Mechanism of phosphorus removal in beneficiation of high phosphorous oolitic hematite by direct reduction roasting with dephosphorization agent

High phosphorous oolitic hematite ore is one of typical intractable iron ores in China, and the conventional beneficiation methods are found to be impracticable to , remove phosphorus from the ore effectively. Better beneficiation index were gotten by direct reduction roasting with dephosphorization agent followed by two stages of grinding and magnetic separation. P content decreases from 0.82% in the raw ore to 0.06% in the magnetic concentrate, and the total iron grade increases from 43.65% to 90.23%, the recovery of iron can reach 87%. Mechanisms of phosphorus removal in the beneficiation of high phosphorous oolitic hematite ore by direct reduction roasting with dephosphorization agent were studied using XRD, SEM and EPMA. The results showed that about 20% of the apatite in the raw ore transferred into phosphorus and volatilized with the gas in the process of reduction roasting, while the rest 80% apatite was not involved in the reaction of generation of phosphorus, and remained as apatite in the roasted products, which was removed to tailings by grinding and magnetic separation. A small amount of phosphorus existed in the magnetic concentrate as apatite. The oolitic texture of raw ore was partly changed during roasting, resulting in the formation of nepheline in the reaction between the dephosphorization agent, SiO2 and Al2O3 in the raw ore, which greatly improved the liberation degree of minerals in the roasted products, and it was beneficial to the subsequent grinding and magnetic separation.

Behaviors of vanadium and chromium in coal-based direct reduction of high-chromium vanadium-bearing titanomagnetite concentrates followed by magnetic separation

The reduction behaviors of FeO·V2O3 and FeO·Cr2O3 during coal-based direct reduction have a decisive impact on the efficient utilization of high-chromium vanadium-bearing titanomagnetite concentrates. The effects of molar ratio of C to Fe n(C)/n(Fe) and temperature on the behaviors of vanadium and chromium during direct reduction and magnetic separation were investigated. The reduced samples were characterized by X-ray diffraction(XRD), scanning election microscopy(SEM) and energy dispersive spectrometry(EDS) techniques. Experimental results indicate that the recoveries of vanadium and chromium rapidly increase from 10.0% and 9.6% to 45.3% and 74.3%, respectively, as the n(C)/n(Fe) increases from 0.8 to 1.4. At n(C)/n(Fe) of 0.8, the recoveries of vanadium and chromium are always lower than 10.0% in the whole temperature range of 1100-1250 °C. However, at n(C)/n(Fe) of 1.2, the recoveries of vanadium and chromium considerably increase from 17.8% and 33.8% to 42.4% and 76.0%, respectively, as the temperature increases from 1100 °C to 1250 °C. At n(C)/n(Fe) lower than 0.8, most of the FeO·V2O3 and FeO·Cr2O3 are not reduced to carbides because of the lack of carbonaceous reductants, and the temperature has little effect on the reduction behaviors of FeO·V2O3 and FeO·Cr2O3, resulting in very low recoveries of vanadium and chromium during magnetic separation. However, at higher n(C)/n(Fe), the reduction rates of FeO·V2O3 and FeO·Cr2O3 increase significatly because of the excess amount of carbonaceous reductants. Moreover, higher temperatures largely induce the reduction of FeO·V2O3 and FeO·Cr2O3 to carbides. The newly formed carbides are then dissolved in the γ(FCC) phase, and recovered accompanied with the metallic iron during magnetic separation.

Effect of reducing agents on reducing atmosphere in coal-based direct reduction of beach titanomagnetite

Beach titanomagnetite(TTM)provides a cheap alternative source of Fe and Ti,but this ore is difficult to process to make suitable concentrates for the blast furnace.Recently studies showed that it is feasible to separate Fe and Ti by coal-based direct reduction.In this study,beach TTM was selected as the research object,the effects of reducing agents on reducing atmosphere in coal-based direct reduction of beach TTM were analyzed,and the role of volatiles was also studied.The results showed that when bitumite and coke were used as reducing agents of TTM,the CO produced from volatiles was involved in the reduction reaction,and the generated CO_(2) provided the raw material for the reaction of TTM.The reduction effect of bitumite was better than that of coke.The reason is that bitumite+TTM had a higher gas generation rate and produced a higher CO partial pressure,while coke+TTM had a lower gas generation rate and produced a lower CO partial pressure.When graphite was used as a reducing agent,there was a solid-solid reaction in the early stage in the reaction.With the continuous accumulation of CO_(2),the Boudouad reaction started and accelerated.Graphite+TTM also produced a higher CO partial pressure.

