碱浸提取碲的工艺研究

运用正交法,对碱浸提取碲的工艺进行了优化.考察了浸出温度、NaOH浓度、液固比、浸出时间对Te浸出率的影响.同时对最佳浸出工艺下主要杂质元素Si、Se、Cu的浸出进行了研究.结果表明:浸出温度、NaOH浓度、液固比、浸出时间对Te的浸出率无显著影响.碱浸提取碲的最佳浸出工艺为:浸出温度95℃,NaOH浓度3moL/L,液固比6∶1,浸出时间3h.此工艺条件下浸出液中Si、Se含量较高,进行后续碲的提取工艺前需除杂.     

水解法提取木聚糖工艺条件的正交实验

玉米芯先经酸预处理,用碱液(NaOH溶液)水解法提取木聚糖.为了提高玉米芯水解法获得木聚糖的提取率,通过单因素实验考察了浸提时间、固液比、碱浓度、浸提温度对木聚糖提取率的影响.通过正交实验对工艺条件进行了优化,最佳工艺条件为:NaOH溶液质量分数6%,固液比1∶21,浸提温度91 ℃,浸提时间94 min.各因素对提取木聚糖影响程度依次为碱质量分数＞浸提温度＞浸提时间＞固液比.木聚糖最佳提取率为玉米芯原料的20.3%.     

克二中区S27储层砂静态碱耗实验研究

选择新疆克拉玛依油田三采试验区T2k1克二中区T2k1的实际储层砂,在30℃、50℃下,按1/2(g/ml),1/5(g/ml),1/10(g/ml)三个固/液比分别与1.6% wt的Na2CO3溶液和1% wt的NaOH溶液反应8 h,120h,360h.监测了反应前后碱液浓度的变化,计算了各个条件下的绝对碱耗量.结果表明:随着固/液比减小和温度升高,储层砂的绝对碱耗量增加;1% wt NaOH的碱耗量大于1.6% wt Na2CO3的碱耗量,但碱剂类型、反应时间、反应温度与绝对碱耗量的关系受固/液比影响大.     

从砷化镓废渣中制备砷酸钠和金属镓的实验研究

通过碱性氧化浸出、冷却结晶、旋流电积等工艺从砷化镓废渣中制备砷酸钠和金属镓,浸出阶段考察了NaOH质量浓度、液固比、浸出时间、浸出温度及H_2O_2加入量等5个因素对砷和镓的浸出率的影响。结果表明:NaOH质量浓度100 g/L、液固比5∶1、浸出时间2 h、浸出温度70℃、H_2O_2和As的摩尔比为1. 2时,镓的浸出率能达到99%以上,砷的浸出率能达到96%以上,浸出液经蒸发浓缩、冷却结晶后可以制得纯度为63. 6%的砷酸钠晶体,重结晶后砷酸钠纯度可达到96. 7%。结晶母液经旋流电积后制得纯度为99. 946%的金属镓。     

正交试验对沸石合成条件的优化

以粉煤灰为原料,通过单因素试验研究NaOH浓度、反应温度、液固比对沸石合成产物的影响,借助正交试验考察影响因素的主次顺序,寻求最优条件组合,沸石的品质以阳离子交换量(CEC值)定性衡量.单因素试验结果表明,沸石合成的较优条件为NaOH浓度2mol/L、反应温度150℃、液固比10mL/g.L16(45)型正交试验表明,各因素的主次顺序为NaOH浓度、反应温度、液固比,较优组合水平为A2B2C2,即NaOH浓度2mol/L、反应温度150℃、液固比10mL/g,此条件下合成沸石的CEC均值为144mmol/100g.     

