A process for removing impurities from medium and low grade collophanite

Abstract

The present invention provides a process for removing impurities from medium- and low-grade collophosphite, which first performs reverse flotation to remove magnesium after grinding, then performs gravity separation, and finally performs reverse flotation to remove silicon to obtain phosphate concentrate. After the powder ore is ground and heated, dolomite is preferentially removed by reverse flotation, reducing the content of fine and micro-grained gangue minerals entering gravity separation, improving the efficiency of spiral chute gravity separation to remove fine-grained iron-aluminum-silicate impurities, greatly reducing the content of fine-grained impurities entering the reverse flotation desiliconization process, greatly reducing the adverse effects of fine and micro-grained particles on reverse flotation desiliconization of cationic collectors, greatly exerting the desiliconization efficiency of cationic collectors, significantly improving flotation efficiency, and finally significantly reducing the content of aluminum, magnesium, and silicate in the phosphate concentrate.

Description

Technical Field

[0001] The invention belongs to the technical field of collophanite beneficiation, and in particular relates to a process for removing impurities from medium- and low-grade collophanite. Background Art

[0002] At present, the main methods for phosphate ore beneficiation include flotation, gravity separation, scrubbing and desludging, chemical leaching, photoelectric separation and combined separation. The most effective method for the separation of collophanite is flotation, but flotation still has problems such as large reagent consumption and high beneficiation cost. Gravity separation has low beneficiation cost and obvious enrichment effect. In order to make full use of the advantages of gravity separation and flotation technology, it is of great significance to study the gravity-flotation combined with efficient impurity removal technology for collophanite.

[0003] With the transformation and upgrading of phosphorus chemical enterprises, especially the increasing demand for superior and first-class phosphate concentrates of high-purity phosphoric acid (purified phosphoric acid), technical bottlenecks have become the main shortcoming restricting the sustained and healthy development of phosphorus chemical industry.

[0004] CN109453891A discloses a combined process of spiral chute refloatation of high-polyphosphite, which uses a spiral chute to re-select phosphate ore, and double reverse flotation to remove magnesium and silicon from the re-selected concentrate to obtain double reverse flotation phosphate concentrate. This process fails to maximize the technical advantages of the combined re-floatation and impurity removal process, reduce the impact of coarse and fine inclusions on re-selection and flotation, and cannot maximize the impurity removal efficiency and improve the concentrate quality. Summary of the invention

[0005] The invention provides a process for removing impurities from medium- and low-grade collophanite, which can significantly improve the ore dressing efficiency and obtain high-grade and high-quality phosphate concentrate.

[0006] The technical solution of the present invention is a process for removing impurities from medium and low-grade collophanite, comprising the following steps:

[0007] S1. Grind the raw collophanite ore until the phosphate rock -0.074mm accounts for 65-70wt%, then add water to make slurry, add it to the flotation tank, add anion collector to carry out magnesium removal operation; the flotation tank contains concentrate, the foam product is tailings, the tailings are scavenged, and the obtained middlings are combined with the concentrate in the flotation tank to form magnesium-removed concentrate;

[0008] S2, the de-magnesiumized concentrate is concentrated to 20-23wt%, and introduced into the first-level spiral chute for sorting to obtain high-concentration coarse-gravity concentrate and low-concentration fine-grained tailings; the fine-grained tailings are directly introduced into the second-level spiral chute for scavenging to obtain middlings and gravity-separated tailings, and the coarse-grained concentrate and middlings are combined into gravity-separated concentrate;

[0009] S3. The re-selected concentrate is diluted with water to 25-30wt%, and then put into the flotation tank. A cationic collector is added for desiliconization. The foam product is the desiliconized tailings, and the product in the flotation tank is the desiliconized concentrate product.

[0010] Furthermore, the P2O5 content in the collophanite ore is 21-26%; MgO≥1%, SiO2≥15%, and sesquioxide 8%≥R2O3(Al2O3+Fe2O3)≥3%.

[0011] Further, when water is added to the slurry in S1, the concentration is 20-25%.

[0012] Furthermore, the anionic collector in S1 is a composite anionic collector, which is specifically prepared by mixing cottonseed fatty acid and OP-10 in a mass ratio of 8-10:1.

[0013] Furthermore, the amount of anion collector added relative to the original ore is 0.25-0.5 Kg / t.

