凝汽机组低品位化供热改造及其应用前景

简要分析了供热过程的最低理论燃料单耗问题,详细论述了常规自然对流散热器所决定的采暖供热温度匹配问题及其对策。并依据基于电量的统一性能评价方法,分析计算了凝汽机组低品位化供热改造的性能指标,以及配合高效末端散热器,适应用户末端散热器温度匹配能够取得的节能效果,并从总能系统的角度进行了阐述。还对这种改造的特点及其应用进行了分析。

都龙超大型锡锌多金属矿床“低品位矿化废石资源”特征与分布规律

云南省马关县都龙锡锌多金属超大型矿床是滇东南3大锡石硫化物矿床之一,铜街—曼家寨(铜曼)矿段是矿区最主要矿段,随着生产规模的扩大,矿山地质资源逐步被消耗,寻找可接替资源是当前矿山紧迫任务。本文研究表明,铜曼矿段内"低品位矿化废石资源"分布广泛,储量巨大,具有重要开采价值。通过该矿段地质勘探与生产的总结,发现"低品位矿化废石资源"与锡锌多金属工业矿体空间分布基本一致,下寒武统新寨岩组第2段和第3段碳酸盐岩夹片岩是最有利的赋矿地层,其中工业主矿体周围、大理岩与片岩层间剥离断层带和断层破碎带是该类资源的主要矿化类型。"低品位矿化废石资源"与工业矿石矿物组成和金属矿物嵌布特征基本一致,仅矿石矿物所占比例相对较低,该类资源以夕卡岩型硫化矿石为主,其次为大理岩型和片岩型硫化矿石;矿石中Sn、Zn和Cu等3个矿化元素均呈较好正相关关系,表明这些有价元素在矿石中共生关系密切。此外,矿区平面上成矿元素分带明显,由西至东,Zn逐渐变低,Sn和Cu则呈现明显升高趋势,而由北至南,呈现Sn-Cu-Zn(北部)→Zn(中部)→Sn-Zn(南部)变化趋势。该研究成果为都龙矿山"低品位矿化废石资源"开采和富集回收提供了新的依据。

Multi-effect distillation system for seawater desalination driven by tidal energy and low grade energy

A multi-effect distillation technology for seawater desalination driven by tidal energy and low grade energy is presented.In the system,tidal energy is utilized to supply power instead of coventional electric pumps during the operation,resulting in the decrease of dependence on steady electric power supply and a reduction in the running costs.According to the technological principle,a testing unit is designed and built.The effects of the feed seawater temperature and the heat source temperature on the unit performance are tested and analyzed.The experimental results show that the fresh water output is 27 kg/h when the heating water temperature is 65 ℃ and the absolute pressure is 25 kPa.The experimental and theoretical analysis results indicate that the appropriate heating water temperature is a key factor in ensuring the steady operation of the system.

Leaching kinetics of low-grade copper ore containing calcium-magnesium carbonate in ammonia-ammonium sulfate solution with persulfate

The leaching kinetics of copper from low-grade copper ore was investigated in ammonia-ammonium sulfate solution with sodium persulfate. The effect parameters of stirring speed, temperature, particle size, concentrations of ammonia, ammonium sulfate and sodium persulfate were determined. The results show that the leaching rate is nearly independent of agitation above 300 r/min and increases with the increase of temperature, concentrations of ammonia, ammonium sulfate and sodium persulfate. The EDS analysis and phase quantitative analysis of the residues indicate that bornite can be dissolved by persulfate oxidization. The leaching kinetics with activation energy of 22.91 kJ/mol was analyzed by using a new shrinking core model （SCM） in which both the interfacial transfer and diffusion across the product layer affect the leaching rate. A semi-empirical rate equation was obtained to describe the leaching process and the empirical reaction orders with respect to the concentrations of ammonia, ammonium sulfate and sodium persulfate are 0.5, 1.2 and 0.5, respectively.

