一种用于酸性条件下胶磷矿反浮选脱硅的新型阳离子捕收剂

针对目前胶磷矿反浮选脱硅阳离子捕收剂的选择性差、仅适用于弱碱性矿浆条件、消泡难等问题,本研究采用主要成分为十六烷基三甲基溴化铵（CTAB）和邻苯二甲酸二丁酯（DBP）的两种化工厂副产品研发了一种新型、经济、高效的胶磷矿反浮选脱硅阳离子捕收剂。在微浮选试验中,当pH值为6–10,新型捕收剂用量为25 mg·L?1时,石英和氟磷灰石纯矿物之间存在显著的可浮性差异。在pH值为6,新型捕收剂用量为0.4 kg·t?1时,对脱镁磷精矿进行粗选脱硅浮选试验,获得了P₂O₅品位为29.33wt%、SiO₂品位为12.66wt%、P₂O₅回收率为79.69wt%的胶磷矿精矿指标。利用FTIR、Zeta电位和接触角测量进行机理研究,研究结果表明新型捕收剂对石英的吸附能力高于氟磷灰石,同时DBP的协同效应也增强了石英和氟磷灰石之间疏水性的差异。泡沫稳定性试验结果表明,新型阳离子捕收剂的最大消泡率可达142.8 mL·min?1,其消泡速率远高于其他常见传统阳离子捕收剂。     

中低品位胶磷矿柱式反浮选试验研究

通过对西南某中低品位胶磷矿矿石性质进行分析,提出利用高效的微细粒分选设备旋流-静态微泡浮选柱进行分选研究;利用现场生产用药剂,考察处理量、药剂制度、循环泵工作压力、充气量等因素对浮选指标的影响,并优化工艺和操作参数条件,在此基础上进行连选试验.研究结果表明:采用一粗一精单反浮选的工艺流程可以获得精矿P2O5品位为30.01％,精矿回收率89.10％.该技术的研究为我国中低品位胶磷矿的分选提供了新的途径.     

α-氯代癸酸磷矿捕收剂的合成及其浮选性能

以癸酸为主要原料,通过氯代和中和反应得到脂肪酸改性产品CH3（CH2）7CHClCOONa,与常规脂肪酸类比较捕收剂W10的浮选效果。结果表明,W10和氯代癸酸钠都有常温浮选效果,两者P2O5回收率均在80％左右,但氯代癸酸钠作为捕收剂时,精矿品位达到30．00％,达到浮选指标,其浮选性能优于W10。     

捕收剂乳化提高浮选性能研究

研究用捕收剂以煤油为主,采用的乳化剂有ＲＭ－１（非离子型）、油酸钠（阴离子型）、十二烷基硫酸钠（阴离子型）。实验室和工业结果均表明,捕收剂乳化以后,可节约用量,而且非离子型的ＲＭ－１乳化剂制取的煤油液具有良好的浮选促进作用和浮选选择性。     

F_(203)-TBP混合捕收剂浮选锡石细泥

用Ｆ＿（２０３）－ＴＢＰ混合物作捕收剂（Ｆ＿（２０３）３是一种含Ｎ，Ｃ，Ｏ的有机化合物新药；ＴＢＰ是磷酸三丁酯代号），单宁作抑制剂，浮选车河选厂锡石细泥闭路试验结果，可从含锡１．３８％的给矿得到品位为３８．５５％，回收率为９１．１３％的锡精矿。Ｆ＿（２０３）的毒性比现有浮选车河锡石细泥的捕收剂的毒性低；药剂费用也可降低２２％～２３％。     

安徽某地石英砂的加温反浮选试验

研究了温度从15℃升高到50℃时,安徽省某地石英砂矿物的反浮选特性。研究表明在矿浆pH值2～2.5时,温度为45℃,石英表面ζ电位的负值是试验温度范围内的最低值,此时长石的表面ζ电位的负值比在25℃时要大,有利于阳离子捕收剂在长石表面的吸附,从而提高长石与石英的浮选分离效果。     

低品位氧化锌矿反浮选研究

针对某地高硅、铁低品位氧化矿的特点,采用反浮进的工艺流程进行了试验.试验证明,该工艺流程可获得锌品位、回收率的锌精矿.     

