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By H A. Challenor

Eyewash safety showers located in remote hot environments may be unsafe to use due to excess water temperature and high risk levels of water borne bacteria such as Legionella. Australian Standard AS 4775 specifies that dispensed water shall be bacteria free and have a maximum temperature of 37.8°C. Legionnaires disease is contracted when water aerosols containing Legionella bacteria are inhaled.Laboratory tests show 30 to 70 times more aerosols are generated by showers operating at 38°C than with showers operating at 25°C. Furthermore, excess temperature in flushing water can cause burn injuries leading to hospitalisation, skin grafts and/or permanent scarring. Excess temperature also increases the rate of reaction to corrosive substances, opens pores in the skin to allow greater dermal penetration and can cause dizziness and fainting by dilating blood vessels and reducing blood pressure.CITATION:Challenor, H A, 2017. Managing excess temperature and high risk levels of Legionella bacteria in eyewash safety showers, in Proceedings Iron Ore 2017, pp 337–346 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 24, 2017

By B Muller

The disposal of tailings has become a greater issue due to the changing economics of mineral extraction. The processing of lower grade deposits and tougher environmental regulations has led to the selection process for tailings disposal options becoming more complicated. Current selection methods are too time consuming or too easily manipulated for inexperienced users to use successfully. This paper examines a different method for the initial selection of a tailings disposal system, using a numerical system that was introduced into the mining method selection field. The numerical method system was compiled by reviewing previous work done on the positive and negative aspects of the tailings disposal methods used by the coal and metallurgical mining industries. Assessing the common trends and design aspects of each of the methods was undertaken and a ranking was designed to explain how well suited the tailings disposal methods are to a situation. The tool successfully analysed sites using common data to indicate methods for disposal that would have provided minimal technical problems, low environmental risk and been practically possible. The tool also proved successful in identifying critical areas that would need further investigation. This new system was tested on some disaster case studies from around the world in order to show that it could have successfully selected a different method for tailings disposal that would have prevented the disaster.

Sep 26, 2011

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Jan 1, 2000

By Kanagarajah Ravishankar

Drilled shaft (caisson) foundations are extensively used in the past. In the central part of New Jersey, the soils consist of silt and clayey silt overlying Triassic age shale bedrock. The shale is friable and generally severely weathered and highly fractured. The shale crumbles readily into small fragments. In general, the shale is encountered in shallow depth. Subsurface investigations indicated that the upper part of the shale is very soft, highly weathered, and able to advance through soil exploration methods. The underlain formation is relatively hard and required rock coring techniques for sapling. Individual core recoveries ranged from 10 to 100 percent and Rock Quality Designations (RQD) vary from 0 to 90 percent. Although several methods available to estimate the axial capacity of drilled shafts, it is a difficult task for the designer choose the suitable method due to self disintegrate nature of the shale. In the recent years several bridge foundations have been constructed using drilled shafts founded on shale rock in the vicinity of the central New Jersey. Based on load tests carried out on drilled shafts during construction, subsurface exploration, and laboratory tests, suitable methods for determining axial capacity of drilled shafts is analyzed. This paper presents the findings of the field tests and suitable methods to determine the axial capacity of drilled shaft founded on soft shale. Key words: Drilled shaft, axial capacity, Osterberg load test, shale.

