铝土矿反浮选新型捕收剂QAS224的应用研究

应用自行研制的新型捕收剂QAS224,对Al₂O₃品位为64%左右、铝硅比为6.1左右的河南某铝土矿矿石进行反浮选脱硅试验,结果在磨矿细度为-0.074 mm占81.24%,矿浆pH为11的条件下,通过1次粗选、2次精选、2次扫选,获得了精矿Al₂O₃品位为67.79%,Al₂O₃回收率为81.72%,铝硅比为9.67的较好指标。试验结果证明QAS224是铝土矿反浮选脱硅的有效捕收剂。     

云南某铝土矿反浮选脱硅试验

云南某铝土矿石为一水硬铝石型高硅高铝低铝硅比矿石,研究了矿石反浮选提铝降硅工艺技术条件。研究结果表明：在磨矿产品细度为-0.074 mm占85%,pH=5.5,组合捕收剂GE-601+E3总用量为150 g/t（质量比为21）,抑制剂OP用量为600 g/t情况下,采用1粗1精3扫、中矿顺序返回闭路反浮选试验流程处理该矿石,室温和8 ℃下反浮选精矿Al₂O₃品位分别为67.65%和67.52%,回收率分别为82.04%和81.77%,铝硅比分别为10.72和10.58,组合捕收剂室温和低温脱硅浮选效果均良好。     

重庆某高硫铝土矿石选矿试验

重庆某高硫铝土矿石中主要含铝矿物为一水硬铝石,主要含硫矿物为黄铁矿,铝硅比仅为4.73。对该矿石脱硫降硅提铝工艺技术条件进行了研究。结果表明,采用1次粗浮选选硫,2粗3精1精扫、中矿顺序返回浮铝土矿流程处理该矿石,最终获得了Al₂O₃品位为65.81%、含硫0.42%、铝硅比为9.09、Al₂O₃回收率为77.88%的铝土矿精矿,副产品硫精矿硫品位为22.13%、回收率为82.69%、Al₂O₃含量为32.94%、回收率为5.15%。     

用新型捕收剂G7浮选一水硬铝石型铝土矿

昆明冶金研究院研发了一种新型、高效的铝土矿正浮选捕收剂G7,为检验其在提高精矿指标、耐低温、易消泡脱水方面的优越性能,以云南某一水硬铝石型铝土矿为研究对象进行了选矿试验。结果表明,Al₂O₃和SiO₂含量分别为67.51%和12.64%、铝硅比为5.34的铝土矿石在磨矿细度为-0.074 mm占90%、矿浆温度为3 ℃和20 ℃的情况下,采用2粗2精1扫、中矿顺序返回流程处理,分别可获得Al₂O₃品位和回收率为75.51%和89.13%,铝硅比为15.63的优质精矿和Al₂O₃品位和回收率为76.25%和90.28%、铝硅比为16.68的优质精矿。精矿指标分析和浮选过程观察表明,G7是一种性能优良、稳定、高效且适应性强的铝土矿浮选捕收剂。     

国外某悬浮焙烧—磁选高铝铁精矿降铝试验

国外某铁矿石经悬浮焙烧-磁选得到的铁精矿TFe含量为64.50%、Al2O3含量为5.95%，主要铁矿物是磁铁矿，少量赤铁矿。为解决其Al₂O₃含量较高的问题，研制了一种新型阴离子捕收剂DTL-1，通过改变浮选温度、矿浆pH值、捕收剂DTL-1用量对悬浮焙烧-磁选后的高铝铁精矿样进行脱铝试验。结果表明：在温度25 ℃、pH值8.0、捕收剂DTL-1用量50 g/t的条件下，采用1粗3精1扫的浮选闭路试验流程，最终获得了TFe品位66.40%，TFe回收率87.81%，Al₂O₃品位4.90%的浮选铁精矿产品。浮选铁精矿产品中粒度较大的磁铁矿颗粒中包裹着细粒的氧化铝矿物，在优先确保铁的回收率前提下，部分含铝铁氧化物进入铁精矿，因而也造成了铁精矿中铝含量较高的情况。     

