CN104911331A - Method for separating vanadium-titanium magnetite concentrate by calcining, oxidation alkali leaching, pickling and magnetic reseparation - Google Patents

Method for separating vanadium-titanium magnetite concentrate by calcining, oxidation alkali leaching, pickling and magnetic reseparation Download PDF

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Publication number
CN104911331A
CN104911331A CN201510320887.7A CN201510320887A CN104911331A CN 104911331 A CN104911331 A CN 104911331A CN 201510320887 A CN201510320887 A CN 201510320887A CN 104911331 A CN104911331 A CN 104911331A
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magnetic
concentrate
content
pickling
alkali
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郭客
陈巍
王忠红
刘晓明
鞠洪钢
赵亮
全名巍
宋仁峰
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Angang Group Mining Co Ltd
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Angang Group Mining Co Ltd
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Abstract

The invention relates to a method for separating vanadium-titanium magnetite concentrate by calcining, oxidation alkali leaching, pickling and magnetic reseparation, which comprises the following steps: adding CaO into vanadium-titanium magnetite concentrate according to the weight ratio of 1:(0.1-0.2), calcining at 800-1400 DEG C, putting the calcining product A in a 5-52 wt% alkali solution, adding an oxidizer, carrying out alkali leaching reaction at 220-330 DEG C for 0.5-2.0 hours to obtain a filtrate and an alkali leaching filter cake B, and carrying out pickling and magnetic reseparation on the filter cake B. The introduced O2 or H2O2 in the oxidation alkali leaching process oxidates S-containing compounds, thereby accelerating the reaction, lowering the reaction temperature and shortening the reaction time. The iron concentrate with the TFe content of 66-70% and the titanium concentrate with the TiO2 content of 70-82% are respectively separated, thereby lowering the contents of TiO2, S, Si, Al and other impurities entering the blast furnace, enhancing the utilization coefficient of the blast furnace and also enhancing the comprehensive utilization ratio of the titanium resources.

Description

Utilize calcining, alkali oxide leaching, pickling and magnetic to weigh and select v-ti magnetite concentrate method again
Technical field
The present invention relates to a kind of ore-dressing technique of v-ti magnetite concentrate, particularly relate to and a kind of utilize calcining, alkali oxide leaching, pickling and magnetic is heavy selects v-ti magnetite concentrate method again.
Background technology
Vanadium titano-magnetite is a kind of complex ore of multiple metallic element, is based on the magnetite of the symbiosis of iron content, vanadium, titanium.And one of v-ti magnetite concentrate product that to be vanadium titano-magnetite obtain through ore dressing, wherein vanadium is composed with isomorph and is stored in titanomagnetite, displacement ferric ion.Titanomagnetite is oikocryst mineral (Fe 3o 4) and chadacryst ore deposit [ulvite 2FeOTiO 2, ilmenite FeOTiO 2, aluminum-spinel (Mg, Fe) (Al, Fe) 2o 4] complex body that formed.Such as, Chinese Panzhihua Region Midi Concentrator v-ti magnetite green ore and the chemistry of the v-ti magnetite concentrate after selecting iron multielement analysis the results are shown in Table 1, and v-ti magnetite green ore and vanadium titano-magnetite concentrate material phase analysis result are respectively in table 2 and table 3.
Table 1 Chinese Panzhihua Region Midi Concentrator raw ore and v-ti magnetite concentrate chemistry multielement analysis result
Element TFe FeO mFe S Fe 2O 3 TiO 2 V 2O 5
Raw ore 29.53 21.36 20.20 0.631 17.70 10.54 0.27 8
Concentrate 54.01 32.42 51.16 0.574 40.97 12.67 0.61
Element SiO 2 Al 2O 3 CaO MgO Co P As
Raw ore 22.80 7.65 6.36 7.23 0.02 0.015 <0.01
Concentrate 3.21 3.30 0.98 2.90 0.02 0.008 <0.010
Table 2 Chinese Panzhihua Region Midi Concentrator v-ti magnetite green ore titanium, iron chemical phase analysis result
Table 3 Chinese Panzhihua Region Midi Concentrator vanadium titano-magnetite concentrate titanium, iron chemical phase analysis result
Vanadium titano-magnetite aboundresources in the world, whole world reserves reach more than 40,000,000,000 tons, and reserves in China reaches 98.3 hundred million tons.In v-ti magnetite ore, iron is mainly composed and is stored in titanomagnetite, the TiO in ore 2main tax is stored in granular ilmenite and titanomagnetite.Generally, the titanium of about 57% is composed and is stored in titanomagnetite (mFeTiO 3nFe 3o 4) in, the titanium of about 40% is composed and is stored in ilmenite (FeTiO 3) in, because vanadium titano-magnetite ore composition is complicated, character is special, and thus the comprehensive utilization of this kind of ore is the international a great problem always thoroughly do not solved.This occurrence characteristics of vanadium titano-magnetite mineral determines the effective separation adopting physical concentration method cannot realize titanium, iron from the source of ore, cause v-ti magnetite ore after physical concentration, iron concentrate grade low (TFe<55%), the titanium in iron ore concentrate enters blast furnace slag (TiO completely at iron manufacturing process 2content reaches more than 22%) form vitreum, TiO 2lose activity and cannot economic recovery, meanwhile, titanium recovery rate is low only has 18%.Therefore sort titanium iron ore by the beneficiation method of physics and greatly reduce the value that titanium and iron utilizes separately.