Jan 1, 2006

By D. E. Spiller

Introduction Precious metals recovery from ore can generally be accomplished using gravity concentration, flotation, and/or hydrometallurgical (leaching) techniques. The objective of this paper is to show why gravity concentration can be an important part of recovery systems that employ leaching as the primary unit operation. A brief discussion of modem gravity concentration equipment is also presented. Discussion Gravity concentration of ores has generated increasing interest in recent years. Reasons for this interest include: • Gravity concentration is environmentally attractive. There is little or no use of reagents. Hence, it is relatively nonpolluting. • The cost of cyanide has continued to increase. Therefore, cost savings may be realized whenever leaching feed tonnage can be reduced by preconcentration. • Compared to flotation and leaching, gravity equipment costs are low per processed ton. Field installation costs for gravity circuits usually are less because many' units are supplied as self-contained modules. Also, the cost required to supply services, particularly power, to a gravity plant site are also less. In situations where preconcentration at coarse particle size is applicable, significant grinding equipment savings may be possible. • Gravity circuit operating costs are also relatively low compared to typical flotation and leaching circuits. Reagents, power, maintenance, and manpower savings in a well-engineered gravity plant may be realized. Again, if grinding is reduced, significant power and steel (media and liners) savings are possible. •In recent years, more efficient gravity concentrating devices have been developed. Benefits to Precious Metal Leaching Gravity beneficiation can complement precious metal leaching in two ways. First, the recovery of coarse liberated values before leaching may reduce leach time requirements and may reduce reagent consumption. Second, gravity preconcentration can reduce the size of a leach plant by decreasing the quantity of material to be leached. Coarse gold and silver have been shown to leach rather slowly. Kameda (1949) and Habashi (1967) have investigated the kinetics of cyanide leaching systems. They concur that in a heterogeneous reaction, the rate of gold and silver dissolution is directly proportional to the surface area. Thus, the instantaneous rate of dissolution for spherical 0.37 mm (400 mesh) gold is theoretically -25 times faster than for the same amount of gold at .841 mm (20 mesh), based on data from Fuerstenau, Chander, and Abouzeid (1979). Conversely, coarse liberated, +.841 mm (20 mesh), gold is more readily recovered by gravity concentration than is fine, -.037 mm (400 mesh) gold. Therefore, it is apparent that the two recovery systems complement one another. Figure 1 data demonstrates the potential synergism. A sample of - 3.327 mm (6 mesh) Nevada gold-bearing ore was cyanide leached using conventional bottle-roll test procedures. Gold extraction was determined as a function of leaching time. A second sample split from the same leach feed material was hand jigged to remove a coarse heavy mineral fraction, including virtually all of the +.210 mm (65 mesh) liberated free gold. This second sample, with the coarse gold and heavy minerals removed, was subjected to an identical cyanide leach procedure. Figure 1 presents the resulting comparative extraction data. Note that the percent gold extraction for the sample containing no +.210 mm (65 mesh) free gold includes the coarse gold recovered by gravity. The data show that the sample containing coarse gold required about 72 hours of leaching time to achieve 80% extraction. This compared to about 22 hours of leaching time for 80% gold recovery from the sample that contained only -.210 mm (65 mesh) free gold. Thus, there was a 69% reduction in leaching time. The improved extraction data is not wholly attributable to coarse gold removal, but rather it was the combination of gold removal and rejection of other heavy mineral cyanide consumers or leach retardants. Further investigation was not warranted at this time. Preconcentration is the second manner in which leaching systems can benefit from gravity concentration. The premise is that preconcentration can reduce the quantity of leach feed, which, in turn, may reduce leaching costs. Figure 2 presents preliminary data developed by CSMRI for US Minerals Exploration (USMX). Centennial Exploration Inc., in agreement with USMX, is proceeding with evaluations to determine the suitability of various processing schemes for recovery of gold values from the Montana Tunnels property. The data shows how the ore can be preconcentrated by gravity techniques to result in a reduced feed tonnage to secondary extraction techniques, presumably flotation or cyanide leaching. Testing has shown that Reichert cones, followed by treating the cone concentrate on spirals, can deliver about 88% gold recovery in about 13% weight, that is, 87% weight rejection. Consequently, fine grinding and reagent costs are attributable to only 13% of the plant feed rate. Cost data is not yet available, but the potential exists for significant cost savings.

Jan 1, 1985

By S. Warren Hobbs

The section of the Northwestern United States that includes the northern Rocky Mountains and adjacent areas to the west and east is one of large mineral production and important mineral potential. The metallic ore deposits vary widely in age and type and their geologic settings are similarly diverse. A brief resume of the region's physiographic characteristics and geology sets the framework for the selected ore districts that are described in following papers and makes it possible to place these ore deposits in their proper reference to space and time.

Jan 1, 1968

By Steve Dillingham

Blasting crews, explosive product drivers, drillers, blasters, supervisors, and salespeople all arrive daily at customer blast sites in vehicles of every type whether passenger cars, pickup trucks, or bulk trucks. The majority of these vehicles are company owned and re p re s e n t a substantial investment and risk for the company. That may explain why the costs resulting from blasting incidents is often surpassed by those involving motor vehicle accidents.

