lithium mica ore was crushed to -2 mm, ground to -74 μm accounting for 71.5%, and water was added to obtain a slurry with a concentration of 33.5 wt%;

[0066] (2) Slurry flotation: sodium carbonate was added to the slurry to adjust the pH to 8.5, and then 0.2 kg / t of the nanoparticle collector prepared in Preparation Example 2 and 0.2 kg / t of fatty alcohol polyoxyethylene ether AEO-9 were added according to the weight of the slurry, and flotation was performed. The flotation air flow rate was 0.8 m³ / min, the stirring speed was 1200 rpm, and the time was 10 min to obtain flotation concentrate and flotation tailings;

[0067] Comparative Example 3:

[0068] The difference between Comparative Example 3 and Example 2 is that: the partially hydrophobic nano-silica prepared in Preparation Example 2 is used to replace the nano-particle collector, and the nano-silica is mixed with the alkyl hydroxamic acid collector to obtain the collector;

[0069] S1: 9 g of nano-silica with a particle size of 80 nm was added to 6 g of deionized water, and then 300 g of a toluene solution containing hexadecyltrimethoxysilane with a concentration of 2 wt% was added, and the mixture was stirred at room temperature for 2 min, and then 6 g of triethylamine was added, and the mixture was stirred at room temperature for 2.5 h, washed with water, washed with alcohol, and vacuum freeze-dried to obtain partially hydrophobic nano-silica;

[0070] S2: Mixing 10 g of partially hydrophobic nano-silica and 1.4 g of an alkyl hydroxamic acid collector to obtain a collector;

[0071] Among them, the molecular formula of the alkyl hydroxamic acid collector is RCONHOH, R=C7.

[0072] (1) Ore preparation: The lithium mica ore was crushed to -2 mm, ground to -74 μm accounting for 71.8%, and water was added to obtain a slurry with a concentration of 33.5 wt%;

[0073] (2) Slurry flotation: Sodium carbonate is added to the slurry to adjust the pH to 8.6. Then, 0.2 kg / t of collector and 0.2 kg / t of fatty alcohol polyoxyethylene ether AEO-9 are added according to the weight of the slurry. The flotation air flow rate is 0.8 m³ / min, the stirring speed is 1200 rpm, and the time is 10 min to obtain flotation concentrate and flotation tailings.

[0074] Comparative Example 4:

[0075] The difference between Comparative Example 4 and Example 2 is that no air is passed during the flotation process;

[0076] (1) Ore preparation: The lithium mica ore was crushed to -2 mm, ground to -74 μm accounting for 71.4%, and water was added to obtain a slurry with a concentration of 33.5 wt%;

[0077] (2) Slurry flotation: sodium carbonate was added to the slurry to adjust the pH to 8.5, and then 0.2 kg / t of the collector prepared in Preparation Example 2 and 0.2 kg / t of fatty alcohol polyoxyethylene ether AEO-9 were added according to the weight of the slurry, and flotation was performed. The stirring speed of the flotation was 1200 rpm and the time was 10 min to obtain a flotation concentrate and a flotation tailing.

[0078] Comparative Example 5:

[0079] The difference between Comparative Example 5 and Example 2 is that no foaming agent is added;

[0080] (1) Ore preparation: The lithium mica ore was crushed to -2 mm, ground to -74 μm accounting for 71.5%, and water was added to obtain a slurry with a concentration of 33.5 wt%;

[0081] (2) Slurry flotation: sodium carbonate was added to the slurry to adjust the pH to 8.7, and then 0.2 kg / t of the collector prepared in Preparation Example 2 was added based on the weight of the slurry. The flotation was performed with an air flow rate of 0.8 m³ / min, a stirring speed of 1200 rpm, and a flotation time of 10 min to obtain a flotation concentrate and a flotation tailing.

[0082] The grades and recoveries of the flotation concentrates and flotation tailings prepared in the examples and comparative examples are shown in Table 1;

[0083] Table 1 Grade and recovery rate of flotation concentrate and flotation tailings

[0084]

[0085] Data analysis: It can be seen from the embodiments and comparative examples that although the lepidolite flotation process provided by the present invention only adopts one flotation process, the grade of the flotation concentrate obtained is not less than 4.52%, and the recovery rate is not less than 85.3%, which is comparable to the traditional "one roughing, multiple refining, and multiple sweeping" process flow. While simplifying the process flow, the grade and recovery rate of the flotation concentrate are guaranteed.

