Flotation recovery method of high-mud type iron lepidolite

Abstract

The invention discloses a flotation recovery method of high-mud type iron lepidolite. The method comprises the following steps: 1) magnetic separation: crushing raw ore, and performing magnetic separation to obtain magnetic concentrate; (2) rapid flotation is conducted, specifically, a regulator GT-T composed of starch, lignosulfonate and sodium pyrophosphate is added into magnetic separation concentrate, after stirring is conducted, a collecting agent GT-B composed of lauryl amine and petroleum sodium sulfonate is added, after stirring is conducted, rough and fine open-circuit flotation is conducted, and high-grade iron lepidolite concentrate 1 is obtained; and 3) tailings re-separation: adding an adjusting agent GT-T into the tailings, stirring, adding a collecting agent GT-B, stirring, adding a flocculating agent GT-X composed of rosin amine polyoxyethylene ether and diesel oil, stirring, and carrying out closed-circuit flotation of one-time roughing, two-time concentration and two-time sweeping to obtain iron lepidolite concentrate 2. According to the method, rapid enrichment of coarse particles and efficient recovery of micro-fine particles are achieved through the magnetic separation-differential flotation technology, so that the separation efficiency of the iron lepidolite and the resource recovery rate are remarkably improved.

Description

Technical Field

[0001] The present invention belongs to the technical field of mineral processing, and specifically relates to a flotation recovery method for high-slime-type zinnwaldite, specifically to a combined beneficiation process for high-slime-type zinnwaldite ore, and particularly to a stepped flotation enhanced recovery method for magnetic separation concentrate. Background Art

[0002] As an important carrier of lithium resources, the efficient recovery of zinnwaldite is of great significance to the development of the new energy industry. High-slime-type zinnwaldite ore generally faces serious problems such as severe slime interference, low synergy efficiency of magnetic separation-flotation, difficulty in recovering fine and ultrafine particles, and insufficient process adaptability in the traditional flotation process. In the existing process, the fine granulation of ore easily leads to the slime covering the surface of target minerals, significantly reducing the flotation selectivity, forcing the excessive use of regulators to suppress slime interference. Although conventional magnetic separation can remove strongly magnetic minerals, the improper matching of magnetic field intensity causes the entrainment loss of weakly magnetic target minerals.

[0003] The recovery of fine and ultrafine particles in the traditional flotation process relies on a single collector, but the adsorption of slime leads to low reagent utilization rate and it is difficult to achieve effective recovery; although the stepwise flotation technology attempts to regulate in segments, due to problems such as unreasonable design of pulp concentration gradient, mismatch between reagent concentration and pulp viscosity, the re-election efficiency of tailings is limited.

[0004] In addition, traditional regulators have a single function and it is difficult to balance slime dispersion and gangue inhibition. The compound collector system is prone to fluctuations in separation effect due to insufficient emulsion stability in high-viscosity pulp. The above problems seriously restrict the efficient development of high-slime-type zinnwaldite resources, and there is an urgent need to achieve technological breakthroughs through optimization of magnetic separation parameters, innovation of differential speed flotation process, and synergistic mechanism of composite reagents. Summary of the Invention

[0005] Based on the differences in surface physical and chemical properties between zinnwaldite and gangue minerals, the present invention realizes the efficient recovery of high-slime-type zinnwaldite through a multi-stage coupling process of magnetic separation pretreatment - differential speed flotation - flocculation enhancement.

[0006] A method for recovering high-slime-type zinnwaldite based on stepwise flotation includes the following steps: 1) Magnetic separation: The raw ore is crushed to a particle size of -2 mm (i.e., less than 2 mm) and then undergoes wet low-intensity magnetic separation to obtain magnetic separation concentrate.

[0007] 2) Fast flotation: The magnetic separation concentrate is added with regulator GT-T and stirred for 3 - 5 min. A small amount of collector GT-B is added, and after stirring for 1 - 2 min, an open-circuit flotation of one roughing and one cleaning is carried out to obtain high-grade zinnwaldite concentrate 1.