[0014] Furthermore, an inhibitor is added before the anion collector is added in S1, and the inhibitor is phosphoric acid, and the amount used relative to the dry weight of the original ore is 6.5-10.5 kg / t; when the tailings are scavenged, the inhibitor phosphoric acid is added at 1.5-2.5 kg / t.

[0015] Furthermore, the ratio of the pitch to the spiral diameter of the first-stage spiral chute in S2 is 0.45-0.6.

[0016] Furthermore, the ratio of the pitch to the spiral diameter of the Class II spiral chute selected for the S2 mid-sweep is 0.36-0.40.

[0017] Furthermore, the cationic collector in S3 is industrial-grade dodecylamine, and the amount added is 0.1-0.3 kg / t relative to the dry weight of the gravity concentrate.

[0018] Furthermore, phosphoric acid as an inhibitor is added before the cationic collector is added in S3, and the amount added is 2.5-3.5 kg / t relative to the dry weight of the gravity separation concentrate.

[0019] The present invention has the following beneficial effects:

[0020] (1) Compared with the conventional process of gravity separation followed by flotation, the process of the present invention adjusts the magnesium removal flotation step to before gravity separation, first performs reverse flotation magnesium removal, preferentially removes carbonate fine-grained gangue, reduces the content of fine-grained gangue minerals entering gravity separation, improves the efficiency of spiral chute gravity separation in removing fine-grained iron-aluminum-silicate impurities, greatly reduces the content of fine-grained impurities in gravity separation concentrate, and avoids the adverse effects of coarse and fine inclusions on reverse flotation desiliconization of cationic collectors, greatly exerts the desiliconization efficiency of cationic collectors, significantly improves flotation efficiency, and obtains high-grade high-quality phosphate concentrate with significantly reduced iron, aluminum, magnesium and silicate contents.

[0021] (2) The process of the present invention is conducive to overcoming the low temperature environment of 5-10°C. Magnesium removal flotation is directly carried out after grinding. Since the ore obtained in the grinding process has a certain temperature, magnesium removal reverse flotation is directly carried out, and the slurry temperature can reach above 10°C, which helps to improve the effect of magnesium removal flotation; if gravity separation is carried out first, it will cause hypothermia, and the temperature needs to be raised by external means during the later magnesium removal flotation, resulting in increased energy consumption.

[0022] (3) After being treated by the process of the present invention, the final phosphate concentrate is of good quality, with the characteristics of high P2O5≧32.5%, low MgO≦0.5%, low SiO2≦14%, and low sesquioxide R2O3 (Al2O3+Fe2O3)≦1.8%. BRIEF DESCRIPTION OF THE DRAWINGS

[0023] Figure 1 It is a process flow chart of the present invention. DETAILED DESCRIPTION

[0024] The embodiments of the present invention will be described in detail below in conjunction with examples, but those skilled in the art will understand that the following examples are only used to illustrate the present invention and should not be considered to limit the scope of the present invention. The pharmaceutical raw materials used in the examples of the present invention are all commercially available.

[0025] Example 1

[0026] In this embodiment, under a low temperature environment of 10 degrees Celsius, the raw ore P2O5 grade is 22.31%, the MgO content is 2.12%, the SiO2 content is 17.35%, and the sesquioxide R2O3 content is 6.25% The collophosphate ore is decontaminated, the grinding slurry temperature is 15 degrees Celsius, and the process flow is shown in FIG. Figure 1 , the specific steps are as follows:

[0027] (1) Reverse flotation magnesium removal operation

[0028] The raw mineral material ground to -0.074mm and accounting for 65wt% is prepared into a slurry concentration of 25wt%, and a demagnesium agent is added to the slurry to adjust the slurry, and demagnesium roughing is performed. The demagnesium roughing agent and its dosage are 6.5kg / t of phosphoric acid, and the anion collector (the mass ratio of cotton oil fatty acid and OP-10 surfactant is 8:1) is used in an amount of 0.25kg / t. The foam product obtained by the demagnesium roughing is fed into the scavenging operation, and the dosage of the demagnesium scavenging agent is 2kg / t of phosphoric acid. The foam product obtained by the reverse flotation scavenging is the demagnesium tailings;