Characterization of semi-coke generated by coal-based direct reduction process of siderite

Solid wastes derived from metallurgical industries pose a significant threat to environment. The utilization and disposal of these solid wastes are the major concern in the world. Semi-coke generated in coal-based direct reduction process of iron ore is a by-product and its suitable utilization is not available so far. In order to handle it properly, the characteristics of this by-product were comprehensively investigated. A series of analysis methods were used to demonstrate its mineral compositions, petrography and physico-chemical properties. The results reveal that the semi-coke has poor washability. The fixed carbon content of semi-coke reaches 76.11% and the gross calorific value is 28.10 MJ/kg, both of which are similar to those of traditional sinter coke breeze. Also, semi-coke ash possesses lower content of SiO2, Al2O3, S and higher content of CaO and MgO, which could improve the strength of sinter ore when partially substituting for coke breeze in sintering. Semi-coke features well-development porous structure and higher reaction activity, which predicts that the sintering speed could be elevated to some extent when employing it as a partial replacement of coke breeze, so the studies further suggest that the potential adverse effect of the high reactivity on sintering process could be weakened by adequately coarsening the semi-coke's particle size.

Preparing high-purity iron by direct reduction?smelting separation of ultra-high-grade iron concentrate

A new process for preparing high-purity iron(HPI)was proposed,and it was investigated by laboratory experiments and pilot tests.The results show that under conditions of a reduced temperature of 1075°C,reduced time of 5 h,and CaO content of 2.5wt%,a DRI with a metallization rate of 96.5%was obtained through coal-based direct reduction of ultra-high-grade iron concentrate.Then,an HPI with a Fe purity of 99.95%and C,Si,Mn,and P contents as low as 0.0008wt%,0.0006wt%,0.0014wt%,and 0.0015wt%,respectively,was prepared by smelting separation of the DRI using a smelting temperature of 1625°C,smelting time of 45 min,and CaO content of 9.3wt%.The product of the pilot test with a scale of 0.01 Mt/a had a lower impurity content than the Chinese industry standard.An HPI with a Fe purity of 99.98wt%can be produced through the direct reduction?smelting separation of ultra-high-grade iron concentrate at relatively low cost.The proposed process shows a promising prospect for application in the future.

Recovery of iron and copper from copper tailings by coal-based direct reduction and magnetic separation

A technique comprising coal-based direct reduction followed by magnetic separation was presented to recover iron and copper from copper slag flotation tailings.Optimal process parameters,such as reductant and additive ratios,reduction temperature,and reduction time,were experimentally determined and found to be as follows：a limestone ratio of 25%,a bitumite ratio of 30%,and reduction roasting at 1473 Kfor 90 min.Under these conditions,copper-bearing iron powders（CIP）with an iron content of 90.11% and copper content of 0.86%,indicating iron and copper recoveries of87.25% and 83.44%respectively,were effectively obtained.Scanning electron microscopy and energy dispersive spectroscopy of the CIP revealed that some tiny copper particles were embedded in metal iron and some copper formed alloy with iron,which was difficult to achieve the separation of these two metals.Thus,the copper went into magnetic products by magnetic separation.Adding copper into the steel can produce weathering steel.Therefore,the CIP can be used as an inexpensive raw material for weathering steel.

红土镍矿浸渣选铁试验研究

针对某含铁51.38%、含硫2.01%的红土镍矿浸渣进行了离心选矿、磁选、磁化焙烧等试验研究。结果表明,磁化焙烧-磁选工艺是该红土镍矿浸渣选铁的高效分选工艺。煤基还原-弱磁选获得了精矿TFe品位62.77%、回收率94.04%的指标;气基还原-弱磁选获得了精矿TFe品位68.77%、回收率95.15%、精矿中杂质硫含量0.12%的指标,可达到铁精矿质量要求;直接还原-弱磁选可获得精矿TFe品位90.89%、回收率88.90%的指标,但存在精矿中硫含量偏高的问题。

难选鲕状赤铁矿焙烧-磁选和直接还原工艺的探讨

针对难选鲕状赤铁矿,在实验室条件下采用磁化焙烧-磁选工艺制取铁精矿和直接还原工艺制取海绵铁,研究了还原时间、温度、还原剂用量等对两种焙烧过程的影响。研究结果表明:采用无烟煤作还原剂,在850℃时焙烧45 min的焙烧矿经过磁选后获得铁精矿品位达到61.60%,回收率达到96.65%的较好指标;采用直接还原在环状装料方式下还原焙烧,采用无烟煤和碳酸钙的混合物为还原剂,1 050℃时焙烧5.0 h,经过磁选得到的海绵铁的品位、金属化率和回收率可分别达到了89%、90%和85%。

