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By W. Kennedy

THROUGHOUT northeastern Texas we have an extensive series of iron-ore deposits, occupying portions of nineteen counties and having approximately an actual ore-covered area of 1000 square miles. The existence of these ores has long been known. Kennedy* says: "Iron-ore is distributed in profusion throughout Texas, being found almost everywhere except in the level region of the coast." Before this, C. Newell† had mentioned the existence of these ores thus: " Iron is found in abundance in the eastern, northern and middle portions of Texas;" and in 1849 Edward Smith‡ said: " We met with immense hills of iron which were said to be, and appeared to be, of excellent quality, in Cass and Titus counties.. . .. It is well known that Texas abounds with iron of first-rate quality." In 1859, Dr. Shumard? reports that detailed surveys had been made in several counties, and says: " Our labors have demonstrated the important fact that we have a vast iron-region in the eastern part of the state, embracing considerable areas in Cass, Harrison, Rusk, Panola, Smith, San Augustine and Shelby counties. The ore-deposits belong to the Tertiary era and consist chiefly of hematites and limonites, of which there are several varieties. We have also found in this district extensive beds of carbonate of iron." Dr. Shumard gives no very exaggerated description of Cass county§ as it then existed when he says : " Cass county alone is capable of supplying a number of furnaces with an abundance of excellent iron-ore for many years." Of Rusk county he says: "Our de-

Jan 1, 1895

By M. C. Malamphy

A general description of the activities and results of bauxite investi-gation by the Bureau of Mines at its bauxite project in Saline and Pulaski Counties, Ark., has been given in Report of Investigations 4251. The factual data obtained during the project are reported in a series of Reports of Investigation beginning with No. 4251, of which this is the sixteenth. The data are accompanied by maps showing the location of holes drilled and the deposits developed. Each area upon which work was done is reported by half townships. Each bauxite deposit investigated was assigned a name and a number for ready reference and identification, and its location and name are shown on the appropriate half-township map. The description of each half township is followed by a description of each deposit developed in that half township. The 14 half-township maps and 77 detail maps of the bauxite deposits in this series of reports give the locations and numbers of all the holes drilled on the project, except the deep holes drilled to test the geophysical anomalies. The locations of the latter holes show on figure 13 in volume 1. The descriptions of the deposits developed are followed by complete detailed logs of the drill holes. Many of the widely spaced exploratory holes did not discover bauxite of minable thickness and grade. The pertinent information obtained from the widely spaced exploratory drill holes has been summarized and tabulated, and a table for each half township is given with the discussion covering that area.

Jan 1, 1948

By P. Groleau, O. Tavchandjian, D. Marcotte

In open pit mining, each small mining unit (SMU) is selected for mill processing based on an estimate of its true grade. The estimation is usually done using analyses of cuttings from blast holes within that SMU. One common estimator used for selection is the arithmetic average of within-SMU samples. This estimator is conditionally biased. The effect of a conditional bias is to overestimate the profit really recovered from the deposit. For this reason, a conditionally unbiased estimator is preferable. Increasing the number of samples generally improves the accuracy of the estimation (and diminishes the conditional bias, if present, of most estimators). However, there exists a practical limit where an additional sample does not improve the estimation sufficiently to cover the expense of making the analysis. This limit is specific to each deposit and depends on geostatistical parameters, economic factors, and mining parameters. This study examines how the number of samples affects the recovered reserves and the resulting profit. Also, the effects on optimal profit of varying the different factors are examined. Globally, the number of samples required increases with the selectivity of the operation. For a given ore deposit, the greater the number of SMUs selected for processing, the fewer samples are required and vice-versa.

Jan 1, 1998

By D. H. Schwarz

The performance of proton exchange membrane (PEM) fuel cells depends on catalyst layer composition and structure. Experimental observations reveal that state-of-the-art catalyst layers consist of microporous agglomerates of carbon-supported catalyst sites bound together by polymer electrolyte. In between the agglomerates are macropores which provide pathways for the transport of gaseous reactants. In this work, three-dimensional, multicomponent and multiphase transport computations are performed using a computational fluid dynamics (CFD) code (FLUENTTM) with a new PEM fuel cell module, which has been further improved by taking into account the detailed composition and structure of the catalyst layers using a multiple thin-film agglomerate model. In this model, it is assumed that thin films of polymer electrolyte and liquid water surround the catalyst sites. The results of CFD computations show that transport limitations associated with the thin film of polymer electrolyte and the presence of liquid water in the catalyst and gas diffusion layers have substantial negative effects on PEM fuel cell performance. Also, the results of computations with variable distributions for a given precious metal loading show that, in general, uniform catalyst loading provides the best fuel cell performance.

