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By H. Matsuda, T. Iwashita, H. Awata, H. Liwei, D. Hirabayashi, Y. Saito, S. Ozawa, Y. Goto

"The present study is concerned with investigating the effect of chemical treatment on enhancing the HCl absorption capacities of coal burnt ash and slag. It was found that the HCl absorption capacity of coal burnt ash discharged from fluidized bed combustor was improved by hydration with water and 5% polyethylene glycol solution. This phenomenon was based on an increase of the specific surface area of treated coal burnt ash. The HCl absorption capacity of hydrated ash with polyethylene glycol solution was almost equivalent to that of standard CaO and Ca(OH)2. In particular, the specific surface area of hydrated ash with polyethylene glycol solution was found to be 19 m2/g, which was more than over eight times larger than that of raw sample. Hydrated slag treated with lON sodium hydroxide solution also showed high HCl absorption capacity, with highly improved surface area. It was found that the absorption capacity of slag for HCl depended on the treatment time, showing the maximum absorption capacity for 12 hours of hydration treatment.IntroductionA large amount of coal burnt ash, slag and dust are discharged from such industrial processes as coal combustion and metal refinery etc., and it becomes an serious problem to recover and diminish these solid wastes because of the shortage of dump sites and subsequent metal elution to the environment. Figure 1 shows the present situation of reutilization of coal burnt ash and slag in Japan [1,2]. It is shown in this figure that these solid wastes are recycled as materials in the fields of earthwork, agriculture and forestry etc. A part of these solid wastes, however, is still dumped to landfill. To decrease this dumped amount, the development of environmentally friendly reutilization technique for these solid wastes is of great importance."

Jan 1, 2000

By William Petruk

The concentrator of the Iron Ore Company of Canada consists of spiral circuits, a cone plant and a magnetics plant with regrind. The proportions of liberated and unliberated grains of iron oxides at each point throughout the concentrator were measured by image analysis, and recoveries were calculated for each particle type by the MATBAL materials balancing program. The results show that spirals and cones recover more than 92% of the liberated iron oxide grains in the size range of 848 to 150 um,, but the recovery drops off rapidly for finer grained particles. The tailings from the spirals and cones are reground and low intensity magnetic separators are used to recover liberated magnetite in the -37.5 um size range. The concentrator does not have a unit that efficiently recovers iron oxide grains in the size range of 150 to 37.5 um Consequently about 25% of the iron oxides in the ore are lost to tailings as liberated. grains in the size range of 150 to 37.5 um

Jan 1, 1989

By Jack S. McPartland

The recent developments on the concentration and recovery of gold from leached gold thiosulfate solutions by solvent extraction, reduction-precipitation, and electrochemical deposition are reviewed in this paper. Alkyl phosphorus esters have been found to extract gold only in alkaline thiosulfate solutions, The recovery of gold increases with thiosulfate concentration in the aqueous phase and with alkyl phosphorus esters concentration in the organic phase with aliphatic hydrocarbon diluents. The addition of ammonia improves the extraction. The extraction of gold with amines in either aliphatic or aromatic diluent decreased in the following order: primary>secondary>tertiary amines. Increase in the extractant concentration at high pH increased the extraction of gold. At pH < 9, the primary amine, RR&apos;CHNH2 with 1&apos;9-23 carbon atoms is the main extractant with TBP as the synergistic reagent and at pH >9, the primary amine shows a synergistic effect on the extraction of gold&apos; with TBP. TheAu(I) ions in thiosulfate solution are reduced by sodium borohydride (SBH) to metallic gold, with particles in the sub micron range, At an SBH/ Au ratio of 6.25, pH = 6, and at room temperature, the I e ppm Au(I) in a solution of ] 0 gil thiosulfate was completely reduced to metallic gold in about 60 minutes.