国外某悬浮焙烧-磁选高铝铁精矿降铝试验

国外某铁矿石经悬浮焙烧-磁选得到的铁精矿TFe含量为64.50%、Al2O3含量为5.95%，主要铁矿物是磁铁矿，少量赤铁矿。为解决其Al₂O₃含量较高的问题，研制了一种新型阴离子捕收剂DTL-1，通过改变浮选温度、矿浆pH值、捕收剂DTL-1用量对悬浮焙烧-磁选后的高铝铁精矿样进行脱铝试验。结果表明：在温度25 ℃、pH值8.0、捕收剂DTL-1用量50 g/t的条件下，采用1粗3精1扫的浮选闭路试验流程，最终获得了TFe品位66.40%，TFe回收率87.81%，Al₂O₃品位4.90%的浮选铁精矿产品。浮选铁精矿产品中粒度较大的磁铁矿颗粒中包裹着细粒的氧化铝矿物，在优先确保铁的回收率前提下，部分含铝铁氧化物进入铁精矿，因而也造成了铁精矿中铝含量较高的情况。     

西南地区某高泥堆积型铝土矿选矿试验研究

西南地区某高泥堆积型铝土矿中Al₂O₃、SiO₂含量分别为47.25%、10.86%，铝硅比为4.35；主要目的矿物为硬水铝石，主要脉石矿物为高岭石，其他金属矿物有赤铁矿、针铁矿以及锐钛矿等。矿石中矿物的共生关系较为复杂，目的矿物嵌布粒度不均，单体解离度不高；铁、钛矿物与铝矿物之间紧密共生，难于实现有效分离。针对该矿石采用全泥正浮选的选矿工艺，以氢氧化钠为调整剂、水玻璃为抑制剂，EMB-506为捕收剂，在磨矿细度为-0.074 mm占70%的条件下，闭路试验可获得Al₂O₃品位为51.62%，SiO₂品位为6.05%，铝硅比为8.53，Al₂O₃回收率为86.66%的浮选精矿。该工艺流程结构及浮选药剂制度简单，所得浮选精矿达到了拜耳法生产氧化铝对给料的要求。     

低品位磷矿浮选试验研究

对某地低品位磷矿进行了浮选试验研究, 探索了磨矿细度、捕收剂(十二胺)用量、抑制剂(硫酸)用量及浮选流程对浮选指标的影响。结果表明, 采用一粗一精一扫、中矿返回粗选的浮选工艺流程, 在磨矿细度为-0.074 mm粒级占78%、粗选十二胺用量500 g/t、硫酸用量18 kg/t, 精选十二胺用量200 g/t、硫酸用量9 kg/t, 扫选硫酸用量14.4 kg/t条件下, 针对P₂O₅品位为22.41%的原矿可获得P₂O₅品位32.47%、P₂O₅回收率83.32%的磷精矿, 实现了P₂O₅的有效富集。     

铝酸盐荧光粉中绿粉与蓝粉的浮选研究

荧光粉中的稀土氧化物具有极高的回收利用价值。为实现细粒级稀土铝酸盐荧光粉绿粉(Ce,Tb)Mg Al₁₁O₁₉和蓝粉Ba Mg Al₁₀O₁₇:Eu²⁺的分离,对其进行了浮选试验,探究了p H值、药剂种类和用量对荧光粉浮选回收效率的影响。试验结果表明在酸性条件下,相比海藻酸钠和硅酸钠,木质素磺酸钠(SL)对蓝粉的抑制效果更好。在p H=5、抑制剂SL用量12500g/t、分散剂碳酸钠用量26500g/t、捕收剂十二烷基硫酸钠用量5000g/t的条件下,可得到回收率为68.04%、稀土氧化物品位15.79%的绿粉精矿和回收率为68.46%、稀土氧化物品位6.85%的蓝粉精矿。     