反浮选法分离粉石英和斜绿泥石及其机理

采用油酸作阴离子捕收剂,用反浮选的方法对粉石英的纯化进行研究,探讨粉石英和斜绿泥石的分离机理以及捕收剂的吸附机理。实验中测定粉石英和斜绿泥石的ζ电位、pH值、油酸浓度、浮选时间对SiO2回收率和纯度的影响;分析粉石英、斜绿泥石的晶体结构和表面性质以及油酸作阴离子捕收剂时的溶液化学。结果表明:在pH为8~10的碱性溶液中,粉石英难与金属阳离子缔合,所以,不与油酸根离子结合;而斜绿泥石解离后层间存在活性金属阳离子,可与油酸根离子发生化学吸附产生疏水作用,因而使用反浮选的方法能有效地实现二者的分离;当pH为9时,斜绿泥石的去除率最高,粉石英的纯度达99.86%,回收率为35%。     

几种捕收剂对低品位硫化镍矿浮选的作用

研究了丁基黄药、丁胺黑药和硫胺酯单独使用与组合使用对低品位硫化镍矿浮选的影响.实验结果表明,单独使用丁基黄药和硫胺酯浮选硫化镍矿时,一次粗选镍精矿品位可达3%以上,但回收率较低;单独使用丁胺黑药时,一次粗选镍精矿品位可达2%以上,回收率最高可达57%.组合用药时,丁胺黑药与硫胺酯组合使用可以有效提高低品位硫化镍矿的浮选回收率.针对丁基黄药、丁胺黑药和硫胺酯单独使用与组合使用进行了红外光谱分析,阐述了组合药剂对硫化镍矿浮选的作用机理.     

钙离子对脂肪酸类捕收剂浮选石英的影响机理

为了探究脂肪酸类捕收剂在浮选石英过程中,钙离子活化石英的作用及机理,进行了浮选试验、动电位检测、红外光谱分析.结果显示石英被浮选需要钙离子的活化,钙离子能增加石英的表面动电位,以及改性脂肪酸捕收剂DWD-3能在被钙离子活化的石英表面发生吸附,且存在化学吸附、氢键吸附;量子力学模拟显示,Ca²⁺,Ca(OH)⁺,OH^-能在石英表面发生吸附,且Ca²⁺吸附作用最强,钙离子活化石英的过程是Ca²⁺优先在石英表面上的O处发生化学吸附,形成被Ca²⁺活化的表面,捕收剂DWD-3以单键氧O_Ⅰ与2个双键O_Ⅱ吸附在被活化的石英表面.     

Improving the sulfidation-flotation of fine cuprite by hydrophobic flocculation pretreatment

Hydrophobic flocculation pretreatment was performed to assess its effect on the recovery of fine cuprite in sulfidation-flotation. The results of the micro-flotation experiment showed that cuprite recovery is related to the particle size, and that an excessive content of fine particles (<18 μm) impacted the recovery of coarse particles. When hydrophobic flocculation pretreatment was used, the recovery of fine cuprite in sulfidation-flotation increased from 60.3% to 86.3% under optimum conditions (pH 9.5; sodium oleate concentration, 2×10-4 mol·L-1; stirring time, 6 min; stirring speed, 1600 r·min-1). The laser particle size analysis and optical microscopy results indicate that hydrophobic flocculation pretreatment effectively reduces the content of fine cuprite, and augments the apparent particle size in the pulp. We performed the Derjaguin-Landau-Verwey-Overbeek (DLVO) theory and extended DLVO theory calculations to further support the interpretation of the results.     

利用微泡浮选柱从浮选尾矿中回收微细粒级白钨矿

针对湖南安化湘安钨业公司白钨浮选尾矿中微细粒级未能在浮选机中有效分选的特点,研究开发一种微泡浮选柱,该浮选柱采用微孔材质发泡,并利用专家系统控制浮选柱关键工作参数.半工业试验获得了较适宜的柱浮选工作参数:表观矿浆速率为0.27 cm/s,表观气体速率为1.35 cm/s.工业试验获得的精矿品位可达24.52%,回收率为43.41%,富集比达35.03.水析试验结果表明:5～10,10～19和19～38 μm 3个粒级的回收率均达到65%以上.试验测得的浮选柱内气泡的Sauter直径为400 μm,仅为机械搅拌浮选机气泡的1/3,气泡直径减小促使浮选速率常数k显著增大,这是浮选柱能有效回收微细粒级白钨矿的主要原因.     

Beneficiation of the gold bearing ore by gravity and flotation

Gold concentration usually consists of gravity separation, flotation, cyanidation, or the combination of these processes. The choice among these processes depends on the mineralogical characterization and gold content of the ore. Recently, the recovery of gold using gravity methods has gained attention because of low cost and environmentally friendly operations. In this study, gold pre-concentrates were produced by the stepwise gravity separation and flotation techniques. The Knelson concentrator and conventional flotation were employed for the recovery of gold. Gold bearing ore samples were taken from Gümüşhane Region, northern east part of Turkey. As a result of stepwise Knelson concentration experiments, a gold concentrate assaying around 620 g/t is produced with 41.4wt% recovery. On the other hand, a gold concentrate about 82 g/t is obtained with 89.9wt% recovery from a gold ore assaying 6 g/t Au by direct flotation.     