Jan 1, 1999

By Gus Van Weert

Solvent extraction of iron from zinc refinery liquors has been investigated before, mainly on the MEHPA / HC1 stripping combination. Leakage of chloride into the zinc electrolyte and the costs of handling the ferric chloride eluate have prevented this option from being adopted. This work explores the stripping of the ferric loaded organic phase with nitric acid to take advantage of the possibility of regenerating up to 6N nitric acid, and clean spheroidal hematite production by autoclave treatment of ferric nitrate solution. The work was carried out with DEHPA (di 2-ethyl-hexyl phosphoric acid) in kerosine. Degradation testing proved the DEHPA to be very stable, when exposed to nitric acid solutions up to 6N and temperatures up to 70" C over a period of five days. Ferric iron was removed completely froma synthetic hot acid leach liquor, irrespective of the DEHPA concentration. Stripping the ferric turned out to be the problematic step. Although the stripping efficiency was considerable better at 50 than at 20° C, it decreased with increasing DEHPA/Fe ration in the organic phase. Eluates of no more than 5-10 g/l ferric nitrate were produced. surprisingly, nitric acid was found to be a less efficient stripping agent than sulphuric at equivalent normality and its efficacy levelled off above 4N. It was concluded that DEHPA does not appear to be the S.X. reagent to take advantage of the low cost autoclave decomposition of ferric nitrate opportunity in zinc hydrometallurgy and that another S.X. reagent, also resistant to nitric acid, but based on weaker compound formation is needed.

Jan 1, 1998

By Takaaki Shibata

The Tamano Refinery has developed two new processes for removing the impurities, arsenic, antimony, and bismuth from an electrolyte. The first process involves bleeding the surface electrolyte in the electro-refining cells. The copper concentration level in the outermost layer is lower than that in any of the other layers, while the impurity levels are almost the same in all layers. Bleeding the surface electrolyte increases the amount of impurities removed in the liberator by approximately 20%. The second process involves adsorbing antimony and bismuth from the electrolyte by means of chelating resin. As the electrolyte passes through the packed column of chelating resin, antimony and bismuth are selectively adsorbed. They are then desorbed by means of hydrochloric acid, and then recovered in the form of a highly concentrated antimony chloride solution by means of evaporation. This process almost doubles the capacity for removing antimony from an electrolyte.

Jan 1, 1987

By Vaughan R. Voller

"Solidification systems from conduction controlled freezing of a pure material through to alloy phase change that include flow and segregation processes are described. The components in a general solidification model-variables, governing transport equations, and local scale relationship-are outlined. A central focus is the presentation of an explicit procedure for coupling between macroscopic and local scale variables.IntroductionSince the seminal work of Byres et al.[!] in the 1940's there has been a steady progress in modeling solidification systems with increasing levels of complexity. The object of this paper is to track the evolution of these solidification models and present a general modeling approach that applicable to a wide range of the solidification systems of current interest.Solidification SystemsIn a broad sense the underlying physics in a solidification process occur at two distinct length scales:1. The macroscopic scale of the process over which the governing transport equations of heat, mass and momentum hold.2. The local scale at the solid-liquid interface where the solidification phenomena are operative."

Jan 1, 1999

By R. A. Sharma

Neodymium metal was successfully produced by the reduction of Nd203 with Na in the presence of CaC12-NaCl melts by the overall reduction reaction Nd203 + 3 CaC12 + 6 Na ? 2 Nd + 3 CaO + 6 NaCl at -750°C. The metal produced was extracted from the salt media by dissolving it in a Nd-Fe alloy pool. The yield (percentage of theoretical neodymium produced) was found to depend on the salt composition and the mass ratio of salt to neodymium in the reactor. A yield of 95% was obtained with salt ratios as low as 3.5. Further decrease in the salt ratio below 3.25 decreased the yield. Stirrer position had little effect on yield; while increasing the temperature and placing fins in the reactor increased the yield. The Nd-Fe alloy produced was of as good quality as that produced using Ca reductant, and was suitable for direct use in producing General Motors MAGNEQUENCH? alloy for permanent magnets.

