CN105013608A - Vanadium-titanium magnetite concentrate re-concentration method achieved through oxidation alkaline leaching, acid pickling, desliming and gravity and magnetism - Google Patents
Vanadium-titanium magnetite concentrate re-concentration method achieved through oxidation alkaline leaching, acid pickling, desliming and gravity and magnetism Download PDFInfo
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Abstract
本发明涉及一种利用氧化碱浸、酸洗、脱泥及重磁再选钒钛磁铁精矿方法,其特征在于包括如下步骤:氧化碱浸将钒钛磁铁精矿,置于碱溶液中,加入氧化剂,然后在220℃~330℃的温度下碱浸反应0.5~2小时,过滤,得滤液和碱浸滤饼A,将碱浸滤饼A加水配制成矿浆,再置于H2SO4溶液中,50~90℃下酸洗5~60分钟,过滤,得滤液和酸浸滤饼B;再将B加水制成质量浓度30~35%的矿浆进行脱泥及重磁联合再选,分别得到TFe含量为64~70%铁精矿和TiO2含量为60~82%钛精矿。其优点是:氧化碱浸中O2或H2O2的引入使含S化合物氧化,加速了反应,降低了反应温度,缩短了反应时间,大大降低能耗和设备投资。
The invention relates to a method for re-selecting vanadium-titanium magnetite concentrate by oxidizing alkali leaching, pickling, desliming and remagnetization, which is characterized in that it comprises the following steps: the vanadium-titanium magnetite concentrate is placed in an alkaline solution by oxidizing alkali leaching, Add an oxidizing agent, then react with alkali leaching at a temperature of 220°C to 330°C for 0.5 to 2 hours, filter to obtain the filtrate and alkali leaching filter cake A, add water to the alkali leaching filter cake A to make a slurry, and then place it in H 2 SO 4 In the solution, acid wash at 50-90°C for 5-60 minutes, filter to obtain the filtrate and acid leaching filter cake B; then add water to B to make a pulp with a mass concentration of 30-35% for desliming and gravity-magnetic re-election. Iron concentrates with TFe content of 64-70% and titanium concentrates with TiO 2 content of 60-82% were obtained respectively. Its advantages are: the introduction of O 2 or H 2 O 2 in the oxidative alkali leaching oxidizes the S-containing compound, accelerates the reaction, reduces the reaction temperature, shortens the reaction time, and greatly reduces energy consumption and equipment investment.
Description
技术领域 technical field
本发明涉及一种钒钛磁铁精矿的选矿工艺,尤其涉及一种利用氧化碱浸、酸洗、脱泥及重磁再选钒钛磁铁精矿方法。 The invention relates to a beneficiation process of a vanadium-titanium magnetite concentrate, in particular to a method for re-selecting a vanadium-titanium magnetite concentrate by using oxidative alkali leaching, pickling, desliming and gravity magnetism.
背景技术 Background technique
钒钛磁铁矿是一种多金属元素的复合矿,是以含铁、钒、钛为主的共生的磁铁矿。而钒钛磁铁精矿是钒钛磁铁矿经过选矿获得的产物之一,其中钒以类质同象赋存于钛磁铁矿中,置换高价铁离子。钛磁铁矿是主晶矿物(Fe3O4)与客晶矿【钛铁晶石2FeO·TiO2、钛铁矿FeO·TiO2、铝镁尖晶石(Mg,Fe)(Al,Fe)2O4】形成的复合体。例如,中国攀枝花地区密地选矿厂钒钛磁铁矿原矿和选铁后的钒钛磁铁精矿化学多元素分析结果见表1,钒钛磁铁矿原矿和钒钛磁铁矿精矿物相分析结果分别见表2和表3。 Vanadium-titanium magnetite is a compound ore of multi-metal elements, which is a symbiotic magnetite mainly containing iron, vanadium and titanium. The vanadium-titanium magnetite concentrate is one of the products obtained from vanadium-titanium magnetite through ore dressing, in which vanadium exists in the titanium magnetite in the same quality and isomorphism to replace high-valent iron ions. Titanomagnetite is the main crystal mineral (Fe 3 O 4 ) and guest crystal ore [ilmenite spar 2FeO·TiO 2 , ilmenite FeO·TiO 2 , aluminum magnesium spinel (Mg,Fe) (Al,Fe ) 2 O 4 ] complexes formed. For example, the chemical multi-element analysis results of vanadium-titanium magnetite raw ore and vanadium-titanium magnetite concentrate after iron selection in Midi Concentrator in Panzhihua, China are shown in Table 1. The vanadium-titanium magnetite raw ore and vanadium-titanium magnetite concentrate The analysis results are shown in Table 2 and Table 3, respectively.
表1 中国攀枝花地区密地选矿厂原矿和钒钛磁铁精矿化学多元素分析结果 Table 1 Chemical multi-element analysis results of raw ore and vanadium-titanium magnetite concentrate of Midi concentrator in Panzhihua, China
表2 中国攀枝花地区密地选矿厂钒钛磁铁矿原矿钛、铁化学物相分析结果 Table 2 The chemical phase analysis results of titanium and iron in the vanadium-titanium magnetite raw ore of Midi concentrator in Panzhihua, China
表3 中国攀枝花地区密地选矿厂钒钛磁铁矿精矿钛、铁化学物相分析结果 Table 3 Chemical phase analysis results of titanium and iron in the vanadium-titanium magnetite concentrate of Midi concentrator in Panzhihua, China
世界上钒钛磁铁矿资源丰富,全世界储量达400亿吨以上,中国储量达98.3亿吨。钒钛磁铁矿石中铁主要赋存于钛磁铁矿中,矿石中的TiO2主要赋存于粒状钛铁矿和钛磁铁矿中。一般情况下,约57%的钛赋存于钛磁铁矿(mFeTiO 3·nFe3O4)中,约40%的钛赋存于钛铁矿(FeTiO3)中,由于钒钛磁铁矿矿石组成复杂,性质特殊,因而这类矿石的综合利用是国际一直未彻底解决的一大难题。钒钛磁铁矿矿物的这种赋存特点决定了采用物理选矿方法无法从矿石的源头实现钛、铁的有效分离,造成钒钛磁铁矿石经物理选矿后,铁精矿品位低(TFe<55%),铁精矿中的钛在炼铁过程完全进入高炉渣(TiO2含量达22%以上)形成玻璃体,TiO2失去了活性而无法经济回收,同时,钛回收率低只有18%。因此用物理的选矿方法选别钛铁矿石大大降低了钛和铁单独利用的价值。 The vanadium-titanium magnetite resources are abundant in the world, with reserves of more than 40 billion tons in the world and 9.83 billion tons in China. Iron in vanadium-titanium magnetite ore mainly occurs in titanomagnetite, and TiO 2 in the ore mainly occurs in granular ilmenite and titanomagnetite. Generally, about 57% of titanium occurs in titanomagnetite (mFeTiO 3 ·nFe 3 O 4 ), and about 40% of titanium occurs in ilmenite (FeTiO 3 ). The ore has complex composition and special properties, so the comprehensive utilization of this kind of ore is a major problem that has not been completely solved in the world. The occurrence characteristics of vanadium-titanium magnetite minerals determine that the effective separation of titanium and iron from the source of the ore cannot be achieved by physical beneficiation methods, resulting in low grade iron concentrates (TFe<55 %), the titanium in the iron concentrate completely enters the blast furnace slag (TiO 2 content is more than 22%) to form a glass body during the ironmaking process, and the TiO 2 loses its activity and cannot be economically recovered. At the same time, the titanium recovery rate is as low as 18%. Therefore, the use of physical beneficiation methods to separate ilmenite ore greatly reduces the value of titanium and iron alone.
中国是世界上第一个以工业规模从复杂钒钛磁铁矿中综合提取铁、钒、钛的国家,但由于一般的物理方法不能从根本上改变铁、钛致密共生的赋存特性,因此,采用通常的重选法、磁选法、浮选法等物理选矿方法进行钛、铁分离,效率低,很难选出品位高而杂质少的钛精矿或铁精矿;同时,TiO2回收效率不高,钒钛磁铁矿原矿经过选矿分离后,约54%的TiO2进入铁精矿,这些TiO2经高炉冶炼后几乎全部进入渣相,形成TiO2含量20~24%的高炉渣;另外,由于铁精矿中的S、Si、Al等杂质含量也过高,上述原因不仅造成冶炼高炉利用系数低、能耗大、钛资源浪费,而且矿渣量大、环境污染严重。 China is the first country in the world to comprehensively extract iron, vanadium and titanium from complex vanadium-titanium magnetite on an industrial scale. , using the usual gravity separation, magnetic separation, flotation and other physical beneficiation methods to separate titanium and iron, the efficiency is low, and it is difficult to select titanium concentrate or iron concentrate with high grade and less impurities; at the same time, TiO 2 The recovery efficiency is not high. After the vanadium-titanium-magnetite raw ore is separated by ore dressing, about 54% of TiO 2 enters the iron ore concentrate. After smelting in the blast furnace, almost all of the TiO 2 enters the slag phase, forming a high Slag; In addition, due to the high content of impurities such as S, Si, and Al in the iron concentrate, the above reasons not only lead to low utilization coefficient of smelting blast furnace, high energy consumption, waste of titanium resources, but also large amount of slag and serious environmental pollution.