Jan 1, 2003

By Sheldon P. Wimpfen

Sand and gravel has been, and will continue to be, our principal construction material. Most of us are familiar with the patterns of past growth. Today we will examine the present status of this vital industry. We shall also attempt to look into the future to see where it is going and to evaluate the constraints that are being imposed. Natural aggregates are among the lowest in value of all bulk industrial minerals. From the processing point of view, sand and gravel production is relatively simple since the products are sized materials obtained from a deposit of satisfactory physical soundness. They can be used either loose or bound to provide bulk, strength, or protection. One of the main uses of coarse aggregate is in concrete as a filling material. As a general rule, the coarse aggregate constitutes about three-quarters of the volume of concrete. The other main use for aggregate is as a roadstone or roadbase to provide protective and stabilizing bulk beneath the road surface. Also it is used in a mixture with petroleum products as the wearing course on top. The sand and gravel industry is geographically widespread. In general, the output of operations is directly proportional to the amount of construction in any area. A wide variety of other aggregates compete with sand and gravel, including crushed stone, blast-furnace slag, clinker, and artificial and natural lightweight materials.

Jan 1, 1972

By R. A. Kipp

THE dry concentration of ores and minerals is a subject on which there is very little engineering literature available. With the advent of flotation, practically all milling research was turned to this development as it held the answer to large-scale economic operations. However, there are many concentration problems that can be more adequately and economically handled by the dry methods that have been developed, and, consequently, I believe it is very advisable for mining men to become fully 'conversant with just what can be done in the dry state. Therefore, today, I should like to give a brief summary of the processes that are available and, in a general way, how they can be a p-p lied. I shall confine my remarks mainly to two dry processes: THE Am FLOTA'l'ION TABLE AND THE ELECTROSTATIC SEPARATOR Dry Magnetic Separators should probably also be included, but I am

Jan 1, 1961

By R. McCulloch

Exploration activity in metallic minerals continued to decline in Montana in 1992 as companies either reduced budgets or left the state. The number of active projects dropped another 50% from 1991. Expenditures were 3% to 6% of 1989's 28.6 million. Many of the major mining companies closed offices in Montana. Although depressed commodity prices have hurried the exodus, continuous changes in environmental regulations, as well as the Mining Law of 1872, without regard to economic effect have had the most serious impact. Many companies indicate that Third World countries present a more stable business environment than the United States. Grassroots projects are nearly non-existent, while late-stage programs continue cautiously. Exploration activity continued at the following projects: Garnet, Pegasus Gold; Southern Cross, Asarco; Elkhorn, Goldfields; 7-Up-Pete, Phelps Dodge and Canyon Resources; Miller Mountain, Pegasus Gold; Sheep Creek, Cominco American Resources; Checkerboard, Kennecott and BHP Utah International; Emigrant Creek, Pegasus Gold; and Crevise, American Copper and Nickel. All mineral activity, whether on private, state or federal land, is regulated by the Montana Department of State Lands under three classes of permits - Small Miner Exclusion Statements (SMES), Exploration Licenses and Operating Permits. A total of 230 Exploration Licenses are active in the state with 26 of those being new. However, there are only 433 projects current in Montana, with many of those in the reclaiming stage. The majority of exploration projects consist of individuals expending time rather than money on properties. The number of mines functioning under operating plans declined by one when W.R. Grace closed its Zonolite operation.

Jan 1, 1993

By J. Svoboda, W. J. C. Venter, R. N. Guest

The use of gravity or magnetic separation followed by upgrading on a shaking table recovered about 25 per cent of the copper and lead minerals, at a combined metal grade of 20 per cent, from the final flotation tailing at the Tsumeb Mine. Magnetic separation recovered the copper minerals more efficiently than the lead minerals; it also recovered finer particles more efficiently than standard gravity separators. However, the grade of the concentrate produced by magnetic separation was not of sufficiently high grade, and further upgrading by gravity separation was necessary. It was decided to install an 'all-gravity' plant, which would have the added advantage of acting as a monitor or 'policeman' on a plant that receives a fairly variable feed.

Jan 1, 1988

By E. Fregeau-Wu

The major obstacle in recycling steelmaking slags to the blast furnace is their phosphorus content. Removal of the phosphorus, which is primarily associated with the silicate and phosphate phases, would allow for greater recycle of these slags for their iron, manganese, and lime contents. Calculations show that separation of the silicates from the oxide phases would remove nearly 90% of the phosphorus from the slag. The variable grain size of the as-received slag made liberation by fine grinding difficult. Therefore, slow-cooling experiments were undertaken to improve the grain size distribution. The grain size distributions were determined using in-situ image analysis. The samples were ground to their apparent liberation size and high gradient magnetic separation was used to separate the magnetic oxides from the nonmagnetic silicates and phosphates. Liberation analysis and modeling was performed on selected separation products for discussion of benefication characteristics.