Jan 1, 1989

By Qingchun Yu, Baoqiang Xu, Yongnian Dai, Bin Yang, Fulong Zhu

"In this paper, the mechanism of the carbothermic reduction-chlorination process to extract aluminum from alumina in vacuum was investigated. The carbon monoxide (CO) content of the gas, phases of the reduction slag, surface morphology of the reduction slag and the product aluminum were measured by means of gas chromatography (GC), XRD, and SEM. The experimental results indicated that the content of CO in the gas decreased, but Al4C3 increased with time at 1773K during the carbothermic reduction process of alumina. AlCl(g) and graphite (C) were produced by the reaction between Al4C3 and AlCh(g) at 1773K, and then aluminum beads generated after the decomposition of AlCl(g). It implied the mechanism of alumina carbothermic reduction-chlorination process in vacuum involves three steps as follows: the carbothermic reduction of A}z03 to produce Al4C3, the chlorination of Al4C3 by AlCh(g) to produce AlCl(g), and the decomposition of AlCl(g) to produce Al(l).IntroductionAluminum is the basic raw material for many industries [1], and it is used extensively because of various excellent properties such as light weight, good malleability, high strength, excellent thermal conduction, high corrosion resistance and so on [2]. Industrial production of aluminum currently includes refining alumina and electrolytic aluminum processes. However, its development has been effected by some factors such as high energy consumption and heavy environmental burdens [3]. In the past several decades, there had been numerous attempts to produce aluminum by new methods. Extracting aluminum from alumina or bauxite directly by carbothermic reduction-chlorination process had been studied [ 4-6]. Although the alumina carbothermic reduction process [7~9] had been studied, the mechanism of the carbothermic reduction-chlorination process had not been systematically elucidated.In order to investigate the mechanism of the carbothermic reduction-chlorination process in vacuum, GC, XRD and SEM techniques will be used to determine the CO content and to characterize the surface morphology and composition of the slags at various temperatures."

Jan 1, 2011

By K. J. Cathro

A hydrometallurgical process is described for the recovery of copper from chalcopyrite concentrates. The process is based on a cupric chloride leach, the concentrate having first been "activated" by heating it with sulphur at a temperature near 400°C. This activation process converts most of the iron of the chalcopyrite to pyrite, which is insoluble in the cupric chloride leachant. The copper is converted to the mineral idaite, which dissolves readily in hot cupric chloride: at 106°C, 98% of the copper is extracted in a three-stage countercurrent leach, the leaching tirqe being 1 hour per stage. The sulphur of the chalcopyrite appears as elemental sulphur, which is recycled for the activation of further concentrate, the final leach residue consisting of pyrite, plus the gangue present in the original concentrate. The copper extracted can be recovered by electrowinning, using a diaphragm cell. Current efficiency (one-electron basis) is about 70% but the cell voltage is only about 1 V, so that power consumption is three times lower than for the conventional sulphate electrowinning process. The leachant is simultaneously regenerated in the anode compartment. The behaviour of minor elements in the concentrate has also been studied.

Jan 1, 1976

By Jianxin Zhang

A process, in which arsenic in the residue is transformed directly into scorodite and copper dissolves into the solution simultaneously, was studied. The process was carried out in acidic ferrous sulfate solution under oxygen pressure. The optimum conditions of the process were determined as follows: the initial H2SO4 concentration of 10 g/L; liquid/solid ratio of 10; Fe/As molar ratio in the feedstock of 1; the addition of calcium lignosulphonate was 1% of sludge; oxygen pressure, temperature and reaction time were 1.5 MPa, 150 °C and 6 h, respectively. Under these conditions, arsenic transformed to crystalline scorodite in size of 5–10 μm, and the transformation efficiency reached 97.60%; meanwhile, the copper was leached with a high efficiency of 93.34%. The stability of the leaching residue was evaluated by TCLP testing, and the results indicated the concentration of arsenic in the leaching solution was extremely low as 0.04 mg/L and met the standard of safe stockpile.

By Joseph Goodwill

1. PURPOSE The objective of this paper is to summarize a study undertaken by the EPRI Center for Metals Production (CMP). The overall purpose of the study was to identify electrolytic opportunities that might merit consideration for joint development by industry and EPRI. 2. SCOPE The electrolytic technologies of interest to this study include: Electrorefining, Electrowinning, and Fused-salt electrolysis. Recovery from primary, secondary, and waste materials were included in this effort; however, we excluded from consideration: Plating and coating operations, such as electrogalvanizing; Mining and milling operations in which electric energy is predominantly used for comminution; Electrothermal processes in which submerged-arc, open electric-arc, and plasma-based smelting processes use electric energy largely to provide a melting or refining ("heating") capability; and Aluminum electrotechnologies since these were the subject of a separate study (CMP 86-2). We recognize that there may be some "gray areas" between these excluded areas and the electrolytic sectors of interest. For this reason, a secondary criterium was used, namely, if the major use of electric energy was to extract or refine the