CN2011100879566公开了“一种钛铁矿的选矿方法”,是将钒钛磁铁矿原矿经磨矿、碱浸预处理、过滤、再磨矿后磁选得到钛精矿和铁精矿的方法。该方法将含铁32.16%和含TiO212.11%的钒钛磁铁矿原矿通过磨矿、碱浸预处理、过滤、再磨矿后磁选处理,形成了含铁59.30%铁精矿和含TiO220.15%的钛精矿。由于该方法是针对钛铁矿原矿而言,原矿SiO2、Al2O3、CaO、MgO等脉石矿物含量高,碱浸的过程将优先发生在SiO2、Al2O3等矿物身上,碱浸过程中形成了与钛相似的碱浸后化合物,碱浸钛铁原矿消耗的NaOH碱量是469Kg/t原矿,成本高;而且钛铁原矿碱浸后形成的钛化合物,与石英等脉石矿物碱浸后形成的硅的化合物,要想在后续的磁选中实现有效分离是十分困难的,这也制约了钛铁原矿碱浸后铁精矿品位和钛精矿品位的提高。同时,该方法采用两次磨矿过程改变矿物表面物理化学性质,增加了该方法的复杂程度和工序成本。总之,用该种方法过程复杂,而且处理过程中碱消耗量大、成本高;同时,无法获得更高品位的铁精矿和钛精矿。 CN2011100879566 discloses "a method for beneficiating ilmenite", which is a method for obtaining titanium concentrate and iron concentrate by magnetically separating vanadium-titanium magnetite raw ore through grinding, alkali leaching pretreatment, filtration, and regrinding. In this method, the vanadium-titanium magnetite raw ore containing 32.16% iron and 12.11% TiO2 is processed through grinding, alkali leaching pretreatment, filtration, and magnetic separation after regrinding to form iron concentrate containing 59.30% iron and iron containing TiO 2 20.15% titanium concentrate. Since this method is aimed at raw ilmenite ore, which contains high gangue minerals such as SiO 2 , Al 2 O 3 , CaO, and MgO, the alkaline leaching process will preferentially occur on minerals such as SiO 2 , Al 2 O 3 , etc. Alkaline leaching compounds similar to titanium are formed during the alkaline leaching process. The amount of NaOH alkali consumed by alkaline leaching of ilmenite raw ore is 469Kg/t raw ore, which is high in cost; It is very difficult to achieve effective separation of silicon compounds formed after alkali leaching of ore minerals in subsequent magnetic separation, which also restricts the improvement of iron concentrate grade and titanium concentrate grade after alkali leaching of ilmenite raw ore. At the same time, the method uses two grinding processes to change the physical and chemical properties of the mineral surface, which increases the complexity and process cost of the method. In short, the process of this method is complicated, and the alkali consumption is large and the cost is high during the treatment process; at the same time, it is impossible to obtain higher-grade iron ore concentrates and titanium concentrates.
CN201310183580.8公开了“一种湿法处理钒钛铁精矿制备钛液的方法”,提出了用盐酸洗分离钛铁的方法。该发明为湿法处理钒钛磁铁精矿制备钛液的方法,包括钒钛磁铁精矿盐酸浸取、熔盐反应、再酸洗、硫酸酸溶、过滤等获得钛液等过程,该方法主要是针对提取钛精矿,其工艺过程复杂,盐酸浸取过程中需用盐酸与铁和钒反应溶解进滤液中,消耗大量盐酸,成本高;同时,熔盐过程中用NaOH与钛和硅反应消耗碱。另外,由于该方法浸取过程中使用了盐酸,盐酸中氯离子对设备腐蚀大,不易工业化生产。该方法主要适用于高钒低铁含量的低贫钒钛磁铁精矿中钛的回收利用。 CN201310183580.8 discloses "a method for wet treatment of vanadium-titanium iron concentrate to prepare titanium liquid", and proposes a method for separating ferrotitanium by washing with hydrochloric acid. The invention is a method for wet treatment of vanadium-titanium magnetite concentrate to prepare titanium liquid, including hydrochloric acid leaching of vanadium-titanium magnetite concentrate, molten salt reaction, re-pickling, sulfuric acid acid dissolution, filtration and other processes to obtain titanium liquid. It is aimed at extracting titanium concentrate. The process is complicated. In the process of hydrochloric acid leaching, hydrochloric acid needs to react with iron and vanadium to dissolve into the filtrate, which consumes a large amount of hydrochloric acid and has high cost. At the same time, NaOH is used to react with titanium and silicon in the molten salt process. Consume alkali. In addition, because hydrochloric acid is used in the leaching process of this method, the chloride ions in the hydrochloric acid corrode the equipment greatly, which is not easy for industrial production. The method is mainly applicable to the recovery and utilization of titanium in low-vanadium-depleted titanium magnetite concentrate with high vanadium and low iron content.
CN201410164253.2 公开了一种“利用碱浸、酸洗、脱泥及重磁联合再选钒钛磁铁精矿方法”,该发明将钒钛磁铁精矿置于纯碱溶液中,碱浸反应0.5~5小时,过滤后于H2SO4溶液中酸洗,再将酸浸滤饼进行脱泥,将脱泥作业得到的沉砂加水配制成矿浆进行重选,再将重选尾矿进行磁选, 分别得到TFe含量为63%~69%铁精矿、TiO2含量为60%~82%钛精矿。该方法实现了对钒钛磁铁精矿进行高效选别,但由于反应中单纯采用碱浸,在280~370℃温度下反应0.5~5小时,化学反应温度较高,时间较长,且反应后SiO2和TiO2含量高达3%,杂质含量较高,致使高炉利用系数降低,增加了炼铁成本;同时,该发明方法中消耗的碱量高达100kg/t精矿,碱耗较高,钛资源利用率不高。 CN201410164253.2 discloses a "method for reselecting vanadium-titanium magnetite concentrate by using alkali leaching, pickling, desliming and gravity-magnetism combination". After 5 hours, pickling in H 2 SO 4 solution after filtering, then desliming the acid leaching filter cake, adding water to the grit obtained from the desliming operation to make slurry for gravity separation, and then magnetic separation for gravity separation tailings , were obtained TFe content of 63% ~ 69% iron concentrate, TiO 2 content of 60% ~ 82% titanium concentrate. This method realizes high-efficiency sorting of vanadium-titanium magnetite concentrate, but because alkali leaching is only used in the reaction, the reaction is carried out at a temperature of 280-370°C for 0.5-5 hours, the chemical reaction temperature is high, the time is long, and after the reaction The content of SiO2 and TiO2 is as high as 3%, and the content of impurities is high, which reduces the utilization coefficient of the blast furnace and increases the cost of ironmaking; at the same time, the amount of alkali consumed in the inventive method is as high as 100kg/t of concentrate, and the alkali consumption is relatively high. Resource utilization is not high.
发明内容 Contents of the invention
为了克服上述选矿方法的不足,本发明所要解决的技术问题是在物理和化学选矿方法有效结合的基础上,提供一种成本低、回收质量和效率高、工艺简单,且操作性好的利用氧化碱浸、酸洗、脱泥及重磁再选钒钛磁铁精矿方法,实现了对钒钛磁铁精矿中钛、铁进行高效分离,提高了入炉前铁品位,减少进入高炉TiO2、S、Si、Al等杂质的含量,提高高炉利用系数,减少高炉渣的排放量,降低了炼铁成本,同时提高TiO2资源综合利用率,减少环境污染。 In order to overcome the deficiencies of the above-mentioned beneficiation methods, the technical problem to be solved by the present invention is to provide a low cost, high recovery quality and efficiency, simple process, and good operability using oxidation based on the effective combination of physical and chemical beneficiation methods. The methods of alkali leaching, pickling, desliming and re-selection of vanadium-titanium magnetite concentrate achieve efficient separation of titanium and iron in vanadium-titanium magnetite concentrate, improve the iron grade before entering the furnace, and reduce the amount of TiO 2 , The content of impurities such as S, Si, and Al can increase the utilization factor of the blast furnace, reduce the discharge of blast furnace slag, reduce the cost of ironmaking, and at the same time improve the comprehensive utilization rate of TiO2 resources and reduce environmental pollution.
