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.

Recovery of iron from waste ferrous sulphate by co-precipitation and magnetic separation

Magnetite concentrate was recovered from ferrous sulphate by co-precipitation and magnetic separation. In co-precipitation process, the effects of reaction conditions on iron recovery were studied, and the optimal reaction parameters are proposed as follows： n（CaO）/n（Fe2＋） 1.4：1, reaction temperature 80 ℃, ferrous ion concentration 0.4 mol/L, and the final mole ratio of Fe3＋ to FJ＋ in the reaction solution 1.9-2.1. In magnetic separation process, the effects of milling time and magnetic induction intensity on iron recovery were investigated. Wet milling played an important part in breaking the encapsulated magnetic phases. The results showed that the mixed product was wet-milled for 20 min before magnetic separation, the grade and recovery rate of iron in magnetite concentrate were increased from 51.41% and 84.15% to 62.05% and 85.35%, respectively.

Flotation-Calcination-Magnetic Separation Hybrid Process for Concentration of Rare Earth Minerals Contained in a Carbonatite Ore

A hybrid process consisting of flotation and magnetic separation has been developed to concentrate multi-phase rare earth minerals associated with a carbonatite ore that contains a significant amount of niobium. The deposit is known to contain at least 15 different rare earth minerals identified as silicocarbonatite, magnesiocarbonatite, ferrocarbonatites, calciocarbonatite, REE/Nb ferrocarbonatite, phosphates and niobates. Although no collector exists to float all the different rare earth minerals, the hydroxamic acid-based collectors have shown adequate efficiency in floating most of these minerals. 92% recovery of total rare earth oxide (TREO) and niobium in 45% mass was possible at d80 of <65 microns grind size. It was also possible to reduce the mass pull to 28%, but TREO and Nb’s recovery dropped to 85%. Calcination of the concentrate followed by quenching and fine grinding to <25 μm allowed upgrading the flotation concentrate by magnetic separation. It was demonstrated that at least 87% TREO and 85% Nb could be recovered in 16% of the feed mass. The paper discusses the overall concept of the flowsheet and the experimental strategies that led to this process.

Efficient enrichment of nickel and iron in laterite nickel ore by deep reduction and magnetic separation

The process of deep reduction and magnetic separation was proposed to enrich nickel and iron from laterite nickel ores.Results show that nickel-iron concentrates with nickel grade of 6.96%,nickel recovery of 94.06%,iron grade of 34.74%,and iron recovery of 80.44% could be obtained after magnetic separation under the conditions of reduction temperature of 1275℃,reduction time of 50 min,slag basicity of 1.0,carbon-containing coefficient of 2.5,and magnetic field strength of 72 kA/m.Reduction temperature and time affected the possibility of deep reduction and reaction progress.Slag basicity affected the composition of slag in burden and the spilling and enriching rate of nickel-iron from a matrix to form nickel-iron particles.Nickel-iron particles were generated,aggregated,and grew gradually in the reduction process.Nickel-iron particles can be effectively separated from gangue minerals by magnetic separation.

Pilot scale test of producing nickel concentrate from low-grade saprolitic laterite by direct reduction-magnetic separation

The enrichment of Ni from a low-grade saprolitic laterite ore,which has been pre-treated by high pressure grinding roller（HPGR） to be 74% passing 0.074 mm and contains 0.92% Ni,18.47% Fe,10.61% MgO and 42.27% SiO2,was conducted by using pelletizing,rotary kiln reduction and magnetic separation process on a semi industrial scale,and the effects of reduction duration,mass ratio of coal to pellets（C/P）,the types of magnetic separator,the sections of grinding-separation and the grinding fineness on the recovery of Ni and Fe were examined.It is shown that nickel concentrate containing 3.13 % Ni and 59.20 % Fe was achieved at recoveries of 84.36 % and 71.51% for Ni and Fe,respectively under the following conditions：reducing at （1120±40） ℃ for 120 min,C/P being 1.0,wet grinding of reduced pellets up to 70%-87% passing 0.074 mm and a magnetic field intensity of 238.8 kA/m during the first section of grinding-magnetic separation,and a grinding fineness of 84%-91% passing 0.045 mm and a magnetic intensity of 39.8 kA/m during the second section of grinding-magnetic separation.The enriched Ni containing concentrate has a low content of S and P,and can be used for further processing to produce high-grade ferronickel alloy.