Jan 1, 1995

By M. Kerr, Petra Lampinen, Jinming Zhang, Petra Brodin, C. Hull, Nils-Johan Bolin, R. Asmatulu

"According to Darcy’s equation, the rate of dewatering varies as inverse square of the surface area of the particles to be dewatered and is proportional to the pressure drop across a filter cake. Thus, the difficulty in dewatering fine mineral concentrates can be overcome by using high-pressure filters. When treating ultrafine particles, however, it is often necessary to further dewater the filter cake using a thermal dryer. In the present work, novel dewatering aids that can reduce the pressure requirement have been developed. The reagents are designed to reduce surface tension, increase particle hydrophobicity and increase capillary radius, all at the same time. They were tested for dewatering mineral and coal concentrates in both laboratory and pilot-scale tests. The results show that the use of the novel dewatering aids can substantially reduce cake moistures and, at the same time, increase throughput. In this communication, the case history of using the reagents commercially at the Garpenberg concentrator is given. The use of the novel dewatering aids allowed the operators to use a vacuum filter to obtain desired cake moistures and obviate the need to use a thermal dryer.INTRODUCTIONDewatering is an integral part of mineral and coal processing operations, as it is necessary to remove water from concentrates (or clean coal) to minimize shipping costs, meet contract specifications, and recycle the process water. In general, the cost of dewatering increases with decreasing particle size, which can be attributed to the difficulty in removing the water trapped in the pores (or capillaries) formed between the particles constituting a filter cake.The various parameters affecting the process of dewatering, including the effect of particle size, can be understood in view of Darcy’s equation (Darcy, 1856):"

Jan 1, 2005

By Manuel Gomez

Primary, low-temperature tars are generally known to alter with time. The effect is believed to occur, in some instances, without the external influence of air, light, or heat. Since one of the processes occurring during extended storage is oxidation, the action of mild oxidation induced by air and heat, in addition to the effect of storage, was studied. The principal effect in aging and oxidation is formation of pitch at the expense of the distillate fraction. The fraction boiling above 270° C. is mainly involved in pitch formation due to storage, whereas all fractions are affected to some degree when tar is heated at 125° C. in the presence of air. If the effects of air are considered, the aromatic and tar-acid fractions seem to be important in pitch formation.

Jan 1, 1956

By Hanley J

Coal winning in the Latrobe Valley exceeds 45 mtpa uses a small number of Bucket Wheel Excavators and high speed conveyor systems. While this equipment offers many advantages in the production cycle, incorrect use can lead to low productivity and risks to the safety of the operators and equipment. The mining equipment is of high capital value and easily damaged when used outside the correct operating range. In the extreme case, this can result in total machine collapse, having major implications for personnel safety as well as production and costs. With an active productivity improvement program in the Latrobe Valley in the last two years resulting in a 50 per cent reduction of personnel, many experienced operators have retired. Mine management are facing the urgent need to rapidly train new operators to maintain high productivity levels and reduce the risk of accidents to personnel and high capital cost plant. With the assistance of the Central Gippsland College of TAFE a Certificate Level course has been developed for open cut operators. The course covers all essential training needs for the Latrobe Valley mines. The training is developed from National Competency Standards, allowing transfer of skills, and is flexible to suit other enterprises. It has been written for self-paced learning with the total job understanding, and to have competency-based learning outcomes, training tasks and assessments built-in to modules. With the introduction of this new training program, management is convinced that the element of risk within the mine will be significantly reduced. For these reasons it is an ideal training model for other enterprises.

Jan 1, 1994

By T. Wang

A new generated NaNO3 ? NaNO2 ? KNO3 ternary thermal energy storage system which has low melting point as 124°C and high specific heat capacity was studied for the short and long?term thermal stability by using the Thermogravimetric Analyzer (TGA). Based on the actual working condition of molten salt in solar two, the simulated cyclic thermal stability measurement was also conducted. The ternary system demonstrates excellent short ? term thermal stability and thermal reliability after cyclic heating and cooling processes. However, in the long ? term as well as the simulated cyclic stability measurement, large amount of weight loss was observed. Moreover, the stepwise long ? term isothermal study reveals that below 440°C the weight change is minimal and this temperature is viewed as the upper limit temperature of thermal stability for this ternary system. To determine the lost phases at high temperature above the upper thermal stability limit, techniques such as XRD and SEM are used to compare the condition of the salt mixture before and after the long ? term thermal stability study. On the basis of the comparison, it can be concluded that the major weight loss was due to the decomposition of sodium nitrite.