Technological conditions and kinetics of leaching copper from complex copper oxide ore

The kinetic behavior of leaching copper from low grade copper oxide ore was investigated. The effects of leaching temperature, H2SO4 concentration, particle size of crude ore and agitation rate on the leaching efficiency of copper were also evaluated. And the kinetic equations of the leaching process were obtained. The results show that the leaching process can be described with a reaction model of shrinking core. The reaction can be divided into three stages. The first stage is the dissolution of free copper oxide and copper oxide wrapped by hematite-limonite ore. At this stage, the leaching efficiency is very fast （leaching efficiency is larger than 60%）. The second stage is the leaching of diffiuent copper oxides, whose apparent activation energy is 43.26 kJ/mol. During this process, the chemical reaction is the control step, and the reaction order of H2SO4 is 0.433 84. The third stage is the leaching of copper oxide wrapped by hematite-limonite and silicate ore with apparent activation energy of 16.08 kJ/mol, which belongs to the mixed control.

Roasting oxidation behaviors of ReS_2 and MoS_2 in powdery rhenium-bearing, low-grade molybdenum concentrate

The oxidation roasting process of molybdenum concentrate has significant advantages in industrial applications.However,utilization of low-grade mineral has many problems because it is more complex than the standard concentrate.In this study,the oxidation behaviors of powdery rhenium-bearing low-grade molybdenum concentrate were investigated through thermodynamic calculation,roasting experiments,thermogravimetric analysis,and phase analysis.The results obtained show that oxidation of MoS2 begins at 450℃,and MoO3 reacts with metal-oxide forming molybdate at 600℃.Finally,MoO3 can be dissolved in ammonia with a maximum content of approximately 80%.The volatile speed of Re was considerably slower than the oxidation speed of MoS2 because the nonvolatile products ReO2 and ReO3 were generated in reactions among MoS2,SO2,and Re2O7.The final volatilization rate of Re was almost 70%.This study determined the problems related to the roasting of low-grade molybdenum concentrate,which lays the scientific foundations for subsequent enhancement of molybdenum and rhenium extraction.

Co-extraction of valuable metals and kinetics analysis in chlorination process of low-grade nickel-copper sulfide ore

To efficiently co-extract Ni and Cu from low-grade nickel-copper sulfide ore,chlorination roasting with NH;Cl followed by a water leaching process was investigated.The results show that 98.4%Ni and 98.5%Cu can be synchronously extracted when the ore particle size is 75-80μm,the roasting time is 2 h,the mass ratio of NH;Cl to ore is 1.6:1 and the roasting temperature is 550°C.The evolution behavior of various minerals was elucidated using X-ray diffraction(XRD)coupled with scanning electron microscopy(SEM).The kinetics of the chlorination process based on the differential thermal and thermogravimetric analysis(DTA-TG)data was analyzed by Kissinger method and Flynn-Wall-Ozawa(FWO)method.The chlorination process of low-grade nickel-copper sulfide ore mainly contains two stages:the decomposition of NH;Cl and the chlorination of ore.The maximum apparent activation energies(Ea)at two stages are determined to be 114.8 and 144.6 kJ/mol,respectively.The condensed product of exhaust gas is determined to be ammonium chloride,which can be recycled as the reactant again,making the process economic and clean.

Improved recovery of a low-grade refractory gold ore using flotation–preoxidation–cyanidation methods

In this work, different flotation–preoxidation–cyanidation methods are considered for treating a lowgrade refractory gold ore. On the one hand, the results of selective flotation show that 22% and 31.1%of total Sb and As, respectively, remained in the final tailings and only about 28% of the total Au remained for further cyanidation processes. On the other hand, in bulk method of flotation the maximum Au recovery of 90.6% achieved after 60 min of flotation at the grind size with K80 of 146 micron. In addition, the bulk flotation method resulted in the concentrate with low concentrations of Sb and As elements. To improve the recovery of low-grade refractory gold ores, flotation should be followed by roasting, biological, or pressure oxidation processes so that the gold could be liberated prior to cyanidation processes. It is also found that the pressure oxidation pre-treatment of the concentrates prior to cyanidation may yield high gold recoveries of over than 83%. In these processes, recoveries are controlled by the temperature and the oxygen partial pressure in the solvent. However, by utilizing the bio-oxidation technique, the oxidation of sulfur to sulfate cannot be completed and, consequently, the gold recovery may be limited to only 72.2%.

