707阴离子交换树脂吸附铌的性能及动力学研究

研究了707阴离子交换树脂从草酸溶液中吸附铌 草酸配合物的特性。试验结果表明,在[(NH4)2C2O4]=0 20mol/L,[K2S2O7]=0 030mol/L,[Nb]=8 0×10-5mol/L,pH=2 0条件下,铌的吸附率有最大值,测得Freundlish常数K=1 69×103,吸附速率常数k25℃=1 73×10-4s-1,吸附活化能Ea=22 1kJ/mol,树脂的饱和容量为每g干树脂123 2mg。     

W-305C树脂吸附铌的性能及动力学研究

研究了W 305C离子交换树脂从草酸溶液中吸附铌草酸配合物的性能。试验结果表明,在[(NH4)C2O4]=0 20mol/L,[K2S2O7]=0 030mol/L,[Nb]=8 0×10-5mol/L,pH=2 0条件下,铌的吸附率有最大值,测得Freundish常数K=3 50×103,吸附率常数k25℃=1 70×10-4s-1,吸附活化能Ea=24 94kJ/mol,铌的饱和容量为每g干树脂143 2mg。     

无取向电工钢的高温塑性变形流动应力

 以指导无取向电工钢热轧工艺为目的,采用Gleeble 1500热模拟试验机进行高温等温压缩,在应变速率为0.01~10s-1和变形温度500~1200℃条件下,对试样进行试验研究。结果表明：随着变形温度的升高,在回复与再结晶过程中发生α-Fe向γ-Fe相的?洌贾挛忍鞅溆αΤ氏帧耙斐！北浠２捎肁rrhenius关系模型,模型参数能很好的与试验结果相吻合。利用模型分别计算得500~800℃时,应力水平因子α=0.0390MPa-1,应力指数n=7.93,结构因子A=1.9×1018 s-1,热变形激活能Q=334.8kJ/mol;1050~1200℃时,应力水平因子α=0.1258MPa-1,应力指数n=5.29,结构因子A=1.0×1028 s-1,热变形激活能Q=769.9kJ/mol。     

钒渣单复焙烧影响因素及热动力学研究

在钙化焙烧提钒新工艺中,对比研究了钒渣单复焙烧的影响因素、热力学及动力学。结果表明:单复焙烧熟料形貌差异较大,单焙烧熟料成块成球,复焙烧熟料疏松;熟料循环焙烧时循环比为0.50时转化率为90%以上;尾渣复焙烧转化率为16.20%,熟料单焙烧转化率为82.38%,说明单焙烧不完全,采用尾渣与熟料混合循环的复焙烧可以显著提高转化率,尾渣循环比R_2/R_1为0.67时转化率为89.55%;单焙烧钙钒比为0.6～0.8,复焙烧具有较低的钙钒比为0.40且可以再次利用单焙烧剩余钙量;单复焙烧4 h转化率分别约为85%、88%～92%;复焙烧温度较单焙烧温度低约30℃,低温下转化率高5.54～7.18个百分点;单复焙烧在870℃时化学反应平衡常数K分别为35.17、132.44;ΔG分别为-33.84 kJ/mol、-46.48 kJ/mol;ΔS分别为-0.220 9 kJ/mol、-0.198 1 kJ/mol;ΔH分别为216.88 kJ/mol、183.92 kJ/mol;钒渣单复焙烧反应受零级化学反应控制,反应活化能E分别为37.36 kJ/mol、30.65 kJ/mol;指前因子A分别为7.32×10~(13) min~(-1)、3.11×10~(13) min~(-1)。     

FeS热分解动力学的热分析

FeS诱发含硫油品自燃的事故受到了业界的日益关注.通过在不同升温速率(2,5,8,10,15℃/min)下的热分析实验,应用模型和非模型拟合研究了FeS的热分解动力学机理,结果表明:FeS受热氧化是FeS与氧气物理吸附、化学吸附和化学反应过程,对FeS的模型拟合结果不稳定,可靠性较差;采用等转化率法得到FeS热分解的表观活化能E=(135.81±8.27)kJ/mol;通过Satava-Sestak方程确定了FeS的受热分解符合成核和生长模型函数A2:g(α)=[-ln(1-α)]1/2,其表观活化能E=148.43 kJ/mol,表观指前因子A=3.82×109K/s.     

还原气氛下亚锡硅酸盐的硫化挥发规律

详细考察了接近工业生产条件的CO——CO_2还原性混合气氛下亚锡硅酸盐的硫化挥发规律,探讨了气氛的氧势对挥发过程的影响,得出了在以FeS 作硫化剂的情况下,加或不加添加剂CaO 时,过程的最佳氧势区分别在Po_2=10~(-6·5)～10~(-5·5)Pa、Po_2=10~(-1·75)～10~(-4·25)Pa(Ca/Sn=1.5时)和Po_2=10~(-5·5)～10~(-4·5)Pa(Ca/Sn=1.0时)。对实验结果进行有关数据处理,则得出了CO/CO_2混合气氛下以FeS 作硫化剂时,添加CaO 或不加时挥发反应的级数分别为n=1.03～1.20和n=1.78～2.18,活化能分别为85.221kJ/mol 和183.485kJ/mol。     