Jan 1, 2012

By T. Iriyama, Y. Fujita, R. Omatsuzawa, T. Nishio, N. Okochi

Sm-Fe-N hard magnetic materials are good candidates for high performance bonded magnets. Rapid-quenching method should be highly favorable for the fabrication of isotropic magnet powder with high coercivity. We report here the magnetic properties of Sm-Fe-N powders produced by a few kinds of rapid-quenching processes. Sm2Fe17 powders were prepared by gas-atomization (GA), rotating-disk atomization (RDA), or melt-spinning (MS) method. The Sm2Fe17 GA powder had dendritic solidification structure, while the RDA powder had micro crystalline structure with size of 3-5 µm. Coercivities (HcI) of their nitrides were 56 (for GA) and 217 ka/m (for RDA). The H cI of nitrided Sm2Fe17 MS powder with wheel velocity (V5) of 5-20 m/s, which had also microcrystalline structure, was 53 to 509 Wm. These coercivity values are not sufficient for bonded magnet applications. From the study of additive effects, it was found that C, C-Si, C-Ga, or C-Al addition decreased the crystal size of quenched Sm2Fe17 powder. Consequently, high coercivity was obtained; the HcI of nitrided Sm2Fe1C1Si1 MS powder with V5 of 10 m/s was 898 kA/m, and 708 kA/m for the nitrided Sm2Fe15C1Si1 RDA powder. The compression-molded bonded magnet made from Sm2Fe18Si0.1Nx MS powder showed a high (BH)max of 82.7 kJ/m3.

Jan 1, 2000

By Farlow C. Davis

Magma Copper Company, San Manuel Division, is a wholly owned subsidiary of the Newmont Mining Corporation and located at San Manuel, Arizona. The ore body is a dessiminated porphry copper deposit in quartz monzonite. The primary copper mineral is chalcopyrite with minor amounts of chalcocite, bornite, and covellite. The molybdenite occurs mainly on fracture faces or else associated with quartz veinlets. Pyrite is the major sulphide mineral in gangue. The feed to the concentrator from the underground mine is approximately 0.720 percent copper and 0.029 percent molybdenite. The concentrator has a design capacity of 62,000 tons of ore daily. After a standard ore treatment circuit consisting of grinding, rougher flotation, regrinding, and cleaning, a bulk concentrate is produced which contains 27 to 29 percent copper and 1.0 to 1.2 percent molybdenite. This concentrate is pumped to the molybdenite plant feed thickener.

Jan 1, 1977

By Charles E. Wait

HAVING had some experience in the use of gasoline-gas in a laboratory, I have been induced by frequent inquiry to present a few hints concerning it, which I hope may be of some value to those who are out of the reach of coal-gas, and who contemplate using this gas. While I will not delay to describe the method by which this gas is made, it may not be out of place to note that, unless frequent charges are put into the carburetter, the gas is variable in richness. This may be partially avoided, it is said, by repeatedly pumping from one cell in the carburetter to another. If a machine is in constant use for many months without re charging, the gas becomes rather poor for illuminating purposes ; but in this condition it may be burned in a Bunsen burner with advan tage. Could the gas be used always in this condition, no difficulty would be experienced, and it would be quite satisfactory for general laboratory purposes. Considerable annoyance, however, is experi enced whenever the machine is filled with gasoline. The Bunsen and other burners then produce a flame too smoky to be used with com fort, and, in fact, with most of them, a full supply of air; turned on at the base of the burner, will not suffice to completely burn off all the carbon. Finding it necessary to adopt some method by which I could get a flame free from unconsumed carbon, I adopted a plan which has in every way proved satisfactory. In addition to the pipe conveying the gas to the table, and ar ranged with a stopcock, there is brought to the table, within a fern inches of the gas-pipe, another pipe, conveying air direct from the pump. This pipe is also arranged with a stop-cock. The air- and gas-pipes are now connected with one pipe by a T joint; and to this last pipe is attached, by rubber tubing, the Bunsen or other burner. I have successfully used this simple system in connection with both small and large Bunsen burners, a burner with three tubes, arid a Fletcher burner; and recently I have added a new still to our equipment, and have been tising with it a Fletcher special high power burner No. 6, to which this system has been applied. With