[0086] It can be seen from Example 2 and Comparative Examples 1-2 that after one flotation, the grades of the flotation concentrates of the alkyl hydroxamic acid collector and the nanoparticle collector are 3.35 and 2.51 respectively, which are much lower than the collector formed by the mixture of the two. In addition, the effect of the nanoparticle collector alone is worse than that of the nanoparticle collector. This is mainly because the alkyl hydroxamic acid collector has high selectivity for lepidolite, while the nanoparticle collector can be adsorbed on the mineral, which increases the adsorption concentration of the agent on the mineral surface on the one hand, and increases the roughness of the mineral surface on the other hand, forming a larger hydrophobic surface. The froth flotation effect is improved, and the combination of the two synergistically improves the flotation effect on lepidolite.

[0087] It can be seen from Example 2 and Comparative Example 3 that the effect of partially hydrophobic nano-silica is lower than that of nanoparticle collector. This is mainly because the nanoparticle collector has higher surface activity than that of partially hydrophobic nano-silica, and the urea group post-grafted on the partially hydrophobic nano-silica has a higher adsorption capacity and can be more effectively adsorbed on the mineral surface.

[0088] It can be seen from Example 2 and Comparative Examples 4-5 that the addition of air and frother during the flotation process helps to improve the flotation effect. This is because the air and frother synergistically promote the generation of bubbles, and the fatty alcohol polyoxyethylene ether helps to stabilize the bubbles and prevent bubble breakage.

[0089] Those skilled in the art should understand that the discussion of any of the above embodiments is merely illustrative and is not intended to imply that the scope of the present invention is limited to these examples. Under the concept of the present invention, the technical features in the above embodiments or different embodiments may be combined, the steps may be implemented in any order, and there are many other variations of the different aspects of the present invention as described above, which are not provided in detail for the sake of simplicity.

Claims

1. A flotation separation method for lepidolite ore, characterized in that: The following steps are involved: (1) Ore preparation: crushing, grinding, and adding water to obtain slurry; (2) Slurry flotation: add a regulator to the slurry to adjust the pH to 8-9, then add a collector and a frother, and flotate to obtain flotation concentrate and flotation tailings; (3) Concentrate processing: Dehydrating, drying and purifying the lithium mica concentrate to obtain high-purity lithium mica; In the step (2), the collector is a mixture of an alkyl hydroxamic acid collector and a nanoparticle collector in a weight ratio of 10:1.2-1.6; the molecular formula of the alkyl hydroxamic acid collector is RCONHOH, R=C6-8; The preparation method of the nanoparticle collector is as follows: S1: Add nano-silica to deionized water, then add a toluene solution containing hexadecyltrimethoxysilane at a concentration of 1wt%-3wt%, stir at room temperature for 2-3min, then add triethylamine, continue stirring at room temperature for 2-3h, wash with water, wash with alcohol, and freeze-dry in vacuum to obtain partially hydrophobic nano-silica; S2: adding part of the hydrophobic nano-silica to a mixed solution of deionized water and ethanol, stirring at room temperature for 2-3 minutes, then adding ureapropyltrimethoxysilane, heating to 50-55°C, stirring for 5-6 hours, washing with water, washing with alcohol, and vacuum freeze-drying to obtain a nanoparticle collector; In the step S1, the weight ratio of nano-silicon dioxide, deionized water, toluene solution containing hexadecyltrimethoxysilane and triethylamine is 0.6-1.2:0.3-1:20-40:0.4-0.8; In the step S2, the weight ratio of partially hydrophobic nano-silica, deionized water, ethanol and ureapropyltrimethoxysilane is 0.6-1.2:20-40:5-15:0.1-0.

3.

2. The flotation separation method of lepidolite ore according to claim 1, characterized in that: In the step (1), the crushing to -2 mm and the grinding to -74 μm account for 68%-73%.

3. The flotation separation method of lepidolite ore according to claim 1, characterized in that: The concentration of the slurry in step (1) is 32wt%-35wt%.

4. The flotation separation method of lepidolite ore according to claim 1, characterized in that: The air flow rate of flotation in step (2) is 0.8-1m 3 / min, stirring speed is 1000-1400rpm, and time is 8-12min.

5. The flotation separation method of lepidolite ore according to claim 1, characterized in that: Calculated by weight of the ore pulp, the amount of the collector added in step (2) is 0.18-0.24 kg / t, and the amount of the frother added is 0.15-0.3 kg / t.

6. The flotation separation method of lepidolite ore according to claim 1, characterized in that: In the step (2), the adjusting agent is sodium carbonate, and the foaming agent is a fatty alcohol foaming agent.

7. The flotation separation method of lepidolite ore according to claim 1, characterized in that: The grade of the flotation concentrate in step (2) is 4.52%-4.68%, and the recovery rate is 85.3%-88.4%.

8. The flotation separation method of lepidolite ore according to claim 1, characterized in that: The grade of the high-purity lepidolite in step (3) is 6.23%-6.42%.

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