Jan 1, 1990

By W. K. Collins

Engineered Scavenger Compounds (ESCs), developed by the U.S. Bureau of Mines, are a novel class of compounds that can selectively recover a desired element from a solid or molten alloy. Lithium titanate (Li2Ti3O7 or L&apos;20-3T&apos;02) is being used as an ESC to recover lithium (Li) from Aluminum-Lithium (Al-Li) alloys. X-ray diffraction measurements have shown that lithium titanate undergoes a phase change from an orthorhombic structure to a hexagonal structure. This change is due to the incorporation of Li in the matrix of the material and the effect of temperature. Although both are metastable, the hexagonal phase that forms during the scavenging of Li from Al-Li alloys appears to be the more stable phase. Recovering Li from the ESC by electrodeposition does not cause the structure to revert to the orthorhombic phase. Both the orthorhombic and the hexagonal structures of Li2Ti3O7 have similar scavenging capacities for Li. This paper proposes a new mechanism for the phase transformation.

Jan 1, 1993

By J. Priede, G. Gerbeth

"Two different instability mechanisms of electromagnetically levitated bodies are analyzed. The first is due to the virtual coupling between the electric current passing through the magnetic system and the variation of position of the body. This mechanism can result in unstable mass center oscillations of a levitated solid body. Another type of instabilities may occur because of the coupling between the motion and the electric currents induced in the body. This effect can cause a spontaneous rotation of the body setting in as the frequency of the alternating magnetic field exceeds certain critical threshold depending on the configuration of the field. This instability can be suppressed by an axial steady magnetic field of strength comparable to that of the levitating field. 1 IntroductionAn alternating magnetic field interacting with eddy currents induced in a conducting body causes repulsion between the body and the coil generating the field. Besides, the eddy currents provide a contactless heating of the body. In such a way metallic samples can be levitated and melted without a contacting crucible. This technique allows to avoid contamination of the melt with container material and so to achieve a considerable super-cooling of the melt before it solidifies. Nowadays electromagnetic levitation is widely used in material research for measuring of the thermophysical properties of molten metals (2) .It has been observed since early fifties when the first experiments on the electromagnetic levitation were done that under certain circumstances solid levitated samples start to rotate or oscillate [l]. Surface oscillations can appear in addition when the sample is melted. Such phenomena are undesirable because they often result in failure of the levitation. Recently we showed that rotational as well as oscillatory instabilities can have electrodynamic origin related to additional eddy currents induced by motion of the body in an alternating magnetic field [3] , [4].In the first part of the present paper we analyze another possible cause of oscillatory instability due to the coupling between the electric current passing trough the magnetic system with variation of the position of the levitated body. The second part of the paper deals with the origin and suppression of rotational instability in a uniform alternating magnetic field."

Jan 1, 1999

By Joo Hyun Park, Soo Sang Park

The distribution behavior of Pb between molten copper and FetO-SiO2 (-CaO, Al2O3) slags was investigated at 1473 K (1200 oC) and p(O2)=10-10 atm in view of the reaction mechanism of Pb dissolution into the slag. The distribution ratio of Pb (LPb) decreases with increasing CaO content (~6 mass pct) irrespective of Fe/SiO2 ratio (1.4~1.7). However, the addition of alumina into a slag with Fe/SiO2=1.5 linearly decreases the LPb, whereas a minimum value is observed at about 4 mass pct Al2O3 at Fe/SiO2=1.3. The log LPb continuously decreases with increasing Fe/SiO2 ratio, and the addition of Al2O3 (5 to 15 mass pct) into the silica-saturated iron silicate slag (Fe/SiO2 < 1.0) yields the highest Pb distribution ratio. The log LPb linearly decreases by increasing the log (Fe3+/Fe2+) value. The Pb distribution ratio increases and the excess free energy of PbO decreases with increasing Cu2O content in the slag. However, from the viewpoint of copper loss into the slag, the silica-saturated system containing small amounts of alumina is strongly recommended to stabilize PbO in the slag phase at a low Cu2O content.

