鲕状赤铁矿磁化焙烧—磁选—反浮选降磷试验

鄂西宁乡式鲕状赤铁矿嵌布粒度极细,SiO2、Al2O3、P等杂质含量高,用其生产的铁精矿很难达到冶炼要求.针对铁品位为43.76%,磷含量为0.84%的鄂西鲕状赤铁矿进行提铁降磷试验研究,通过对磁化焙烧温度、磁化焙烧时间、还原煤的配比等影响因素的条件试验,确定在焙烧时间60 min,焙烧温度750℃,还原煤11%(质量比)的最佳焙烧条件.焙烧产品磨矿至-0.038 mm占80.54%、用永磁选机进行弱磁选,获得了铁品位54.10%、铁回收率93.19%、磷含量0.80%的粗铁精矿.进行反浮选药剂制度试验,得到了铁品位58.95%、铁综合回收率80%、磷含量0.50%的铁精矿,其最佳浮选药剂制度为NaOH 750 g/t,淀粉800 g/t,石灰500 g/t,RA-715 750 g/t,G310 107.73 g/t,浮选温度30℃.在此浮选制度下,进行一粗一精试验,精选石灰和捕收剂用量减半,可得铁品位59.87%,磷含量降至0.28%,综合回收率71.08%,综合试验结果表明,本文探索的工艺流程具有很大的可行性,能够为鲕状赤铁矿的选矿利用提供参考.     

鲕状赤铁矿选矿与脱磷技术动态

鲕状赤铁矿铁品位高、储量大,因其特殊的结构且含磷较高,成为国内外选矿技术的难点之一.对鲕状赤铁矿的选矿工艺从直接入选、磁化焙烧-磁选、深度还原3类进行了系统归纳和总结,并介绍了3类工艺的分选机理.指出直接入选能耗低,污染少,但对磨矿细度有要求,矿石泥化严重;磁化焙烧-磁选是目前处理鲕状赤铁矿行之有效的办法,精矿指标较好,但磷含量较高;深度还原法处理鲕状赤铁矿,能够进一步提高精矿指标,但能耗高、工艺复杂.此外,从物理法、化学法和微生物法3个方面评述了鲕状赤铁矿有害杂质磷的脱除,建议结合技术经济来考察和分析其脱磷效果.     

冷却方式对焙烧鲕状赤铁矿磨矿性能影响

鲕状赤铁矿有用矿物嵌布粒度细,磨矿成本高,属于典型难选铁矿石。磁化焙烧-磁选工艺是分选此类难选铁矿石的有效方法,研究了不同冷却方式对磁化焙烧矿的磨矿性能产生的影响.鲕状赤铁矿进行磁化焙烧后分别隔绝空气密闭冷却、水淬及空气中自然冷却,进行粒度筛析和磨矿试验.发现焙烧后矿石粒径变粗,从原矿的平均粒径为0.069 9mm至密闭冷却、水冷和自然冷却焙烧矿的0.088 2mm、0.084 3mm、0.087 0mm.相同磨矿条件下,原矿、密闭冷却焙烧矿、水冷焙烧矿和自然冷却焙烧矿-0.045mm含量分别为84.89%、83.89%、76.51%、77.14%.表明磁化焙烧使鲕状赤铁矿变得更为难磨,密闭冷却时磨矿效果最佳,自然冷却次之,水冷最差.     

河北某鲕状(菱)赤铁矿选矿试验研究

对河北某地含铁品位38.57%的鲕状(菱)赤铁矿进行了选矿试验研究,考察了该矿石的工艺矿物学特征,重点研究了采用磁选、浮选、磁化焙烧.弱磁选等选别工艺的分选效果,试验结果表明磁化焙烧-弱磁选工艺是分选此类难选铁矿石的有效方法.在温度750℃,焙烧时间80min,煤粉配比5%的最佳焙烧条件下,焙烧矿经弱磁选可以获得精矿铁品位为59.94%.回收率84.87%的良好指标,并通过XRD分析对磁化焙烧的反应机理进行了初步的探讨.     

低品位铁矿石流化焙烧-磁选提质试验研究

针对我国低品位铁矿石嵌布粒度极细,成分复杂,难提难选的现况,运用循环流化床和磁选管进行劣质铁矿石的流化焙烧 磁选试验研究,试验采用CO、N2的混合气体营造还原性气氛（其中CO体积分数为10%）,将粒径为1 mm以下的新疆某低品位铁矿石（原矿铁品位为9.63%）于850 ℃焙烧10 min,得到强磁性的磁铁矿,将焙烧产物破碎细磨（磨至200 目以下占75%）,利用湿式磁选管在71.66 kA/m的磁场强度下进行弱磁选抛尾,可以得到铁精矿品位为46.25%,全铁回收率为25.52%的选矿指标.研究表明,运用循环流化床焙烧-弱磁选的方法提质铁矿石,可以有效地减少焙烧时间,在保证选矿达标的基础上,有效地降低生产周期.     