Jan 1, 2005

By Jacob T. Wainwright

This stove has been designed to meet the requirements of a fur nave that must be operated with either a reducing or a neutral flame ; and more particularly to make feasible the operating of re duction-furnaces with the cheaper grades of gas-fuel, which has heretofore been impracticable. In general, cheap gas-fuels result from incomplete combustion, and, being suited for further combustion, answer very well for heat ing purposes; and high temperatures may be attained with them, when they are used in conjunction with regenerative stoves. But, in operating a reduction-furnace, a high temperature is not the only requirement. It is also necessary that the fuel shall not be cansumed beyond a certain degree, in order to insure a re ducing flame for the reduction-process ; consequently, it becomes necessary to dispense with combustion as much as possible previous to the use of the fuel-gases in the reduction-furnace. Moreover, these gases must be preheated to a high degree, in order that a sufficient temperature may be maintained in the furnace; and, since this preheating cannot be accomplished by a previous, partial combustion of the gases themselves, it follows that the stoves for pre heating these gases should be heated to as high degree as it is pos sible to attain with an inferior fuel. This result can best be accom plished by complete combustion of the heating gases, and is further facilitated by preheating all of the air and gas used for this com plete combustion. It is more particularly desirable to preheat the air, since the combustible gases used may generally be obtained directly from the furnace, or gas-producer. The reducing gases, escaping from a gas-fuel reduction-furnace, are capable of considerable further dilution with air to secure com plete combustion, and consequently may be utilized for heating the stove to a high temperature, provided the air be sufficiently preheated.

Jan 1, 1889

By G. T. Fisher

The U.S. Bureau of Mines conducted research on calcium ferrite slags to determine the distribution of Pb and minor elements (Ag, Cd, Zn, and Co) between Pb bullion and slag. These fundamental data are required to evaluate nonferrous smelting with calcium ferrite slags as a substitute for conventional iron silicate slags. Determinations were made in 25-cm3-capacity MgO crucibles at 1,250° C for 2 h in a CO-CO2 atmosphere with an O2 potential that would retard the formation of magnetite. Calcium ferrite slag composition variables studied were from 16 to 32 wt pct CaO in 2-pct increments and Si02 ranging from 1 to 5 wt pct in 1-pct increments with equal additions of MgO. For all compositions studied, the retention of Pb in the calcium ferrite slag was less than 1 wt pct, which compares favorably to about 6 wt pct Pb in conventional iron silicate slags used in smelting of Pb concentrates. Minor element distribution was insensitive to changes in slag composition in the range studied. However, slag composition had significant impact on the melting and fluidity of the slag.

Jan 1, 2010

By Ryokichi Shimpo

Much amount of sludge is produced from the wastewater treatments in surface finishing industry. There are many problems for the abandonment of the sludge because of the high water content and heavy metals included. It is often observed that hexavalent chromium, which is toxic, is generated in the course of heat treatments of the sludge for reducing its amount by dispersing water. In this study, conditions of the sludge calcinations for reducing water contents have been examined. Maximum hexavalent chromium generation was observed in the temperature range around 773 K, at which most of the crystallization water decomposed from the sludge. When the heating temperature became higher than 1173 K, the amount of the hexavalent chromium extracted from the calcined sludge became as low as 10 mg/kg sludge. When the sludge was heated with iron oxides, FeO or Fe203, iron ferrite has been synthesized which may be used as a magnetic material. The amounts of hexavalent chromium generated in the heat treatment with iron oxides were not much different from those without iron oxides.

