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By Dean Brox

Tunnel boring machines (TBMs) have been used for the construction of various tunnels for mining projects for the purpose of access, conveyance of ore and waste, drainage, exploration, water supply and water diversion. Several mining projects have seen the successful and economic beneficial use of TBMs, and there is an increasing awareness of the benefits of TBMs for mining projects. Key technical considerations for the use of TBMs for the construction of tunnels for mining projects include geological issues (rock type, rock alteration, rock strength, rock abrasivity, durability, ground water inflows), depth of cover and the potential for overstressing/rockbursts, site access and terrain, portal locations, TBM constraints, minimum tunnel size, tunnel support requirements, contractor and labor experience, and project schedule demands. TBMs offer key project schedule benefits with high rates of progress, which is a unique economic advantage recognized for mining projects where timing is of the essence in order to advance mining development and achieve early startup for operations. A thorough evaluation of the anticipated ground conditions should be performed as part of the decision to adopt a TBM for a mining project. Minimum specifications for TBMs are necessary in order to prevent the incorrect type of TBM from being used. Appropriate geotechnical investigations are necessary in order to provide key information on the anticipated subsurface conditions for careful planning and design of the proposed tunnel alignment. The overall site and geological conditions associated with some projects may suggest that it is not appropriate to use TBMs for some mining projects due to a perceived higher risk.

Jan 1, 2013

By Goran Mathisson

PLASMARED is a process in which one well established and one new advanced technology have been merged into a practical tool for production of high quality direct reduced iron at lowest energy consumption and costs. Before describing the process, we have to reiterate the simple fundamentals of making metallized iron from iron ores: The main components necessary are: - Iron oxide - Carbon and/or hydrogen - Heat

Jan 1, 1982

By Miguel Angel Ibarra

This paper describes the results of a digital computer simulation study in which the impact of the proposed truck dispatching on the overall mine productivity was investigated for an operating open pit mine. In the study, the existing operation was first simulated for a one-month period, using both the actual time study data and the computed cycle times. By adjusting certain input parameters to the simulation program, the actual one month production was first duplicated through simulation. Afterwards, the operation was again simulated in a dispatch mode. Comparison of simulation results under the dispatch mode, with the initial base case simulation results, showed a definite improvement in overall productivity with dispatching; i.e., approximately 10% gain for the operation. As expected, the results also showed that the extent of possible improvement did vary with the particular pit configuration being investigated.

Jan 1, 1980

By Ronald L. Wiegel

The United States produces approximately 40 mil- lion tons of phosphate rock per year. Most of this rock is converted to diammonium phosphate and used as a fertilizer component for the domestic and international agricultural market. Much of this phosphate rock production is concentrated in the southeastern United States, with about half coming from washing and screening operations and half from flotation separation from sand and clay. This paper summarizes current operating practice in the industry and touches on some of the emerging technologies required to solve the problems associated with a diminishing high- grade reserve and the need to implement dolomite/francolite separations on a commercial scale.

Jan 1, 1999

By K. L. Lamb

At the Sierrita mine, 40 km (25 miles) south of Tucson, Arizona, Duval Corporation is rapidly gaining operating experience with its new Portable Crushing and Conveying System, (PCCS). As of November 23, 1983, over 4.5 million tons (5.0 million st) of ore have passed through the system. Mining cost per ton for the months of July, August and September was 83% below the average for the first five months of 1983. Although all the cost reductions experienced cannot be attributed to the Portable Crusher, it was a major contributing factor. The Sierrita ore deposit has several natural topographic and geologic conditions which are not favorable to mining. It is not a sidehill deposit and therefore no low-level exit points exist to facilitate truck haulage. The ore is very hard. Quartz diorite, hardest of the three major rock types, averages 172-207 MPa (25-30 000 psi) unconfined compressive strength and may run as high as 345 MPa (50 000 psi). Blasting in this rock produces oversize material which runs approximately 20% 51 centimeters (20 inches). The ore is low grade, averaging slightly better than .30% contained copper and .035% contained molybdenum, with small amounts of silver. These conditions, plus high design production tonnages, make the Portable Crushing and Conveying System (PCCS) practical for Sierrita. Without crushing and conveying, truck haulage would be prohibitively expensive. A typical, October, 1983, ore haul from the bottom of the pit would involve a vertical lift of 260 meters (850 ft) and a horizontal distance of 3 860 meters (12 200 ft) were not in-pit crushing and conveying available. The corresponding cycle time would be nearly 32 minutes.