[0008] 3) Re-election of tailings: Add water to the tailings to a concentration of 20 - 25%, add the regulator GT-T, and stir for 3 - 5 min. Add the collector GT-B and stir for 3 - 5 min. Add the flocculant GT-X, and after stirring for 3 - 5 min, conduct a closed-circuit flotation of one roughing, two cleanings, and two scavengings to obtain the lepidolite concentrate 2.

[0009] Preferably, in the high-slime-type lepidolite, the proportion of particles with a size of -20μm is 20 - 40%, for example, 30%.

[0010] Preferably, the magnetic field strength in step 1 is 0.1 - 0.3 T.

[0011] Preferably, the magnetic separation process flow is one roughing and one scavenging.

[0012] Preferably, the stirring speed for flotation is 1800 - 2200 rpm, specifically 1900 - 2100 rpm.

[0013] Preferably, the aeration time for flotation is 10 - 15 s.

[0014] Preferably, the regulator GT-T is compounded from starch, lignosulfonate, and sodium pyrophosphate in a mass ratio of 1:2:(2 - 4).

[0015] Preferably, the collector GT-B consists of dodecylamine and sodium petroleum sulfonate in a mass ratio of 2:(4 - 7).

[0016] Preferably, the flocculant GT-X consists of rosin amine polyoxyethylene ether and diesel in a mass ratio of 1:(1 - 1.5).

[0017] Preferably, the pulp concentration in step 2 is 30 - 40%.

[0018] Preferably, the concentration of the regulator GT-T in step 2 is 20 - 50 g / t.

[0019] Preferably, the concentration of the collector GT-B in step 2 is 30 - 70 g / t.

[0020] Preferably, the pulp concentration in step 3 is 20 - 25%.

[0021] Preferably, the concentration of the regulator GT-T in step 3 is 70 - 100 g / t.

[0022] Preferably, the concentration of the collector GT-B in step 3 is 150 - 200 g / t.

[0023] Preferably, the concentration of the flocculant GT-X in step 3 is 50 - 100 g / t.

[0024] Among them, the component ratios and concentrations of the regulator GT-T, the collector GT-B, and the flocculant GT-X are all determined through a large number of experiments.

[0025] In a specific embodiment, in step 2), water is added to the magnetic separation concentrate to a pulp concentration of 35%, 40 g / t of the regulator GT-T is added, and it is stirred for 3 minutes, then 70 g / t of the collector GT-B is added. After stirring for 2 minutes, one roughing and one cleaning are carried out to obtain high-grade lepidolite concentrate 1.

[0026] In step 3), water is added to the tailings to a concentration of 25%, 50 g / t of the regulator GT-T is added, and it is stirred for 3 minutes. 200 g / t of the collector GT-B is added and stirred for 3 minutes; 80 g / t of the flocculant GT-X is added and stirred for 3 minutes, and then a closed-circuit flotation of one roughing, two cleanings, and two scavengings is carried out. The obtained concentrates are combined to obtain lepidolite concentrate 2.

[0027] The core principle of the present invention includes: First, the lepidolite mineral is preferentially separated by a magnetic separation device with a weak magnetic field intensity, while avoiding excessive entrainment loss of strong magnetic separation. Secondly, in the two-stage flotation gradient regulation, in the first-stage rapid flotation under the conditions of high pulp concentration and low reagent dosage, the compound collector is preferentially adsorbed on the surface of coarse-grained lepidolite, and its hydrophobic chain quickly spreads to form a hydrophobic film to achieve rapid enrichment of high-grade concentrate. Strengthen the collision and aggregation of coarse-grained lepidolite, cooperate with the regulator with a specific proportion compound to achieve selective inhibition of gangue minerals and pulp dispersion, and then utilize the synergistic effect of cationic and anionic collectors to quickly form a stable foam layer under high-speed stirring conditions to preferentially recover coarse-grained lepidolite. In the re-selection stage of the tailings, by reducing the pulp concentration and increasing the reagent dosage, combined with the bridging effect of the flocculant, it promotes the formation of hydrophobic aggregates of fine-grained lepidolite and the oil phase, breaking through the Stokes floating rate limit of fine-grained minerals. The synergistic effect of each component in the regulator GT-T (starch inhibits gangue, lignosulfonate disperses pulp, sodium pyrophosphate chelates harmful ions) can synchronously optimize the pulp dispersibility and selective adsorption interface, while the cationic and anionic compound of the collector GT-B (electrostatic adsorption of dodecylamine and micelle solubilization of petroleum sulfonate) strengthens the hydrophobic stability of the surface of the target mineral. The flocculant GT-X realizes the hydrophobic aggregation of fine particles and efficient capture of bubbles through the directional adsorption of rosin amine polyoxyethylene ether and the physical bridging of diesel. Combining multiple cleanings and scavengings in the closed-circuit process, the efficient enrichment of fine-grained minerals is achieved.