Jan 1, 2003

By R. E. Morgan, J. F. Barkley

"INTRODUCTION In city smoke-abatement work it is customary to require approval of the sizing of fuel-burning equipment. This is to insure that, with reasonable attention, the equipment can be operated essentially smokelessly, having such features as ample furnace volume and air supply. To answer the many inquiries that come to the Bureau of Mines as to what dimensions are acceptable and standard, the requirements of a number of smoke-abatement ordinances have been tabulated.CONCLUSIONSStudy of these tabulations shows that even after many years of fuel-burning practice and experience there is no common standard. Variations in the requirements for the same type of equipment under the same conditions of operation are great. This causes considerable difference in installation costs. Over-all costs being affected, the competitive positions of different types of fuel-burning equipment are also affected. It is apparent that there is a lack of information as to proper dimensions that is commonly accepted as authoritative.DISCUSSIONTable 1 gives the minimum furnace-setting heights required by various cities for mechanical underfeed stokers installed under H.R.T. boilers. Figure 1 shows a typical installation. The dimension given for each case is the distance from the dead plate of the stoker to bottom of boiler shell, unless otherwise noted. As shown in the table, different cities use different size-classification systems; it is not possible to make complete, exact, direct comparisons. Table la gives a comparison for various cities, showing the approximate percentage difference in H.R.T. furnace-setting heights required from those recommended by the Stoker Manufacturers' Association arbitrarily chosen to serve as a basis of comparison. The percentages are approxi¬mate because they were calculated as if all heights as given were measured from the dead plate of the stoker. In two cases a lower height is permitted, in a number of cases the height may be the same, and in a number of cases the height must be greater. The percentage increase goes to 100 in one case. Lack of uniformity is quite evident."

Mar 1, 1950

By M. L. Requa

IN the fall of 1910, the Nevada Petroleum Co., operating in the Coalinga field in California, determined to drill a number of wells with rotary tools, in order to prove conclusively the relative value of the rotary as compared with the standard rig. At that time, the rotary was but little known in California and its proposed introduction met with considerable criticism on the part of the operators and quiet opposition among the standard drilling crews. There were few men available who were competent to handle a rotary outfit, and these few had had but little experience in California fields. Machines as then built were much inferior to those now used, being lighter in construction and of a poorer quality of material. The shells encountered at depths between 1,700 and 2,000 ft. seemed to be too hard to permit of successful drilling, and at these depths the rig was changed over to standard tools and by them completed. Because of the heavy depreciation, the time lost in converting from rotary to standard, and the comparatively small profit, it was concluded that unless in the future there was some material improvement in rotary rigs, nothing would be gained by drilling with rotary tools upon this property. Little or no drilling has been done upon the property since that time, but in the meantime a large number of men have been trained in the use of the rotary; in fact, many standard-tool men have abandoned standard drilling and, starting in as roustabouts, have become thoroughly competent rotary operators. The improvements in the machinery have been such as to remove many of the objections and the rotary drilling of to-day is in every way superior to that of 1910. In contrast, a well drilled recently by the Kern Trading & Oil Co., in the east side Coalinga field may be cited. The first 2,500 ft. was drilled in 24 days and it is confidently expected that the water string will be set with a rotary at 3,400 ft.-a record obviously far beyond that made in 1910 by the Nevada Petroleum Co.

Jan 2, 1915

By G. A. Sutton

This paper, describes a case example from San Manuel mine, where in situ copper leaching of the supergene zone was done over a 14 year period (1986–1999) in conjunction with underground and open-pit mining operations of a copper porphyry deposit. In 1993–1994, testing of well fields was done to determine the amount of copper that would be recovered from in situ leaching of the supergene zone, which was composed primarily of chrysocolla and copper-clay minerals. It was determined that the leach fluid would reach 70% (‘sweep efficiency’) of the copper minerals and that 75% of them would be leached for a recovery of 52% of the acid soluble copper. However, the final results from Zone 6, which had the longest leach history, showed that approximately 42% of the acid soluble copper was recovered, indicating that the `sweep efficiency’ was closer to 55%. Comparison of drill-core assay data from two holes, one drilled before (ARM158) and the other after (SMO9607) leaching, also indicate that approximately 42% of the acid soluble copper was recovered. The discrepancy between the predicted (52%) and actual (42%) recovery is probably related to a combination of factors, including difficulty in maintaining fluid saturation, channeling effects and gypsum precipitation. INTRODUCTION In situ copper leaching at the San Manuel mine, Arizona, was initiated in 1986. It was recognized that some of the copper oxide resource was not going to be economically recoverable by conventional open pit and heap leach methods using solvent extraction and electrowinning (SXEW). A high stripping ratio would be required to access part of the resource. At the time, in situ leaching was becoming popular in the uranium industry but the approach was virtually unknown in the copper industry, especially not at a large scale.