110~*树脂吸附铜的行为研究

研究了Cu2+在110*树脂上的吸附行为。结果表明:在pH=4.19的HAc-NaAc缓冲溶液中,110*树脂吸附Cu2+效果最佳,静态饱和吸附容量为240mg/g;用1.0~2.0mol/LHCl溶液洗脱,洗脱率达100%;表观速率常数k298=1.55×10-4s-1,表观活化能Ea=37.2kJ/mol;等温吸附服从Freundlich经验式;吸附热力学参数ΔH=14.8kJ/mol,ΔS=52.0J/(mol·K),ΔG=-0.7kJ/mol。用化学和红外光谱法确定了吸附机制为化学吸附。     

亚锡硅酸盐硫化挥发动力学研究

本研究利用差重分析法考察了在惰性氩气氛下,FeS作硫化剂时亚锡硅酸盐硫化挥发过程的动力学。研究结果表明,挥发过程速率符合Usno=KNsno,其中K值随温度升高而升高,反应级数和对应的活化能值分别为n=1.95～2.27,E=195.95kJ/mol(不加CaO时)和n=1.02～1.45,E=86.07kJ/mol(添加CaO时)。活化能值表明,不加CaO时挥发(过程表现为化学反应步骤所控制,加入相当量的CaO后过程则转化为扩散步骤所控制。     

氨基膦酸树脂吸附铈的研究

研究了铈(Ⅲ)在氨基磷酸树脂上的吸附行为。试验结果在pH5.0时,静态饱和吸附容量为197mg/g干树脂;用2mol.L-1HCl洗脱,洗脱率为98.1%;表观速率常数k298=2.71×10-5s-1;等温吸附服从Freundlich经验式;吸附反应中的△Ho=50.0kJ.mol-1,△So=286J.mol-1.K-1,△Go=-35.1kJ.mol-1。树脂功能基与Ce3+的配比为2∶1;并用红外光谱探讨了树脂与铈的成键。     

W-305C树指吸附钽的性能及动力学研究

研究了 W 3 0 5 C树脂从草酸溶液中吸附钽草酸配合物的性能。结果表明 ,在 [(NH4) C2 O4]=0 .1 5 mol/L,[K2 S2 O7]=0 .0 2 8mol/ L,[Ta]=8.0× 1 0 -5mol/ L,p H=2 .0条件下 ,钽的吸附率有最大值 ,Freundlish常数k=3 .48× 1 0 3 ,吸附常数 K2 98=1 .72× 1 0 -4,吸附活化能 Ea=2 5 .70 k J/ mol,树脂对钽的饱和容量为每克干树脂 1 5 3 .5 2 mg。     

硫化铜精矿综合利用的研究

研究了采用硫化铜精矿生产硫酸铜、亚硫酸铵和铁红的方法,试验考察了焙烧温度、浸出酸度、浸出温度和浸出时间等因素对浸出率和浸出液铁含量的影响,确定了最佳工艺条件。技术指标：浸出率大于96％,浸出三价铁含量2-3/L,一次结晶的硫酸铜达到了GB437－80中一级硫酸铜的标准。试验证明,本工艺在生产上是可行的。     

硫化锌精矿空气氧化硫酸浸出的动力学研究

硫化锌精矿的浸出受精矿粒度、酸度、反应温度、催化剂加入量、浸出时间等诸多因素的影响。从动力学的角度分析整个浸出流程,研究综合浸出的最优化条件,使锌的浸出率可达98%以上,并建立了符合混合控制的数学模型:1-(1-α)1/3=8.07×10-7D-1[H+]-0.47[Fe2+]0.14exp(-11072/RT)t+B,通过Arrhenius经验公式,求得反应活化能为11.0727 kJ/mol。     

Fluidized Bed Selective Oxidation-Sulfation Roasting of Nickel Sulfide Concentrate: Part II. Sulfation Roasting

The fluidized bed sulfation roasting process followed by water leaching was investigated as an alternative process to treat nickel sulfide concentrate for nickel production. The effects of several roasting parameters, such as the sulfation gas flow rate, roasting temperature, the addition of Na₂SO₄, and the roasting time, were studied. 79 pct Ni, 91 pct Cu, and 95 pct Co could be recovered with minimal dissolution of Fe of 4 pct by water leaching after two-stage oxidation-sulfation roasting under optimized conditions. The sulfation roasting mechanism was investigated, showing that the outermost layer of sulfate melt and the porous iron oxide layer create a favorable sulfation environment with high partial pressure of SO₃. Sulfation of the sulfide core was accompanied by the conversion of the sulfide from Ni_1?x S to Ni₇S₆ as well as inward diffusion of the sulfation gas.     