Jan 1, 1984

By V. Rastogi

Acid rock drainage is the result of exposure of sulfidic materials, such as pyrites and sulfidic ores, to air and water. The oxidation of these sulfides is catalyzed by more than a million times by the bacteria Thiobacillus jerrooxidans and close families. Inhibition of these bacteria can mitigate 98% or more of the acid loading generated. This can be achieved by the use of certain surfactants that are selectively bactericidal to the lithotropic acidophilic bacteria. Bactericides are used in two ways. As periodic or continuous treatments for prevention of acidification of material actively handled and moved, and in controlled release form for reclamation where treatment sets up a change in site microbiology to replace Thiobacillus jerrooxidans with heterotrophic bacteria and creates a stable site without need for further treatments. Use of bactericides begins with laboratory testing of the site material to qualify the site for bactericide treatment and dosage determination. Site specific treatments are designed for active sites and reclamation. Once acid production is controlled, solubilization of metals is correspondingly reduced. Results and economics of use from active and reclamation sites are presented.

Jan 1, 1995

By Gordon M. Pugh, David G. Rasmussen

INTRODUCTION Variations of cut-and-fill stoping are the most com¬monly used methods for mining narrow, nearly vertical veins in the United States. These methods offer the best results for maximum recovery with minimum dilution of ore from veins that are irregular along either the strike or dip. The vein is developed along the strike length in 61 to 122-m (200 to 400-ft) blocks or stopes and vertical lifts of 48 to 61 m (160 to 200 ft). The level development is either on the vein or parallel to the vein with crosscuts intersecting the vein at right angles. Chutes, manways, and haulageways are established in the center of the stopes at the lower level of the ore block. Mining then proceeds by blasting down horizon¬tal cuts of the vein in both directions from the chutes. As the blasting continues to the outer limits of the stope, the broken ore is slushed into chutes. When a complete horizontal cut has been made, the service raise, including chutes, is raised to a height of a few feet from the back and the stope is filled with either sand fill or waste rock. Then the mining cycle is re¬peated until the stope reaches the level above. LEVEL DEVELOPMENT PRACTICES FOROVERHAND CUT-AND-FILL STOPING The two basic methods used are: the development of the haulage drift within the vein, or development of the haulage drift paralleling the vein in the foot or hang¬ing wall with crosscuts intersecting the vein perpendicu¬larly at 61 to 122-m (200 to 400-ft) intervals. The il¬lustrated methods of development show some of the variations which can be used to develop a stope block. Development for cut-and-fill stoping does not require the more elaborate schemes outlined for shrinkage or sublevel long-hole stoping. Many factors must be considered before deciding on a particular development plan and usually the nature of the vein or host rock will dictate which specific method to use. If the vein is relatively straight, then develop¬ment on the vein is feasible. Excessive ground pressures will necessitate the leaving of a sill pillar (Fig. la) and highly fractured or "heavy" ground will require timber¬ing (Fig. 1b). Development on the vein has the advan¬tage of reducing capital outlay and waste handling prob¬lems since the mined rock has value. But removal of the sill pillar is costly and time-consuming. Also, where no sill pillar is left, excessive timbering and subsequent level repair is a drawback. Development in either the footwall or hanging wall is the more common procedure and offers greater flexi¬bility in overall development of consecutive stoping blocks by partially disassociating the level development from stoping (Fig. 2). Chutes, manways, and service compartments com¬prise the next stage of development and are established in a variety of ways. They provide access to the stope, ventilation ducts, and muck removal. Several different development systems are in common use and average costs for each have been developed. Raise borers are used to establish ventilation and ac¬cess from level to level. They are normally reamed with 1.5 or 1.8-m (5 or 6-ft) cutterhead and may either be left bald or lined with a 1.2-m (4-ft) diam corrugated liner plate, depending upon ground conditions. The use of an Alimak raise climber is another effec¬tive method of driving a raise from level to level. The size usually is 2 m (7 ft) square. This will accommodate square-set timber which is later installed. This method of raising is not common, although it is a valuable tech¬nique especially when level spacing is several hundred feet.