Jan 1, 2019

By Ian James McDonald

Each year, approximately 400 million tons of blast furnace slag is produced worldwide with a tapping temperature of around 1,500°C. It is normally used as a substitute for cement clinker or as an aggregate material in road construction. Currently, the slag is granulated in wet-granulation plants using large volumes of water and to date, it has not been possible to utilise the remnant heat energy of the molten slag, which amounts to approximately 1.8 GJ of energy per ton. However, in an R&D project currently underway led by Primetals Technolgies and industrial partners, a new dry atomising technology is being investigated to use air to cool molten slag and recover the lost heat energy. The resultant pelletised slag fulfils the same criteria as wet-granulated slag for use in the cement industry. Phase 1 of the project has now been completed where a technical plant was set up at the University of Leoben in 2012. A series of dry-slag granulation campaigns were carried out using remelted blast furnace slag. The elevated offgas temperatures and the quality of the slag product as verified by the FEhS Building Material Institute have shown the process suitability as an industrial application and the decision was taken to escalate the project from large laboratory scale to a full size pilot plant. The phase 2 development of this plant is now underway and is scheduled for installation at the site of an industrial partner in early 2016 where full slag flow will be fed directly to the plant from Blast Furnace ‘slag runners. This paper will show the development path taken to date and the planned route to our goal in 2016 of being the first to industrialise the ‘game changing’ process of dry Slag Granulation with Heat Recovery from the slag.

Aug 16, 2017

By W. C. Meyer

The concept of using powdered coal-in-oil-mixtures (COM) as a composite fuel to conserve oil resources and relieve import supply problems is being vigorously promoted by the Department of Energy via the sponsoring of various demonstration units (1). The idea of COM as a useful alternative fuel has been explored on and off for over 100 years (2). The approach must be capable of grinding, mixing, storing, and combusting COM with presumably only minor equipment changes on oil-fired boilers. DOE estimates that 56 000 m3 of oil (350 000 barrels) could be spared daily if a 50-50 COM was used to generate electricity. Of course, the ability of COM suppliers to prepare and deliver a stable suspension of COM at reasonable economic costs must be achieved if the concept is to be successful. This report reviews the various methods used to attain stable coal-oil mixtures. Various physical-chemical principles associated with each method of stabilization are also discussed.

Jan 1, 1981

By D Johansson, C Yi

Blasting techniques are widely used to fragment rock masses into smaller pieces. Numerical modelling is an efficient method employed by many researchers in the blasting field. It is difficult for a conventional continuum-based approach such as the finite element method (FEM) to model the rock fragmentation by blasting and the expansion work on the rock by explosive and its detonation products. In this paper, the particle blast method (PBM) was employed to model the behaviour of the detonation and a bonded particle model (BPM) was used to model the brittle material to be blasted. A mortar cylinder with a centrally placed hole for the explosive was modelled and the results were compared to the experimental data. The blast process from crack initiation to fragment formation was analysed. The influence of coupling ratio on fragmentation was also investigated.CITATION:Yi, C and Johansson, D, 2015. Discrete element modelling of blast fragmentation of a mortar cylinder, in Proceedings 11th International Symposium on Rock Fragmentation by Blasting, pp 793–798 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 24, 2015

By F. D. Osgood, A. C. Fieldner, W. A. Selvig

In connection with investigations being conducted by the Bureau of Mines regarding pulmonary disease which is prevalent among metal miners in certain districts , a large number of rock- dust samples have been collected , and have been examined in the Bureau's laboratories . The earlier work has been described in Bureau of Mines Technical Paper 105 , " Pulmonary disease among miners in the Joplin district , Missouri , and its relation to rock dust in mines " by A. J. Lanza and Edwin Higgins , 1915 , and in Bureau of Mines Bulletin 132 , " Siliceous dust in relation to pulmonary disease among miners in the Joplin district , Missouri , " by Edwin Higgins , A.J. Lanza , F.B.Laney , and Geo . S. Rice , 1919. The methods of collecting and examining dust samples have been greatly improved during the investigation , much valuable information having been made available by similar investigations conducted by the Miners ! Phthisis Prevention Committee , Union of South Africa . The latter are described in the " General report of the Miners ' Phthisis Prevention Committee , Johannesburg, Union of South Africa , 1916 " , and " Final report of the Miners ' Phthisis Prevention Committee , Johannesburg, Union of South Africa , 1919. " The investigations by the Bureau of Mines and the South African committee show that by prolonged inhalation of air containing minute particles of sharp and insoluble rock dust , the lungs retain sufficient dust particles to cause decided injury to the lung tissue , the injured lung becoming very susceptible to the development of tuberculosis .