煤种类及添加剂对红土镍矿选择性直接还原的影响规律

以含Ni 1.49%,Fe 34.69%的红土镍矿为研究对象,采用煤基直接还原法选择性还原镍铁矿物,研究并分析了焙烧过程中还原剂和添加剂种类及用量、焙烧温度以及焙烧时间对镍铁选择性还原的影响规律。结果表明:以宁夏烟煤为还原剂,NCS为添加剂,1 200℃焙烧50 min,磁选得到镍铁产品中含镍9.51%,镍的回收率为84.04%,镍铁回收率差为54.49%。通过X射线衍射(XRD)、扫描电镜(SEM)及X射线能谱分析(EDS)等测试手段分析了磁选镍铁产品中镍铁的存在形式,结果表明:红土镍矿直接还原过程中铁矿物大部分被还原成浮士体,仅有少部分铁矿物被还原成金属铁,并与镍矿物还原金属镍形成铁纹石和镍纹石,实现了红土镍矿中镍铁的选择性还原。

硫酸渣直接还原焙烧提铁降硫影响因素研究

对硫酸渣直接还原过程中脱硫剂的种类与用量、焙烧温度、不同升温方式、焙烧时间等因素对焙烧过程的影响进行了研究。确定的最佳工艺条件为:还原剂用量为30%,脱硫剂SH用量为15%,随炉升温,焙烧温度为1 200℃,焙烧时间为60 min。在此条件下可得到铁品位为90.13%、铁回收率为95.06%、硫含量为0.040%的还原铁。X射线衍射和扫描电镜分析结果表明,硫酸渣中的硫与脱硫剂反应生成了硫化钙,且焙烧温度的升高和随炉升温都有利于硫化钙的生成;通过磨矿-磁选方法可将硫化钙与金属铁分离,从而达到脱硫目标。

从内蒙古某高硫铁尾矿中回收铁的研究

内蒙古某硫铁矿属以硫为主、伴生低品位铜锌的复杂硫化矿石,经浮选流程产生了铁品位为17.75%、硫含量为5.87%的高硫铁尾矿。针对此高硫铁尾矿进行了磁选、摇床、磁选-反浮选和直接还原焙烧-磁选等一系列提铁降硫的探索试验研究。结果表明,采用常规选矿方法很难达到理想的分选效果;而采用直接还原焙烧-磁选方法可获得铁品位为93.57%、硫含量为0.39%、对弱磁精矿的回收率为82.01%的直接还原铁产品,为有效提高资源综合利用率提供了新的途径。

流化床处理粉铁矿工艺研究

评述了流化床处理粉铁矿的工艺,根据流化床反应器在工业中的应用实际,分别评述了流化床磁化焙烧生产铁精矿,流化床预热粉铁矿并低预还原度还原后添加到铁浴熔池,流化床生产直接还原铁(DRI)和碳化铁(Fe3C)作电炉炼钢原料,多级流化床生产较高预还原度DRI供熔融气化炉终还原冶炼铁水等工艺中流化床反应器的应用。

某氰化尾渣直接还原焙烧—磁选选铁试验

对铁品位高达48.05%的某氰化尾渣进行了直接还原焙烧—磁选试验研究。结果表明,在烟煤2用量为20%,还原温度为1 150℃,还原时间为40 min,还原产物磨矿细度为-0.074 mm占78%,1次弱磁选磁场强度为87.56 kA/m情况下获得了铁品位为94.11%、回收率为90.14%的还原铁粉。该工艺不仅高效地富集了铁,而且使硫等含量较高的有害成分富集到了尾矿中。

高磷鲕状赤铁矿直接还原同步脱磷新脱磷剂

为开发高磷鲕状赤铁矿,采用直接还原焙烧的方法对含TFe品位为43.58%,磷含量0.83%的鄂西某宁乡式高磷鮞状赤铁矿进行系统研究。通过X线衍射仪(XRD)和扫描电镜(SEM)对提铁降磷机理进行研究。研究结果表明:在配合使用NCP和新脱磷剂TS2种脱磷剂的条件下,可以获得TFe品位为91.35%、铁回收率为85.12%、磷含量为0.081%的直接还原铁粉。原矿加入脱磷剂焙烧后,含磷矿物的物相并没有发生变化,仍以氟磷石的形式存在,通过细磨磁选实现提铁降磷。加入的脱磷剂有助于破坏鲕状结构,使金属铁颗粒与脉石颗粒的接触面变得平滑、清晰,改善高金属铁和脉石的解离条件,同时脱磷剂还能促进中间产物铁橄榄石的还原。