微硅粉湿法制备水玻璃的热力学及工艺研究

研究了微硅粉湿法制备水玻璃的热力学及溶出工艺,考察了反应时间、反应温度、硅碱配比、液固比对SiO_2浸出率和水玻璃模数的影响。结果表明,利用微硅粉湿法制备水玻璃是可行的。其最佳工艺条件为:反应温度180℃、原料硅碱配比为2.5、液固比10 mL/g、反应时间40 min,在此工艺条件下,SiO_2浸出率为77.1%,水玻璃模数为2.3。     

高取代度羧甲基松木纤维素制备工艺优化及表征

以松木纤维素为原料,氯乙酸钠为醚化剂,采用浓碱预处理,醚化过程中两次加碱法制备高取代度羧甲基松木纤维素。以单因素实验的方法对反应条件进行优化,探讨了浓碱预处理及醚化两阶段中碱浓度、温度、处理时间、固液比及醚化剂用量各因素对产品取代度的影响。结果表明其制备最佳工艺为:浓碱预处理为NaOH质量分数40%,固液比1g∶35mL,温度30℃,时间1.5h;醚化阶段工艺为NaOH的用量2g,氯乙酸钠的用量4.3g,固液比1g∶20mL。第一阶段加入质量分数50%的碱剂和70%的氯乙酸钠,温度35℃、时间1.5h,第二阶段加入剩余的碱剂和醚化剂,温度75℃、时间2h,在此条件下制得取代度高达1.237的羧甲基松木纤维素,并采用红外光谱和XRD对产物结构进行了表征。     

高取代度羧甲基松木纤维素制备工艺优化及表征

以松木纤维素为原料,氯乙酸钠为醚化剂,采用浓碱预处理,醚化过程中两次加碱法制备高取代度羧甲基松木纤维素。以单因素实验的方法对反应条件进行优化,探讨了浓碱预处理及醚化两阶段中碱浓度、温度、处理时间、固液比及醚化剂用量各因素对产品取代度的影响。结果表明其制备最佳工艺为：浓碱预处理为NaOH质量分数40%,固液比1g∶35mL,温度30℃,时间1.5h;醚化阶段工艺为NaOH的用量2g,氯乙酸钠的用量4.3g,固液比1g∶20mL。第一阶段加入质量分数50%的碱剂和70%的氯乙酸钠,温度35℃、时间1.5h,第二阶段加入剩余的碱剂和醚化剂,温度75℃、时间2h,在此条件下制得取代度高达1.237的羧甲基松木纤维素,并采用红外光谱和XRD对产物结构进行了表征。     

渣油废催化剂回收制备拟薄水铝石

采用渣油废催化剂为原料,考察了焙烧温度,液固比及pH等因素对金属回收率的影响;通过正交实验考察了原料配比、焙烧温度、焙烧时间、液固比及水浸时间对氧化铝转化率的影响,并采用N2物理吸附⁃脱附、XRD和SEM进行表征。结果表明,在焙烧温度800 ℃和液固比为5∶1时,钼和钒的浸出率可以达到94.0%以上;控制pH为1时,钼酸回收率可以达到97.6%;通过正交实验,确定了氧化铝回收条件：原料配比1.5,焙烧温度900 ℃,焙烧时间3 h,液固比5∶1,水浸时间15 min。     

碱熔融-水热法利用粉煤灰合成沸石的研究

以粉煤灰为原料,采用Na2CO3碱熔融-水热法合成沸石产品.通过正交实验方案,考察NaOH浓度、晶化温度、晶化时间和固液比等实验因素对合成沸石产率和种类的影响,优化合成沸石条件.通过极差分析得知:NaOH浓度对合成沸石的产率和种类均有重要影响,当NaOH=2mol/L时,合成沸石的产率可达56%,沸石的种类有钙十字沸石、方沸石、丝光沸石和浊沸石.     

Kinetics of nickel leaching from roasting-dissolving residue of spent catalyst with sulfuric acid

Sulfuric acid leaching process was applied to extract nickel from roasting-dissolving residue of a spent catalyst, the effect of different parameters on nickel extraction was investigated by leaching experiments, and the leaching kinetics of nickel was analyzed. The experimental results indicate that the effects of particle size and sulfuric acid concentration on the nickel extraction are remarkable; the effect of reaction temperature is mild; while the effect of stirring speed in the range of 400–1 200 r/min is negligible. Decreasing particle size or increasing sulfuric acid concentration and reaction temperature, the nickel extraction efficiency is improved. 93.5% of nickel in residue is extracted under suitable leaching conditions, including particle size (0.074–0.100) mm, sulfuric acid concentration 30% (mass fraction), temperature 80 °C, reaction time 180 min, mass ratio of liquid to solid 10 and stirring speed 800 r/min. The leaching kinetics analyses shows that the reaction rate of leaching process is controlled by diffusion through the product layer, and the calculated activation energy of 15.8 kJ/mol is characteristic for a diffusion controlled process. Foundation item: Project (50574101) supported by the National Natural Science Foundation of China; Project (2003UDBEA00C020) supported by the Collaborative Project of School and Province of Yunnan Province, China     