织金含稀土磷矿石反浮选试验研究

根据矿石性质,采用自制的白云石类脉石矿物高效捕收剂AB对织金含稀土磷矿石进行1粗1精反浮选试验,在磨矿细度为-0.075 mm占82.80%、AB用量粗选和精选分别为350 g/t和100 g/t,抑制剂磷酸用量粗选和精选分别为12 kg/t和2 kg/t的条件下,获得了P₂O₅品位为31.20%、P₂O₅回收率为84.62%、RE₂O₃含量为0.133%、RE₂O₃回收率为92.98%、MgO含量为1.07%的磷精矿,为综合性开发该含稀土磷矿石资源提供了经济可行的技术依据。     

组合捕收剂低温浮选某铅锌尾矿中萤石的试验研究

某铅锌浮选尾矿中,含有CaF2 21.13%,品位较高,具有一定的经济价值。为实现该尾矿中萤石资源的综合回收利用,分别进行了Na2CO3用量试验、抑制剂配比试验、单一捕收剂试验、组合捕收剂配比试验、组合捕收剂用量试验及浮选温度试验等条件试验。条件试验结果表明,当Na2CO3用量为400g/t,Na2SiO3用量为500g/t,腐殖酸钠用量为300g/t,组合捕收剂中油酸钠(OL)与石油磺酸钠(SPS)比例为2:1时,组合捕收剂用量为200g/t时,能取得较好的萤石浮选指标,且浮选结果受温度影响较小。在上述试验确立的药剂制度的基础上,进行了闭路试验,最终获得的精矿中萤石品位为品位97.32%,回收率为78.08%。     

Flotation of rare earth minerals from silicate–hematite ore using tall oil fatty acid collector

The flotation of rare earth (RE) minerals (i.e. xenotime, monazite-(Nd), RE carbonate mineral) from an ore consisting mainly of silicate minerals (i.e. primary silicate minerals and nontronite clay) and hematite was investigated using tall oil fatty acids (Aero 704, Sylfat FA2) as collector. The RE minerals are enriched with Fe. The effects of tall oil fatty acid dosage, pH, temperature, and conventional depressants (sodium lignin sulfonate, sodium metasilicate, sodium fluoride, sodium metasilicate and sodium fluoride, and soluble starch) were determined at grinding size of P80 = 63 μm. At this grinding size, the grain size of the RE minerals ranges from 2 to 40 μm, percentage liberation is 9–22%, and percentage association with nontronite and quartz is 30–35%. Results indicated that Sylfat FA2 at 22450 g/t concentration was the more efficient tall oil fatty acid collector at natural pH (pH 7) to basic pH (pH 10.0–11.5). Flotation at the room temperature (25 °C) gave higher selectivity than 40 °C temperature flotation. The results on the effect of depressants showed similar selectivity curves against the gangues SiO₂, Al₂O₃, and Fe₂O₃ suggesting that the chemical selectivity of the depressants has been limited by the incomplete liberation of the RE minerals in the feed sample. High recoveries at 76–84% (Y + Nd + Ce)₂O₃ but still low (Y + Nd + Ce)₂O₃ grade at 2.1% in the froth were obtained at flotation conditions of 63 μm, 25 °C, pH 10.5, 1,875 g/ton sodium metasilicate and 525 g/ton sodium fluoride or 250 g/ton soluble starch as depressant for the silicates and hematite, and 22,450 g/t Sylfat FA2 as collector for the RE minerals (initial (Y + Nd + Ce)₂O₃ feed grade = 0.77%). The recoveries of gangue SiO₂, Al₂O₃, and Fe₂O₃ in the froth were low at 25–30%, 30–37%, and 30–36%, respectively. The mineralogical analysis of a high grade froth and its corresponding tailing product showed that the RE minerals have been concentrated in the froth while the primary silicate minerals and hematite have been relatively concentrated in the tailing. However, the clay minerals, primary silicate minerals, and hematite still occupy the bulk content of the froth. This suggests that incomplete liberation of the RE minerals led to the poor grade result, supporting likewise the selectivity curve results by the different depressants. This study showed that liberation is important in achieving selective separation.     