基于还原焙烧的某海滨钛磁铁矿的钛铁分离

为了高效分离印尼某高铁钛、低硫磷海滨钛磁铁矿中的钛铁,实现资源的充分利用,采用直接还原焙烧—磨矿—弱磁选工艺对该试样进行了还原焙烧工艺技术条件研究,并对确定条件下的焙烧产物进行了不同磨矿细度下的钛铁分离验证试验。结果表明,添加剂NCS对铁还原和钛铁分离有促进作用;在烟煤A用量(与试样的质量比)为30%、NCS用量为11%、还原焙烧温度为1 250℃、还原焙烧时间为60 min、磨矿细度为-43μm占69.02%、弱磁选磁场强度为151 kA/m的情况下,可获得TFe品位为93.74%、回收率为95.91%、TiO2品位为0.45%的还原铁粉,实现了钛铁的高效分离;钛在尾矿中的富集为后续回收钛创造了条件。

某难选铁矿石直接还原焙烧磁选研究

对某含铁品位为28.82%,含磷0.35%的难选铁矿石进行了直接还原焙烧磁选研究。研究了焙烧温度、还原剂用量、焙烧时间、助溶剂用量、磨矿粒度以及磁场强度对直接还原铁品位和回收率的影响。在还原剂用量为30%,助溶剂QK用量为20%,焙烧温度为1 200℃,焙烧时间为30 min,一段磨矿粒度为-43μm粒级含量达到95%以上,二段磨矿粒度为-30μm粒级含量达到100%,一段磁选场强为111.5 kA/m,二段磁选场强为95.5 kA/m的条件下,可以获得品位为90.94%,回收率为82.67%的直接还原铁。

煤的挥发分对高磷鲕状赤铁矿直接还原焙烧的影响

在相同还原焙烧条件和等量添加剂下,系统研究了煤基还原剂中的挥发分对鄂西高磷鲕状赤铁矿在直接还原焙烧过程中对含铁和磷矿物的影响.通过XRD及SEM分析对煤基中挥发分的作用机理进行初步阐述.研究结果表明,煤的挥发分有利于铁回收率的提高,对降低铁产品中磷不利.经过类高温干馏的两种煤种的原煤作为还原剂时,焙烧产品新生成脉石矿物不同,并且原矿中鲕绿泥石和石英等会与含铁矿物发生复杂的化学反应,原矿中含磷矿物大部分仍以氟磷灰石形式存在,但少部分含磷矿物仍会参与还原焙烧的反应.

高磷鲕状赤铁矿直接还原同步脱磷研究

对含铁品位为43.58%、含磷0.83%的鄂西某宁乡式高磷鲕状赤铁矿进行了直接还原焙烧脱磷试验研究。研究了焙烧温度、还原剂用量、焙烧时间、脱磷剂用量对直接还原铁指标的影响。在还原剂用量17.5%,TS用量50%,NCP用量2.5%,焙烧时间60 min,一段磨矿粒度为-0.074 mm粒级占89.56%,磁选磁场强度为87.58 kA/m;二段磨矿粒度为-0.025 mm粒级占100%,磁选磁场强度为87.58 kA/m时可得到铁品位91.58%,回收率84.96%,磷品位0.049%的直接还原铁磁选精矿。

某高磷鲕状赤铁矿磷赋存状态及还原焙烧脱磷研究

采用浸出、电渗析等试验方法研究了尼日利亚高磷鲕状赤铁矿中磷的赋存状态并采用XRD和SEM分析添加脱磷剂Na2CO3直接还原焙烧产物的特性.结果表明,含磷矿物主要是呈微细粒包裹体嵌布在铁矿物鲕粒的裂隙或孔洞中的纤磷钙铝石(CaAl3(OH)6(HPO4)(PO4))、蓝磷铝铁矿(FeAl2(PO4)2OH·6H2O)以及均匀分散在铁矿物中的磷.通过添加脱磷剂Na2CO3的还原焙烧磁选可实现高效铁磷分离.磷通过两种方式去除:一部分含磷矿物与金属铁分离,存在于脉石矿物中,通过磨矿磁选可以有效去除,一部分含磷矿物与Na2CO3反应生成溶于水的磷矿物,最终使得还原铁产品中的磷含量降低.

某难选褐铁矿直接还原焙烧-磁选工艺研究

对某难选褐铁矿进行了直接还原焙烧-磁选工艺研究。进行了焙烧温度、焙烧时间以及还原剂添加量的条件试验,以及焙烧样品的多种磁选流程对比试验。在原料粒度-2 mm,焙烧温度1 150℃,CaCO3用量为矿量的15%,煤添加量为矿量的25%,盖煤量为球团质量的33%,保温时间2 h,一段磨矿粒度为-0.045 mm粒级占97%,一次粗选场强79 kA/m、两次精选场强45 kA/m时,矿物焙烧金属化率95.24%,铁精矿品位80.61%,回收率88.58%。