Transformation and leaching kinetics of silicon from low-grade nickel laterite ore by pre-roasting and alkaline leaching process

The mineralogical phase transformation of a low-grade nickel laterite ore during pre-roasting process and the extraction of silicon during alkaline leaching process were investigated.The results indicate that the reaction activity of nickel ores is effectively improved by pre-roasting at650°C for2h,because of the transformation of lizardite into magnesium olivine and protoenstatite.When finely ground ore samples(44-61μm)pre-roasted firstly react with sodium hydroxide solution(60g/L)with a solid/liquid ratio of1:5at140°C for120min,the extraction of silicon can reach89.89%,and the other valuable elements of magnesium,iron and nickel are accumulated in the solid residues.The leaching kinetics of nickel laterite ore can be described successfully by the diffusion through the product layer control model.The activation energy is calculated to be11.63kJ/mol and the kinetics equation can be expressed as1-3(1-x)2/3+2(1-x)=13.53×10-2exp[-11.63/(RT)]t.

Recovery of Zn, Pb, Fe and Si from a low-grade mining ore by sulfidation roasting-beneficiation-leaching processes

To recover Zn, Pb, Fe and Si from a low-grade mining ore in the Lanping basin, Yunnan Province, China, a novel technology using the roasting with pyrite and carbon followed by beneficiation and hydrochloric acid leaching was proposed. Firstly, several factors such as pyrite dosage, roasting temperature, carbon powder dosage, holding time and particle size affecting on the flotation performance of Zn(Pb) and magnetic separation performance of Fe were simultaneously examined and the optimum process parameters were determined. A flotation concentrate, containing 17.46% Zn and 3.93% Pb, was obtained, and the Zn and Pb recoveries were 86.04% and 69.08%, respectively. The obtained flotation tailing was concentrated by a low-intensity magnetic separator. The grade of iron increased from 5.45% to 43.45% and the recovery of iron reached 64.87%. Hydrochloric acid leaching was then carried out for the magnetic separation tailing and a raw quartz concentrate containing 81.05% SiO2 was obtained. To further interpret the sulfidation mechanism of smithsonite, surface morphology and component of the sample before and after reactions were characterized by XRD and EPMA-EDS. The aim was to achieve the comprehensive utilization of the low-grade mining ore.

Cooperative removal of Mn^(2+),NH_(4)^(+)-N,PO_(4)^(3-)-P and F−from electrolytic manganese residue leachate and phosphogypsum leachate

Electrolytic manganese residue leachate(EMRL)contains plenty of Mn^(2+) and NH_(4)^(+)-N,and phosphogypsum leachate(PGL)contains large amounts of PO_(4)^(3-)-P and F^(-).Traditional methods of EMRL and PGL discharge could seriously damage the ecological environment.In this study,an innovative method for cooperative removal Mn^(2+),NH_(4)^(+)-N,PO_(4)^(3-)-P,F^(-)from PG and POFT was studied.The result showed that Mn^(2+),PO_(4)^(3-)-P and F^(-)were mainly removed in forms of Mg_(3)Si_(4)O_(10)(OH)_(2),Mn_(3)O_(4),Mn_(3)(PO_(4))_(2),Mg_(3)(PO_(4))_(2),CaSO_(4)·2H_(2)O,MnF_(2),MnOOH and Ca_(2)P_(2)O_(7)·2H_(2)O,when LG-MgO was used to adjust the pH value of the system to 9.5,and the volume ratio of EMRL and PGL was 1:4,as well as reaction for 1 h at 25℃.NH_(4)^(+)-N was mainly removed by struvite precipitate,when the molar ratio of N:Mg:P was 1:3:2.4.The concentrations of Mn^(2+),NH_(4)^(+)-N and F^(-)were lower than the integrated wastewater discharge standard.The concentration of PO_(4)^(3-)-P decreased from 254.20 mg/L to 3.21 mg/L.This study provided a new method for EMRL and PGL cooperative harmless treatment.