鲕状赤铁矿“磁化焙烧-晶粒长大-磁选”新工艺研究

基于宣龙式鲕状赤铁矿嵌布粒度极细、结构复杂等特点,进行了磁化焙烧-晶粒长大-磁选新工艺研究。在焙烧温度为800℃,煤粉配比10%,焙烧时间45min的条件下,使赤铁矿还原焙烧成磁铁矿,经过弱磁选,可得到铁精矿品位62.5%,回收率85.5%的良好选矿技术指标。通过一系列观测手段及相关理论说明,证实了磁铁矿晶粒能够长大。     

硫酸渣磁化焙烧—磁选提铁降硫

硫酸渣铁品位为55.08％,其中有害元素硫的含量为1.3％.为高效利用硫酸渣,必须提高铁含量、降低硫磷等有害元素.硫酸渣试样直接进行弱磁选,得到铁精矿品位60.54％,精矿回收率仅为54.46％,采用磁化焙烧-弱磁选的方法来进行选铁试验,通过对磁化焙烧时间、磁化焙烧温度、还原剂的质量配比等条件试验,确定了在焙烧时间40 min,焙烧温度750℃,还原剂10％的最佳焙烧条件.焙烧矿磨矿至-0.074 mm 97.02％,用弱磁选管进行磁选的最佳试验条件,在此焙烧条件下,进行一粗一精的磁选,获得了铁品位64.57％,精矿回收率86.99％,硫含量降低到0.13％.     

高磷鲕状赤褐铁矿离析焙烧提铁降磷研究

针对重庆桃花高磷鲕状赤褐铁矿中,有害元素P的质量分数较高为1.17%,有85.90%的P分布于褐铁矿中,其余以胶磷矿形式产出,提出了离析焙烧-弱磁选工艺实现提铁降磷.矿石与氯化剂、还原剂混匀后置入焙烧炉中进行离析焙烧,铁从弱磁性矿物转变为强磁性矿物后,焙烧矿采用弱磁选回收铁.结果表明:焙烧矿中产生了以磁铁矿(Fe3O4)、金属铁(Fe)为主的新矿相及少量的氧化亚铁(FeO)新矿相,实现了铁矿物与磷矿物的有效分离;在离析焙烧温度950℃、焦炭用量20%、废盐用量45%、离析焙烧时间60min、弱磁选磁场强度H=0.12T、弱磁选磨矿细度小于0.038mm占95%的综合工艺条件下,得到了Fe的质量分数为71.65%,P的质量分数为0.17%,Fe回收率为87.92%的铁精矿分选指标,提铁降磷效果显著.     

磁化矿石颗粒模型及磁选过程分析

基于磁选过程中颗粒尺寸、磁场强度和磁选精矿品位三者之间的关系,建立磁化矿石颗粒模型,对其进行理论分析与计算,确定最佳磁场强度,并进行磁化矿石的磁选研究。结果表明：在配煤量4%（质量分数）,焙烧温度850℃,焙烧时间60 min,磨矿细度-0.074 mm占60%（质量分数）,磁场强度为40 mT的条件下,得到铁品位57.7%（质量分数）,铁回收率90.3%（质量分数）的铁精矿,较好地实现了铁精矿的富集和回收。     

鲕状赤铁矿悬浮焙烧的粒度影响研究

研究了+100,100~74,74~45,-45μm 4个不同粒级对重庆巫山某鲕状赤褐铁矿悬浮焙烧过程的影响,并通过扫描电镜及能谱分析、磁性分析、X射线衍射分析、热力学分析等手段对其机理进行研究.结果表明:悬浮焙烧前后,矿物颗粒的微观形貌及嵌布特征并未发生明显变化;悬浮焙烧后,鲕状赤褐铁矿中的弱磁性铁矿物可以转变为强磁性铁矿物,除-45μm粒级外,其他粒级物料的比磁化率和磁化强度显著提高,且物料颗粒越细,其比磁化率和磁化强度就越高.鲕状赤褐铁矿石中的赤铁矿转变成磁铁矿,当物料粒度在-45μm时,有FeO相存在,说明发生了过还原反应.在一定温度条件下,赤铁矿很容易被还原为磁铁矿,且有稳定的Fe3O4相存在.因此,控制好焙烧物料的粒度对于获得较好的悬浮焙烧质量产品以及提高焙烧效率至关重要.     

Effect of coal levels during direct reduction roasting of high phosphorus oolitic hematite ore in a tunnel kiln

The effect of coal levels on phosphorus removal from a high phosphorus oolitic hematite ore after direct reduction roasting have been investigated. Raw ore, coal, and a dephosphorization agent were mixed and the mixture was then roasted in a tunnel kiln. The roasted products were treated by two stages of grinding followed by magnetic separation. XRD and SEM–EDS examination of the products was used to analyze differences in the roasted products. The results show that coal is one of the most important factors affecting the direct reduction roasting process. When the inner coal levels increased from 0% to 15% the iron grade decreased linearly from 94.94% to 88.81% and the iron recovery increased from 55.94% to 92.94%. At the same time the phosphorus content increased from 0.045% to 0.231%. Increasing the inner coal levels also caused more hematite to be reduced to metallic iron but the oolitic structure of the roasted product was preserved in the presence of high coal loading. The phase of the phosphorus in raw ore was not changed after direct reduction roasting. The effect of coal on the phosphorus content in the H-concentrate arises from changes in the difficulty of mechanically liberating the metallic iron from the phosphorus bearing minerals.     