电镀污泥中铬的无害化处理及动力学分析

以某表面处理工业园电镀废水处理污泥为研究对象,以铬浸出率为指标,通过对重金属的浸出,分步回收达到无害化、资源化的目的.将污泥干燥、研磨,在不同浓度硫酸溶液中浸出,控制浸出时间、浸出温度和搅拌速率;浸出完成后抽滤使浸出液与残渣分离.采用正交试验法,确定对铬浸出效果影响因素的顺序为:硫酸浓度>搅拌速度>浸出时间>固液比.通过单因素优化试验,结果显示:当浸出温度为25 ℃、固液比为1∶15、浸出时间为20 min、搅拌速率为800 r/min、硫酸体积分数为30%时,铬的浸出率最高.最后用黄钠铁矾法除铁,用焦亚硫酸钠还原六价铬,用氢氧化钠分步沉淀铬、镍重金属,锌则继续留在溶液中.电镀污泥的浸铬实验的浸出动力学研究结果表明硫酸作为浸出剂的反应级数为1,反应的速率常数为:k=0.053 2e-4.52/RT.     

氨浸置换法资源化废杂铜的研究

采用碳铵浸取-置换沉积新方法资源化、无害化回收废杂铜.考察浸取时间、氨水浓度、碳酸铵用量、催化剂用量对浸取效果的影响,并讨论还原剂用量、置换时间及搅拌速度对沉积过程的影响.结果表明:在(NH4)2CO3:6.5 g,NH3.H2O:6 mol/L、浸取催化剂:12 mL、浸取时间3 h的最优条件下,铜、锌的浸取率分别达到93.12%、95.03%,而伴生元素留在滤渣中.二段置换沉积过程中,在最佳工艺条件下,制备出产品纯度与附加值高的铜粉和七水硫酸锌.铜的回收率达到98.3%.该工艺具有操作简单、生产效率高、成本低、无污染等特点.     

Zinc leaching from electric arc furnace dust in alkaline medium

Physical and chemical properties of electric arc furnace (EAF) dust from Tianjin seamless Pipe Company were measured and analyzed. The zinc leaching tests in alkaline medium were carried out under variation of leaching agent concentration, leaching temperature, leaching cumulative time and solid-to-liquid ratio. The thermodynamics and kinetics of the zinc leaching process were also analyzed. The results show that the EAF dust contains 10% (mass fraction) zinc and the median particle size is 0.69 μm. The zinc recovery of 73.4% is obtained under the condition of 90 °C, 6 mol/L NaOH, and 60 min leaching time. With the increase of concentration of NaOH and the cumulative time, zinc leaching will be significantly increased. The kinetics study demonstrates that the leaching reaction is chemically controlled and the reaction activation energy is 15.73 kJ/mol.     

电镀污泥中铜的浸出工艺及其动力学

以硫酸为浸出剂,对某表面处理工业园电镀废水处理污泥中的铜做了浸出试验研究.将污泥干燥、研磨,X射线衍射和X射线能谱仪分析表明污泥中含铜19.03%.采用单因素优化试验探讨了固液比、反应时间、浸出温度、硫酸质量分数、搅拌速度对铜浸出率的影响.结果表明:当硫酸质量分数为20%,固液比为1∶10,搅拌速率为700r/min时,在20℃下反应40min,铜的浸出率可达97%以上;根据未反应核收缩模型,对硫酸浸铜过程的动力学机理进行了研究,结果表明:硫酸浸铜过程的控制步骤为固体膜扩散控制,其反应级数为0.828 2,浸出活化能为11.809kJ/mol.研究为含铜电镀污泥安全处置提供理论依据.     