China is first country comprehensively extracting iron, vanadium, titanium with technical scale from complicated vanadium titano-magnetite in the world, but due to general physical method fundamentally can not change iron, the tax of the fine and close symbiosis of titanium deposits characteristic, therefore, adopt the physical concentration methods such as common gravity separation method, magnetic method, flotation process to carry out titanium, iron is separated, efficiency is low, is difficult to select of high grade and the ilmenite concentrate that impurity is few or iron ore concentrate; Meanwhile, TiO 2organic efficiency is not high, v-ti magnetite green ore after Mineral separation, the TiO of about 54% 2enter iron ore concentrate, these TiO 2after blast-furnace smelting, almost all enter slag phase, form TiO 2the blast furnace slag of content 20 ~ 24%; In addition, because the foreign matter contents such as S, Si, the Al in iron ore concentrate are also too high, above-mentioned reason not only causes that steelmaking furnace utilization coefficient is low, energy consumption is large, titanium resource waste, and amount of slag is large, environmental pollution is serious.
CN2011100879566 discloses " a kind of beneficiation method of ilmenite ", is the magnetic separation after ore grinding, alkaline pretreatment, filtration, again ore grinding of v-ti magnetite green ore is obtained the method for ilmenite concentrate and iron ore concentrate.The method is by iron content 32.16% with containing TiO 2the v-ti magnetite green ore of 12.11%, by magnetic separation process after ore grinding, alkaline pretreatment, filtration, again ore grinding, defines iron content 59.30% iron ore concentrate and contains TiO 2the ilmenite concentrate of 20.15%.Because the method is for ilmenite raw ore, raw ore SiO 2, Al 2o 3, the gangue mineral content such as CaO, MgO is high, the process of alkali leaching preferentially will occur in SiO 2, Al 2o 3with it mineral, define the alkali leaching rear compound similar to titanium in the dipped journey of alkali, the NaOH alkali number of alkali leaching ferrotianium raw ore consumption is 469Kg/t raw ore, and cost is high; And the titanium compound formed after the leaching of ferrotianium raw ore alkali, the compound of the silicon formed after soaking with the gangue mineral alkali such as quartz, it is very difficult for wanting to realize effectively being separated in follow-up magnetic separation, and this also constrains the raising of the rear iron concentrate grade of ferrotianium raw ore alkali leaching and ilmenite concentrate grade.Meanwhile, the method adopts twice grinding process to change mineral surface physicochemical property, adds complexity and the process cost of the method.In a word, by this kind of procedure complexity, and in treating processes, quantity of alkali consumption is large, cost is high; Meanwhile, more high-grade iron ore concentrate and ilmenite concentrate cannot be obtained.
CN201310183580.8 discloses " a kind of wet processing sefstromite concentrate prepares the method for titanium liquid ", proposes the method with salt pickling separating titanium iron.This invention is the method that wet processing v-ti magnetite concentrate prepares titanium liquid, comprise v-ti magnetite concentrate hydrochloric acid leaching, molten salt react ion, again pickling, sulfuric acid solution, filtration etc. and obtain the processes such as titanium liquid, the method is mainly for extraction ilmenite concentrate, its complex technical process, need in hydrochloric acid leaching process to react with hydrochloric acid and iron and vanadium to dissolve in filtrate, consume a large amount of hydrochloric acid, cost is high; Meanwhile, NaOH and titanium and pasc reaction is used to consume alkali in fused salt process.In addition, owing to employing hydrochloric acid in the method leaching process, in hydrochloric acid, chlorion is large to equipment corrosion, not easily suitability for industrialized production.The method is mainly applicable to the recycling of titanium in the low poor v-ti magnetite concentrate of high vanadium low iron content.
CN201410164486.2 discloses one and " utilizes calcining, alkali leaching, pickling and magnetic reconnection to close and select v-ti magnetite concentrate method again ", v-ti magnetite concentrate is at high temperature calcined by this invention, then calcination product is placed in alkaline solution, alkali leaching reaction 0.5 ~ 5 hour, in H after filtration 2sO 4pickling in solution, then acidleach filter cake is carried out magnetic separation and gravity treatment, obtaining TFe content is respectively 65% ~ 69% iron ore concentrate, TiO 2content is the ilmenite concentrate of 70% ~ 82%.The method achieve and efficiently sort v-ti magnetite concentrate, but adopt alkali leaching due to simple in reaction, react 0.5 ~ 5 hour at 300 ~ 370 DEG C of temperature, chemical reaction temperature is higher, and the time is longer, and SiO after reaction 2and TiO 2content is up to 3%, and foreign matter content is higher, causes the capacity factor of a blast furnace to reduce, adds ironmaking cost; Meanwhile, the alkali number consumed in this inventive method is up to 100kg/t concentrate, and alkaline consumption is higher, and titanium resource utilization ratio is not high.
Summary of the invention
In order to overcome the deficiency of above-mentioned beneficiation method, technical problem to be solved by this invention is on the basis that physics and chemistry beneficiation method effectively combines, there is provided a kind of cost low, reclaim quality and efficiency is high, technique is simple, and the utilization calcining of good operability, alkali oxide leaching, pickling and the heavy method selecting v-ti magnetite concentrate again of magnetic, achieve and high efficiency separation is carried out to titanium, iron in v-ti magnetite concentrate, improve into stokehold Iron grade, reduce and enter blast furnace TiO 2, the impurity such as S, Si, Al content, improve the capacity factor of a blast furnace, reduce the quantity discharged of blast furnace slag, reduce ironmaking cost, improve TiO simultaneously 2comprehensive resource utilization rate, reduces environmental pollution.