Jan 1, 2014

By S. H. Castro

The El Teniente molybdenite plant (Chile), with a nominal capacity of 3,200 tpd of bulk Cu-Mo concentrate, produced approximately 3,385 metric tonnes of Mo during 1998 by using the phosphate Nokes process (LR-744 reagent). This is the only Chilean moly plant which traditionally has depressed copper and iron sulfides with LR 744 reagent, which is a hydrolysis product of P2S5 and NaOH. El Teniente&apos;s ore Mo grade is 0.02%Mo and is concentrated to 0.45%-0.50% Mo in bulk Cu-Mo flotation. Subsequently, molybdenite is floated by selective flotation to depress Cu and Fe sulfides, such as chalcopyrite, bornite, digenite and pyrite, to produce typically a 49%-50%Mo molybdenite concentrate. An optimisation program during the last two years was undertaken to improve molybdenite recoveries. Current molybdenite plant technology at El Teniente and use of phosphate Nokes in the process process is reported. Use of phosphate Nokes reagent has increased average Mo recovery from 70.5% in 1977 to a mean 74.7% for the first 4 months of 1999. The global Mo recovery, including bulk flotation, incresed from 42.3% in 1997 to 46.8% in 1998, with a projection of 55.5% for 1999. Mo production per year has increased from 3,181 tons in 1997 to 3,385 tons in 1998 and a projected 4,000 tons for 1999.

Jan 1, 1999

By Zhongning Shi, Ramana G. Reddy, Min Li, Zhaowen Wang

Electrodeposition of cobalt from cobalt tetrafluoroborate (Co(BF4)2) was investigated using 1- butyl-3-methylimidazolium tetrafluoroborate (BMIMBF4) ionic liquid. The experiments were conducted at 353 K. Chronoamperometry experiments confirm that the electrodeposition of cobalt on tungsten electrode proceeds via three-dimensional instantaneous nucleation with diffusion-controlled growth process. The average diffusion coefficient of Co(II) was found to be 7.5 × 10-8 cm2s-1 at 353 K, which is in good agreement with the estimated value from cyclic voltammetry. The electrochemical deposit was characterized using SEM-EDS and XRD methods. The SEM image shows formation of a dense and compact deposit at - 0.75 V. The EDS and XRD analysis confirm that the obtained deposit is pure cobalt metal.

Jan 1, 2015

By Animesh Jha

"Fine ceramic powders (<500 nm) exhibit exceptional physical and mechanical properties in engineered structural ceramics. The production of fine powders, in particular the non-oxide ceramics, via a cheaper route than the organic solvent route has been rather elusive. This paper examines the physical chemistry of sub-oxide vapour-phase reduction reaction for the nucleation of non-oxide ceramic phase. Well known vapour species eg SiO and BO in the production of technical ceramic powders (SiC, BN) are particularly discussed for understanding the nucleation process of SiC and BN ceramic phases respectively. The regimes of partial pressures and temperatures are particularly 'identified. The calculated nucleation rate as a function of the temperature is compared with the experimental results on powder morphology. The production of amorphous and nanocrystalline h-BN powders is discussed in the context of substrate structure and thermodynamic parameters.IntroductionThe physical and mechanical properties of ceramic components are strongly dependent on the powder particle size and morphology used for their fabrication. For example the strength (Os) of the structural ceramic components increases with decreasing average grain-size (d) in accordance with the Hall-Petch relationship: Os = 0""0 + k/(d)1/2. Here k is an empirical constant which can be established from the graphical relationship. However as the particle size approaches submicrometer-size range, the dependence of strength on the particle size becomes more complex [1]. The design of ceramic components based on nanometer-size particles have recently become an important field which can offer ceramic materials with properties comparable to their metal counterparts. Several materials based on these concepts have been developed and reported by Niihara [2]. For example, the Al2O 3-SiC nanocomposite, having average B-SiC grain size 0.3-2um with intragranular SiC particulates of 5-10 nm has unusually high strengths (1100 MN/m2) at low and elevated temperatures combined with enhanced fracture toughness (> 4 MN-ml.5) which is an unusual combination in structural ceramics. Besides mechanical properties, as the particle size of ceramic powders approaches the nanometer-size range, the charge transport properties also improve by many times. The effect of particle size on the electrical conductivity of Si3N4-TiN composites has been well documented [3.4]. The size of ceramic particulates without doubt plays a pivotal role in determining the properties of ceramic components."