提高含砷铜金精矿焙烧-氰化工艺金、银、铜回收率的试验研究

提出了一种提高含砷铜金精矿焙烧-氰化工艺金、银、铜回收率的新方法。该方法是将金精矿加入硫化钠后进行焙烧预处理,可有效地提高金、银、铜的回收率。试验结果表明,金、银、铜的浸出率分别提高8．22％,57．43％,7．82％,且不影响制酸和电解铜工艺。     

Combined processing technology of the Udokan sulfide copper concentrate

The combined processing variant of the Udokan sulfide copper concentrate involving low-temperature roasting with potassium chloride and the subsequent two-stage leaching of the obtained sinter consisting of copper chloride and potassium salts is described. At the first stage, water-soluble salts are leached (potassium sulfate and chloride) with water and, at the second stage, copper chloride is leached with sulfuric acid. The final products are copper vitriol and potassium fertilizer. The parameters of the main stages of the suggested technology are optimized.     

An application of acetonitrile leaching and disproportionation.: Refining segregated copper from roasted concentrates and ores

Pure copper with > 99% recovery has been obtained on a laboratory scale from a variety of copper sulfide concentrates by the following steps. An oxidative roast at 800–900°C to remove sulfur; reduction of the calcine, preferably but not necessarily under segregation roasting conditions at 650–750°C, to generate particulate copper; screening, in the case of segregation roasting, to partially separate from magnetite the over-size carbon which is coated with copper, gold and silver; selective dissolution in acetonitrile-water of the copper from both fractions; then thermal disproportionation of the copper(I) sulfate solution to remover pure copper powder. At least 80% of the silver and > 98% of the copper is recovered by this new concept. Cyanidation of leach residues recovers > 99% of the copper, > 90% of the silver and 80% of the gold, without interference from the iron in the residue. The method has been applied to the product of a segregation roast of refractory copper ores (TORCO process), to the product of a double roast of copper concentrates (Opie-Coffin process) and to the product of a non-segregation reductive roast of a dead roasted concentrate (USBM process). It is also applicable to calcines reduced in a blast furnace.Successful scale up could result in a low cost process for producing copper from copper concentrates. The energy requirements promise to be less than 6000 kJ as 25 psig steam per kg copper, if effective use of steam from the exothermic roasts can be achieved.     

锌阳极泥—硫化锌精矿中锰锌分离的焙烧参数研究

通过锌阳极泥与硫化锌混合焙烧—热水浸出工艺对锌阳极泥中的锰进行了提取,研究了焙烧气氛、物料配比、焙烧温度、焙烧时间等对锰提取率的影响。结果表明：锌阳极泥与硫化锌精矿质量比1∶1.5,焙烧温度760℃,焙烧时间2 h,浸出温度80℃,浸出时间2 h,液固比5∶1,锰的浸出率在80%以上。     

Characteristics of electrochemical reactions accompanying chlorinating annealing of copper sulfide concentrates

The mechanism for reactions accompanying the low-temperature roasting of a copper sulfide concentrate with potassium and sodium chlorides without the formation of sulfurous gases and their evolution into the atmosphere is suggested. The obtained cake is leached with water under pH correction in the vicinity of 6. Water-soluble sulfates transfer into the solution, while copper and iron hydroxides remain in the cake. The further leaching of the cake by an aqueous sulfuric acid at the final pH 2.5–3.0 makes it possible to completely transfer copper into the solution, while iron completely remains in the solid phase. The recovery of copper into the solution under the optimal annealing conditions (t = 450°C, τ = 1 h) and the above conditions of double leaching is no lower than 95%. Chlorinating annealing at temperatures below and above 450°C results in a decrease in the recovery of copper into the final product of copper sulfate. The found maximum is explained.     

某复杂银精矿提金、银工艺研究

采用一段焙烧-酸浸-氰化工艺处理某复杂银精矿,结果表明:在焙烧温度923 K,焙烧时间2 h,酸浸反应液固比1.5∶1,反应pH 值为0.8～1.0,反应温度368 K,反应时间1.5 h,氰化反应液固比2∶1,反应pH 值为10～11,NaCN 浓度1.5 ‰,反应时间48 h 条件下,氰化浸出时Au、Ag 的浸出率分别为72.01 %、18.41 %,尾渣银含量355 g/t.在复杂银精矿与其它矿样按一定比例重新配矿后,采用相同试验条件,氰化时Au、Ag 的浸出率分别提高24.89 %、15.66 %,尾渣中银含量降低了223.35 g/t.      

硫精矿有价金属综合回收技术研究与经济评价

玻利维亚某矿区选锡尾矿经浮选产出含银硫精矿,其中含金1.03×10-6、银178.7×10-6、铜0.53%,硫化物包裹金73.45%,硫化银73.63%。为综合回收该硫精矿的有价元素,采用“添加剂焙烧-烟气制酸-酸浸回收铜-氰化回收金、银-尾渣出售”工艺。试验结果表明,当增加添加剂焙烧时,金、银、铜的回收率均大幅提高,其最终浸出率分别为80.07%、90.32%和89.55%。初步经济分析结果表明,该含银硫精矿吨矿生产总价值为1 471.51元,不计原料费用时吨矿生产总成本为1 228.13元。该含银硫精矿的处理工艺实现了二次资源的可持续利用,为企业节能减排提供了一条新途径。