[0028] In addition, in terms of reagents, the compounding of starch and lignosulfonate in the regulator can simultaneously inhibit silicate minerals such as quartz and disperse slime, and sodium pyrophosphate adjusts the pulp pH to a weakly alkaline environment; in the collector, dodecylamine adsorbs on the surface of lepidolite through electrostatic action, and sodium petroleum sulfonate enhances hydrophobic association through hydrogen bonding; in the flocculant, rosin amine polyoxyethylene ether provides hydrophobic groups, and diesel forms an oil bridge to promote the growth of flocs. The three form a three-stage enhanced separation system of "dispersion - collection - flocculation".

[0029] The present invention systematically solves the problem of separating highly slimey lepidolite through magnetic-floatation synergistic separation, adapting the pulp concentration and reagent gradient, and multi-component interface regulation technology. The process optimizes the pulp concentration, reagent system, and flow field conditions in stages, breaking through the technical bottleneck that traditional flotation is difficult to balance fine particle dispersion, selective collection, and floc stability.

[0030] The present invention realizes the rapid enrichment of coarse particles and the efficient recovery of fine particles through the magnetic separation - differential speed flotation process, significantly improving the separation efficiency and resource recovery rate of lepidolite. In addition, the dosage of the reagents in the present invention is low, reducing the beneficiation cost and environmental pollution. Detailed implementation manners

[0031] The technical solutions of the present invention will be further described in detail below in combination with specific embodiments.

[0032] Example 1 This example uses a high-slime type lepidolite ore in Hunan. The original ore has a Li2O grade of 0.30 and a slime content of 30% (-20 microns). The specific steps and reagent system are as follows: 1) Magnetic separation: The original ore is crushed to a particle size of -2 mm (i.e., less than 2 mm), and magnetite is removed by wet low-intensity magnetic separation (magnetic field intensity 0.2 T). After a one-roughing and one-scavenging process (the original ore enters the magnetic separator to obtain concentrate and tailings, and the tailings then enter the magnetic separator to obtain scavenger concentrate and tailings), the concentrates from the two times are combined to obtain the magnetic separation concentrate.

[0033] 2) Rapid flotation: Water is added to the magnetic separation concentrate to a pulp concentration of 35%, and 40 g / t of regulator GT-T is added and stirred for 3 min. A small amount of 70 g / t of collector GT-B is added, and after stirring for 2 min, an open-circuit flotation of one-roughing and one-cleaning (the original ore enters the flotation machine to obtain concentrate and tailings, and the concentrate then enters the flotation machine to obtain concentrate and tailings) is carried out. The concentrates from the two times are combined to obtain high-grade lepidolite concentrate 1.

[0034] 3) Re-election of tailings: Add water to the tailings to a concentration of 25%, add 50 g / t of regulator GT-T, and stir for 3 min. Add 200 g / t of collector GT-B and stir for 3 min. Add 80 g / t of flocculant GT-X, and after stirring for 3 min, conduct a closed-circuit flotation of one roughing, two cleanings, and two scavengings (the raw ore enters the flotation machine, the rough concentrate and tailings are obtained through roughing, the concentrate is further subjected to two cleanings to obtain the concentrate and middlings, the tailings enter the flotation machine for two scavengings to obtain the scavenged concentrate and tailings, and the middlings are in a closed-circuit cycle). The obtained concentrates are combined to obtain lepidolite concentrate 2.