Effects of aluminosilicate minerals on copper-molybdenum flotation from Sarcheshmeh porphyry ores

The presence of aluminosilicate minerals adversely affects the floatability of copper during processing of Sarcheshmeh porphyry ores. In this work, ores containing different types and quantities of aluminosilicate minerals, as well as copper sulfides (i.e., chalcopyrite and chalcocite), were examined in the laboratory to determine how aluminosilicate minerals affect the floatability of copper and to find ways to decrease their unwanted effects. It was shown that ore types that contain muscovite and vermiculite have the greatest effect on the floatability of copper sulfides. To overcome this problem, two reagent strategies were used: (a) dextrin at 100, 200, and 600 g/ton and (b) a mixture of dextrin, sodium silicate, and sodium hexametaphosphate (20%, 40%, and 40% by weight, respectively) at 200 and 300 g/ton. The reagent mixture (case b) at a dosage of 200 g/ton improved the copper recovery and decreased the grade and the recovery of Al₂O₃ and SiO₂ in the flotation concentrate. As an alternative solution, flotation at reduced pulp density (24%) improved the grade and the recovery of copper and molybdenum and decreased the grade and recovery of Al₂O₃ and SiO₂ in comparison to the currently used pulp density (29%) in the Sarcheshmeh flotation plant.     

Beneficiation of diasporic-bauxite ore by selective flocculation with a polyacrylate flocculant

Beneficiation of a diasporic-bauxite ore by selective flocculation with sodium polyacrylate (PAS) was studied. Flocculation behaviors of diaspore, kaolinite, illite and pyrophyllite in the presence of PAS were also investigated by sedimentation tests. PAS shows good selectivity for the flocculation separation of diaspore from silicates in the range of pH 8–12. A concentrate with Al₂O₃/SiO₂ ratio 8.9 and the recovery of Al₂O₃ 87.0% was obtained from a diasporic-bauxite ore with Al₂O₃/SiO₂ ratio 5.7 with the addition of sodium carbonate 5 kg/t and PAS 7 g/t. Micro-photogram, zeta potential and FTIR measurements were then performed to study the mechanisms. It was observed that the zeta potentials of diaspore shifted to negative value strongly and the isoelectric point (IEP) of diaspore was also changed from about pH 6–4, and the flocs of diaspore were larger and denser than that of kaolinite in the presence of PAS. A new band at 1640 cm⁻¹ was found in the infrared spectrum of the diaspore after interaction with PAS, and this new band matched the characteristic peak at 1624 cm⁻¹ in the infrared spectrum of the reaction products of PAS and AlCl₃. The adsorption of PAS on diaspore is concluded to be chemisorption caused by the chemical reaction between carboxylate group and the exposed Al sites on diaspore surface.     

Preparation of hydrolyzate of hogwash oil (HHO) and its application in separating diaspore from kaolinite

Hydrolyzate of hogwash oil (HHO) was prepared at ambient pressure using hogwash oil that was collected from the dietary industry as a raw material by means of non-pressurized multistage hydrolyzation. Before hydrolyzation, the collected hogwash oil was subject to pre-treatment that included deposition edulcoration, acidification degumming and deodorization by digestion as well as vacuum dewatering. Components of the synthesized products were detected by gas chromatography and the results indicated that unsaturated fatty acids are the major components.Using HHO as a collector, the flotation behavior of diaspore and kaolinite was investigated. The reaction mechanisms of diaspore and kaolinite with HHO are discussed based on a zeta potential test. The results showed that a concentrate with a mass ratio of Al₂O₃ to SiO₂ of 14.42 and an Al₂O₃ recovery of 81.80% can be obtained at a feed mass ratio of Al₂O₃ to SiO₂ of 5. This indicated that the HHO is a potential collector for separating diaspore from aluminosilicate minerals in bauxite. The results also suggested a possible way to reuse discarded hogwash oil to avoid environmental pollution.     