Jan 1, 1982

By Yuanhong Qi, KUN CUI, Tao Fu, XUNFU WANG, Yingyi Zhang, Jie Wang, Junmao Qie

Indonesia is the world’s largest supplier of nickel ore, with its nickel laterite resources mainly distributed in the Moluccas and Sulawesi regions. However, there is scant research on the mineralogical characteristics of nickel laterite deposited in southeast Sulawesi, Indonesia. This study aims to investigate the mineral phase structure, chemical analysis, thermal properties and element distribution characteristics of the middle and lower part of the rotten rock zone located in the southeast Sulawesi ophiolite belt.

By Peter Kreisman

This paper and four other papers on the MEQB Regional Copper-Nickel Study, presented today at the 1980 SME-AIME Fall Meeting are based on four years of extensive inter-disciplinary research and impact analysis supported largely by the Legislative Commission on Minnesota Resources. This study produced over 180 technical reports, the results of which are summarized in an Executive Summary available from the Minnesota State Planning Agency, St. Paul, Minnesota. Those wishing to examine the many scientific, engineering, environmental, social or economic issues addressed by the study are encouraged to obtain a copy of the executive summary. The executive summary includes a bibliography of all technical reports produced by the Regional Copper-Nickel Study and describes the five-volume, 26 chapter Regional Copper-Nickel Study Report.

Jan 1, 1980

By Ashish Ranjan Kumar, Barbara Arnold, Luis Sanchez Gonzalez

Mining operations adopt a variety of measures to combat dust underground. These include ventilation systems, water sprays, physical barriers, and personal protective equipment (PPE). Respirators are a class of PPE that can provide clean particulate- free air to the users. Powered Air-Purifying Respirators (PAPRs), used commonly in medical and other sectors, are not prevalent in the mining industry. PAPRs use P100 filters that have 99.97% dust capture efficiency. We present our results from computational fluid dynamics (CFD) simulations performed to understand the impact of a PAPR on airflow and dust particle concentration in the vicinity of a miner. These results show that donning PAPRs can lower the miners’ exposure to respirable coal dust in their workplace.

Feb 1, 2025

By John C. Stankus, Xiaoting Li

"Underground structures such as square/arch steel sets, king beam structure, etc. are widely used for the roof support in underground mines. In designing and calculating the required carrying capacities of these underground structures, how to identify the potential rock loads on the structures over long term is a challenging question for ground control engineers. Based on design practices of both surface and underground structures, the authors have designed and applied appropriate load factors in underground structural design projects to accommodate different rock types, entry/intersection dimensions, roof geologic conditions, availability of roof bolts and water, etc. Classification of load (reduction) factors for large-span underground intersections are presented for reference. Success of designed underground structures in effectively controlling roof stability over years indicates the applicability of the load factor concept for underground structural design, and its same significance as of the load factor specifications in ASCE/SEI 7-10, ANSI/AWC NDS (2015), and AASHTO LRFD Bridge Design Specifications (2012). The load factor concept will facilitate ground control engineers to appropriately utilize popular structural design software, such as RISA-3D or -2D to efficiently design and analyze underground structures.LOAD FACTORS IN STRUCTURE DESIGN APPLICATIONSIn structures, loads cause stresses, deformations, and displacements. Assessment of their effects is carried out by the methods of structural analysis with the help of load factors. Loads can be increased or decreased by load factors to meet the requirement that design strength be higher than maximum loads. For example, in designing a staircase, a dead load factor may be 1.2 times the weight of the structure, and a live load factor may be 1.6 times the maximum expected live load based on the Load and Resistance Factor Design (LRFD) method. These two “factored loads” are combined (added) to determine the “required strength” of the staircase. Some typical load (adjustment) factors used for structural design are as follows:"