In order to realize object of the present invention, technical scheme of the present invention is achieved in that
One of the present invention utilizes calcining, alkali oxide leaching, pickling and magnetic to weigh and selects v-ti magnetite concentrate method again, it is characterized in that comprising the steps:
1) calcine
Be 50% ~ 55%, TiO by TFe content range 2content range is 10% ~ 15%, SiO 2content is 3% ~ 6%, Al 2o 3content is 3% ~ 6%, the ratio of the v-ti magnetite concentrate of S content >0.5% 1:0.1 ~ 0.2 by weight adds CaO, at the temperature of 800 DEG C ~ 1400 DEG C, carry out calcining 20 ~ 60 minutes, forms calcination product A;
2) alkali oxide leaching
Calcination product A in step 1) is placed in the alkaline solution that mass concentration is 5% ~ 52%, adds oxygenant, then at the temperature of 220 DEG C ~ 330 DEG C, alkali soaks reaction 0.5 ~ 2.0 hour, and obtain filtrate and alkali leaching cake B, described filtrate feeds recovery and processing system;
3) pickling
By step 2) in alkali leaching cake B add water and make the ore pulp that solid-liquid mass ratio is 1:1 ~ 10, then be placed in the H that mass concentration is 1% ~ 10% 2sO 4in solution, under 50 DEG C ~ 90 DEG C conditions, pickling 5 ~ 60 minutes, filters pickling reactant, and obtain filtrate and acidleach filter cake C, described filtrate feeds recovery and processing system;
4) magnetic-gravity separation
The ore pulp making mass concentration 30% ~ 34% that added water by acidleach filter cake C in step 3) carries out magnetic separation, obtains magnetic concentrate D and magnetic tailing E respectively;
The ore pulp making mass concentration 36% ~ 41% that added water by magnetic tailing E again carries out gravity treatment, respectively gravity concentrate F and gravity tailings G, described magnetic concentrate D to be TFe content range be 66% ~ 70% final iron ore concentrate, gravity concentrate F is TiO 2content range is the final ilmenite concentrate of 70% ~ 82%, and gravity tailings G is SiO 2content is the true tailings of 57 ~ 62%.
Described alkaline solution is any one in the NaOH aqueous solution, the KOH aqueous solution or NaOH and KOH mixed aqueous solution.
Described oxygenant is O 2or H 2o 2, described O 2add-on is 20 ~ 120psi, H 2o 2add-on is 50 ~ 200kg/t to ore deposit.
Described magnetic separation adopts the drum magnetic separator of 0.12T ~ 0.15T to carry out magnetic separation.
Described magnetic separation adopts the magnetic dewater cone of 0.03T ~ 0.05T to carry out magnetic separation.
Described magnetic separation adopts the drum magnetic separator of 0.12T ~ 0.15T and 0.03T ~ 0.05T magnetic dewater cone to carry out two stages of magnetic separation respectively.
Described gravity treatment adopts the spiral chute of ￠ 0.6 ~ ￠ 1.2 meters to carry out gravity treatment.
Advantage of the present invention is:
Calcination utilizes CaO part to replace alkali lye consumption in the dipped journey of alkali, decreases the consumption 20% ~ 30% of NaOH or KOH in follow-up alkali leaching operation; Because CaO price is NaOH price 1/5 ~ 1/6, be 1/20 of KOH price, therefore greatly can reduce production cost.
The process of alkali oxide leaching has carried out chemical reaction to elements such as Ti, S, Si, Al in v-ti magnetite concentrate, defines corresponding salt, makes the iron in v-ti magnetite concentrate change the form of ferric oxide into.With v-ti magnetite concentrate unlike, SiO in ilmenite raw ore 2and Al 2o 3content far away higher than SiO in v-ti magnetite concentrate 2and Al 2o 3content, wherein SiO in ilmenite raw ore 2>20%, Al 2o 3>7%, SiO in v-ti magnetite concentrate 2<6%, Al 2o 3<6%.In alkali leaching ilmenite raw ore process, because the process of alkali leaching preferentially will occur in SiO 2, Al 2o 3deng on mineral, alkali is made to soak v-ti magnetite concentrate more less than alkali leaching ilmenite raw ore alkali consumption, simultaneously O 2introducing make containing S compound oxidation, oxidized FeTiO 3, accelerate reaction, reduce temperature of reaction, shorten the reaction times, better effects if, greatly reduce energy consumption and facility investment.Such as, after calcining, when soaking with NaOH alkali oxide, the alkali number that the present invention consumes is less than 50kg/t concentrate, reduces more than 9 times than the alkali number 469kg/t raw ore of alkali leaching raw ore consumption.Than not passing into O 2alkali leaching consume alkali number reduce 10kg/t concentrate; Meanwhile, O 2introducing alkali is soaked temperature of reaction is minimum is down to 220 DEG C, the reaction times is less than 2 hours.
Acid cleaning process has dissolved oxysalt and the sulfide such as Ti, Si, Al after alkali leaching effectively, makes it to dissociate with iron ore concentrate.In addition because the present invention adopts sulfuric acid to carry out pickling, reaction conditions is gentle, and little to equipment corrosion, cost is low, is more conducive to suitability for industrialized production.
Add magnetic reconnection and close ore dressing, make iron concentrate grade bring up to 66% ~ 70% by 50% ~ 55%, in isolated iron ore concentrate, S content significantly reduces simultaneously, is down to is less than 0.10%, SiO by more than 0.50% 2content is down to less than 1%, Al by 3% ~ 6% 2o 3content is down to less than 1.8%, TiO by 3% ~ 6% 2content is down to less than 6% by more than 12%; Meanwhile, TiO can also be obtained 2content is the ilmenite concentrate of 70% ~ 82%.
Integrated use calcining of the present invention, alkali oxide leaching, pickling and magnetic reconnection close the method process v-ti magnetite concentrate of ore dressing, achieve titanium in v-ti magnetite concentrate, iron high efficiency separation, reduce and enter blast furnace TiO 2, the impurity such as S, Si, Al content, improve the capacity factor of a blast furnace, reduce the quantity discharged of blast furnace slag, reduce ironmaking cost, for subsequent smelting creates better condition, improve the comprehensive utilization ratio of titanium resource simultaneously.
Accompanying drawing explanation
Fig. 1 is present invention process schema.
Fig. 2 is the process flow sheet that the present invention adopts two stages of magnetic separation.
Fig. 3 is the process flow sheet that the present invention adopts another embodiment of two stages of magnetic separation.
Fig. 4 is the process flow sheet that the present invention adopts another embodiment of two stages of magnetic separation.
