CN104174484A - Desiliconization processing method for bauxite flotation tailings - Google Patents
Desiliconization processing method for bauxite flotation tailings Download PDFInfo
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- CN104174484A CN104174484A CN201410331868.XA CN201410331868A CN104174484A CN 104174484 A CN104174484 A CN 104174484A CN 201410331868 A CN201410331868 A CN 201410331868A CN 104174484 A CN104174484 A CN 104174484A
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Abstract
The invention relates to a desiliconization processing method for bauxite flotation tailings. The method comprises the following steps: performing graded concentration on bauxite flotation tailings slurry with the concentration of 10-20 percent to separate to-be-floated coarse-fraction component slurry; adding a collecting agent into the to-be-floated coarse-fraction component slurry; performing flotation to obtain rough concentrate and waste tailings; cleaning and grading the obtained rough concentrate to obtain concentrate. Due to the adoption of the method, waste bauxite flotation tailings serve as raw materials, the concentrate of which the A/S (alumina-silica ratio) exceeds 6 is recovered, the A/S of the discharged tailings of bauxite is about 1.20, the ore consumption is reduced, and the ore dressing recovery percentage is improved.
Description
Technical field
The present invention relates to a kind of processing method of alumyte flotation tailings desiliconization, belong to alumyte flotation tailings machining field.
Background technology
In fact bauxite refer to industrial can utilize, and take the general designation that gibbsite, two diaspores, boehmite or diaspore be the ore that essential mineral was formed.The application of bauxite has metal and nonmetal two aspects, is the optimum feed stock of producing metallic aluminium, is also topmost application, and its consumption accounts for the more than 90% of world's bauxite total output.Though bauxite is little at the consumption proportion aspect nonmetal, purposes is very extensive.
Mineral dressing and silicon removal of bauxite, has found an effective approach for the reasonable utilization of China's middle-low bauxite resource, has effectively improved the utilization rate of bauxite resource.
At present, bauxite direct-flotation desiliconisation has been realized industrial applications in alumina producing, alumina silica ratio (A/S) adopts selective milling-aggregative flotation desilication process at 5～6 bauxite, has obtained that concentrate A/S is greater than 10, mine tailing A/S is the floatation indicators of 1.5 left and right.
But along with the sharply decline of bauxite head grade, the degree of dissociation of ore is had relatively high expectations, grindability and the washability variation of while ore, alumyte flotation tailings Al
2o
3content is up to more than 42%, and mine tailing A/S is more than 1.60, and recovery rate in ore-dressing is less than 75%.
The alumina industry of China at least produces the alumyte flotation tailings of 1,000 ten thousand tons every year according to statistics, because mine tailing is loose, needs and cement, flyash mix ability condensation cure, therefore after tailing solidifying, produced more solid waste, the stacking of these solid waste need to take a large amount of soils, and the mine tailing that heap puts away also has dam break risk.
Summary of the invention
Technical problem to be solved by this invention is to provide a kind of processing method of alumyte flotation tailings desiliconization, can reduce ore deposit consume by method of the present invention, improves recovery rate in ore-dressing, has solved deficiency of the prior art.
The technical scheme that the present invention solves the problems of the technologies described above is as follows: a kind of method that reduces alumyte flotation tailings alumina silica ratio, comprise the following steps: the flotation tailing slurries that are 10%～20% by concentration (mass fraction) adopt the concentrated operation of classification to separate coarse fraction component slurries to be selected, in coarse fraction component slurries to be selected, add collecting agent, after flotation, obtain again rough concentrate and abandon mine tailing, the rough concentrate obtaining is obtained to concentrate after cleaning classification.
On the basis of technique scheme, the present invention can also do following improvement.
Further, the concentrated operation of described classification adopts hydrocyclone, and the component to be selected separating is+0.045mm coarse fraction component slurries that the solid content of described coarse fraction component slurries is controlled at 25%～50%.
Hydrocyclone is to utilize centrifugal force to accelerate the classifying equipoment of ore particle sedimentation, it needs pressure to ore deposit, therefore consumption of power is large, but floor space is little, low price, treating capacity is large, classification efficiency is high, can obtain very thin overflow product, be used for the classifying equipoment in second segment closed circuit grinding.
Further, that described collecting agent is oleic acid is saponified, aphthenic acids, benzene first hydroximic acid in mass ratio 20:1:1 ratio after mixing, obtain, the addition of described collecting agent is coarse fraction component slurries per ton 200～2000g.