为了实现本发明的目的,本发明的技术方案是这样实现的: In order to realize the purpose of the present invention, technical scheme of the present invention is achieved like this:
本发明的一种利用氧化碱浸、酸洗、脱泥及重磁再选钒钛磁铁精矿方法,其特征在于包括如下步骤: A method for re-selecting vanadium-titanium magnetite concentrate utilizing oxidative alkali leaching, pickling, desliming and re-magnetization of the present invention is characterized in that it comprises the following steps:
1)氧化碱浸 1) Oxidative alkaline leaching
将TFe含量范围为50%~55%, TiO2 含量范围为10%~15%,SiO2含量为3%~6%、Al2O3含量为3%~6%、S含量>0.5%的钒钛磁铁精矿,置于质量浓度为5%~52%的碱溶液中,加入氧化剂,然后在220℃~330℃的温度下碱浸反应0.5~2小时,将反应物进行过滤,得滤液和碱浸滤饼A,所述的滤液给入回收处理系统; The TFe content ranges from 50% to 55%, the TiO2 content ranges from 10% to 15%, the SiO2 content ranges from 3% to 6%, the Al2O3 content ranges from 3 % to 6%, and the S content > 0.5%. Vanadium-titanium magnetite concentrate is placed in an alkali solution with a mass concentration of 5% to 52%, and an oxidizing agent is added, and then alkaline leaching reaction is performed at a temperature of 220°C to 330°C for 0.5 to 2 hours, and the reactant is filtered to obtain a filtrate With alkali leaching filter cake A, described filtrate is fed into recovery treatment system;
2)酸洗 2) pickling
将步骤1)中的碱浸滤饼A加水制成固液质量比为1:1~10的矿浆,再置于质量浓度为1%~10%的H2SO4溶液中,50~90℃条件下酸洗5~60分钟,将酸洗反应物进行过滤,得滤液和酸浸滤饼B,所述的滤液给入回收处理系统; Add water to the alkali leaching filter cake A in step 1) to make a slurry with a solid-to-liquid mass ratio of 1:1-10, and then place it in a H 2 SO 4 solution with a mass concentration of 1%-10%, at 50-90°C pickling under conditions for 5 to 60 minutes, and filtering the pickling reactants to obtain the filtrate and acid leaching filter cake B, and the filtrate is fed into the recovery treatment system;
3)脱泥 3) Desliming
将步骤2)中的酸浸滤饼B加水配制成质量浓度为21%~25%的矿浆,用∮3.0~5.0米的脱泥斗进行脱泥作业,得到沉砂C和溢流D; Add water to the acid leaching filter cake B in step 2) to prepare a slurry with a mass concentration of 21% to 25%, and use a ∮ 3.0 to 5.0 meter desliming bucket to perform desliming operations to obtain sand settling C and overflow D;
4) 重磁再选 4) Gravity and magnetic re-election
将步骤3)中的沉砂C加水配制配成质量浓度36%~41%的矿浆进行重选,得重选精矿E和重选尾矿F;再将重选尾矿F加水配制成质量浓度30%~34%的矿浆进行磁选, 得磁选精矿G和磁选尾矿H; Add water to the grit C in step 3) to make a slurry with a mass concentration of 36% to 41% for gravity separation to obtain gravity concentrate E and gravity tailings F; then add water to the gravity separation tailings F to make mass The slurry with a concentration of 30% to 34% is subjected to magnetic separation to obtain magnetic separation concentrate G and magnetic separation tailings H;
所述的重选精矿E与磁选精矿G合并为TFe含量范围为64%~70%的最终铁精矿,所述的磁选尾矿H与溢流D合并为TiO2含量范围为60%~82%的最终钛精矿。 The gravity separation concentrate E and the magnetic separation concentrate G are merged into the final iron concentrate with a TFe content ranging from 64% to 70%, and the magnetic separation tailings H and the overflow D are combined into a TiO2 content ranging from 60% to 82% of the final titanium concentrate.
所述的碱溶液为NaOH水溶液、KOH水溶液或NaOH和KOH混合水溶液中的任意一种。 The alkaline solution is any one of NaOH aqueous solution, KOH aqueous solution or mixed aqueous solution of NaOH and KOH.
所述的氧化剂为氧气或H2O2,氧气加入量为20~120psi、H2O2加入量为50~200kg/t给矿。 The oxidizing agent is oxygen or H 2 O 2 , the amount of oxygen added is 20-120 psi, and the amount of H 2 O 2 added is 50-200 kg/t of ore .
所述的重选采用¢0.6~¢1.2米的螺旋溜槽进行重选。 The re-election adopts the spiral chute of ¢0.6~¢1.2m for re-election.
所述的磁选采用0.12T~0.15T的筒式磁选机进行磁选。 The magnetic separation adopts a drum type magnetic separator of 0.12T-0.15T for magnetic separation.
所述的磁选采用0.03T~0.05T的磁力脱水槽进行磁选。 The magnetic separation adopts a 0.03T-0.05T magnetic dehydration tank for magnetic separation.
所述的磁选分别采用0.12T~0.15T的筒式磁选机和0.03T~0.05T磁力脱水槽进行两段磁选。 The magnetic separation adopts a 0.12T-0.15T drum magnetic separator and a 0.03T-0.05T magnetic dehydration tank to carry out two-stage magnetic separation.
本发明的优点是: The advantages of the present invention are:
氧化碱浸的过程对钒钛磁铁精矿中Ti、S、Si、Al等元素进行了化学反应,形成了相应的盐,使钒钛磁铁精矿中的铁转变为氧化铁的形式。与钒钛磁铁精矿不同的是,钛铁矿原矿中SiO2含量>20%和Al2O3含量>7%远远高于钒钛磁铁精矿中SiO2含量<6%和Al2O3含量<6%,在碱浸钛铁矿原矿过程中,由于碱浸的过程将优先发生在SiO2 、Al2O3等矿物上,使得碱浸钒钛磁铁精矿比碱浸钛铁矿原矿碱用量更少,同时O2的引入使含S化合物氧化,氧化了FeTiO3,加速了反应,降低了反应温度,缩短了反应时间,效果更好,大大降低能耗和设备投资。例如,用NaOH氧化碱浸时,本发明消耗的碱量小于90kg/t精矿,比碱浸原矿消耗的碱量469kg/t原矿降低了5.2倍以上,比未通入氧气的碱浸消耗的碱量降低了10kg/t精矿;O2的引入使碱浸反应温度最低降至220℃,反应时间小于2小时。 The process of oxidative alkali leaching chemically reacts Ti, S, Si, Al and other elements in the vanadium-titanium magnetite concentrate to form corresponding salts, which converts the iron in the vanadium-titanium magnetite concentrate into the form of iron oxide. Different from vanadium-titanium magnetite concentrate, the SiO 2 content > 20% and Al 2 O 3 content > 7% in ilmenite raw ore are much higher than the SiO 2 content < 6% and Al 2 O 3 content in vanadium-titanium magnetite concentrate. 3 content <6%, in the process of alkali leaching ilmenite raw ore, since the alkali leaching process will preferentially occur on SiO 2 , Al 2 O 3 and other minerals, the alkali leaching vanadium-titanium magnetite The amount of raw ore alkali is less, and the introduction of O 2 oxidizes the S-containing compound, oxidizes FeTiO 3 , accelerates the reaction, reduces the reaction temperature, shortens the reaction time, and has better effect, greatly reducing energy consumption and equipment investment. For example, when using NaOH oxidation alkali leaching, the amount of alkali consumed by the present invention is less than 90kg/t concentrate, which is more than 5.2 times lower than the amount of alkali consumed by alkali leaching raw ore 469kg/t raw ore . The amount of alkali is reduced by 10kg/t of concentrate; the introduction of O 2 makes the alkali leaching reaction temperature drop to 220°C, and the reaction time is less than 2 hours.
酸洗过程有效地溶解了碱浸后的Ti、Si、Al等含氧酸盐和硫化物,使之与铁精矿解离。另外由于本发明采用硫酸进行酸洗,反应条件温和,对设备腐蚀小,成本低,更利于工业化生产。 The pickling process effectively dissolves the oxo acid salts and sulfides such as Ti, Si, Al after alkali leaching, and dissociates them from the iron concentrate. In addition, since the present invention uses sulfuric acid for pickling, the reaction conditions are mild, the equipment is less corroded, the cost is low, and it is more favorable for industrialized production.
脱泥过程按矿物的粒度和比重分级,碱浸后生成的钛化合物比磁铁矿物的粒度细,比重小,钛、铁的比重差异较大,实现了钛、铁的有效分离。 The desliming process is classified according to the particle size and specific gravity of the minerals. The titanium compound produced after alkaline leaching has a finer particle size and a smaller specific gravity than the magnetite minerals. The specific gravity of titanium and iron is quite different, realizing the effective separation of titanium and iron.