Preparation of sodium manganate from low-grade pyrolusite by alkaline predesilication-fluidized roasting technique

Low concentration alkaline leaching was used for predesilication treatment of low-grade pyrolusite. The effects of initial NaOH concentration, liquid-to-solid ratio, leaching temperature, leaching time and stirring speed on silica leaching rate were investigated and the kinetics of alkaline leaching process was studied. The results show that silica leaching rate reached 91.2% under the conditions of initial NaOH concentration of 20%, liquid-to-solid ratio of 4：1, leaching temperature of 180 ℃, leaching time of 4 h and stirring speed of 300 r/min. Shrinking-core model showed that the leaching process was controlled by the chemical surface reaction with activation energy Ea of 53.31 k J/mol. The fluidized roasting conditions for preparation of sodium manganate were optimized by the orthogonal experiments using the desiliconized residue. The conversion rate of sodium manganate was obtained to be 89.7% under the conditions of silica leaching rate of 91.2%, NaOH/MnO2 mass ratio of 3：1, roasting temperature of 500 ℃ and roasting time of 4 h, and it increased with the increase of silicon leaching rate.

Bioleaching of low-grade copper sulfide ores by Acidithiobacillus ferrooxidans and Acidithiobacillus thiooxidans

The grown conditions of Acidithiobacillus ferrooxidans and Acidithiobacillus thiooxidans were investigated,and then experiments were conducted to research the bioleaching behaviors of crude ore of copper sulfide and hand-picked concentrates of chalcopyrite by Acidithiobacillus ferrooxidans and Acidithiobacillus thiooxidans.The experimental results show that the bacteria grow best when the temperature is(30±1) °C and the pH value is 2.0.The bacteria concentration is 2.24×107 mL-1 in this condition.It is found that the copper extraction yield is affected by the inoculum size and the pulp density and the extraction yield increases as the inoculum size grows.The bioleaching rates reach their highest point in sulfide copper and chalcopyrite with a pulp density of 5% and 10%,respectively.Column flotation experiments of low-grade copper sulfide ores show that the bioleaching recovery reaches nearly 45% after 75 days.

Beneficiation of a low grade limestone sample

Pilot scale column flotation studies were conducted on a low grade siliceous limestone ore. Silica content was reduced to less than 13g in the concentrate so that it became satisfactory for use in the paper or rubber industries. The limestone sample was crystalline and constituted primarily of calcite that contained quartz, feldspar, pyroxene, and biotite as gangue minerals. Quartz is the major silicate gangue whereas feldspar, pyroxene, and biotite exist in minor to trace quantities. Traces of pyrite were also observed within the sample. A reverse flotation process was adopted where the silicate gangue minerals were floated using two different commercial cationic collectors： Chem-750 F or Floatamine-D. The studies clearly suggest it is possible to produce a limestone concentrate assaying around 96-97% CaCO3 containing less than 1 % Si02. The effect of feed flow rate, percent solids, froth depth, and wash water on the grade and recovery of the CaC03 concentrate is discussed.

A new approach for recycling arsenic and tin from low-grade tin middlings using a self-sulfurization roasting

Massive amounts of low-grade tin middlings have been produced from tin tailings,in which arsenic and tin are worthy to be recycled.Owing to high sulfur content in these tin middlings,a novel self-sulfurization roasting was proposed to transform,separate and recover arsenic and tin in this research.There was no extra curing agent to be added,which decreased the formation of pollutant S-containing gas.The self-sulfurization process involved a two-stage roasting of reduction followed by sulfurization.First in reduction roasting,FeAsS decomposed to FeS and As and the As then transformed to As_(4)(g)and As_(4)S_(4)(g),via which the arsenic was separated and recovered.The arsenic content in the first residue could be decreased to 0.72 wt.%.Accompanied with it,the FeS was firstly oxidized to Fe_(1−x)S and then to SO_(2)(g)by the coexisted Fe_(2)O_(3),and finally reduced and combined with the independent Fe_(2)O_(3)to form Fe_(1−x)S.In the followed sulfurization roasting,the Fe_(1−x)S sulfurized SnO_(2)to SnS(g),due to which tin could be recovered and its content in the second residue decreased to 0.01 wt.%.This study provided an efficient method to separate and recover arsenic and tin from low-grade tin middlings.