Collecting agent has physical absorption, chemisorbed and surface chemical reaction in the effect of mineral surfaces.The absorption of collecting agent and mineral floating behavior have substantial connection.In certain collecting agent concentration range, along with drug concentration improves, adsorbance increases, and flotation recovery rate significantly rises; Concentration reach mutually on duty after, the amplitude that the rate of recovery improves with concentration and adsorbance diminishes; During collecting agent excessive concentration, adsorbance also can continue to increase, but flotation recovery rate no longer raises, decline even on the contrary.Therefore, in floatation process, to correctly grasp the consumption of collecting agent, to obtain optimum efficiency.
Oleic acid is a kind of undersaturated aliphatic acid, is formed in the bodies of aminal and plant, and chemical formula C
18h
34o
2(or CH
3(CH
2)
7cH=CH (CH
2)
7cOOH), the addition of oleic acid hydrogenation is obtained to stearic acid, two key transisomers of oleic acid are called elaidic acid.Oleic acid is added to alkali and carry out saponification, alkali generally adopts NaOH or potassium hydroxide, and obtaining the saponified of oleic acid is sodium oleate soap or oleic acid potash soap.Benzene first hydroximic acid, Chinese another name: benzo hydroxamic acid, molecular formula: C7H7NO2, molecular weight: 137.14, physicochemical properties: fusing point 126-130 ℃, water-soluble 22g/L (6 ℃).Aphthenic acids, molecular formula: C
7h
10o
2molecular weight: 126.15, be generally dark-brown oily liquids, after refining, it is transparent faint yellow or orange liquid, be commonly used for raw material and the coating drier of producing naphthenate, also anticorrosion for timber, cable, has spread effect to eyes, respiratory tract and skin, just in case contact eyes, are used immediately a large amount of clear water to rinse and send doctor diagnosis and treatment.
Further, described rough concentrate cleans classification and adopts concentrator.
Concentrator is the solid-liquid separating equipment based on gravitational settling effect, is generally by concrete, timber or metal welding fishplate bar and builds up the band cone cylindrical shape shallow slot at the end as structural material.The underflow ore pulp that concentrator can be 45%～55% by gravitational settling simmer down to solid content by the ore pulp containing being heavily admittedly 10%～20%, by the effect that is installed on the rake of microrunning in concentrator (1/3～1/5r/min), the underflow ore pulp of thickening is drawn off by the underflow opening bottom concentrator.Concentrator top produces more peace and quiet clarified solution (overflow), by the annular chute at top, is discharged.Concentrator divides and mainly contains three kinds by its kind of drive, and wherein first two is more common: 1. center driving, this type of concentrator diameter is less conventionally, generally at 24 meters with interior in the majority; 2. periphery running roller driving type, more common big-and-middle-sized concentrator, because it is gained the name by the transmission of transmission dolly, diameter is conventionally 53 meters of left and right, also has 100 meters; 3. periphery tooth bar driving type, this kind of base diameter be more than 53 meters, but now used less.Concentrator is widely used in the Workplace that hydrometallurgy, ore dressing plant, chemical plant etc. need solid-liquid concentration and separation, its diameter 3～100m, the degree of depth 2～4m.
The invention has the beneficial effects as follows: adopt method of the present invention, using discarded alumyte flotation tailings as raw material, reclaiming product is that A/S reaches more than 6 concentrate, and make the outer last person ore deposit A/S of bauxite in 1.20 left and right, reduced ore deposit consumption, improved recovery rate in ore-dressing.Adopt in addition method of the present invention, can from the lower bauxite of A/S, floatingly select the concentrate that A/S is 7 left and right, can reduce the minimum use alumina silica ratio of bauxite, expanded the source that produces alumina raw material.
China's bauxite resource faces exhaustion at present, uses the method, can optimize China's bauxite resource preparation, increases the service efficiency of bauxite; And use method of the present invention, and greatly reducing outer last person ore deposit amount, the corresponding discarded tailing that reduced is stacked land area and the mine tailing of occupying and is stacked the environmental pollution causing.
Accompanying drawing explanation
Fig. 1 is process flow diagram of the present invention.
The specific embodiment
Below in conjunction with accompanying drawing, principle of the present invention and feature are described, example, only for explaining the present invention, is not intended to limit scope of the present invention.