再加上重磁联合选矿,使铁精矿品位由50%~55%提高到64%~70%,同时分离出的铁精矿中S含量大幅降低,由0.50%以上降至小于0.10%,SiO2含量由3%~6%降至1%以下,Al2O3含量由3%~6%降至1.8%以下,TiO2含量由12%以上降至6%以下;同时,还可以得到TiO2含量为60%~82%的钛精矿。 Coupled with the combination of gravity and magnetic beneficiation, the grade of iron concentrate is increased from 50% to 55% to 64% to 70%. At the same time, the S content in the separated iron concentrate is greatly reduced from more than 0.50% to less than 0.10%. The content of SiO 2 is reduced from 3% to 6% to less than 1%, the content of Al 2 O 3 is reduced from 3% to 6% to less than 1.8%, and the content of TiO 2 is reduced from more than 12% to less than 6%. At the same time, it can also be obtained Titanium concentrate with a TiO 2 content of 60% to 82%.
本发明综合运用氧化碱浸、酸洗、脱泥及重磁联合再选钒钛磁铁精矿,实现了钒钛磁铁精矿中钛、铁高效分离,减少进入高炉TiO2、S、Si、Al等杂质的含量,提高高炉利用系数,减少高炉渣的排放量,降低了炼铁成本,为后续冶炼创造了更好的条件,同时提高了钛资源的综合利用率。 The present invention comprehensively utilizes oxidative alkali leaching, pickling, desliming and gravity-magnetism combined re-selection of vanadium-titanium magnetite concentrate, realizes efficient separation of titanium and iron in vanadium-titanium magnetite concentrate, and reduces TiO 2 , S, Si, Al entering blast furnace The content of other impurities can increase the utilization factor of blast furnace, reduce the discharge of blast furnace slag, reduce the cost of ironmaking, create better conditions for subsequent smelting, and improve the comprehensive utilization rate of titanium resources.
附图说明 Description of drawings
图1是本发明工艺流程图。 Fig. 1 is a process flow diagram of the present invention.
图2是本发明采用两段磁选的工艺流程图。 Fig. 2 is the process flow diagram of the present invention adopting two-stage magnetic separation.
图3是本发明采用两段磁选的另一个实施例的工艺流程图。 Fig. 3 is a process flow diagram of another embodiment of the present invention using two-stage magnetic separation.
具体实施方式 Detailed ways
下面结合附图对本发明的具体实施方式做进一步说明: The specific embodiment of the present invention will be further described below in conjunction with accompanying drawing:
实施例1 : Example 1:
如图1所示。 As shown in Figure 1.
1)氧化碱浸 1) Oxidative alkaline leaching
将TFe含量为51.0%, TiO2 含量为14.1%,SiO2含量为4.57%、Al2O3含量为4.79%、S含量0.77%的钒钛磁铁精矿,置于质量浓度为20%的NaOH碱溶液中,通入22psi的O2,然后在230℃的温度下碱浸反应45分钟,将反应物进行过滤,得滤液和碱浸滤饼A,NaOH消耗量53.1 kg/t给矿,所述的滤液给入回收处理系统,其化学反应式为: Put the vanadium-titanium magnetite concentrate with a TFe content of 51.0%, a TiO 2 content of 14.1%, a SiO 2 content of 4.57%, an Al 2 O 3 content of 4.79%, and a S content of 0.77% in NaOH with a mass concentration of 20%. In the alkaline solution, 22psi of O2 was introduced, and then the alkali leaching reaction was carried out at a temperature of 230°C for 45 minutes, and the reactant was filtered to obtain the filtrate and the alkali leaching filter cake A, and the NaOH consumption was 53.1 kg/t for the ore. Said filtrate is fed into the recovery treatment system, and its chemical reaction formula is:
4mFeTiO3+8NaOH +mO2 2mFe2O3↓+4Na2O·(TiO2)m↓+4H2O m≥1 4mFeTiO 3 +8NaOH +mO 2 2mFe 2 O 3 ↓+4Na 2 O·(TiO 2 ) m ↓+4H 2 O m≥1
pFe3O4·q(FeO·TiO2) +2rNaOHpFe3O4↓+qFeO↓+ (Na2O)r·(TiO2)q↓+rH2O pFe 3 O 4 ·q(FeO·TiO 2 ) +2rNaOH pFe 3 O 4 ↓+qFeO↓+ (Na 2 O) r ·(TiO 2 ) q ↓+rH 2 O
Al2O3+2NaOH 2NaAlO2 + H2O Al 2 O 3 +2NaOH 2NaAlO 2 + H 2 O
tSiO2+2NaOH Na2O·(SiO2)t↓+ H2O tSiO 2 +2NaOH Na 2 O·(SiO 2 ) t ↓+ H 2 O
3FeS2+6NaOH 3FeS↓+Na2SO3+2Na2S+3H2O 3FeS 2 +6NaOH 3FeS↓+Na2SO3 + 2Na2S + 3H2O
4FeS2 + 11O2 2Fe2O3 + 8SO2 4FeS 2 + 11O 2 2Fe 2 O 3 + 8SO 2
4FeO+O2 2Fe2O3 4FeO+O 2 2Fe 2 O 3
2SO2+O2+4NaOH 2Na2SO4+ 2H2O 2SO 2 +O 2 +4NaOH 2Na 2 SO 4 + 2H 2 O
2)酸洗 2) pickling
将步骤1)中的碱浸滤饼A加水制成质量固液比为1∶2的矿浆,再置于质量浓度为3%的H2SO4溶液中,75℃酸洗10分钟,将酸洗反应物进行过滤,得滤液和酸浸滤饼B,所述的滤液给入回收处理系统,其化学反应式为: Add water to the alkali leaching filter cake A in step 1) to make a slurry with a mass solid-to-liquid ratio of 1: 2 , then place it in a H2SO4 solution with a mass concentration of 3%, pickle at 75°C for 10 minutes, and remove the acid Wash reactant and filter, get filtrate and acid leaching filter cake B, described filtrate is fed into recovery treatment system, and its chemical reaction formula is:
Na2O·(TiO2) x +H+ (H2O)·(TiO2) x↓+Na+ Na 2 O·(TiO 2 ) x +H + (H 2 O)·(TiO 2 ) x ↓+Na +
Na2O·(SiO2)t+H+ (H2O)·(SiO2) t↓+Na+ Na 2 O·(SiO 2 ) t +H + (H 2 O)·(SiO 2 ) t ↓+Na +
NaAlO2 +4H+ Al3+ +Na++2H2O NaAlO 2 +4H + Al 3+ +Na + +2H 2 O
3)脱泥 3) Desliming
将步骤2)中的酸浸滤饼B加水配制成质量浓度为22%的矿浆,用∮3.0米的脱泥斗进行脱泥作业,得到沉砂C和溢流D; Add water to the acid leaching filter cake B in step 2) to prepare a slurry with a mass concentration of 22%, and use a ∮3.0m desliming bucket to perform desliming operations to obtain sand settling C and overflow D;
4)重磁再选 4) Gravity and magnetic re-election
将步骤3)中的沉砂C加水配制配成质量浓度37%的矿浆给入¢1.2米的螺旋溜槽进行重选,得重选精矿E和重选尾矿F; Add water to the grit C in step 3) to make a slurry with a mass concentration of 37%, and put it into a ¢1.2-meter spiral chute for gravity separation to obtain gravity separation concentrate E and gravity separation tailings F;
再将重选尾矿F加水配制成质量浓度32%的矿浆进行给入0.15T的筒式磁选机进行磁选, 得磁选精矿G和磁选尾矿H; Add water to the gravity separation tailings F to prepare a slurry with a mass concentration of 32% and feed it into a 0.15T drum magnetic separator for magnetic separation to obtain magnetic separation concentrate G and magnetic separation tailings H;
所述的重选精矿E与磁选精矿G合并为TFe含量为64.1%的最终铁精矿的,其中SiO2含量为0.94%、Al2O3 含量为1.27%、S含量为0.02%;所述的磁选尾矿H与溢流D合并为TiO2含量为61.8%的钛精矿。 The gravity separation concentrate E and the magnetic separation concentrate G are combined into a final iron concentrate with a TFe content of 64.1%, wherein the SiO2 content is 0.94 % , the Al2O3 content is 1.27%, and the S content is 0.02% ; The magnetic separation tailings H and the overflow D are merged into TiO 2 content of 61.8% titanium concentrate.
实施例2: Example 2:
如图1所示。 As shown in Figure 1.