Embodiment
Below in conjunction with accompanying drawing, the specific embodiment of the present invention is described further:
As shown in Figure 1.
Embodiment 1:
1) calcine
Be 51.8%, TiO by TFe content 2content is 13.9%, SiO 2content is 4.62%, Al 2o 3content is 5.04%, the v-ti magnetite concentrate of S content 0.88%, and the ratio of 1:0.1 adds CaO by weight, at the temperature of 800 DEG C, carry out calcining 50 minutes, and form calcination product A, its chemical equation is:
CaO+ TiO 2 CaTiO 3
2) alkali oxide leaching
Calcination product A in step 1) is placed in the NaOH alkaline solution that mass concentration is 28%, passes into the O of 26psi 2, then at the temperature of 300 DEG C, alkali soaks reaction 0.5 hour, and filtered by reactant, obtain filtrate and alkali leaching cake B, NaOH consumption 35 kg/t is to ore deposit, and described filtrate feeds recovery and processing system, and its chemical equation is:
4mFeTiO 3+8NaOH +mO 2 2mFe 2O 3↓+4Na 2O·(TiO 2) m↓+4H 2O m≥1
pFe 3O 4·q（FeO·TiO 2) +2rNaOH pFe 3O 4↓+qFeO↓+ (Na 2O) r·(TiO 2) q↓+rH 2O
Al 2O 3+2NaOH 2NaAlO 2+ H 2O
tSiO 2+2NaOH Na 2O·(SiO 2) t↓+ H 2O
3FeS 2+6NaOH 3FeS↓+Na 2SO 3+2Na 2S+3H 2O
4FeS 2+ 11O 2 2Fe 2O 3 + 8SO 2
4FeO+O 2 2Fe 2O 3
2SO 2+O 2+4NaOH 2Na 2SO 4+ 2H 2O
3) pickling
By step 2) in alkali leaching cake B add water and make the ore pulp that quality solid-to-liquid ratio is 1:8, then be placed in the H that mass concentration is 10% 2sO 4in solution, under the condition of 70 DEG C of pickling 15 minutes, filtered by pickling reactant, obtain filtrate and acidleach filter cake C, described filtrate fed recovery and processing system, and its chemical equation is:
Na 2O·(TiO 2) x +H + (H 2O)·（TiO 2) x↓+Na +
Na 2O·（SiO 2） t+H + (H 2O)·（SiO 2) t↓+Na +
NaAlO 2+4H + Al 3++Na ++2H 2O
4) magnetic reconnection closes ore dressing
It is that the drum magnetic separator of 0.12T carries out magnetic separation that the ore pulp of making mass concentration 30% of being added water by acidleach filter cake C in step 3) feeds field intensity, respectively magnetic concentrate D and magnetic tailing E;
The spiral chute that the ore pulp of making mass concentration 36% of being added water by magnetic tailing E again feeds ￠ 0.9 meter carries out gravity treatment, respectively gravity concentrate F and gravity tailings G, described magnetic concentrate D to be TFe content be 66.1% final iron ore concentrate, wherein SiO 2content is 0.50%, Al 2o 3content is 1.77%, S content is 0.01%; Described gravity concentrate F is TiO 2content is the final ilmenite concentrate of 70.3%, and gravity tailings G is SiO 2content is the true tailings of 58.5%.
Embodiment 2:
1) calcine
Be 54.0%, TiO by TFe content 2content is 11.5%, SiO 2content is 3.39%, Al 2o 3content is 4.32%, the v-ti magnetite concentrate of S content 0.67%, and the ratio of 1:0.1 adds CaO by weight, at the temperature of 900 DEG C, carry out calcining 45 minutes, and form calcination product A, its chemical equation is with embodiment 1;
2) alkali oxide leaching
Calcination product A in step 1) is placed in the NaOH alkaline solution that mass concentration is 37%, passes into the O of 42psi 2, then at the temperature of 280 DEG C, alkali soaks reaction 1.0 hours, and filtered by reactant, obtain filtrate and alkali leaching cake B, NaOH consumption 37 kg/t is to ore deposit, and described filtrate feeds recovery and processing system, and its chemical equation is with embodiment 1;
3) pickling
By step 2) in alkali leaching cake B add water and make the ore pulp that quality solid-to-liquid ratio is 1:6, then be placed in the H that mass concentration is 8% 2sO 4in solution, under the condition of 90 DEG C, pickling 25 minutes, filters pickling reactant, and obtain filtrate and acidleach filter cake C, described filtrate feeds recovery and processing system, and its chemical equation is with embodiment 1;
4) magnetic reconnection closes ore dressing
It is that the magnetic dewater cone of 0.03T carries out magnetic separation that the ore pulp of making mass concentration 31% of being added water by acidleach filter cake C in step 3) feeds field intensity, respectively magnetic concentrate D and magnetic tailing E;
The spiral chute that the ore pulp of making mass concentration 37% of being added water by magnetic tailing E again feeds ￠ 0.6 meter carries out gravity treatment, respectively gravity concentrate F and gravity tailings G, described magnetic concentrate D to be TFe content be 67.2% final iron ore concentrate, wherein SiO 2content is 0.57%, Al 2o 3content is 1.60%, S content is 0.01%; Described gravity concentrate F is TiO 2content is the final ilmenite concentrate of 72.4%, and gravity tailings G is SiO 2content is the true tailings of 58.8%.
Embodiment 3:
As shown in Figure 2.