[0029] (2) Removal of iron-aluminum silicate by gravity separation

[0030] The underflows of the single reverse flotation roughing and scavenging operations are combined and concentrated into a pulp concentration of 20wt%, and fed into a spiral chute with a pitch to spiral diameter ratio of 0.45 for separation to obtain a gravity separation coarse particle concentrate and a gravity separation fine particle tailings. The gravity separation fine particle tailings slurry is fed into a spiral chute with a pitch to spiral diameter ratio of 0.36 for scavenging to obtain gravity separation mides and gravity separation tailings. The gravity separation coarse particle concentrate and the gravity separation mides are combined to obtain a coarse phosphate concentrate;

[0031] (3) Reverse flotation desiliconization operation

[0032] The crude phosphate concentrate is prepared into a slurry concentration of 30wt% and fed into the desiliconization reverse flotation. The reverse flotation reagents and dosages are 2.5kg / t phosphoric acid and 0.2kg / t dodecylamine to obtain the final concentrate.

[0033] (4) Through the above-mentioned separation operations, a phosphate concentrate with a P2O5 grade of 32.6%, a sesquioxide R2O3 content of 1.15%, a SiO2 content of 9.49%, a MgO content of 0.44%, a yield of 53.36%, and a concentrate recovery rate of 77.98% was obtained.

[0034] Example 2

[0035] In this example, the raw ore P2O5 grade is 24.25%, MgO content is 3.1%, SiO2 content is 20.35%, and sesquioxide R2O3 content is 4.25% of the collophosphate ore is removed under a low temperature environment of 5 degrees Celsius. The pulp temperature after grinding is 11 degrees Celsius. The specific operation steps are as follows:

[0036] (1) Reverse flotation magnesium removal operation

[0037] The raw mineral material ground to -0.074mm and accounting for 67wt% is prepared into a slurry concentration of 25wt%, and a demagnesium agent is added to the slurry to adjust the slurry, and demagnesium roughing is performed. The demagnesium roughing agent and its dosage are 8kg / t of phosphoric acid, and the anion collector (the mass ratio of cotton oil fatty acid and OP-10 surfactant is 9:1) is used in an amount of 0.4kg / t. The foam product obtained by the demagnesium roughing is fed into the scavenging operation, and the dosage of the demagnesium scavenging agent is 2kg / t of phosphoric acid. The foam product obtained by the reverse flotation scavenging is the demagnesium tailings;

[0038] (2) Removal of iron-aluminum silicate by gravity separation

[0039] The underflows of the single reverse flotation roughing and scavenging operations are combined and concentrated into a pulp concentration of 20wt%, and fed into a spiral chute with a pitch to spiral diameter ratio of 0.6 for separation to obtain a gravity separation coarse particle concentrate and a gravity separation fine particle tailings. The gravity separation fine particle tailings slurry is fed into a spiral chute with a pitch to spiral diameter ratio of 0.36 for scavenging to obtain gravity separation mid-particles and gravity separation tailings. The gravity separation coarse particle concentrate and the gravity separation mid-particles are combined to obtain a coarse phosphate concentrate;

[0040] (3) Reverse flotation desiliconization operation

[0041] The crude phosphate concentrate is prepared into a slurry concentration of 30wt% and fed into the desiliconization reverse flotation. The reverse flotation reagents and dosages are 2.5kg / t phosphoric acid and 0.15kg / t dodecylamine to obtain the final concentrate.

[0042] (4) Through the above-mentioned separation operation, a phosphate concentrate with a P2O5 grade of 34%, a sesquioxide R2O3 content of 1.45%, a SiO2 content of 12.35%, a MgO content of 0.5%, a yield of 50%, and a concentrate recovery rate of 70.1% was obtained.

[0043] Example 3

[0044] In this example, under a low temperature environment of 8 degrees Celsius, the collophosphate ore with a P2O5 grade of 26%, a MgO content of 1.8%, a SiO2 content of 24%, and a sesquioxide R2O3 content of 6.8% was removed. The slurry temperature after grinding was 13 degrees Celsius. The specific operation steps are as follows:

[0045] (1) Reverse flotation magnesium removal operation

[0046] The raw mineral material ground to -0.074mm and accounting for 70wt% is prepared into a slurry concentration of 25wt%, and a demagnesium agent is added to the slurry to adjust the slurry, and demagnesium roughing is performed. The demagnesium roughing agent and its dosage are 7.5kg / t of phosphoric acid, and the anion collector (the mass ratio of cotton oil fatty acid and OP-10 surfactant is 10:1) is used in an amount of 0.3kg / t. The foam product obtained by the demagnesium roughing is fed into the scavenging operation, and the dosage of the demagnesium scavenging agent is 2kg / t of phosphoric acid. The foam product obtained by the reverse flotation scavenging is the demagnesium tailings;