1) 氧化碱浸 1) Oxidative alkaline leaching
将TFe含量为51.6%, TiO2 含量为13.9%,SiO2含量为4.35%、Al2O3含量为4.33%、S含量0.57%的钒钛磁铁精矿,置于质量浓度为27%的NaOH碱溶液中,通入75psi的O2,然后在315℃的温度下碱浸反应1.0小时,将反应物进行过滤,得滤液和碱浸滤饼A,NaOH消耗量83.4 kg/t给矿,所述的滤液给入回收处理系统,其化学反应式同实施例1; The vanadium-titanium magnetite concentrate with a TFe content of 51.6%, a TiO 2 content of 13.9%, a SiO 2 content of 4.35%, an Al 2 O 3 content of 4.33%, and a S content of 0.57% was placed in NaOH with a mass concentration of 27%. In the alkaline solution, 75psi of O2 was introduced, and then the alkaline leaching reaction was carried out at a temperature of 315°C for 1.0 hour, and the reactant was filtered to obtain the filtrate and the alkaline leaching filter cake A, and the NaOH consumption was 83.4 kg/t for the ore. Described filtrate is given into recovery treatment system, and its chemical reaction formula is with embodiment 1;
2)酸洗 2) pickling
将步骤1)中的碱浸滤饼A加水制成质量固液比为1:4的矿浆,再置于质量浓度为5%的H2SO4溶液中,80℃酸洗60分钟,将酸洗反应物进行过滤,得滤液和酸浸滤饼B,所述的滤液给入回收处理系统,其化学反应式同实施例1; Add water to the alkali leaching filter cake A in step 1) to make a slurry with a mass solid-to-liquid ratio of 1:4, then place it in a H 2 SO 4 solution with a mass concentration of 5%, pickle it at 80°C for 60 minutes, and remove the acid Wash reactant and filter, get filtrate and acid leaching filter cake B, described filtrate is fed into recovery treatment system, and its chemical reaction formula is with embodiment 1;
3)脱泥 3) Desliming
将步骤2)中的过滤后的钛铁精矿B加水配制成质量浓度为21%的矿浆,用∮5.0米的脱泥斗进行脱泥作业,得到沉砂C和溢流D; Add water to the filtered ilmenite concentrate B in step 2) to prepare a slurry with a mass concentration of 21%, and use a ∮5.0m desliming hopper to perform desliming operations to obtain sand settling C and overflow D;
4)重磁再选 4) Gravity and magnetic re-election
将步骤3)中的沉砂C加水配制配成质量浓度36%的矿浆给入¢0.6米的螺旋溜槽进行重选,得重选精矿E和重选尾矿F; Add water to the grit C in step 3) to prepare a slurry with a mass concentration of 36%, and feed it into a 0.6-meter spiral chute for gravity separation to obtain gravity concentration ore E and gravity separation tailings F;
再将重选尾矿F加水配制成质量浓度33%的矿浆进行给入0.12T的筒式磁选机进行磁选, 得磁选精矿G和磁选尾矿H; Then add water to the gravity separation tailings F to prepare a slurry with a mass concentration of 33% and feed it into a 0.12T drum magnetic separator for magnetic separation to obtain magnetic separation concentrate G and magnetic separation tailings H;
所述的重选精矿E与磁选精矿G合并为TFe含量为69.9%的最终铁精矿,其中SiO2含量为0.31%、Al2O3 含量为0.27%、S含量为0.01%;所述的磁选尾矿H与溢流D合并为TiO2含量为65.2%的钛精矿。 The gravity separation concentrate E and the magnetic separation concentrate G are combined into a final iron concentrate with a TFe content of 69.9%, wherein the SiO2 content is 0.31 % , the Al2O3 content is 0.27%, and the S content is 0.01%; The magnetic separation tailings H and the overflow D are combined into a titanium concentrate with a TiO content of 65.2%.
实施例3 : Example 3:
如图1所示。 As shown in Figure 1.
1) 氧化碱浸 1) Oxidative alkaline leaching
将TFe含量为52.6%, TiO2 含量为12.2%,SiO2含量为3.35%、Al2O3含量为5.37%、S含量0.71%的钒钛磁铁精矿,置于质量浓度为33%的NaOH碱溶液中,加入59kg/t给矿的H2O2,然后在255℃的温度下碱浸反应1.5小时,将反应物进行过滤,得滤液和碱浸滤饼A,NaOH消耗量85.5 kg/t给矿,所述的滤液给入回收处理系统,其化学反应式为: The vanadium-titanium magnetite concentrate with a TFe content of 52.6%, a TiO 2 content of 12.2%, a SiO 2 content of 3.35%, an Al 2 O 3 content of 5.37%, and a S content of 0.71% was placed in NaOH with a mass concentration of 33%. In the alkaline solution, add 59kg/t of H 2 O 2 fed to the ore , and then perform alkaline leaching reaction at a temperature of 255°C for 1.5 hours, and filter the reactant to obtain the filtrate and alkaline leaching filter cake A, and the NaOH consumption is 85.5 kg/ t feeds the ore, and the described filtrate is fed into the recovery treatment system, and its chemical reaction formula is:
2mFeTiO3+4NaOH+mH2O2 mFe2O3↓+2Na2O·(TiO2)m↓+(m+2)H2O m≥1 2mFeTiO 3 +4NaOH+mH 2 O 2 mFe 2 O 3 ↓+2Na 2 O·(TiO 2 ) m ↓+(m+2)H 2 O m≥1
pFe3O4·q(FeO·TiO2) +2rNaOHpFe3O4↓+qFeO↓+ (Na2O)r·(TiO2)q↓+rH2O pFe 3 O 4 ·q(FeO·TiO 2 ) +2rNaOH pFe 3 O 4 ↓+qFeO↓+ (Na 2 O) r ·(TiO 2 ) q ↓+rH 2 O
Al2O3+2NaOH 2NaAlO2 + H2O Al 2 O 3 +2NaOH 2NaAlO 2 + H 2 O
tSiO2+2NaOH Na2O·(SiO2)t↓+ H2O tSiO 2 +2NaOH Na 2 O·(SiO 2 ) t ↓+ H 2 O
3FeS2+6NaOH 3FeS↓+Na2SO3+2Na2S+3H2O 3FeS 2 +6NaOH 3FeS↓+Na2SO3 + 2Na2S + 3H2O
2FeS2 + 11H2O2 Fe2O3 +4SO2+ 11H2O 2FeS 2 + 11H 2 O 2 Fe 2 O 3 +4SO 2 + 11H 2 O
2FeO+H2O2 Fe2O3 + H2O 2FeO+H 2 O 2 Fe 2 O 3 + H 2 O
SO2+H2O2+2NaOH Na2SO4+ 2H2O SO 2 +H 2 O 2 +2NaOH Na 2 SO 4 + 2H 2 O
2)酸洗 2) pickling
将步骤1)中的碱浸滤饼A加水制成质量固液比为1:3.2的矿浆,再置于质量浓度为7.5%的H2SO4溶液中,50℃酸洗55分钟,将酸洗反应物进行过滤,得滤液和酸浸滤饼B,所述的滤液给入回收处理系统,其化学反应式同实施例1; Add water to the alkali leaching filter cake A in step 1) to make a pulp with a mass solid-to-liquid ratio of 1:3.2, then place it in a H 2 SO 4 solution with a mass concentration of 7.5%, pickle it at 50°C for 55 minutes, and remove the acid Wash reactant and filter, get filtrate and acid leaching filter cake B, described filtrate is fed into recovery treatment system, and its chemical reaction formula is with embodiment 1;
3)脱泥 3) Desliming
将步骤2)中的过滤后的钛铁精矿B加水配制成质量浓度为24%的矿浆,用∮5.0米的脱泥斗进行脱泥作业,得到沉砂C和溢流D; Add water to the filtered ilmenite concentrate B in step 2) to prepare a slurry with a mass concentration of 24%, and use a ∮5.0-meter desliming bucket to perform desliming operations to obtain sand settling C and overflow D;
4)重磁再选 4) Gravity and magnetic re-election
将步骤3)中的沉砂C加水配制配成质量浓度41%的矿浆给入¢0.9米的螺旋溜槽进行重选,得重选精矿E和重选尾矿F; Add water to the grit C in step 3) to prepare a slurry with a mass concentration of 41%, and put it into a 0.9-meter spiral chute for gravity separation to obtain gravity separation concentrate E and gravity separation tailings F;
再将重选尾矿F加水配制成质量浓度31%的矿浆给入场强为0.03T的磁力脱水槽进行磁选, 得磁选精矿G和磁选尾矿H; Then add water to the gravity separation tailings F to prepare a slurry with a mass concentration of 31%, and give it to a magnetic dehydration tank with an entry field strength of 0.03T for magnetic separation to obtain magnetic separation concentrate G and magnetic separation tailings H;
所述的重选精矿E与磁选精矿G合并为TFe含量为66.2%的最终铁精矿,其中SiO2 含量为0.55%、Al2O3 含量为0.91%、S含量为0.02%,所述的磁选尾矿H与溢流D合并为TiO2含量为63.2%的钛精矿。 The gravity separation concentrate E and the magnetic separation concentrate G are combined into a final iron concentrate with a TFe content of 66.2%, wherein the SiO2 content is 0.55 % , the Al2O3 content is 0.91%, and the S content is 0.02%. The magnetic separation tailings H and the overflow D are combined into a titanium concentrate with a TiO content of 63.2%.