Application of multi-stage dynamic magnetizing roasting technology on the utilization of cryptocrystalline oolitic hematite: A review

A large number of studies have shown that oolitic hematite is an iron ore that is extremely difficult to utilize because of its fine disseminated particle size, high harmful impurity content and oolitic structure. To recover iron from oolitic hematite, we developed a novel multistage dynamic magnetizing roasting technology. Compared with traditional magnetizing roasting technologies, this novel technology has the following advantages: firstly, the oolitic hematite is dynamically reduced in a multi-stage roasting furnace, which shortens the reduction time and avoids ringing and over-reduction; secondly, the novel dynamic magnetizing roasting technology has strong raw material adaptability, and the size range of raw materials can be as wide as 0–15 mm; thirdly, the roasting furnace adopts a preheating-heating process, and the low-calorific value blast furnace gas can be used as the fuel and reductant, which greatly reduces the cost. The actual industrial production data showed that the energy consumption in the roasting process can be less than 35 kg of standard coal per ton of raw ore. The iron grade of the concentrate and iron recovery reached 65% and 90%, respectively.     

Magnetizing roasting mechanism and effective ore dressing process for oolitic hematite ore

Magnetizing roasting of oolitic hematite ore from western Hubei Province was investigated.The mechanism for reduction roasting of oolitic hematite ore was discussed and analyzed.It is found that flash magnetizing roasting-magnetic separation process is a promising approach for the processing of oolitic hematite ore from western Hubei Province.     

高磷铁矿碳热还原同步脱磷的实验研究

为探索高磷铁矿的有效利用途径,对高磷鲕状赤铁矿进行碳热还原同步脱磷实验研究,在含碳球团中添加CaO和Na2C O3作为脱磷剂,采用D T A-T G- M S综合热分析、X R D、SE M、E DS等方法分别对高磷鲕状赤铁矿的碳热还原过程以及还原产物进行分析. 结果表明,添加适量的CaO和Na2C O3可以显著提高脱磷率;在1 573 K、Na2C O3添加量为2 %、含碳球团碱度为1 .2的条件下,高磷鲕状赤铁矿能够被快速还原成含磷0 .09 %、含碳4 .6 %的碳饱和铁,脱磷率达到95 %;生铁中碳过饱和后以片状石墨的形态析出,生铁中的磷以夹杂物Ca3（P O4）2和Na2Ca4（P O4）2SiO4的形式存在     

Separation of hematite from banded hematite jasper(BHJ) by magnetic coating

The separation of iron oxide from banded hematite jasper(BHJ) assaying 47.8% Fe, 25.6% Si O2 and 2.30%Al2O3 using selective magnetic coating was studied. Characterization studies of the low grade ore indicate that besides hematite and goethite,jasper, a microcrystalline form of quartzite, is the major impurity associated with this ore. Beneficiation by conventional magnetic separation technique could yield a magnetic concentrate containing 60.8% Fe with 51% Fe recovery. In order to enhance the recovery of the iron oxide minerals, fine magnetite, colloidal magnetite and oleate colloidal magnetite were used as the coating material. When subjected to magnetic separation, the coated ore produces an iron concentrate containing 60.2% Fe with an enhanced recovery of56%. The AFM studies indicate that the coagulation of hematite particles with the oleate colloidal magnetite facilitates the higher recovery of iron particles from the low grade BHJ iron ore under appropriate conditions.     

Evaluation of reduction roasting and magnetic separation for upgrading Mn/Fe ratio of fine ferromanganese

Low grade ferromanganese tailing was subjected to different mineralogical techniques, reduction roasting and magnetic separation to establish whether commercially acceptable manganese qualities and high Mn/Fe ratio could be obtained, and also to determine the best processing route for beneficiating this ore. The main manganese mineral within the feed sample is birnessite, with minor amounts of pyrolusite and todorokite. Size by assay analysis conducted presented a result with a yield of about 35.75% and Mn grade of 27.63% to coarse (?3.35 + 1 mm) and yield of 20.24% and Mn grade of 27.71% to (?1.18 + 0.50 mm) fraction. Two-stage high induced magnetic separations at 16,000 and 11,000 G produces Mn grades with similar grade to that obtained from the ferromanganese feed sample. Reduction roasting followed by magnetic separation on ?1.18 + 0.50 mm at 1000 G recovered 72.31% Mn with a grade of 58.44% Mn, 2.52% Fe and 3.29% Si at Mn/Fe ratio of 23.22. This study reveals the influence of roasting in converting the hematite and goethite to magnetite and the response of the roasted fraction to magnetic separation.