[0035] Among them, the regulator GT-T is compounded from starch, lignosulfonate, and sodium pyrophosphate according to a mass ratio of 1:2:2.

[0036] The collector GT-B is composed of dodecylamine and sodium petroleum sulfonate in a mass ratio of 2:5.

[0037] The flocculant GT-X is composed of rosin amine polyoxyethylene ether and diesel in a mass ratio of 1:1.5.

[0038] The test results are shown in Table 1. It can be seen that the recovery rate of Li2O in lepidolite concentrate 1 is 68.27%, the recovery rate of Li2O in lepidolite concentrate 2 is 21.91%, the total recovery rate of Li2O is 90.17%, and the loss rate of Li2O in the tailings is only 9.83%. Among them, the recovery rate of Li2O / % = concentrate yield × grade / (raw ore yield × grade).

[0039] Table 1. Test results

[0040] Comparative Example 1 Taking a high-slime-type lepidolite ore in Hunan in Example 1 as the raw ore, using the same reagent system as in Example 1, the influence of the magnetic separation process on the flotation index of lepidolite was investigated.

[0041] The test results are shown in Table 2. It can be seen that without the magnetic separation operation, the recovery rate of Li2O in lepidolite concentrate 1 is only 32.60%, the recovery rate of Li2O in lepidolite concentrate 2 is 4.70%, the total recovery rate of Li2O is 37.30%, and the loss rate of Li2O in the tailings is 62.70%.

[0042] Table 2. Flotation test results without magnetic separation operation

[0043] Comparative Example 2 Taking a high-slime-type lepidolite ore in Hunan in Example 1 as the raw ore, using the same technological process as in Example 1, the influence of different reagent systems on the flotation index of lepidolite was investigated.

[0044] The test results are shown in Table 3. It can be seen that the effective enrichment of coarse and fine-grained lepidolite cannot be achieved by using collectors, regulators or inhibitors other than those of the present invention.

[0045] Table 3. Test Results of Reagent Types

[0046] Example 2 In this example, a high-slime-type lepidolite ore in Jiangxi (with a large slime content in the ore) was used. The grade of Li2O in the original ore was 0.25, and the slime content was 35% (-20 microns). The specific steps and reagent regime are as follows: 1) Magnetic separation: The original ore was crushed to a particle size of -2 mm, and magnetite was removed by wet low-intensity magnetic separation (magnetic field intensity: 0.2 T) to obtain magnetic separation concentrate.

[0047] 2) Rapid flotation: Water was added to the magnetic separation concentrate to a pulp concentration of 35%, 50 g / t of regulator GT-T was added, and it was stirred for 3 min. 80 g / t of collector GT-B was added, and after stirring for 2 min, an open-circuit flotation of one roughing and one cleaning was carried out to obtain high-grade iron-lithium mica concentrate 1.

[0048] 3) Tailings re-election: Water was added to the tailings to a concentration of 20%, 60 g / t of regulator GT-T was added, and it was stirred for 3 min. 200 g / t of collector GT-B was added, and it was stirred for 3 min. 80 g / t of flocculant GT-X was added, and after stirring for 3 min, a closed-circuit flotation of one roughing, two cleanings and two scavengings was carried out to obtain iron-lithium mica concentrate 2.

[0049] Among them, the regulator GT-T is compounded from starch, lignosulfonate and sodium pyrophosphate according to a mass ratio of 1:2:3.

[0050] The collector GT-B is composed of dodecylamine and sodium petroleum sulfonate according to a mass ratio of 2:6.

[0051] The flocculant GT-X is composed of rosin amine polyoxyethylene ether and diesel according to a mass ratio of 1:1.2.

[0052] The test results are shown in Table 4. It can be seen that the recovery rate of Li2O in the finally obtained iron-lithium mica concentrate 1 is 65.15%, the recovery rate of Li2O in the iron-lithium mica concentrate 2 is 21.28%, and the total recovery rate of Li2O is 86.43%. The loss rate of Li2O in the tailings is only 13.57%.