May 1, 1920

By G. H. K. Schenck, H. W. Robinson

Private weather forecasting services available on a contract basis can significantly reduce the uncertainty associated with weather-induced changes in mineral production and demand. Such services offer the mineral producer the opportunity for tactical and strategic response to weather conditions. There are at least 50 private meteorological consulting services as well as numerous individual meteorology consultants. Fees for services range from $150-$600 per day for one man to over $20,000 per month where a complex forecasting or modification program is undertaken.

Jan 1, 1971

By A. G. McGregor

In the Southwest a number of large steel chimneys discharge the gases from the copper smelting furnaces. Some of these chimneys show no deterioration after twenty years, others show serious deterioration after tour years service. A steel stack 20 ft. 7 1/2 in. (6.3 m.) in diameter by 279 ft. (85 m.) high, Fig. 1, used for roaster gases only at the Calumet & Arizona Mining CO.'s plant at Douglas, Ariz., was lined to the top of the bell, 40 ft. (12 m.) from the base, with common brick placed on end, lime and cement mortar being used. Above the bell, the stack was lined with 4 by 4 by 8 in. (10 by 10 by 20 cm.) building tile, placed on end, making a lining 4 in. thick. The tiles were supported on angle-iron rings, riveted to the inside of the chimney shell about 15 ft. apart. The chimney was put into use in the latter part of 1913. Four years later, considerable of the tile lining had fallen out of place and mixed with the flue dust at the bottom of the chimney. About a year later, several flattened and distorted sheets were noticed about 70 ft. below the top. An investigation showed that the thickness of these distorted sheets, which was originally 1/4in., had been reduced one half. One year later, or six years after the chimney went into use, several holes, a square foot and smaller in area, appeared near the flat spots mentioned. Soon afterwards, when the chimney was shut down for repairs, it was found that most of the tile lining for 125 ft, from the top had fallen out and that the steel shell was badly corroded; but where the lining remained in place the shell was undamaged. Many of the tile which had fallen out of place had become soft or flaky and crumbling, others were as hard as when placed. An analysis of the hard and soft tile is shown on p. 2. These tile were of a fireclay mixture, and were used instead of common brick on account of lower cost. The steel forming the upper 125 ft. (38 m.) was replaced, and a 4 1/4-in. (11-cm.) lining of firebrick was laid from the bell to the top of the chimney. The brick were placed on end, wedges and keys being used. After each 8-in. course was set, keyed, and wedged tight, thin mortar —water, sodium silicate, and finely ground silica—was worked into the joints. The brick are supported on angle-iron rings, every 15 ft., in the same way as the tile.

Jan 1, 1921

Pillars are used in various forms in underground mining. For several decades pillar design has been an issue in both soft and hard rock underground mining. The major use of pillars is in room-and-pillar type mining. Pillars are critical structures in the production levels of block cave mines. With most mines leaning towards the use of block caving for ore extraction today, pillar design has become even more important. Empirical pillar design is popular. The use of numerical modeling is becoming more common in the design of underground mine pillars. A critical review of both the empirical and numerical modeling approaches to the design of underground mine pillars reveal some fundamental flaws. In general, pillar design approaches often assume the pillars are vertical, and are subjected to pure compression. Pillar failure criteria are then based on this assumption. Most papers on numerical modeling for pillar stability assessment use vertical pillars. This paper shows that where pillars are not subjected to pure compression the empirical pillar design approaches are inadequate, and that when using numerical models for the assessment of pillar performance, more care is needed in the determination of the average pillar stresses. The practice of relating field scale pillar strength to laboratory scale intact rock strength determined from conventional laboratory compression tests in empirical design equations is misleading when pillars are inclined or subjected to compressive and shear stresses. New data is added to the hard rock empirical pillar stability chart, and it is shown that it is misleading to conclude that failure of pillars with width to height ratios greater than 2.3 should not be anticipated. New empirical pillar design charts are presented for pillars at various degrees of inclination for different ko-ratios.

Aug 1, 2013

By Jerry A. Steding

How many times have you heard "You can't drive piles because of the vibration."? You hear these words come out of the mouths of owners, engineers, and contractors. When a piling contractor hears this, he can almost always do a psychological profile of the individual uttering these words of wisdom. If the owner makes the statement, we know the odds are about 95% that the owner owns the building next door or he wouldn't care what happens to it; or the guy next door has threatened to sue the owner if he even so much as walks across the site of the proposed new building. If the owner owns the building next door to the new construction you can also be just as certain the old structure has undergone settlement. If the statement comes from the mouth of the contractor, it is usually just a case of him being inexperienced.