Jan 1, 1996

By Shenghui Guo, Ying Lei, Wei Li, Jinhui Peng, Libo Zhang, Yu Li

"Abstract: In this work, low temperature carbothermic reduction of Panzhihua low grade ilmenite after ball milling were investigated. The ball milling of ilmenite and graphite were processed in a ball mill. The milling process can attenuate the particle of ilmenite, and the ilmenite characteristic peak were broadened and weakened. The reduction temperature and behaviors are studied on the thermo gravimetric apparatus; lower temperature and higher reaction rate were obtained for the reduction of ilmenite with graphite after milled. The reduction temperature and times of ilmenite decreased to 1000-1025°C and 40-60min via milled for 4 or 8h; Low down the reduction temperature can eliminate the solution of Mg2+ and FeTi03 in a certain degree; the reduced ilmenite has an obvious loose structure, can favored the subsequent procedure such as leaching.1 IntroductionThere is about 9.66 billion tons of Vanadic Titan-magnetite deposited in Panzhihua, China. In this reserve, there is about 8.73x108 t ofTi02, Accounting for 90.54% of the Chinese reserves, and 35.17% of the international reserves and ranking first in the world. [JJ Titanium mineral deposits major classified into Rock type and Placer type. Rock type is the igneous mineral, which have great content of FeO (Relative to Fe20 3) and gangue. The rock textures are compact and almost are complex ore. Rock type Titanium mineral major are ilmenite and titan magnetite ore, difficult to separate the metal ore and gangue. l.21 Thus, the Ti02 grade of Rock type ilmenite much lower than the Placer type (Table land 11). Panzhihua ilmenite are the typical Rock-type, due to the higher contents of MgO, CaO and Si02 and the complex mineralogy in ilmenite in Panzhihua region, making its direct reduction progressively more difficult. Mechanical activation, usually carried out by energetic milling or ball milling, was reported to significantly decreased the reduction temperature and accelerate the reaction rate of reduction of ilmenite with carbon, silicon and Magnesium. f3·(>J It reported that the process temperature were decreased to 600-1000°C and the annealing time were reduced to 30-120min after milled for 50-400 h. Shorter milling time involved was imported for else procedure such as direct leaching. 171 Hardly any research had referred to the carbothermic reduction reactivity of ilmeite milled for a short time as 1-1 Oh. In the present work, the aim is to investigate the reactivity and reaction temperature of carbothermic reduction of Panzhihua ilmenite. The Phase composition and micro morphology of samples are characterized."

Jan 1, 2011

By Paul S. Korinko, Simona E. Hunyadi Murph

To address a radiological contamination event, a materials and process project was initiated. During the course of this project, processes and materials were developed that successfully captured zinc vapors physically and chemically, respectively. However, the physical traps were placed in an undesirable area and it was determined that the chemically captured zinc reactions was reversible. Consequently, advanced nanomaterials are being developed to capture zinc vapors. The current state of the art of the science and engineering of these materials and their efficacy for zinc capture and retention is discussed in this paper.

Jan 1, 2015

By Devon A. Hannan

Thermodynamic modeling calculations were undertaken for the recovery of mercury from HgTe using a vacuum retort reactor. However, discrepancies were noted in the available thermodynamic data which introduced uncertainty in the model results. Consequently, a series of transpiration experiments were first undertaken on the thermal decomposition of HgTe into an inert carrier gas stream according to the following reaction: HgTe(s) Hg(g) + Te(s) From the measured weight loss of the HgTe samples, the Gibhs free energy change for the above volatilization reaction was determined as a function of temperature. The experimentally determined Gibbs free energy values were found to be approximately 30% larger (more posi-tive) than the values available in databases such as HSC Chemistry. By conducting a second-law analysis on the experimental data, the heat of formation of HgTe at 298.15 K and the en-tropyat 298.15 K were determined, which then formed the basis for a new database for the modeling calculations. The results of this study showed that HgTe was more stable at elevated temperatures than was generally indicated in the literature. The new database is given in a for-mat for direct insertion in HSC Chemistry. Using the new database, modeling calculations were conducted to examine the effects of the re-tort pressure and the molar ratio of purge gas to HgTe on the volatilization of mercury from HgTe. A diagram is presented that shows the minimum temperature for 100% volatilization of mercury from HgTe as a function of retort pressure for different molar ratios of purge gas (N2) to HgTe. For example, it was found that the minimum temperature for 100% volatilization could be decreased by 175 K, at a retort pressure of 0.05 bar, if the molar ratio ofN2 to HgTe is increased from 1.0: 1 to 100.0: I.

Jan 1, 2003