Effect of reductant type on the embedding direct reduction of beach titanomagnetite concentrate

Iron and titanium were recovered from beach titanomagnetite(TTM) concentrate by embedding direct reduction and magnetic separation. The reduction products and the effects of the reductant type and reduction temperature on the reduction behavior were investigated. The results showed that the reduction of TTM concentrate was strongly related to the gasification reactivity of the reductant. Bitumite presented a better product index than wheat-straw biochar and coke, mainly because the gasification reactivity of bitumite was better than that of the other reductants. In addition, high temperatures were not beneficial to embedding direct reduction because of the emergence of a molten phase and iron-joined crystals, which in turn reduced the diffusion rate of the reducing gas and impeded the reduction reaction in the central area of the roasted briquette. The use of bitumite as the reductant at a C/Fe molar ratio of 1.4 and a reduction temperature of 1200°C for 120 min resulted in direct-reduction iron powder assaying 90.28 wt% TFe and 0.91 wt% TiO_2 with an iron recovery of 91.83% and titanium concentrate assaying 46.01 wt% TiO_2 with a TiO_2 recovery of 91.19%. Titanium existed mainly in the form of anosovite and ilmenite in the titanium concentrate.

Study on Scandium Separation from Rare Earth Ore in Yunnan Province

Chemical components, main mineral content and mineral composition of rare earth ore in Yunnan Province was measurated by the analysis of the spectrum and the chemical components. The study shows that main metals mineral in the rare earth ore are magnetite, tatanomagnetite, limonite; less metals mineral are ilmenite, hematite; some minim minerals were iron pyrites, zircon, scheelite, and so on. Main nonmetals mineral are quartz, feldspar(plagioclase, K-feldspar); less nonmetals mineral are hopfnerite, biotite, titanite; some minim minerals are kaolinite and dolomite. Ilmenite has the highest content of Sc as 175 g·t -1, next is titanite as 81.2 g·t -1. Based on this result, A new method of extracting Sc is put forward. The technological flowsheet of separating Sc of low-intensity magnetic separation,tabing, gravity concentrate, high-gradient magnetic separation, and electrostatic separation was prepared. Amplified experiment obtained Sc concentrate with Sc content of 148.54 g·t -1, the yield of 7.92%,recovery of 69.20%, at the same time, a Fe concentrate with the grade of 63.88% and the yield of 5.91% is obtained.

Characterization and processing of talc-magnesite from the Zinelbulak deposit

The characteristics of talc-magnesite from the Zinelbulak deposit(Uzbekistan) were investigated via X-ray diffraction, differential thermal analysis,infrared spectroscopy and optical microscopy.The mineralogical composition of the Zinelbulak talc-magnesite consists of 52 wt.%talc,43 wt.%carbonates and 5 wt.%of the iron-containing minerals magnetite,siderite and chlorite.Petrographic analysis confirmed the presence of carbonates in two forms:magnesite and breunnerite.Grindability tests revealed that talc and magnesite particles are completely separated after a grinding process carried out for 10～12 min.The distribution of the yield of talc and magnesite,as a function of the particle size,shows an irregular feature in that a comparatively coarser sample(>0.1 mm) is richer in magnesite and poor in talc while a comparatively finer sample(<0.1 mm) has a composition poorer in magnesite.The dressability of the Zinelbulak talc-magnesite was tested using conventional gravity concentration,flotation and electromagnetic separation.Gravity concentration was found to be the most economic initial process for the complete separation of magnesium carbonate and talc.Subsequent flotation and magnetic separation techniques could further increase the yield of high quality magnesite and talc.Refractory samples prepared by heating the separated magnesite at 1600℃for 2 h met the State Standards for refractory materials.

EXPERIMENTAL STUDIES ON FELDSPAR ORE CONCENTRATION

A series of Fe2O3-elimination experiments were conducted on feldspar samples from Tangshan Stone-powder Plant. These experimental methods include scrubbing desliming, flotation, rod milling and high gradient magnetic separation. Some technical factors of feldspar concentration and a new technological flow-sheet of ceramics raw material concentration were put forward.

Dixon水脂分离序列R_(2)*值评估肝铁浓度及鉴别肝结节的研究进展

肝脏肿瘤性病变等多种疾病均可引起铁浓度的改变。MR Dixon水脂分离序列可以通过测量R_(2)*值精准量化人体肝脏及肝内结节的铁浓度,从而判别结节性质及分化程度。介绍Dixon水脂分离序列的成像原理,并就肝铁浓度评估、肝内良恶性结节的鉴别及恶性结节分化程度评估方法的研究进展进行综述。

程潮铁精矿带式磁选提质试验研究

针对氢基还原竖炉所需高品位铁精矿原料短缺问题,对程潮铁精矿进行磁滚筒精选、弱磁选—磁选柱精选以及带式磁选机精选提质研究。结果表明,在磨矿细度为-0.037 mm占85%时,采用两段带式磁选流程,可以得到TFe品位为70.16%、产率为87.33%、TFe回收率为92.51%的精矿产品;优化带式磁选参数并额外新增一排清洗水出水口进行磁选,得到精矿TFe品位为70.78%、产率为45.71%,尾矿TFe品位为62.40%、产率为54.29%。带式磁选精选2个流程均能获得TFe品位符合要求的精矿产品,且额外新增一排清洗水出水口的流程得到的尾矿产品可以直接利用。本研究可为国内优质磁铁精矿提质制备高品位铁精矿提供参考。