1) calcine
Be 50.2%, TiO by TFe content 2content is 15.1%, SiO 2content is 4.02%, Al 2o 3content is 5.31%, the v-ti magnetite concentrate of S content 0.75%, and the ratio of 1:0.1 adds CaO by weight, at the temperature of 1000 DEG C, carry out calcining 60 minutes, and form calcination product A, its chemical equation is with embodiment 1;
2) alkali oxide leaching
Calcination product A in step 1) is placed in the NaOH alkaline solution that mass concentration is 51%, adds 80kg/t to ore deposith 2o 2, then at the temperature of 265 DEG C, alkali soaks reaction 1.0 hours, and filtered by reactant, obtain filtrate and alkali leaching cake B, NaOH consumption 41 kg/t is to ore deposit, and described filtrate feeds recovery and processing system, and its chemical equation is:
2mFeTiO 3+4NaOH+mH 2O 2 mFe 2O 3↓+2Na 2O·(TiO 2) m↓+(m+2)H 2O m≥1
pFe 3O 4·q（FeO·TiO 2) +2rNaOH pFe 3O 4↓+qFeO↓+ (Na 2O) r·(TiO 2) q↓+rH 2O
Al 2O 3+2NaOH 2NaAlO 2+ H 2O
tSiO 2+2NaOH Na 2O·(SiO 2) t↓+ H 2O
3FeS 2+6NaOH 3FeS↓+Na 2SO 3+2Na 2S+3H 2O
2FeS 2+ 11H 2O 2 Fe 2O 3 +4SO 2+ 11H 2O
2FeO+H 2O 2 Fe 2O 3+ H 2O
SO 2+H 2O 2+2NaOH Na 2SO 4+ 2H 2O
3) pickling
By step 2) in alkali leaching cake B add water and make the ore pulp that quality solid-to-liquid ratio is 1:7, then be placed in the H that mass concentration is 7% 2sO 4in solution, under the condition of 80 DEG C of pickling 40 minutes, filtered by pickling reactant, obtain filtrate and acidleach filter cake C, described filtrate fed recovery and processing system, and its chemical equation is with embodiment 1;
4) magnetic reconnection closes ore dressing
It is that the magnetic dewater cone of 0.03T carries out a stages of magnetic separation that the ore pulp of making mass concentration 32% of being added water by acidleach filter cake C in step 3) feeds field intensity, obtains an a stages of magnetic separation concentrate D1 and stages of magnetic separation mine tailing E1 respectively; Two stages of magnetic separation are carried out to the drum magnetic separator that the ore pulp employing field intensity of a stages of magnetic separation concentrate D1 mass concentration 31% is 0.13T, obtains two stages of magnetic separation concentrate D2 and two stages of magnetic separation mine tailing E2 respectively;
The spiral chute that the ore pulp of making mass concentration 37% of being added water by two stages of magnetic separation ore deposit tail E2 again feeds ￠ 0.6 meter carries out gravity treatment, respectively gravity concentrate F and gravity tailings G, two described stages of magnetic separation concentrate D2 to be TFe content be 68.9% final iron ore concentrate, wherein SiO 2content is 0.32%, Al 2o 3content is 1.09%, S content is 0.01%; Described gravity concentrate F is TiO 2content is the final ilmenite concentrate of 77.4%, and a described stages of magnetic separation mine tailing E1 and gravity tailings G merges into SiO 2content is the true tailings of 58.1%.
Embodiment 4:
As shown in Figure 3.
1) calcine
Be 51.0%, TiO by TFe content 2content is 14.1%, SiO 2content is 3.28%, Al 2o 3content is 4.72%, the v-ti magnetite concentrate of S content 0.80%, and the ratio of 1:0.15 adds CaO by weight, at the temperature of 1100 DEG C, carry out calcining 35 minutes, and form calcination product A, its chemical equation is with embodiment 1;
2) alkali oxide leaching
Calcination product A in step 1) is placed in the NaOH alkaline solution that mass concentration is 19%, passes into the O of 92psi 2, then at the temperature of 310 DEG C, alkali soaks reaction 1.5 hours, and filtered by reactant, obtain filtrate and alkali leaching cake B, NaOH consumption 32 kg/t is to ore deposit, and described filtrate feeds recovery and processing system, and its chemical equation is with embodiment 1;
3) pickling
By step 2) in alkali leaching cake B add water and make the ore pulp that quality solid-to-liquid ratio is 1:9, then be placed in the H that mass concentration is 6% 2sO 4in solution, under the condition of 60 DEG C, pickling 50 minutes, filters pickling reactant, and obtain filtrate and acidleach filter cake C, described filtrate feeds recovery and processing system, and its chemical equation is with embodiment 1;
4) magnetic reconnection closes ore dressing
It is that the magnetic dewater cone of 0.03T carries out a stages of magnetic separation that the ore pulp of making mass concentration 32% of being added water by acidleach filter cake C in step 3) feeds field intensity, obtains an a stages of magnetic separation concentrate D1 and stages of magnetic separation mine tailing E1 respectively; Two stages of magnetic separation are carried out to the drum magnetic separator that the ore pulp employing field intensity of a stages of magnetic separation concentrate D1 mass concentration 33% is 0.13T, obtains two stages of magnetic separation concentrate D2 and two stages of magnetic separation mine tailing E2 respectively;
The spiral chute that the ore pulp of making mass concentration 38% of being added water by two stages of magnetic separation mine tailing E2 again feeds ￠ 1.2 meters carries out gravity treatment, obtain gravity concentrate F and gravity tailings G respectively, two described stages of magnetic separation concentrate D2 and gravity concentrate F merge into the final iron ore concentrate that TFe content is 67.9%, wherein SiO 2content is 0.35%, Al 2o 3content is 1.11%, S content is 0.01%; Described gravity tailings G is TiO 2content is the final ilmenite concentrate of 81.2%, and a described stages of magnetic separation mine tailing E1 is SiO 2content is the true tailings of 58.2%.
Embodiment 5:
As shown in Figure 4.