Jan 1, 1886

By R. W. Bartlett

The rates of oxidation of tungsten have been determined at temperatures between 1320" and 3170°C and oxygen pressures to 1 amn using a surface -recession measurement technique. Above approximately 2000°C and 10-6 atm the rate is independent of temperature and can be calculated from gas collision theory assuming a constant reaction probability, e, of 0.06. Oxygen molecules react at surface sites where oxygen atoms have previously chemisorbed. This provides a direct pressure dependence at low pressures but at high pressures tungsten oxide molecule s form an adjacent gas boundary layer which lowers the PO2 at the tungsten surface. A correction for this effect using free-convection theory fits the rate data over the entire oxygen-pressure range from 10-8 to 1 atrn as well as data using O2-A mixtures. Below 10-6 atrn and above 2000°C, e decreases with increasing temperature because of desorption of oxygen atoms. Below 2000°C the rate decreases with decreasing temperature at all oxygen pressures following an apparent activation energy of 42 kcal per mole and depending on (Po2)n with n varying between 0.55 and 0.80. MOST of the previous tungsten oxidation studies have employed gravimetric methods and have been limited to temperatures below 1000°C where the weight loss associated with evaporation of tungsten oxides is negligible compared with the weight gain from oxidation.' At higher temperatures, oxygen-consumption rates have been determined from pressure measurements, usually at constant flow rates, by Langmuir,2 Eisinger,3 Becker, Becker, and Brandes,4 and Anderson.5 The sensitivity of this method decreases with increasing pressure and, with the exception of Langmuir's work, these investigations were confined to pressures below 10-6 atm. Above approximately 1300°C, depending on the oxygen pressure, the rate of oxide evaporation is greater than the oxide-formation rate and the recession of the tungsten surface can be measured optically without interference from an oxide layer. This was first done by Perkins and crooks6 who heated tungsten rods in air pressures from 1 to 40 torr at temperatures between 1300" and 3000°C. The present investigation of the oxidation kinetics of tungsten at high temperatures emphasizes oxygen pressures from 10-6 to 1 atm. This is the range of interest for earth atmosphere re-entry applications of tungsten for which little data were previously available. APPARATUS The apparatus is a modification of the type used by Perkins and crooks.' Ground tungsten seal rods, 6 in. long by 0.125 in. diam, were mounted vertically between two water-cooled electrodes, one fixed and the other having free vertical travel. The movable counter-weighted electrode is prevented from undergoing horizontal displacement by three sets of runners mounted at 120-deg intervals. Electrical contact is made by means of a water-cooled mercury pool. A 24-in. vacuum bell jar having a volume of approximately 267 liters was used as the reaction chamber with the sample holder mounted in the middle of the chamber. Power was supplied from an 800-amp dc variable power supply. Temperature readings were made by means of a Latronics automatic two-color recording pyrometer. With this instrument, corrections for emissivity are not necessary provided the spectral emissivi-ties at two closely spaced wavelengths are equal. Supporting measurements were made with a micro-optical pyrometer corrected for emissivity of bare tungsten and window absorptivity. The micro-optical pyrometer was calibrated against a National Bureau of Standards calibrated tungsten lamp and both pyrometers were periodically checked against the melting points of tungsten and molybdenum using the oxidation apparatus. Above 10-6 atm, pressures were measured with an Alphatron gage calibrated against a McCleod gage. At 10-6 atm, a hot-filament ionization gage was employed. A magnified image of the self-illuminated tungsten rod was formed using a 360-mm objective lens mounted outside the bell jar. When the experiment exceeded 1 hr, the image was focused on a ground-glass plate about 10 ft from the tungsten rod at about X8 and the recession of the thickness of this image was monitored with a Gaertner cathe-tometer. When faster rates were encountered, a 35-mm time-lapse cinecamera with a telephoto lens and bellows extension was substituted for the ground-glass plate and cathetometer. Diameter recession rates were determined from the photograph image projected on the screen of an analytical film reader. EXPERIMENTAL PROCEDURE After installing the rod in the apparatus and cleaning it with acetone, the system was evacuated to 5 1 x 10-5 torr. Before oxygen was introduced,

Jan 1, 1964