Jan 1, 2019

By N. T. Mirafuente

The device called "Dynastat" which is described in this paper provides in situ borehole measurements of dynamic and static parameters for use in a wide range of design applications. In the dynamic mode, the Oynastat allows measurement of shear and compressional wave velocities. From these velocities in situ dynamic Young's modulus and other dynamic elastic parameters can be calculated. In the static mode, the device allows measurement of load versus deformation from which in situ static Young's modulus can be determined. Information on in situ horizontal stress fields can also be estimated from the load deformation curves. In both mining and civil engineering applications, in situ measurements of material properties are indispensable for accurate evaluation of design problems. Laboratory dynamic and static test measurements in comparison to in situ measurements are often misleading due to difficulties in obtaining undisturbed samples and in accounting for effects of discontinuities, etc., for mass properties. The authors contend that among in situ borehole measurements, the method employed with the Dynastat is the most accurate and detailed technique of measurement. The Dynastat employs the crosshole method of shear and compressional wave velocity measurements and also performs as a borehole jack to obtain load-deformation relationships of materials at exactly the same location at depth in boreholes.

Jan 1, 1976

By R. K. Singhal

In 1990, Canada produced 69.4 million tonnes of coal of which some 65 million tonnes was produced by surface mining methods. The two Western Canadian provinces, Alberta and British Columbia produce approximately 55 million tonnes of coal by surface mining. These two provinces also provide the bulk of the metallurgical coal for export. The geological conditions in the Western provinces particularly in the mountain region are complex. The coals are friable and the seams are steeply inclined, sheared, intensely folded and faulted. There may be as many as 15 to 20 coal seams present with true thickness ranging from 2.5m to over 8m. Most of the Canadian surface mines are large (over one million tonnes per annum) using modern mining equipment such as shovels, high capacity trucks and draglines. At present the industry is going through another of its up and down cycles but has continued to make impressive gains in productivity. This has been achieved through greater mechanization, better management practices and a will to succeed in a highly competitive world market. This paper presents an insight into the current status of the Canadian Coal Industry and essentially updates the information given by the authors at previous SME meetings.

Jan 1, 1992

By Elham Najafi, Sina Ghassa, Ata Akcil, Hadi Abdollahi, Sandeep Panda, Zohreh Boroumand

The present study investigates the influence of ferrous iron (as FeSO4) and ferric iron (as Fe2 (SO4)3), and pyrite (FeS2) on the ability of bacterial leaching of a high-grade sulfide Zn–Pb ore. In this regard, shake flask experiments were carried out at 5% (w/v) pulp density of the ore sample (having 40.7% Zn and 12.4% Pb initial metal content) using a consortium of mesophilic iron and sulfur-oxidizing acidophiles. A concentration of 0.04 mol/L of ferric iron in the leaching media was found to be optimum for zinc extraction without affecting growth of the microorganisms. Under this concentration, the dissolution of Zn, Pb, Cd, and As was found to be 57%, 0.2%, 0.03%, and 9.9% in 25 days. Using ferrous iron in the media, 0.16 mol/L of Fe2+ was found to be the optimum concentration for efficient bacterial growth and metal dissolution (54.6% Zn, 0.08% Pb, 0.03% Cd, and 10.2% As) from the sample in 25 days. On the other hand, using pyrite as the source of energy for bacterial growth, an initial 12-day lag period was observed when compared to the effect of ferrous iron in the media. Under the optimum concentration (test with 0.24 mol/L iron in the form of pyrite), the dissolution of Zn, Pb, Cd, and As was found to be 39.8%, 0.1%, 0.03%, and 10% in 25 days. The surface chemistry analysis indicated formation of a sulfur layer over the particle surface that hindered reagent diffusion and affected metal recovery through bioleaching.