实施例4 : Example 4:
如图1所示。 As shown in Figure 1.
1) 氧化碱浸 1) Oxidative alkaline leaching
将TFe含量为53.7%, TiO2 含量为11.8%,SiO2含量为4.15%、Al2O3含量为4.98%、S含量0.69%的钒钛磁铁精矿,置于质量浓度为40%的NaOH碱溶液中,通入105psi的O2,然后在280℃的温度下碱浸反应2.0小时,将反应物进行过滤,得滤液和碱浸滤饼A,NaOH消耗量88.6 kg/t给矿,所述的滤液给入回收处理系统,其化学反应式同实施例1; The vanadium-titanium magnetite concentrate with a TFe content of 53.7%, a TiO 2 content of 11.8%, a SiO 2 content of 4.15%, an Al 2 O 3 content of 4.98%, and a S content of 0.69% was placed in NaOH with a mass concentration of 40%. In the alkaline solution, 105 psi of O 2 was introduced, and then the alkaline leaching reaction was carried out at 280°C for 2.0 hours, and the reactant was filtered to obtain the filtrate and the alkaline leaching filter cake A, and the NaOH consumption was 88.6 kg/t for the ore. Described filtrate is given into recovery treatment system, and its chemical reaction formula is with embodiment 1;
2)酸洗 2) pickling
将步骤1)中的碱浸滤饼A加水制成质量固液比为1: 6的矿浆,再置于质量浓度为6.5%的H2SO4溶液中,70℃酸洗25分钟,将酸洗反应物进行过滤,得滤液和酸浸滤饼B,所述的滤液给入回收处理系统,其化学反应式同实施例1; Add water to the alkali leaching filter cake A in step 1) to make a slurry with a mass solid-to-liquid ratio of 1: 6 , then place it in a H2SO4 solution with a mass concentration of 6.5%, pickle at 70°C for 25 minutes, and remove the acid Wash reactant and filter, get filtrate and acid leaching filter cake B, described filtrate is fed into recovery treatment system, and its chemical reaction formula is with embodiment 1;
3)脱泥 3) Desliming
将步骤2)中的过滤后的钛铁精矿B加水配制成质量浓度为24%的矿浆,用∮3.0米的脱泥斗进行脱泥作业,得到沉砂C和溢流D; Add water to the filtered ilmenite concentrate B in step 2) to make a slurry with a mass concentration of 24%, and use a ∮3.0m desliming bucket to perform desliming operations to obtain sand settling C and overflow D;
4)重磁再选 4) Gravity and magnetic re-election
将步骤3)中的沉砂C加水配制配成质量浓度40%的矿浆给入¢1.2米的螺旋溜槽进行重选,得重选精矿E和重选尾矿F;将重选尾矿F加水形成质量浓度30%的矿浆给入场强为0.05T的磁力脱水槽进行磁选,得到磁选精矿G和磁选尾矿H; Add water to the grit C in step 3) to prepare a slurry with a mass concentration of 40% and put it into a ¢1.2m spiral chute for gravity separation to obtain gravity concentrate E and gravity tailings F; Add water to form an ore slurry with a mass concentration of 30%, and conduct magnetic separation in a magnetic dehydration tank with an entry field strength of 0.05T to obtain magnetic separation concentrate G and magnetic separation tailings H;
所述的重选精矿E与磁选精矿G合并为TFe含量为67.0%的最终铁精矿,其中SiO2 含量为0.62%、Al2O3 含量为0.71%、S含量为0.01%;所述的磁选尾矿H与溢流D合并为TiO2含量为71.3%的钛精矿。 The gravity separation concentrate E and the magnetic separation concentrate G are combined into a final iron concentrate with a TFe content of 67.0%, wherein the SiO2 content is 0.62 % , the Al2O3 content is 0.71%, and the S content is 0.01%; The magnetic separation tailings H and the overflow D are combined into a titanium concentrate with a TiO2 content of 71.3%.
实施例5 : Example 5:
如图2所示。 as shown in picture 2.
1)氧化碱浸 1) Oxidative alkaline leaching
将TFe含量为54.1%, TiO2 含量为11.2%,SiO2含量为3.89%、Al2O3含量为5.05%、S含量0.72%的钒钛磁铁精矿,置于质量浓度为25%的KOH碱溶液中,通入120psi的O2,然后在220℃的温度下碱浸反应2.0小时,将反应物进行过滤,得滤液和碱浸滤饼A,KOH消耗量87.8 kg/t给矿,所述的滤液给入回收处理系统,其化学反应式为: The vanadium-titanium magnetite concentrate with a TFe content of 54.1%, a TiO 2 content of 11.2%, a SiO 2 content of 3.89%, an Al 2 O 3 content of 5.05%, and a S content of 0.72% was placed in KOH with a mass concentration of 25%. In the alkaline solution, 120psi of O 2 was introduced, and then the alkali leaching reaction was carried out at 220°C for 2.0 hours, and the reactant was filtered to obtain the filtrate and the alkali leaching filter cake A, and the KOH consumption was 87.8 kg/t for the ore. Said filtrate is fed into the recovery treatment system, and its chemical reaction formula is:
4mFeTiO3+8KOH+mO2 2mFe2O3↓+4K2O·(TiO2)m↓+4H2O m≥1 4mFeTiO 3 +8KOH+mO 2 2mFe 2 O 3 ↓+4K 2 O·(TiO 2 ) m ↓+4H 2 O m≥1
pFe3O4·q(FeO·TiO2) +2rKOHpFe3O4↓+qFeO↓+ (K2O)r·(TiO2)q↓+rH2O pFe 3 O 4 ·q(FeO·TiO 2 ) +2rKOH pFe 3 O 4 ↓+qFeO↓+ (K 2 O) r ·(TiO 2 ) q ↓+rH 2 O
Al2O3+2KOH 2KAlO2 + H2O Al 2 O 3 +2KOH 2KAlO 2 + H 2 O
tSiO2+2KOH K2O·(SiO2)t↓+ H2O tSiO 2 +2KOH K 2 O·(SiO 2 ) t ↓+ H 2 O
3FeS2+6KOH3FeS↓+K2SO3+2K2S+3H2O 3FeS 2 +6KOH 3FeS↓+K 2 SO 3 +2K 2 S+3H 2 O
4FeS2 + 11O2 2Fe2O3 + 8SO2 4FeS 2 + 11O 2 2Fe 2 O 3 + 8SO 2
4FeO+O2 2Fe2O3 4FeO+O 2 2Fe 2 O 3
2SO2+O2+4KOH 2K2SO4+ 2H2O 2SO 2 +O 2 +4KOH 2K 2 SO 4 + 2H 2 O
2)酸洗 2) pickling
将步骤1)中的碱浸滤饼A加水制成质量固液比为1:8.3的矿浆,再置于质量浓度为5%的H2SO4溶液中,85℃酸洗30分钟,将酸洗反应物进行过滤,得滤液和酸浸滤饼B,所述的滤液给入回收处理系统,其化学反应式为: Add water to the alkali leaching filter cake A in step 1) to make a slurry with a mass solid-to-liquid ratio of 1:8.3, then place it in a H2SO4 solution with a mass concentration of 5 %, pickle it at 85°C for 30 minutes, and remove the acid Wash reactant and filter, get filtrate and acid leaching filter cake B, described filtrate is fed into recovery treatment system, and its chemical reaction formula is:
K 2O·(TiO2) x +H+ (H2O)·(TiO2) x↓+ K + K 2 O·(TiO 2 ) x +H + (H 2 O)·(TiO 2 ) x ↓+ K +
K2O·(SiO2)t+H+(H2O)·(SiO2) t↓+ K + K 2 O·(SiO 2 )t+H+ (H 2 O)·(SiO 2 ) t↓+ K +
KAlO2 +4H+ Al3+ +K++2H2O KAlO 2 +4H+ Al3+ +K++2H 2 O
3)脱泥 3) Desliming
将步骤2)中的过滤后的钛铁精矿B加水配制成质量浓度为23%的矿浆,用∮5.0米的脱泥斗进行脱泥作业,得到沉砂C和溢流D; Add water to the filtered ilmenite concentrate B in step 2) to prepare a slurry with a mass concentration of 23%, and use a ∮5.0-meter desliming bucket to perform desliming operations to obtain grit C and overflow D;
4)重磁再选 4) Gravity and magnetic re-election
将步骤3)中的沉砂C加水配制配成质量浓度41%的矿浆给入¢0.6米的螺旋溜槽进行重选,得重选精矿E和重选尾矿F;将重选尾矿F加水形成质量浓度33%的矿浆给入场强为0.12T的筒式磁选机进行一段磁选,得到一段磁选精矿G1和一段磁选尾矿H1;再将浓度为32%的一段磁选精矿G1用场强为0.03T的磁力脱水槽进行二段磁选,分别得二段磁选精矿G2和二段磁选尾矿H2; Add water to the grit C in step 3) to prepare a slurry with a mass concentration of 41%, and put it into a 0.6-meter spiral chute for gravity separation to obtain gravity concentrate E and gravity tailings F; Add water to form an ore slurry with a mass concentration of 33%, and conduct a stage of magnetic separation on a drum magnetic separator with an entry field strength of 0.12T to obtain a stage of magnetic separation concentrate G1 and a stage of magnetic separation tailings H1; then a stage of magnetic separation with a concentration of 32% Concentrate G1 uses a magnetic dehydration tank with a field strength of 0.03T to carry out the second-stage magnetic separation to obtain the second-stage magnetic separation concentrate G2 and the second-stage magnetic separation tailings H2;
所述的重选精矿E与磁选精矿G2合并为TFe含量为68.7%的最终铁精矿,其中SiO2含量为0.30%、Al2O3 含量为0.34%、S含量为0.01%;所述的磁选尾矿H1与H2合并为最终的尾矿,溢流D为TiO2含量为81.8%的最终钛精矿。 The gravity separation concentrate E and the magnetic separation concentrate G2 are merged into a final iron concentrate with a TFe content of 68.7%, wherein the SiO2 content is 0.30 % , the Al2O3 content is 0.34%, and the S content is 0.01%; The magnetic separation tailings H1 and H2 are merged into the final tailings, and the overflow D is the final titanium concentrate with a TiO2 content of 81.8%.