酒钢悬浮焙烧磁铁矿选矿工艺优化研究

对酒钢悬浮焙烧磁铁矿进行了磨矿-选矿工艺优化研究。采用110 mT低场强磁选机预提精、中矿再磨后采用磁选富集抛尾、富集精矿经阳离子反浮选,最终获得TFe品位60.06%、SiO_(2)含量5.17%、金属回收率84.27%的综合精矿;与原工艺指标相比,精矿TFe品位提高了1.96个百分点,金属回收率提高了1.38个百分点,SiO_(2)含量降低了1.06个百分点。

酒钢肃南宏兴矿业铜尾矿中有价元素回收试验研究

酒钢肃南宏兴矿业铜矿石中伴生有价元素Fe、Ba,Fe主要以菱铁矿的形式赋存、Ba以重晶石的形式赋存,浮选选铜后有价元素Fe、Ba富集于尾矿中。试验确定铜尾矿有价元素回收工艺为强磁选-悬浮焙烧-磨矿-弱磁选回收铁精矿,强磁选尾矿浮选回收重晶石。采用此工艺,在铜浮选尾矿含Fe 10.8%、BaO 4.7%的情况下,试验回收铁精矿含Fe 56.4%,回收重晶石含BaSO 488.79%。由于Fe回收只需新建强磁选和脱水工序,强磁选回收铁矿粉进酒钢选矿厂现有悬浮磁化焙烧选别系统即可,投资小、易实施。

弓长岭某磁铁矿高效制备超级铁精矿研究

超级铁精矿作为一种高附加值的新型材料,具有巨大的发展潜力。弓长岭某磁铁矿TFe品位45.62%,SiO_(2)是其主要的脉石成分,含量为33.21%,有害元素P、S含量较低。矿石中的铁主要赋存在磁铁矿中,占全铁的95.05%。矿石中磁铁矿粒度较粗,主要分布在+0.074mm,分布率为82.37%。磁铁矿主要以单体形式产出,部分微细粒石英以包裹、反包裹和细脉状嵌布于磁铁矿中,较难完全解离。为实现该矿石的高值化利用,开展了超级铁精矿制备工艺研究。结果表明,采用阶段磨矿—阶段磁选—反浮选工艺处理该磁铁矿石,在一段磨矿细度为-0.074mm占65%、二段磨矿细度为-0.025mm占90%、反浮选工艺中粗选和精选的捕收剂用量均为25g/t的条件下,可以获得TFe品位72.35%、回收率为81.03%、SiO_(2)含量为0.17%、酸不溶物为0.19%、其他杂质含量微量的高品质超级铁精矿,以及TFe品位71.37%、回收率为6.07%的高纯铁精矿和TFe品位60.26%、回收率为6.71%的普通铁精矿,为磁铁矿的高附加值和梯级化利用提供了技术依据。

河北某普通磁铁矿制备超纯铁精矿试验研究

河北某普通磁铁矿TFe品位为65.25%,矿石性质结构简单,具有制备超纯铁精矿的潜力。研究采用多元素及X射线衍射图、物相分析等方法对原矿进行了工艺矿物学研究,并在此基础上对其进行了提纯试验。结果表明,原矿经过弱磁选粗选后,在磨矿细度-0.038 mm占85%的条件下经弱磁选再选、磁选柱精选得到TFe品位为71.31%的磁选柱精矿以及TFe品位68.12%、产率为3.32%的磁选柱铁尾矿。通过进一步考察药剂制度和工艺流程对铁矿精矿品位、回收率等选别指标的影响,确定了合适的药剂制度。而后磁选柱精矿经1粗3精反浮选降硅工艺试验流程,最终可获得含TFe品位71.95%、综合回收率为80.50%的超纯铁精矿,浮选尾矿TFe品位68.17%符合普通铁精矿标准。通过对选别产品进行试样化学成分分析及残余药剂测定,进一步证明该工艺流程可以实现超纯铁精矿的制备。该工艺在抛尾率为10.79%条件下,将原矿样的73.04%转化为超纯铁精矿,对这一地区超纯铁精矿的制备具有重要的指导意义,也为国内其他地区磁铁矿制备超纯铁精矿的研究提供了一定的参考价值。