1) calcine
Be 53.1%, TiO by TFe content 2content is 11.5%, SiO 2content is 3.47%, Al 2o 3content is 4.48%, the v-ti magnetite concentrate of S content 0.60%, and the ratio of 1:0.15 adds CaO by weight, at the temperature of 1250 DEG C, carry out calcining 22 minutes, and form calcination product A, its chemical equation is with embodiment 1;
2) alkali oxide leaching
Calcination product A in step 1) is placed in the NaOH alkaline solution that mass concentration is 12%, passes into the O of 118psi 2, then at the temperature of 240 DEG C, alkali soaks reaction 2.0 hours, and filtered by reactant, obtain filtrate and alkali leaching cake B, NaOH consumption 30kg/t is to ore deposit, and described filtrate feeds recovery and processing system, and its chemical equation is:
4mFeTiO 3+8KOH+mO 2 2mFe 2O 3↓+4K 2O·(TiO 2) m↓+4H 2O m≥1
pFe 3O 4·q（FeO·TiO 2) +2rKOH pFe 3O 4↓+qFeO↓+ (K 2O) r·(TiO 2) q↓+rH 2O
Al 2O 3+2KOH 2KAlO 2+ H 2O
tSiO 2+2KOH K 2O·（SiO 2） t↓+ H 2O
3FeS 2+6KOH 3FeS↓+K 2SO 3+2K 2S+3H 2O
4FeS 2+ 11O 2 2Fe 2O 3 + 8SO 2
4FeO+O 2 2Fe 2O 3
2SO 2+O 2+4KOH 2K 2SO 4+ 2H 2O
3) pickling
By step 2) in alkali leaching cake B add water and make the ore pulp that quality solid-to-liquid ratio is 1:3, then be placed in the H that mass concentration is 5% 2sO 4in solution, under the condition of 75 DEG C, pickling 30 minutes, filters pickling reactant, and obtain filtrate and acidleach filter cake C, described filtrate feeds recovery and processing system, and its chemical equation is:
K 2O·(TiO 2) x +H + (H 2O)·（TiO 2) x↓+ K +
K 2O·（SiO 2） t+H + (H 2O)·（SiO 2) t↓+ K +
KAlO 2+4H + Al 3++K ++2H 2O
4) magnetic reconnection closes ore dressing
It is that the drum magnetic separator of 0.13T carries out a stages of magnetic separation that the ore pulp of making mass concentration 34% of being added water by acidleach filter cake C in step 3) feeds field intensity, obtains an a stages of magnetic separation concentrate D1 and stages of magnetic separation mine tailing E1 respectively; Two stages of magnetic separation are carried out to the magnetic dewater cone that the ore pulp employing field intensity of a stages of magnetic separation concentrate D1 mass concentration 32% is 0.03T, obtains two stages of magnetic separation concentrate D2 and two stages of magnetic separation mine tailing E2 respectively;
The spiral chute that the ore pulp of making mass concentration 41% of being added water by two stages of magnetic separation mine tailing E1 and E2 again feeds ￠ 1.2 meters carries out gravity treatment, obtain gravity concentrate F and gravity tailings G respectively, two described stages of magnetic separation concentrate D2 to be TFe content be 68.2% final iron ore concentrate, wherein SiO 2content is 0.47%, Al 2o 3content is 1.52%, S content is 0.01%; Described gravity concentrate F is TiO 2content is the final ilmenite concentrate of 75.2%, and described gravity tailings G is SiO 2content is the true tailings of 60.2%.
Embodiment 6:
1) calcine
Be 54.6%, TiO by TFe content 2content is 10.7%, SiO 2content is 4.25%, Al 2o 3content is 3.94%, the v-ti magnetite concentrate of S content 0.60%, and the ratio of 1:0.15 adds CaO by weight, at the temperature of 1300 DEG C, carry out calcining 30 minutes, and form calcination product A, its chemical equation is with embodiment 1;
2) alkali oxide leaching
Calcination product A in step 1) is placed in the NaOH alkaline solution that mass concentration is 6%, adds 199kg/t to ore deposith 2o 2, then at the temperature of 330 DEG C, alkali soaks reaction 2.0 hours, and filtered by reactant, obtain filtrate and alkali leaching cake B, NaOH consumption 48kg/t is to ore deposit, and described filtrate feeds recovery and processing system, and its chemical equation is:
2mFeTiO 3+4KOH+mH 2O 2 mFe 2O 3↓+2K 2O·(TiO 2) m↓+(m+2)H 2O m≥1
pFe 3O 4·q（FeO·TiO 2) +2rKOH pFe 3O 4↓+qFeO↓+ (K 2O) r·(TiO 2) q↓+rH 2O
Al 2O 3+2KOH 2KAlO 2+ H 2O
tSiO 2+2KOH K 2O·(SiO 2) t↓+ H 2O
3FeS 2+6KOH 3FeS↓+K 2SO 3+2K 2S+3H 2O
2FeS 2+ 11H 2O 2 Fe 2O 3 +4SO 2+ 11H 2O
2FeO+H 2O 2 Fe 2O 3+ H 2O
SO 2+H 2O 2+2KOH K 2SO 4+ 2H 2O
3) pickling
By step 2) in alkali leaching cake B add water and make the ore pulp that quality solid-to-liquid ratio is 1:2, then be placed in the H that mass concentration is 3% 2sO 4in solution, under the condition of 90 DEG C, pickling 10 minutes, filters pickling reactant, and obtain filtrate and acidleach filter cake C, described filtrate feeds recovery and processing system, and its chemical equation is with embodiment 5;
4) magnetic reconnection closes ore dressing
It is that the drum magnetic separator of 0.13T carries out a stages of magnetic separation that the ore pulp of making mass concentration 33% of being added water by acidleach filter cake C in step 3) feeds field intensity, obtains an a stages of magnetic separation concentrate D1 and stages of magnetic separation mine tailing E1 respectively; Two stages of magnetic separation are carried out to the magnetic dewater cone that the ore pulp employing field intensity of a stages of magnetic separation concentrate D1 mass concentration 31% is 0.03T, obtains two stages of magnetic separation concentrate D2 and two stages of magnetic separation mine tailing E2 respectively;
The spiral chute that the ore pulp of making mass concentration 41% of being added water by two stages of magnetic separation mine tailing E1 and E2 again feeds ￠ 0.6 meter carries out gravity treatment, respectively gravity concentrate F and gravity tailings G; Two described stages of magnetic separation concentrate D2 to be TFe content be 69.8% final iron ore concentrate, wherein SiO 2content is 0.41%, Al 2o 3content is 1.22%, S content is 0.01%; Described gravity concentrate F is TiO 2content is the final ilmenite concentrate of 80.3%, and described gravity tailings G is SiO 2content is the true tailings of 61.0%.