实施例6 : Embodiment 6:
如图3所示。 As shown in Figure 3.
1)氧化碱浸 1) Oxidative alkaline leaching
将TFe含量为50.7%, TiO2 含量为14.5%,SiO2含量为4.38%、Al2O3含量为4.96%、S含量0.81%的钒钛磁铁精矿,置于质量浓度为10%的KOH碱溶液中,加入102kg/t给矿的H2O2,然后在295℃的温度下碱浸反应1.0小时,将反应物进行过滤,得滤液和碱浸滤饼A,KOH消耗量78.6kg/t给矿,所述的滤液给入回收处理系统,其化学反应式为: The vanadium-titanium magnetite concentrate with a TFe content of 50.7%, a TiO 2 content of 14.5%, a SiO 2 content of 4.38%, an Al 2 O 3 content of 4.96%, and a S content of 0.81% was placed in KOH with a mass concentration of 10%. In the alkaline solution, add 102kg/t of H 2 O 2 to the ore , and then perform alkaline leaching reaction at a temperature of 295°C for 1.0 hour, and filter the reactant to obtain the filtrate and alkaline leaching filter cake A, and the KOH consumption is 78.6kg/ t feeds the ore, and the described filtrate is fed into the recovery treatment system, and its chemical reaction formula is:
2mFeTiO3+4KOH+mH2O2 mFe2O3↓+2K2O·(TiO2)m↓+(m+2)H2O m≥1 2mFeTiO 3 +4KOH+mH 2 O 2 mFe 2 O 3 ↓+2K 2 O·(TiO 2 ) m ↓+(m+2)H 2 O m≥1
pFe3O4·q(FeO·TiO2) +2rKOHpFe3O4↓+qFeO↓+ (K2O)r·(TiO2)q↓+rH2O pFe 3 O 4 ·q(FeO·TiO 2 ) +2rKOH pFe 3 O 4 ↓+qFeO↓+ (K 2 O) r ·(TiO 2 ) q ↓+rH 2 O
Al2O3+2KOH 2KAlO2 + H2O Al 2 O 3 +2KOH 2KAlO 2 + H 2 O
tSiO2+2KOH K2O·(SiO2)t↓+ H2O tSiO 2 +2KOH K 2 O·(SiO 2 ) t ↓+ H 2 O
3FeS2+6KOH 3FeS↓+K2SO3+2K2S+3H2O 3FeS 2 +6KOH 3FeS↓+K 2 SO 3 +2K 2 S+3H 2 O
2FeS2 + 11H2O2 Fe2O3 +4SO2+ 11H2O 2FeS 2 + 11H 2 O 2 Fe 2 O 3 +4SO 2 + 11H 2 O
2FeO+H2O2 Fe2O3 + H2O 2FeO+H 2 O 2 Fe 2 O 3 + H 2 O
SO2+H2O2+2KOH K2SO4+ 2H2O SO 2 +H 2 O 2 +2KOH K 2 SO 4 + 2H 2 O
2)酸洗 2) pickling
将步骤1)中的碱浸滤饼A加水制成质量固液比为1:9的矿浆,再置于质量浓度为8%的H2SO4溶液中,60℃酸洗35分钟,将酸洗反应物进行过滤,得滤液和酸浸滤饼B,所述的滤液给入回收处理系统,其化学反应式同实施例5; Add water to the alkali leaching filter cake A in step 1) to make a pulp with a mass solid-to-liquid ratio of 1:9, then place it in an 8 % H2SO4 solution, pickle at 60°C for 35 minutes, and remove the acid Wash reactant and filter, get filtrate and acid leaching filter cake B, described filtrate is fed into recovery treatment system, and its chemical reaction formula is with embodiment 5;
3)脱泥 3) Desliming
将步骤2)中的过滤后的钛铁精矿B加水配制成质量浓度为23%的矿浆,用∮3.0米的脱泥斗进行脱泥作业,得到沉砂C和溢流D; Add water to the filtered ilmenite concentrate B in step 2) to prepare a slurry with a mass concentration of 23%, and use a ∮ 3.0-meter desliming hopper to perform desliming operations to obtain sand settling C and overflow D;
4)重磁再选 4) Gravity and magnetic re-election
将步骤3)中的沉砂C加水配制配成质量浓度37%的矿浆给入¢0.9米的螺旋溜槽进行重选,得重选精矿E和重选尾矿F;将重选尾矿F加水形成质量浓度32%的矿浆用场强为0.05T的磁力脱水槽进行一段磁选,得到一段磁选精矿G1和一段磁选尾矿H1;再将浓度为34%的一段磁选精矿G1给入场强为0.15T筒式磁选机进行二段磁选,分别得二段磁选精矿G2和二段磁选尾矿H2; Add water to the grit C in step 3) to prepare a slurry with a mass concentration of 37% and put it into a ¢0.9-meter spiral chute for gravity separation to obtain gravity concentrate E and gravity tailings F; Add water to form a pulp with a mass concentration of 32% and use a magnetic dehydration tank with a field strength of 0.05T to perform a stage of magnetic separation to obtain a stage of magnetic separation concentrate G1 and a stage of magnetic separation tailings H1; G1 conducts the second-stage magnetic separation to the drum magnetic separator with an entrance field strength of 0.15T, and obtains the second-stage magnetic separation concentrate G2 and the second-stage magnetic separation tailings H2;
所述的重选精矿E与磁选精矿G2合并为TFe含量为65.5%的最终铁精矿,其中SiO2含量为0.42%、Al2O3 含量为0.27%、S含量为0.01%;所述的磁选尾矿H1与H2合并为最终的尾矿,溢流D为TiO2含量为67.1%的最终钛精矿。 The gravity separation concentrate E and the magnetic separation concentrate G2 are merged into a final iron concentrate with a TFe content of 65.5%, wherein the SiO2 content is 0.42 % , the Al2O3 content is 0.27%, and the S content is 0.01%; The magnetic separation tailings H1 and H2 are merged into the final tailings, and the overflow D is the final titanium concentrate with a TiO2 content of 67.1%.
实施例7 : Embodiment 7:
如图3所示。 As shown in Figure 3.