Jan 1, 1988

By H. P. Allen

THE chloridizing mill of the Standard Reduction Co. is located about 75 miles south of Salt Lake City on the Tintic branch of the Denver & Rio Grande Western R. R. and 12 miles from the Tintic Standard mine. The daily capacity is 200 tons of a siliceous, low-grade, silver-lead ore from this property. It has operated continuously since it was started in January, 1921. The process consists essentially of a chloridizing roast followed by a percolating leach with a nearly saturated solution of common salt, acidified with sulfuric acid, the precipitation of silver on sponge copper and of copper and lead on tin-plate cuttings. The precipitates are shipped to a smelter. Some of the general ideas involved are said to have been used by Augustin in England, in 1840. A number of textbooks treat of the subject, especially the chloridizing roast followed by a leach with sodium hyposulfite or amalgamation. The process was revived in this district by Theo. P. Holt, N. C. Christensen, the Bureau of Mines, and others.

Jan 8, 1925

By Richard L. Stein

INTRODUCTION Increasing a miner's chance of surviving a disaster requires advance planning for emergencies, adequate training of personnel, and the provision of proper survival equipment. It is extremely important that fires and toxic gases be detected early, and that warning systems such as stench, audible and visual alarms, and communications systems be provided to allow timely escape. However, situations do occur where timely warning is not provided and immediate escape is not possible, regardless of the precautions that are taken. Under such circumstances, a combination of new per¬sonal protective devices and old techniques can provide protection that is adequate to save a miner's life. SELF-RESCUE EQUIPMENT Self-rescuers are devices which provide a short-term supply of respirable air under the potentially lethal con¬ditions following a mine disaster. Functioning either as filtration devices for the elimination of carbon monoxide (CO) or as oxygen supplies, self-rescuers allow an endangered miner a limited amount of time in which to effect an escape or reach a place of temporary safety. Carbon-Monoxide Self-Rescuers For almost 50 years, filtration self-rescuers have protected miners' lives following explosions or fires. Designed to eliminate carbon monoxide from inhaled air, the first of these devices was approved in the 1920s and has evolved into the two units that are approved today-the Mine Safety Appliance (MSA) W-65 and the Draeger 810. Both of these units rely on the cata¬lytic conversion of very toxic carbon monoxide to rela¬tively safe levels of carbon dioxide (CO,). As shown in Figs. 1 and 2, both of these filtration self-rescuers operate in the same manner. On inhalation, the mine air passes through both coarse and fine filters that remove the dust, preventing the dust from coating the chemical beds or entering the miner's mouth. The air then passes over a drying agent that removes water vapor; this is required to prevent the water vapor from poisoning the catalyst. Subsequently, the dried air passes over the Hopcalite catalyst, where carbon monoxide is converted to carbon dioxide. Then, the air passes through a heat exchanger that reduces the temperature of the inhaled air. Finally, the air is inhaled by the user. Air exhaled by the user passes through the heat ex¬changer and exits the device through an expiratory valve. Both units of this type are subject to certain limita¬tions: 1) The units do not protect the user against oxygen¬deficient air; when air is inhaled through this type of self-rescuer, the device only removes the carbon mon¬oxide. If the mine air contains less than 15% oxygen, anoxia is inevitable. Symptoms of anoxia include dizzi¬ness, shortness of breath, quickened pulse, and deeper and more rapid respiration while the victim is at rest. During heavy exertion such as would be expected dur¬ ing escape efforts, a 15% oxygen level can cause loss of consciousness. 2) The units do not protect against excessive levels of carbon dioxide. Available data on carbon monoxide and carbon dioxide concentrations following explosions or fires are limited, but there are indications that the carbon monoxide concentration can rise to 2% and the carbon dioxide concentration can reach 5 or 6% of the mine air (by volume). In some situations, the in¬haled air could contain up to 6 or 7% carbon dioxide. At a carbon dioxide concentration higher than 2%, breathing patterns can be affected adversely. At a con¬centration of 6 or 7%, the effects include severe respira¬tory distress, with unconsciousness resulting from exer¬tion such as that needed for escape activities. 3) The units do not protect against high inhalation temperatures. The catalytic oxidation of carbon mon¬oxide to carbon dioxide is an exothermic reaction that evolves a large amount of heat. As the carbon monoxide concentration increases, the temperature of the inhaled air also increases. Available data show that above 1.5% carbon monoxide, the air inhaled through some self¬rescuers can be as high as 90°C (194°F). At those

Jan 1, 1982