[0047] (2) Removal of iron-aluminum silicate by gravity separation

[0048] The underflows of the single reverse flotation roughing and scavenging operations are combined and concentrated into a pulp concentration of 23wt%, and fed into a spiral chute with a pitch to spiral diameter ratio of 0.45 for separation to obtain a gravity separation coarse particle concentrate and a gravity separation fine particle tailings. The gravity separation fine particle tailings slurry is fed into a spiral chute with a pitch to spiral diameter ratio of 0.4 for scavenging to obtain gravity separation mides and gravity separation tailings. The gravity separation coarse particle concentrate and the gravity separation mides are combined to obtain a coarse phosphate concentrate.

[0049] (3) Reverse flotation desiliconization operation

[0050] The crude phosphate concentrate is prepared into a slurry concentration of 30wt% and fed into the desiliconization reverse flotation. The reverse flotation reagents and dosages are 3kg / t of phosphoric acid and 0.3kg / t of dodecylamine to obtain the final concentrate.

[0051] (4) Through the above-mentioned separation operation, a phosphate concentrate with a P2O5 grade of 32.5%, a sesquioxide R2O3 content of 1.7%, a SiO2 content of 14%, a MgO content of 0.48%, a yield of 55%, and a concentrate recovery rate of 68.75% was obtained.

[0052] In order to fully understand the technical advantages of the present invention in the removal of impurities from medium and low-grade collophanite, we selected the phosphate ore of Example 1 and conducted a gravity-flotation joint selection test.

[0053] Comparative Example 1: In this case, except that the slurry temperature of the flotation demagnesium operation was maintained at 10 degrees Celsius, the other operating conditions were exactly the same as those of Example 1.

[0054] Through the above-mentioned separation operations, the phosphate concentrate obtained has a P2O5 grade of 31.44%, a sesquioxide R2O3 content of 2.3%, a SiO2 content of 11.7%, a MgO content of 1.25%, a yield of 50.32%, and a concentrate recovery rate of 70.91%.

[0055] Comparative Example 2: In this case, gravity selection is first performed, i.e. grinding to -0.074 mm The raw ore material accounting for 65wt% is prepared into a slurry concentration of 20wt%, and is fed into a spiral chute with a pitch to spiral diameter ratio of 0.45 for sorting to obtain coarse-gravity separation concentrate and fine-gravity separation tailings. The fine-grained tailings slurry of gravity separation is fed into a spiral chute with a pitch to spiral diameter ratio of 0.36 for scavenging to obtain gravity separation medium and gravity separation tailings. The coarse-gravity separation concentrate is combined with the gravity separation medium to obtain a coarse phosphate concentrate; the coarse phosphate concentrate is prepared into a slurry concentration of 25wt%, and after the temperature is stabilized to 15 degrees Celsius by a heating rod, a demagnesiumizing agent is added to the slurry to adjust the slurry, and a demagnesiumizing roughing is carried out. The foam product obtained from the demagnesiumizing roughing is fed into a scavenging operation, and the foam product obtained from the reverse flotation scavenging is the demagnesiumizing tailings. The underflows of the reverse flotation roughing and scavenging operations are combined to form a slurry concentration of 30wt%, and fed into a desiliconization reverse flotation to obtain a final concentrate. The operating conditions for magnesium removal roughing and silicon removal reverse flotation are exactly the same as those in Example 1.

[0056] Through the above-mentioned separation operations, the phosphate concentrate obtained has a P2O5 grade of 30.88%, a sesquioxide R2O3 content of 2.8%, a SiO2 content of 13.45%, a MgO content of 0.5%, a yield of 53.25%, and a concentrate recovery rate of 73.71%.

[0057] Comparative Example 3: The operating conditions in this case are basically the same as those in Comparative Example 2, with the only difference being that no temperature increase measures are taken before the demagnesium operation.

[0058] Through the above-mentioned separation operations, the phosphate concentrate obtained has a P2O5 grade of 30.13%, a sesquioxide R2O3 content of 3.4%, a SiO2 content of 14.25%, a MgO content of 1.43%, a yield of 55%, and a concentrate recovery rate of 74.28%.