1)氧化碱浸 1) Oxidative alkaline leaching
将TFe含量为52.1%, TiO2 含量为13.0%,SiO2含量为4.26%、Al2O3含量为4.85%、S含量0.84%的钒钛磁铁精矿,置于NaOH质量浓度为22%和KOH质量浓度为5%的混合碱溶液中,通入89psi的O2,然后在260℃的温度下碱浸反应75分钟,将反应物进行过滤,得滤液和碱浸滤饼A,NaOH消耗量32.6kg/t给矿、KOH消耗量49.3 kg/t给矿,所述的滤液给入回收处理系统,其化学反应式同实施例1及实施例5; The content of TFe is 52.1%, the content of TiO 2 is 13.0%, the content of SiO 2 is 4.26%, the content of Al 2 O 3 is 4.85%, and the vanadium-titanium magnetite concentrate with S content of 0.84% is placed in NaOH mass concentration of 22% and Into the mixed alkaline solution with KOH mass concentration of 5%, 89psi of O 2 was introduced, and then the alkali leaching reaction was carried out at 260°C for 75 minutes, and the reactant was filtered to obtain the filtrate and alkali leaching filter cake A, NaOH consumption 32.6kg/t ore feeding, KOH consumption 49.3 kg/t ore feeding, the filtrate is fed into the recovery treatment system, and its chemical reaction formula is the same as that of embodiment 1 and embodiment 5;
2)酸洗 2) pickling
将步骤1)中的碱浸滤饼A加水制成质量固液比为1:9的矿浆,再置于质量浓度为9%的H2SO4溶液中,90℃酸洗20分钟,将酸洗反应物进行过滤,得滤液和酸浸滤饼B,所述的滤液给入回收处理系统,其化学反应式同实施例1及实施例5; Add water to the alkali leaching filter cake A in step 1) to make a pulp with a mass solid-to-liquid ratio of 1 :9, then place it in a H2SO4 solution with a mass concentration of 9%, pickle at 90°C for 20 minutes, and remove the acid Wash reactant and filter, get filtrate and acid leaching filter cake B, described filtrate is fed into recovery treatment system, and its chemical reaction formula is with embodiment 1 and embodiment 5;
3)脱泥 3) Desliming
将步骤2)中的过滤后的钛铁精矿B加水配制成质量浓度为22%的矿浆,用∮3.0米的脱泥斗进行脱泥作业,得到沉砂C和溢流D; Add water to the filtered ilmenite concentrate B in step 2) to prepare a slurry with a mass concentration of 22%, and use a ∮ 3.0-meter desliming bucket to perform desliming operations to obtain grit C and overflow D;
4)重磁再选 4) Gravity and magnetic re-election
将步骤3)中的沉砂C加水配制配成质量浓度40%的矿浆给入¢0.9米的螺旋溜槽进行重选,得重选精矿E和重选尾矿F;将重选尾矿F加水形成质量浓度31%的矿浆用场强为0.05T的磁力脱水槽进行一段磁选,得到一段磁选精矿G1和一段磁选尾矿H1;再将浓度为31%的一段磁选精矿G1给入场强为0.15T筒式磁选机进行二段磁选,分别得二段磁选精矿G2和二段磁选尾矿H2; Add water to the grit C in step 3) to prepare a pulp with a mass concentration of 40% and put it into a 0.9-meter spiral chute for gravity separation to obtain gravity concentrate E and gravity tailings F; Add water to form a pulp with a mass concentration of 31%. Use a magnetic dehydration tank with a field strength of 0.05T to perform a stage of magnetic separation to obtain a stage of magnetic separation concentrate G1 and a stage of magnetic separation tailings H1; G1 conducts the second-stage magnetic separation to the drum magnetic separator with an entrance field strength of 0.15T, and obtains the second-stage magnetic separation concentrate G2 and the second-stage magnetic separation tailings H2;
所述的重选精矿E与磁选精矿G2合并为TFe含量为66.4%的最终铁精矿,其中SiO2含量为0.45%、Al2O3 含量为0.29%、S含量为0.01%;所述的磁选尾矿H1与H2合并为最终的尾矿,溢流D为TiO2含量为68.2%的最终钛精矿。 The gravity separation concentrate E and the magnetic separation concentrate G2 are merged into a final iron concentrate with a TFe content of 66.4%, wherein the SiO2 content is 0.45 % , the Al2O3 content is 0.29%, and the S content is 0.01%; The magnetic separation tailings H1 and H2 are merged into the final tailings, and the overflow D is the final titanium concentrate with a TiO2 content of 68.2%.
实施例8 : Embodiment 8:
如图3所示。 As shown in Figure 3.
1)氧化碱浸 1) Oxidative alkaline leaching
将TFe含量为52.6%, TiO2 含量为12.7%,SiO2含量为4.46%、Al2O3含量为5.05%、S含量0.74%的钒钛磁铁精矿,置于NaOH质量浓度为20%和KOH质量浓度为6%的混合碱溶液中,加入192kg/t给矿的H2O2,然后在265℃的温度下碱浸反应70分钟,将反应物进行过滤,得滤液和碱浸滤饼A,NaOH消耗量30.6kg/t给矿,KOH消耗量46.8kg/t给矿,所述的滤液给入回收处理系统,其化学反应式同实施例3及实施例6; The vanadium-titanium magnetite concentrate with TFe content of 52.6%, TiO2 content of 12.7%, SiO2 content of 4.46%, Al2O3 content of 5.05%, and S content of 0.74% was placed in NaOH mass concentration of 20% and Add 192kg /t of H 2 O 2 to the mixed alkali solution with a KOH mass concentration of 6%, and then perform an alkali leaching reaction at a temperature of 265°C for 70 minutes, and filter the reactant to obtain a filtrate and an alkali leaching filter cake A, NaOH consumption 30.6kg/t feeds ore, KOH consumption 46.8kg/t feeds ore, and described filtrate is fed into recovery treatment system, and its chemical reaction formula is the same as embodiment 3 and embodiment 6;
2)酸洗 2) pickling
将步骤1)中的碱浸滤饼A加水制成质量固液比为1:8.5的矿浆,再置于质量浓度为9.5%的H2SO4溶液中,80℃酸洗30分钟,将酸洗反应物进行过滤,得滤液和酸浸滤饼B,所述的滤液给入回收处理系统,其化学反应式同实施例1及实施例5; Add water to the alkali leaching filter cake A in step 1) to make a slurry with a mass solid-to-liquid ratio of 1 :8.5, then place it in a H2SO4 solution with a mass concentration of 9.5%, pickle it at 80°C for 30 minutes, and remove the acid Wash reactant and filter, get filtrate and acid leaching filter cake B, described filtrate is fed into recovery treatment system, and its chemical reaction formula is with embodiment 1 and embodiment 5;
3)脱泥 3) Desliming
将步骤2)中的过滤后的钛铁精矿B加水配制成质量浓度为22%的矿浆,用∮3.0米的脱泥斗进行脱泥作业,得到沉砂C和溢流D; Add water to the filtered ilmenite concentrate B in step 2) to prepare a slurry with a mass concentration of 22%, and use a ∮ 3.0-meter desliming bucket to perform desliming operations to obtain grit C and overflow D;
4)重磁选再 4) Gravity and magnetic separation
将步骤3)中的沉砂C加水配制配成质量浓度40%的矿浆给入¢0.9米的螺旋溜槽进行重选,得重选精矿E和重选尾矿F;将重选尾矿F加水形成质量浓度31%的矿浆用场强为0.05T的磁力脱水槽进行一段磁选,得到一段磁选精矿G1和一段磁选尾矿H1;再将浓度为31%的一段磁选精矿G1给入场强为0.15T筒式磁选机进行二段磁选,分别得二段磁选精矿G2和二段磁选尾矿H2; Add water to the grit C in step 3) to prepare a pulp with a mass concentration of 40% and put it into a 0.9-meter spiral chute for gravity separation to obtain gravity concentrate E and gravity tailings F; Add water to form a pulp with a mass concentration of 31%. Use a magnetic dehydration tank with a field strength of 0.05T to perform a stage of magnetic separation to obtain a stage of magnetic separation concentrate G1 and a stage of magnetic separation tailings H1; G1 conducts the second-stage magnetic separation to the drum magnetic separator with an entrance field strength of 0.15T, and obtains the second-stage magnetic separation concentrate G2 and the second-stage magnetic separation tailings H2;
所述的重选精矿E与磁选精矿G2合并为TFe含量为67.2%的最终铁精矿,其中SiO2含量为0.47%、Al2O3 含量为0.89%、S含量为0.01%;所述的磁选尾矿H1与H2合并为最终的尾矿,溢流D为TiO2含量为60.8%的最终钛精矿。 The gravity separation concentrate E and the magnetic separation concentrate G2 are combined into a final iron concentrate with a TFe content of 67.2%, wherein the SiO2 content is 0.47 % , the Al2O3 content is 0.89%, and the S content is 0.01%; The magnetic separation tailings H1 and H2 are merged into the final tailings, and the overflow D is the final titanium concentrate with a TiO2 content of 60.8%.
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