Embodiment 7:
1) calcine
Be 52.4%, TiO by TFe content 2content is 11.9%, SiO 2content is 4.07%, Al 2o 3content is 4.65%, the v-ti magnetite concentrate of S content 0.66%, and the ratio of 1:0.2 adds CaO by weight, at the temperature of 1400 DEG C, carry out calcining 20 minutes, and form calcination product A, its chemical equation is with embodiment 1;
2) alkali oxide leaching
Calcination product A in step 1) is placed in NaOH mass concentration be 18% and KOH mass concentration be the mixed ammonium/alkali solutions of 9%, pass into the O of 50psi 2, then at the temperature of 220 DEG C, alkali soaks reaction 1.0 hours, is filtered by reactant, obtain filtrate and alkali leaching cake B, NaOH consumption 30kg/t is to ore deposit KOH consumption 15 kg/t to ore deposit, and described filtrate feeds recovery and processing system, and its chemical equation is with embodiment 1 and embodiment 5;
3) pickling
By step 2) in alkali leaching cake B add water and make the ore pulp that quality solid-to-liquid ratio is 1:3, then be placed in the H that mass concentration is 2% 2sO 4in solution, under the condition of 80 DEG C, pickling 60 minutes, filters pickling reactant, and obtain filtrate and acidleach filter cake C, described filtrate feeds recovery and processing system, and its chemical equation is with embodiment 1 and embodiment 5;
4) magnetic reconnection closes ore dressing
It is that the drum magnetic separator of 0.13T carries out a stages of magnetic separation that the ore pulp of making mass concentration 33% of being added water by acidleach filter cake C in step 3) feeds field intensity, obtains an a stages of magnetic separation concentrate D1 and stages of magnetic separation mine tailing E1 respectively; Two stages of magnetic separation are carried out to the magnetic dewater cone that the ore pulp employing field intensity of a stages of magnetic separation concentrate D1 mass concentration 31% is 0.03T, obtains two stages of magnetic separation concentrate D2 and two stages of magnetic separation mine tailing E2 respectively;
The spiral chute that the ore pulp of making mass concentration 41% of being added water by two stages of magnetic separation mine tailing E1 and E2 again feeds ￠ 0.6 meter carries out gravity treatment, respectively gravity concentrate F and gravity tailings G; Two described stages of magnetic separation concentrate D2 to be TFe content be 68.5% final iron ore concentrate, wherein SiO 2content is 0.35%, Al 2o 3content is 1.10%, S content is 0.01%; Described gravity concentrate F is TiO 2content is the final ilmenite concentrate of 74.8%, and described gravity tailings G is SiO 2content is the true tailings of 61.9%.
Embodiment 8:
1) calcine
Be 52.0%, TiO by TFe content 2content is 12.5%, SiO 2content is 3.87%, Al 2o 3content is 4.65%, the v-ti magnetite concentrate of S content 0.69%, and the ratio of 1:0.2 adds CaO by weight, at the temperature of 1300 DEG C, carry out calcining 30 minutes, and form calcination product A, its chemical equation is with embodiment 1;
2) alkali oxide leaching
Calcination product A in step 1) is placed in NaOH mass concentration be 20% and KOH mass concentration be the mixed ammonium/alkali solutions of 7%, add 56kg/t to ore deposith 2o 2, then at the temperature of 230 DEG C, alkali soaks reaction 1.0 hours, is filtered by reactant, obtain filtrate and alkali leaching cake B, NaOH consumption 31.5kg/t is to ore deposit KOH consumption 12.8 kg/t to ore deposit, and described filtrate feeds recovery and processing system, and its chemical equation is with embodiment 3 and embodiment 6;
3) pickling
By step 2) in alkali leaching cake B add water and make the ore pulp that quality solid-to-liquid ratio is 1:5, then be placed in the H that mass concentration is 3% 2sO 4in solution, under the condition of 80 DEG C, pickling 50 minutes, filters pickling reactant, and obtain filtrate and acidleach filter cake C, described filtrate feeds recovery and processing system, and its chemical equation is with embodiment 1 and embodiment 5;
4) magnetic reconnection closes ore dressing
It is that the drum magnetic separator of 0.13T carries out a stages of magnetic separation that the ore pulp of making mass concentration 33% of being added water by acidleach filter cake C in step 3) feeds field intensity, obtains an a stages of magnetic separation concentrate D1 and stages of magnetic separation mine tailing E1 respectively; Two stages of magnetic separation are carried out to the magnetic dewater cone that the ore pulp employing field intensity of a stages of magnetic separation concentrate D1 mass concentration 31% is 0.03T, obtains two stages of magnetic separation concentrate D2 and two stages of magnetic separation mine tailing E2 respectively;
The spiral chute that the ore pulp of making mass concentration 40% of being added water by two stages of magnetic separation mine tailing E1 and E2 again feeds ￠ 0.6 meter carries out gravity treatment, respectively gravity concentrate F and gravity tailings G; Two described stages of magnetic separation concentrate D2 to be TFe content be 67.3% final iron ore concentrate, wherein SiO 2content is 0.65%, Al 2o 3content is 1.21%, S content is 0.01%; Described gravity concentrate F is TiO 2content is the final ilmenite concentrate of 72.5%, and described gravity tailings G is SiO 2content is the true tailings of 61.1%.