Techniques of copper recovery from Mexican copper oxide ore

Mexican copper ore is a mixed ore containing mainly copper oxide and some copper sulfide that responds well to flotation. The joint techniques of flotation and leaching were studied. The results indicate that an ore containing 19.01% copper could be obtained at a recovery ratio of 35.02% by using sodium sulfide and butyl xanthate flotation. Over 83.33% of the copper oxide can be recovered from the tailings by leaching in suitable conditions, such as 1 h stirring at a temperature around 25 ℃ with a mixing speed of 500 r/min, an H2SO4 concentration of 1.0 mol/L and a mass ratio of the ore-slurry-liquid to solid (mL/mS) of 3. The overall yield of refined ore after flotation and leaching is over 89.18% of the copper, which is much better than sole flotation or leaching. A copper product containing more than 99.9% copper was obtained by using the process: flotation-agitation leaching-solvent extraction-electro-winning.     

Molybdenum removal from copper ore concentrate by sodium hypochlorite leaching

The removal of molybdenum from a copper ore concentrate by sodium hypochlorite leaching was investigated. The results show that leaching time, liquid to solid ratio, leaching ternperature, agitation speed, and sodium hypochlorite and sodium hydroxide concentrations all have a significant effect on the removal of molybdenum. The optimum process operating parameters were found to be: time, 4 h: sodium hydroxide concentration, 10%; sodium hypochlorite concentration, 8%; liquid to solid ratio, 10:1; temperature, 50℃; and,agitation speed, 500 r/min. Under these conditions the extraction of molybdenum is greater than 99.9% and the extraction of copper is less than 0.01%. A shrinking particle model could be used to describe the leaching process. The apparent activation energy of the dissolution reaction was found to be approximately 8.8 kJ/mol.     

Electric-oxidation extraction of molybdenite concentrate in alkaline NaCl electrolyte

The dissolution of molybdenite concentrate in NaCl electrolyte was investigated. The results show that the dissolution rate increases with the increase in liquid-to-solid ratio, stirring speed, NaCl concentration and temperature. When the liquid-to-solid ratio is 30:1, stirring speed is 400 r/min, concentration of NaCl is 4 mol/L at pH=9 and room temperature, the leaching efficiency of molybdenite concentrate will reach 99.5% in 240 min. Molybdenite concentrate cannot be electro-oxidized directly on the anode. The kinetic studies show that the dissolution of molybdenite concentrate is represented by shrinking core model with diffusion through a porous product layer of element sulfur, and the apparent activation energy for the dissolution reaction is 8.56 kJ/mol.     

硼铁矿中硼镁铁的硫酸法共浸出研究

提出一种采用硫酸酸浸硼铁矿使其中的硼、镁和铁元素共同浸出的方法．硫酸酸浸硼铁矿时,主要是与矿物中的硼镁石[Mg（BO2）（OH）]、磁铁矿[Fe3O4]、蛇纹石[Mg3Si2O5（OH）4]反应．通过热力学分析,验证采用硫酸共溶硼铁矿中的硼、镁和铁元素的可行性,并考察硫酸浓度、液固比、酸浸时间和酸浸温度对酸浸的影响,确定硫酸酸浸硼铁矿的最佳工艺条件：硫酸的质量分数30％,液固比（质量比）8：1,浸出温度90℃,浸出时间120min,搅拌速度大约100r／min．在最佳浸出条件下,硼铁矿中的硼、镁和铁元素的浸出率分别达到99．0％,91．0％,92．9％以上,达到了硼铁矿中硼、镁和铁元素共浸出的目的．     

砷灰的提纯工艺

采用湿法-火法对粗砷进行提纯,在氨水溶液浸出浓度为5～6 mol/L,液固比为1.7～2.2,浸出温度为50～65℃,浸出时间为3～4 h的工艺条件下,砷的浸出率可达95%以上,浸出液经蒸发结晶得到亚砷酸氢铵,亚砷酸氢铵经烘干后在500℃下分解得白砷产品,纯度可达99.5%以上,浸出剂氨经蒸发得到回收。该工艺无三废排放,原料消耗少,具有较好的经济和环境效益。