铁精矿提品降钛试验研究

北方某选厂铁精矿,Fe品位为64.44%,但含钛较高,达4.18%(以TiO2计),严重影响了其销售及经济效益。为此,本文针对该铁精矿的性质特点,对其开展了提品降钛的试验研究。首先,对铁精矿进行细磨以尽可能实现单体解离,磨矿细度确定为–0.025 mm粒级占90%;然后,采用弱磁–反浮联合流程,在最佳条件下(磁场强度1 400 Oe,改性淀粉用量1 100 g/t,石灰用量1 100 g/t,以油酸钠为捕收剂、用量50 g/t),获得了满意的技术指标,最终铁精矿Fe品位高达69.57%,回收率达89.49%,TiO2含量大幅降低至1.48%,很好地实现了提品降钛的目的和要求。

袁家村微细粒难选磁赤混合铁矿石提铁降硅试验研究

袁家村铁矿选矿厂原生产工艺流程获得的铁精矿TFe品位仅65%左右、SiO_(2)含量达4%~5%,难以满足市场对高品质铁精矿的需求。为此,在详细分析生产流程中混合磁选精矿性质的基础上,采用卧式搅拌磨机细磨—弱磁选+强磁选—反浮选工艺流程开展了提铁降硅试验研究。对TFe品位42.44%、SiO_(2)含量35.42%的混合磁选精矿,在磨矿细度为-0.045 mm占97%情况下,全流程试验获得了产率53.87%、TFe品位67.87%、SiO_(2)含量1.96%、TFe回收率86.15%的高品质铁精矿;磨矿细度为-0.045 mm占94%情况下,适当增大捕收剂RA用量,可获得产率54.76%、TFe品位67.14%、SiO_(2)含量2.09%、TFe回收率86.63%的高品质铁精矿。研究结果可作为选矿厂提铁降硅工艺流程优化决策依据,对类似性质难选铁矿石的高效开发利用具有参考价值。

云南金厂河铁矿石高效节能选矿工艺试验

云南金厂河铁矿为合理开发利用全铁品位30.47%的铁矿石,确定有效的选矿工艺流程,针对矿石中有用铁矿物嵌布粒度较细,且与脉石矿物紧密共生,导致矿石难以高效抛废的问题,在碎选工艺引入了高压辊磨超细碎选择性碎解技术,并与常规破碎(-10 mm)湿式磁选预选抛尾工艺进行了对比试验。试验结果表明:在磁性铁回收率相近的情况下,采用高压辊磨—湿式预选工艺抛出的尾矿量明显提高,高压辊磨湿式预选精矿分别采用两段阶段磨矿—单一磁选工艺和两段阶段磨矿—磁选—磁选柱分选工艺,最终获得的分选指标相近,基于降本增效及节能减排的目标,推荐高压辊磨—湿式磁选—两段阶段磨矿—磁选—磁选柱分选工艺,作为该铁矿石的原则选矿工艺流程。

复杂难选铁矿阶段磨矿-磁选-浮选联合选矿工艺

以云南滇西某复杂难选铁矿为研究对象,采用X-荧光光谱分析(XRF)、化学多元素分析、X射线衍射(XRD)等手段对该矿石进行分析。结果表明:该铁矿所含回收价值最高的金属元素为铁,矿石中的铁元素主要以磁铁矿形式存在,硫化矿主要包括磁黄铁矿、黄铁矿和闪锌矿等,脉石矿物主要为蛇纹石和滑石等。基于该矿的矿石性质,采用阶段磨矿阶段选别、磁选和浮选脱硫的联合选矿工艺对铁矿石进行选别,通过研究磁场强度、磨矿细度、活化剂种类、捕收剂种类以及起泡剂种类等工艺参数对铁矿石磁选提质及浮选脱硫的影响规律,获得最佳工艺参数条件并依据获得的最佳工艺条件开展闭路试验,最终获得S品位为21.90%、S回收率为68.52%的硫精矿,以及TFe品位为61.56%、TFe回收率为67.41%、S含量为0.67%的铁精矿。结果可为该铁矿石资源高效利用提供支撑。

微细粒铁尾矿浮选回收萤石工艺试验研究

某选铁尾矿细泥中含少量萤石矿物资源。基于工艺矿物学分析,本试验提出萤石磁浮联合工艺,实现了微细粒选铁尾矿中低品位萤石的高效回收。结果表明,铁尾矿-0.038 mm占比76.97%,经磁场强度636.8 kA/m强磁选可抛除42.89%磁性脉石矿物,CaF_(2)品位由13.91%提至21.04%;采用碳酸钠作为调整剂,酸化水玻璃为抑制剂,改性脂肪酸作为捕收剂进行浮选,最终获得产率为7.81%、CaF_(2)品位为90.72%、CaF_(2)回收率为50.94%的萤石精矿产品。