As shown in Figure 1, a kind of method that reduces alumyte flotation tailings alumina silica ratio, comprise the following steps: the flotation tailing slurries that are 10%～20% by concentration (mass fraction) adopt the concentrated operation of classification to sub-elect to be selected+0.045mm coarse fraction component slurries, its solid content is 25%～50%, send into flowsheet of slurry agitation groove, to the collecting agent that adds 200～2000g/ ton-coarse fraction component in coarse fraction component slurries to be selected, after being uniformly mixed, after flotation, obtain again rough concentrate and abandon mine tailing, the rough concentrate obtaining is cleaned through cleaning waterpower, after concentrator separation, underflow is concentrate, concentrator overflow, hydrocyclone overflow, abandoning mine tailing send together and abandons mine tailing deep-cone settling groove and carry out Separation of Solid and Liquid.
Described collecting agent be oleic acid saponified, aphthenic acids, benzene first hydroximic acid in mass ratio 20:1:1 ratio after mixing, obtain.
Below in conjunction with embodiment, the present invention is further illustrated.
Embodiment 1:
Adopt hydrocyclone to carry out classification in concentration 10% alumyte flotation tailings slurries concentrated, obtain solid content and be 25%+0.045mm coarse fraction component slurries, send into flowsheet of slurry agitation groove, collecting agent simultaneously that add on request 2000g/ ton-coarse fraction component, described collecting agent is sodium oleate soap, aphthenic acids, benzene first hydroximic acid in mass ratio 20:1:1 ratio obtains after mixing, after being uniformly mixed, send to and carry out flotation operation, through flotation operation, obtain rough concentrate and abandon mine tailing, rough concentrate cleans through waterpower, after concentrator separation, underflow is concentrate, concentrator overflow, hydrocyclone overflow, abandoning mine tailing send together and abandons mine tailing deep-cone settling groove and carry out Separation of Solid and Liquid.
Mineral processing index: concentrate A/S is 6.36; Abandoning mine tailing A/S is 1.21.
Embodiment 2:
Adopt hydrocyclone to carry out classification in concentration 15% alumyte flotation tailings slurries concentrated, obtain solid content and be 40%+0.045mm coarse fraction component slurries, send into flowsheet of slurry agitation groove, collecting agent simultaneously that add on request 1000g/ ton-coarse fraction component, described collecting agent is oleic acid potash soap, aphthenic acids, benzene first hydroximic acid in mass ratio 20:1:1 ratio obtains after mixing, after being uniformly mixed, send to and carry out flotation operation, through flotation operation, obtain rough concentrate and abandon mine tailing, rough concentrate cleans through waterpower, after concentrator separation, underflow is concentrate, concentrator overflow, hydrocyclone overflow, abandoning mine tailing send together and abandons mine tailing deep-cone settling groove and carry out Separation of Solid and Liquid.
Mineral processing index: concentrate A/S is 7.49; Abandoning mine tailing A/S is 1.15.
Embodiment 3:
Adopt hydrocyclone to carry out classification in concentration 20% alumyte flotation tailings slurries concentrated, obtain solid content and be 50%+0.045mm coarse fraction component slurries, send into flowsheet of slurry agitation groove, collecting agent simultaneously that add on request 200g/ ton-coarse fraction component, described collecting agent is sodium oleate soap, aphthenic acids, benzene first hydroximic acid in mass ratio 20:1:1 ratio obtains after mixing, after being uniformly mixed, send to and carry out flotation operation, through flotation operation, obtain rough concentrate and abandon mine tailing, rough concentrate cleans through waterpower, after concentrator separation, underflow is concentrate, concentrator overflow, hydrocyclone overflow, abandoning mine tailing send together and abandons mine tailing deep-cone settling groove and carry out Separation of Solid and Liquid.
Mineral processing index: concentrate A/S is 7.01; Abandoning mine tailing A/S is 1.18.
As seen from the above-described embodiment, adopt method of the present invention, can be using discarded alumyte flotation tailings as raw material, reclaiming product is that A/S reaches more than 6 concentrate, and makes the outer last person ore deposit A/S of bauxite in 1.20 left and right, has reduced ore deposit consumption, has improved recovery rate in ore-dressing.
The foregoing is only preferred embodiment of the present invention, in order to limit the present invention, within the spirit and principles in the present invention not all, any modification of doing, be equal to replacement, improvement etc., within all should being included in protection scope of the present invention.