[0059] Examples 1-3 show that the process can overcome the low temperature of 5-10 degrees Celsius. By using grinding to produce warm slurry above 10 degrees Celsius, a suitable flotation environment is created for fatty acid demagnesium agents, so that reverse flotation can remove fine-grained dolomite as much as possible. Then, by utilizing the difference in specific gravity between apatite and iron-aluminum impurities, the physical property difference of apatite (3.1-3.2) being larger and iron-aluminum silicate (2.4-2.7) being smaller, the magnesium concentrate is re-selected through a spiral chute, and the coarse-grained useful mineral apatite is enriched into the re-selection concentrate, and the fine-grained iron-aluminum impurities are enriched into the re-selection tailings. For the re-selection concentrate with less impurity content, relatively low-cost reverse flotation desiliconization is used to obtain better indicators. Example 1 and Comparative Example 1, and Comparative Example 2 and Comparative Example 3 show that at low temperatures below 10 degrees Celsius, the flotation performance of fatty acid reagents will be inhibited. Example 1 and Comparative Example 2 show that the present process has more advantages than the prior gravity separation process, can avoid the adverse effects of coarse and fine inclusions of minerals in the flotation slurry on flotation, has good separation effect, can effectively reduce impurities such as magnesium oxide, iron, and aluminum silicate in the concentrate, and has low mineral processing costs.

[0060] The above-mentioned embodiments are only preferred technical solutions of the present invention and should not be regarded as limitations of the present invention. The embodiments and features in the embodiments of the present application can be arbitrarily combined with each other without conflict. The protection scope of the present invention shall be the technical solutions recorded in the claims, including the equivalent replacement solutions of the technical features in the technical solutions recorded in the claims. That is, equivalent replacement improvements within this scope are also within the protection scope of the present invention.

Claims

1. A process for removing impurities from medium and low grade collophanite, characterized in that , including the following steps: S1. Grind the collophanite ore until the phosphate rock -0.074mm accounts for 65-70wt%, then add water to make slurry, add it to the flotation tank, and add anion collector to remove magnesium; the flotation tank contains concentrate, the foam product is tailings, the tailings are scavenged, and the obtained middlings are combined with the concentrate in the flotation tank to form a demagnesified concentrate; the anion collector is a composite anion collector, which is made of cotton oil fatty acid and OP-10 mixed in a mass ratio of 8-10:1; S2, the de-magnesiumized concentrate is concentrated to 20-23wt%, and introduced into the first-level spiral chute for sorting to obtain high-concentration coarse-gravity concentrate and low-concentration fine-grained tailings; the fine-grained tailings are directly introduced into the second-level spiral chute for scavenging to obtain middlings and gravity-separated tailings, and the coarse-grained concentrate and middlings are combined into gravity-separated concentrate; S3. The gravity concentrate is diluted with water to 25-30wt%, and then fed into the flotation tank. A cationic collector is added for desiliconization. The foam product is the desiliconized tailings. The cationic collector is industrial-grade dodecylamine. The amount added is 0.1-0.3kg / t relative to the dry weight of the gravity concentrate. The product in the flotation tank is the desiliconized concentrate product.

2. The process according to claim 1, characterized in that: The P2O5 content in the collophanite ore is 21-26%, MgO≥1%, SiO2≥15%, and sesquioxide 8%≥R2O3(Al2O3+Fe2O3)≥3%.

3. The process according to claim 1, characterized in that: When water is added to the slurry in S1, the concentration is 20-25%.

4. The process according to claim 1, characterized in that: The amount of anionic collector added relative to the dry weight of the original ore is 0.25-0.5 Kg / t.

5. The process according to claim 1, characterized in that: In S1, an inhibitor is added before the anion collector is added. The inhibitor is phosphoric acid, and the amount used relative to the dry weight of the original ore is 6.5-10.5 kg / t; when the tailings are swept, the inhibitor phosphoric acid is added at 1.5-2.5 kg / t.

6. The process according to claim 1, characterized in that: The ratio of the pitch to the spiral diameter of the first-level spiral chute in S2 is 0.45-0.

6.

7. The process according to claim 1, characterized in that: The S2 mid-sweep uses a Class II spiral chute with a pitch to spiral diameter ratio of 0.36-0.

40.

8. The process according to claim 1, characterized in that: Phosphoric acid as an inhibitor is added before the addition of the cationic collector in S3, and the amount added is 2.5-3.5 kg / t relative to the dry weight of the gravity separation concentrate.

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