Flotation of pyrochlore from low grade carbonatite gravity tailings with benzohydroxamic acid

The beneficiation of a low grade niobium tailings by flotation using benzohydroxamic acid (BHA) as a collector and sodium hexametaphosphate (SHMP) as a modifying agent has been studied. A concentrate containing approximately 3% Nb₂O₅ (with an upgrade ratio of 12) was obtained with nearly 80% Nb recovery in 6.5% of the mass after only one rougher flotation stage. After four rougher stages, a final concentrate containing 1.54% Nb₂O₅ with 93% Nb recovery in ∼15% of the mass was obtained. A Box–Behnken experimental design was employed to analyze the effects of collector dose and dispersant dose and pulp temperature and to model the concentrate grade and niobium recovery during rougher–scavenger flotation. The results indicate that temperature has no significant effect on recovery at the 95% or 90% confidence level. Statistical insignificance of the temperature effect has been attributed to the collector mechanism, which essentially does not require BHA species beyond a maximum monolayer capacity that is potentially achievable within the solubility limit at room temperature. SHMP plays an important role in the selective recovery of pyrochlore. Its dispersing function is distinguished from its depressing function in the rougher stages through a positive interaction term together with BHA in both the recovery model and the grade model. The main negative effect of SHMP on niobium recovery emerges in the scavenger stages as a result of depressing action (e.g., targeting primarily the middlings). The effect of reagent dose and temperature on the grade and recovery are presented using 3D graphs from statistical models developed. Negative quadratic terms of the statistical models are discussed in reference to possible overdosing of SHMP on the middlings (involving pyrochlore and gangue) as well as a postulated change of orientation mode of BHA, also due to its overdosing.     

BK425和HA-1对磷矿反浮选脱镁脱铝效果影响研究

某难选磷矿由于其脉石矿物与胶磷矿嵌布关系复杂,为了获得合格的磷精矿,用脱镁捕收剂BK425和脱铝捕收剂HA-1对原矿进行反浮选试验。当原矿磨矿细度为-74μm占70%时,采用反浮选脱镁(一次粗选、一次扫选)—反浮选脱铝(一次粗选、两次精选)全开路流程,可以获得P_2O_5品位31.14%,Al_2O_3含量2.60%,MgO含量0.92%,P_2O_5回收率79.54%的磷精矿。     

承德地区某超贫钒钛磁铁矿干抛尾矿磷回收试验研究

以承德地区某钒钛磁铁矿干抛尾矿为研究对象,采用浮选方法,系统研究干抛尾矿磨矿细度、充气量、抑制剂用量、pH值、捕收剂用量、粗选时间对磷浮选效果的影响。结果表明,干抛尾矿在磨矿细度为-0.074 mm占55%条件下,以水玻璃为抑制剂、DB-1为捕收剂、经过"一次粗选、两次精选、两次扫选"的浮选闭路试验,可以获得磷精矿产率为2.35%、P_2O_5品位为38.34%,P_2O_5回收率为91.94%的浮选指标,实现了该干抛尾矿中磷的高效回收。     

Research on the recycling of valuable metals in spent Al2O3-based catalyst

A new technology was developed to recover multiple valuable elements in the spent Al₂O₃-based catalyst by X-ray phase analysis and exploratory experiments. The experiment results showed: In the condition of roasting temperature of 750 °C and roasting time of 30 min, mol ratio of Na₂O: Al₂O₃ 1.2, the leaching rate of alumina, vanadium and molybdenum in the spent catalyst is 97.2%, 95.8% and 98.9%, respectively. Vanadium and molybdenum in sodium aluminate solution can be recovered by barium hydroxide and barium aluminate, the precipitation rate of vanadium and molybdenum is 94.8% and 92.6%. Al(OH)₃ is prepared from sodium aluminate solution with carbonation decomposition process, and the purity of Al₂O₃ is 99.9% after calcinations, the recovery of alumina can reach 90.6% in the whole process. The Ni–Co concentrate was leached by sulfuric acid, a nickel recovery of 98.2% and over 98.5% cobalt recovery was obtained respectively under the experimental condition of 30% (w/w) H₂SO₄, 80 °C, reaction time 4 h, liquid:solid ratio (8:1) by weight, stirring rate of 800 rpm.