Claims (4)
1. the processing method of an alumyte flotation tailings desiliconization, it is characterized in that, comprise the following steps: the alumyte flotation tailings slurries that are 10%～20% by concentration adopt the concentrated operation of classification to separate coarse fraction component slurries to be selected, in coarse fraction component slurries to be selected, add collecting agent, after flotation, obtain again rough concentrate and abandon mine tailing, the rough concentrate obtaining is obtained to concentrate after cleaning classification.
2. the processing method of alumyte flotation tailings desiliconization according to claim 1, it is characterized in that, the concentrated operation of described classification adopts hydrocyclone, and the component to be selected separating is+0.045mm coarse fraction component slurries that the solid content of described coarse fraction component slurries is controlled at 25%～50%.
3. the processing method of alumyte flotation tailings desiliconization according to claim 1, it is characterized in that, described collecting agent be oleic acid saponified, aphthenic acids, benzene first hydroximic acid in mass ratio 20:1:1 ratio after mixing, obtain, the addition of described collecting agent is coarse fraction component slurries per ton 200～2000g.
4. according to the processing method of alumyte flotation tailings desiliconization described in claims 1 to 3 any one, it is characterized in that, described rough concentrate cleans classification and adopts concentrator.
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Cited By (3)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN105344487A (en) * | 2015-11-19 | 2016-02-24 | 中国地质科学院郑州矿产综合利用研究所 | Regulator for bauxite direct flotation and using method thereof |
| CN107051711A (en) * | 2017-04-10 | 2017-08-18 | 中国铝业股份有限公司 | A kind of method that mineral processing tailing of bauxite is selected again |
| CN111282708A (en) * | 2020-02-14 | 2020-06-16 | 中国铝业股份有限公司 | Method for graded desulfurization, decarburization and desilication of acidified high-sulfur bauxite |
Citations (5)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN1324696A (en) * | 2000-05-19 | 2001-12-05 | 中南工业大学 | Bauxite dressing method |
| CN1369328A (en) * | 2001-02-13 | 2002-09-18 | 中南大学 | Ore dressing process for desiliconizing bauxite |
| CN1389300A (en) * | 2001-06-04 | 2003-01-08 | 中南大学 | Ore dressing and desiliconizing process of bauxite |
| CN1403205A (en) * | 2001-09-11 | 2003-03-19 | 北京矿冶研究总院 | Bauxite dressing method |
| CN101176859A (en) * | 2007-12-17 | 2008-05-14 | 中国铝业股份有限公司 | Method for ore dressing and desilicating mixed type bauxite |
-
2014
- 2014-07-14 CN CN201410331868.XA patent/CN104174484A/en active Pending
Patent Citations (5)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN1324696A (en) * | 2000-05-19 | 2001-12-05 | 中南工业大学 | Bauxite dressing method |
| CN1369328A (en) * | 2001-02-13 | 2002-09-18 | 中南大学 | Ore dressing process for desiliconizing bauxite |
| CN1389300A (en) * | 2001-06-04 | 2003-01-08 | 中南大学 | Ore dressing and desiliconizing process of bauxite |
| CN1403205A (en) * | 2001-09-11 | 2003-03-19 | 北京矿冶研究总院 | Bauxite dressing method |
| CN101176859A (en) * | 2007-12-17 | 2008-05-14 | 中国铝业股份有限公司 | Method for ore dressing and desilicating mixed type bauxite |
Cited By (6)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN105344487A (en) * | 2015-11-19 | 2016-02-24 | 中国地质科学院郑州矿产综合利用研究所 | Regulator for bauxite direct flotation and using method thereof |
| CN105344487B (en) * | 2015-11-19 | 2018-05-08 | 中国地质科学院郑州矿产综合利用研究所 | Regulator for bauxite direct flotation and using method thereof |
| CN107051711A (en) * | 2017-04-10 | 2017-08-18 | 中国铝业股份有限公司 | A kind of method that mineral processing tailing of bauxite is selected again |
| CN107051711B (en) * | 2017-04-10 | 2019-09-13 | 中国铝业股份有限公司 | A kind of method that mineral processing tailing of bauxite selects again |
| CN111282708A (en) * | 2020-02-14 | 2020-06-16 | 中国铝业股份有限公司 | Method for graded desulfurization, decarburization and desilication of acidified high-sulfur bauxite |
| CN111282708B (en) * | 2020-02-14 | 2021-07-20 | 中国铝业股份有限公司 | Method for graded desulfurization, decarburization and desilication of acidified high-sulfur bauxite |
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