[0053] Table 4. Test Results 。

Claims

1. A method for recovering high-slime-type zinnwaldite based on step-by-step flotation, characterized in that It includes the following steps: 1) Magnetic separation: After the raw ore is crushed, magnetic separation is carried out to obtain magnetic separation concentrate; 2) Fast flotation: The magnetic separation concentrate is added with regulator GT-T composed of starch, lignosulfonate, and sodium pyrophosphate. After stirring, collector GT-B composed of dodecylamine and sodium petroleum sulfonate is added. After stirring, open-circuit flotation with one roughing and one cleaning is carried out to obtain high-grade lepidolite concentrate 1; 3) Tailings re-selection: The tailings are added with regulator GT-T. After stirring, collector GT-B is added. After stirring, flocculant GT-X composed of rosin amine polyoxyethylene ether and diesel is added. After stirring, closed-circuit flotation with one roughing, two cleanings, and two scavengings is carried out to obtain lepidolite concentrate 2.

2. The method for recycling high-mud-type zinnwaldite according to claim 1, wherein, In step 1), the raw ore is crushed to a particle size below 2 mm, and magnetic separation concentrate is obtained through wet weak magnetic separation. More specifically, the magnetic field intensity used is 0.1 - 0.3 T.

3. The method for recovering high-slime-type lepidolite as claimed in claim 1, wherein in step 1), the sample with a particle size below 20 microns in the high-slime-type lepidolite accounts for 20 - 40%, for example, 30%; the magnetic separation process flow is one roughing and one scavenging.

4. The method for recycling high-mud-type zinnwaldite as claimed in claim 1, wherein, In step 2), regulator GT-T is added and stirred for 3 - 5 min; collector GT-B is added, and after stirring for 1 - 2 min, open-circuit flotation with one roughing and one cleaning is carried out.

5. The method for recovering high-slime-type zinnwaldite according to claim 1, wherein In step 3), water is added to the tailings to a concentration of 20 - 25%, regulator GT-T is added and stirred for 3 - 5 min; collector GT-B is added and stirred for 3 - 5 min; flocculant GT-X is added and stirred for 3 - 5 min, and then closed-circuit flotation with one roughing, two cleanings, and two scavengings is carried out.

6. The method for recycling high-slime-type zinnwaldite according to claim 1, wherein The flotation stirring speed is 1800 - 2200 rpm, specifically 1900 - 2100 rpm, for example, 1990 rpm; the flotation aeration time is 10 - 15 s.

7. The method for recycling high-slime-type zinnwaldite as claimed in claim 1, wherein, The regulator GT-T is compounded from starch, lignosulfonate, and sodium pyrophosphate according to a mass ratio of 1:2:(2 - 4); the collector GT-B is composed of dodecylamine and sodium petroleum sulfonate with a mass ratio of 2:(4 - 7); the flocculant GT-X is composed of rosin amine polyoxyethylene ether and diesel according to a mass ratio of 1:(1 - 1.5).

8. The method for recycling high-slime-type zinnwaldite as claimed in claim 1, wherein, The pulp concentration in step 2) is 30 - 40%, the concentration of regulator GT-T is 20 - 50 g / t, and the concentration of collector GT-B is 30 - 70 g / t.

9. The method for recycling high-mud-type zinnwaldite as claimed in claim 1, wherein, The pulp concentration in step 3) is 20 - 25%; the concentration of regulator GT-T is 70 - 100 g / t; the concentration of collector GT-B is 150 - 200 g / t; the concentration of flocculant GT-X is 50 - 100 g / t.

10. The method for recovering high-slime-type lepidolite as claimed in claim 9, wherein ​ In step 3), water is added to the tailings to a concentration of 25%, 50 g / t of regulator GT-T is added, and the mixture is stirred for 3 min. 200 g / t of collector GT-B is added and stirred for 3 min; 80 g / t of flocculant GT-X is added, and after stirring for 3 min, closed-circuit flotation with one roughing, two cleanings, and two scavengings is carried out, and the obtained concentrates are combined to obtain lepidolite concentrate 2.

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