基于离散元理论的自然崩落法放矿细颗粒渗移过程研究

为探讨自然崩落法放矿过程中废石细颗粒的渗移规律及诱因,利用离散元软件EDEM分别从粗细颗粒数量比、直径比及矿岩含水率三个方面对细颗粒的渗移过程进行模拟,并设置标志颗粒对不同区域的相邻粗细颗粒进行跟踪。研究结果表明：放矿提前贫化率随粗细颗粒数量比的增大不断降低;废石细颗粒渗移速率随粗细颗粒直径比的增大而增大,且受含水率影响显著;相邻位置的细颗粒的渗移速度大于粗颗粒的下降速度。研究结果为进一步探讨矿石损失贫化、确定采场结构参数及优化放矿管理制度提供理论支持。     

Magnetic and flotation studies of banded hematite quartzite (BHQ) ore for the production of pellet grade concentrate

To identify and establish beneficiation techniques for banded hematite quartzite (BHQ) iron ore, a comprehensive research on BHQ ore treatment was carried out. The BHQ ore was assayed as 38.9wt% Fe, 42.5wt% SiO₂, and 1.0wt% Al₂O₃. In this ore, hematite and quartz are present as the major mineral phases where goethite, martite, and magnetite are present in small amounts. The liberation of hematite particles can be enhanced to about 82% by reducing the particle size to below 63 μm. The rejection of silica particles can be obtained by magnetic and flotation separation techniques. Overall, the BHQ ore can be enriched to 65.3wt% Fe at 61.9% iron recovery. A flowsheet has been suggested for the commercial exploitation of the BHQ ore.     

Kinetic approach to the study of froth flotation applied to a lepidolite ore

The number of published studies related to the optimization of lithium extraction from low-grade ores has increased as the demand for lithium has grown. However, no study related to the kinetics of the concentration stage of lithium-containing minerals by froth flotation has yet been reported. To establish a factorial design of batch flotation experiments, we conducted a set of kinetic tests to determine the most selective alternative collector, define a range of pulp pH values, and estimate a near-optimum flotation time. Both collectors (Aeromine 3000C and Armeen 12D) provided the required flotation selectivity, although this selectivity was lost in the case of pulp pH values outside the range between 2 and 4. Cumulative mineral recovery curves were used to adjust a classical kinetic model that was modified with a non-negative parameter representing a delay time. The computation of the near-optimum flotation time as the maximizer of a separation efficiency (SE) function must be performed with caution. We instead propose to define the near-optimum flotation time as the time interval required to achieve 95%–99% of the maximum value of the SE function.     

混合嗜酸硫杆菌浸出低品位磷矿的正交实验研究

以某矿酸性矿土中分离的嗜酸氧化亚铁硫杆菌和嗜酸氧化硫硫杆菌形成的混合菌为浸磷菌种、黄铁矿为能源物质、无磷无铁9K培养基为浸矿培养基,对混合菌浸出低品位磷矿石(w(P2O5)为22.8%)的浸磷条件进行了单因素优化和正交实验研究.结果表明,最优浸磷的适宜条件是,矿浆浓度为15 g/L,菌种体积浓度为15%,初始pH值为1.5,磷的浸出率为51.07%.     

Kinetics of chemical leaching of chalcopyrite from low grade copper ore: behavior of different size fractions

The kinetics of the chemical leaching of copper from low grade ore in ferric sulfate media was investigated using the constrained least square optimization technique. The experiments were carried out for different particle sizes in both the reactor and column at constant oxidation-reduction potential (E_h), pH values, and temperature. The main copper mineral was chalcopyrite. About 40% of Cu recovery is obtained after 7 d of reactor leaching at 85℃ using -0.5 mm size fraction, while the same recovery is obtained at 75℃ after 24 d. Also, about 23% of Cu recovery is obtained after 60 d of column leaching for +4–-8 mm size fraction whereas the Cu recovery is as low as about 15% for +8–-12.7 and +12.7–-25 mm size fractions. A 4-stage model for chalcopyrite dissolution was used to explain the observed dissolution behaviors. The results show that thick over-layers of sulphur components cause the parabolic behavior of chalcopyrite dissolution and the precipitation of Fe3+ plays the main role in chalcopyrite passivation. In the case of coarse particles, transformation from one stage to another takes a longer time, thus only two stages including the initial reaction on fresh surfaces and S0 deposition are observed.