Claims (7)

1. utilize calcining, alkali oxide leaching, pickling and magnetic to weigh and select a v-ti magnetite concentrate method again, it is characterized in that comprising the steps:
1) calcine
Be 50% ~ 55%, TiO by TFe content range 2content range is 10% ~ 15%, SiO 2content is 3% ~ 6%, Al 2o 3content is 3% ~ 6%, the ratio of the v-ti magnetite concentrate of S content >0.5% 1:0.1 ~ 0.2 by weight adds CaO, at the temperature of 800 DEG C ~ 1400 DEG C, carry out calcining 20 ~ 60 minutes, forms calcination product A;
2) alkali oxide leaching
Calcination product A in step 1) is placed in the alkaline solution that mass concentration is 5% ~ 52%, adds oxygenant, then at the temperature of 220 DEG C ~ 330 DEG C, alkali soaks reaction 0.5 ~ 2.0 hour, and obtain filtrate and alkali leaching cake B, described filtrate feeds recovery and processing system;
3) pickling
By step 2) in alkali leaching cake B add water and make the ore pulp that solid-liquid mass ratio is 1:1 ~ 10, then be placed in the H that mass concentration is 1% ~ 10% 2sO 4in solution, under 50 DEG C ~ 90 DEG C conditions, pickling 5 ~ 60 minutes, filters pickling reactant, and obtain filtrate and acidleach filter cake C, described filtrate feeds recovery and processing system;
4) magnetic-gravity separation
The ore pulp making mass concentration 30% ~ 34% that added water by acidleach filter cake C in step 3) carries out magnetic separation, obtains magnetic concentrate D and magnetic tailing E respectively;
The ore pulp making mass concentration 36% ~ 41% that added water by magnetic tailing E again carries out gravity treatment, respectively gravity concentrate F and gravity tailings G, described magnetic concentrate D to be TFe content range be 66% ~ 70% final iron ore concentrate, gravity concentrate F is TiO 2content range is the final ilmenite concentrate of 70% ~ 82%, and gravity tailings G is SiO 2content is the true tailings of 57 ~ 62%.
2. utilization according to claim 1 calcining, alkali oxide leaching, pickling and magnetic is heavy selects v-ti magnetite concentrate method again, is characterized in that described alkaline solution is any one in the NaOH aqueous solution, the KOH aqueous solution or NaOH and KOH mixed aqueous solution.
3. utilization calcining according to claim 1, alkali oxide leaching, pickling and magnetic weigh and select v-ti magnetite concentrate method again, it is characterized in that described oxygenant is O 2or H 2o 2, described O 2add-on is 20 ~ 120psi, H 2o 2add-on is 50 ~ 200kg/t to ore deposit.
4. utilization calcining according to claim 1, alkali oxide leaching, pickling and magnetic weigh and select v-ti magnetite concentrate method again, it is characterized in that described magnetic separation adopts the drum magnetic separator of 0.12T ~ 0.15T to carry out magnetic separation.
5. utilization calcining according to claim 1, alkali oxide leaching, pickling and magnetic weigh and select v-ti magnetite concentrate method again, it is characterized in that described described magnetic separation adopts the magnetic dewater cone of 0.03T ~ 0.05T to carry out magnetic separation.
6. utilization calcining according to claim 1, alkali oxide leaching, pickling and magnetic weigh and select v-ti magnetite concentrate method again, it is characterized in that described described magnetic separation adopts the drum magnetic separator of 0.12T ~ 0.15T and 0.03T ~ 0.05T magnetic dewater cone to carry out two stages of magnetic separation respectively.
7. utilization calcining according to claim 1, alkali oxide leaching, pickling and magnetic weigh and select v-ti magnetite concentrate method again, it is characterized in that described described gravity treatment adopts the spiral chute of ￠ 0.6 ~ ￠ 1.2 meters to carry out gravity treatment.
CN201510320887.7A 2015-06-12 2015-06-12 Method for separating vanadium-titanium magnetite concentrate by calcining, oxidation alkali leaching, pickling and magnetic reseparation Pending CN104911331A (en)

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Citations (5)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
US20010051120A1 (en) * 1997-10-17 2001-12-13 Marcelo De Matos Process for the production titanium concentrate having a chemical composition similar to ilmenite from highly impure anatase ores
CN102181626A (en) * 2011-04-08 2011-09-14 北京矿冶研究总院 Beneficiation method of ilmenite
CN102787194A (en) * 2012-08-27 2012-11-21 攀枝花学院 Method for preparing titanium-rich material by directly reducing molten slag from vanadium-titanium-ferrum concentrate
CN103526051A (en) * 2013-09-26 2014-01-22 攀钢集团攀枝花钢铁研究院有限公司 Method for separating iron, vanadium and titanium from schreyerite
CN103962226A (en) * 2014-04-23 2014-08-06 鞍钢集团矿业公司 Method for recleaning of vanadium-titanium magnetite concentrates through calcination, alkaline leaching, acid pickling and combination of magnetic separation and gravity concentration

Patent Citations (5)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
US20010051120A1 (en) * 1997-10-17 2001-12-13 Marcelo De Matos Process for the production titanium concentrate having a chemical composition similar to ilmenite from highly impure anatase ores
CN102181626A (en) * 2011-04-08 2011-09-14 北京矿冶研究总院 Beneficiation method of ilmenite
CN102787194A (en) * 2012-08-27 2012-11-21 攀枝花学院 Method for preparing titanium-rich material by directly reducing molten slag from vanadium-titanium-ferrum concentrate
CN103526051A (en) * 2013-09-26 2014-01-22 攀钢集团攀枝花钢铁研究院有限公司 Method for separating iron, vanadium and titanium from schreyerite
CN103962226A (en) * 2014-04-23 2014-08-06 鞍钢集团矿业公司 Method for recleaning of vanadium-titanium magnetite concentrates through calcination, alkaline leaching, acid pickling and combination of magnetic separation and gravity concentration

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