By H. Huang, Y. Takasaki, R. N. Gow

"The presence of enargite (Cu3AsS4) is problematic in gold processing not only because it is environmentally hazardous due to its arsenic content but also because it is refractory, causing both cyanide and oxygen consumption to increase. Selective leaching and treatment of the arsenic would prove advantageous prior to applying conventional gold leaching methods. The reactivity of enargite samples from Butte in Montana and Quiruvilca in Peru were spectroelectrochemically studied under alkaline conditions – pH range of 8-13 – using Raman spectroscopy and cyclic voltammetry. Raman spectra of the surface were collected during and after potentially cycled voltammograms in order to determine and compare surface species against theoretical EH-pH diagrams constructed from mass-balanced thermodynamic calculations using STABCAL software. Under slightly oxidizing conditions, covellite (CuS) peaks were found in a short matter of time, above ~100mV vs. SHE for pH 9-13, suggesting arsenic leaching specifically occurred. At longer conditioning times, elemental sulfur was observed and tended to passivate the surface. The presence of elemental sulfur suggests that copper was eventually leached from the surface. By operating above pH 12, under slightly reducing conditions near -300mV vs. SHE, arsenic can be selectively leached as thioarsenate (AsS4 3-) or thiosulfate (AsO4 3-) without sulfur formation.IntroductionEnvironmental limitations on the pyrometallurgical processing of ores and concentrates bearing arsenic and other hazardous elements often outweigh those in hydrometallurgical processing, particularly when minerals like enargite (Cu3AsS4) are present. The hydrometallurgical processing of enargite-bearing materials can be broken down into two categories: the co-dissolution of arsenic and copper under acidic conditions, and the selective leaching of arsenic under alkaline conditions. The selective leaching of arsenic would preferably leave behind copperrich sulfide phases, typically chalcocite (Cu2S) and covellite (CuS), which then become more amenable to conventional copper (and gold) processing. Furthermore, selective leaching often employs sulfide (S2-), bisulfide (HS-) or polysulfides (Sx 2-) to first extract the arsenic as AsS4 3- and AsS3 3- (Delfini et al., 2003; Anderson and Twidwell, 2008; Tongamp et al., 2010a,b). However, hypochlorite leaching (Viñals et al., 2003; Curreli et al., 2005; Mihajlovic et al., 2007), ammoniacal leaching (Gajam and Raghavan, 1983), and finegrinding to mechanically activate the enargite surface (Balaz and Achimovicova, 2006) have been investigated to increase the normally slow reaction kinetics. Because gold ores are becoming more refractory due to increasing arsenic contents as a result of enargite, these methods are being examined more closely by the gold-mining industry."

Jan 1, 2015

By Fiona Woody

Lime and calcium products are essential building blocks to the modern world, used to purify drinking water, preserve foods, control pollution, improve crop yields, stabilize soil, pave roads and support the production of steel, cement, glass and many other industrial processes. While lime is essential to society, it also has a sizable carbon footprint. Approximately 30 percent of the scope 1 greenhouse-gas emissions generated during lime production are related to fuel combustion, while the remaining 70 percent are created by the unavoidable chemical reactions during calcination.

Feb 1, 2026

By T. F. Tomb

Although the number of operating longwalls (approximately 100) has remained relatively constant over the last five years, longwall production levels have significantly increased and have in some cases more than doubled. As a result of the increase in production, control of respirable dust continues to be a problem. Therefore, the Mine Safety and Health Administration (MSHA), Pittsburgh Health Technology Center, Dust Division, initiated a study to investigate the availability and adequacy of technology for controlling dust on longwall mining operations. The study involved 23 longwall mining operations in three different Coal Mine Safety and Health Districts. Each operation was visited for a full shift. Respirable dust samples were collected on face personnel and at fixed locations along the face. Parameters in place to control dust were noted with special emphasis being placed on obtaining quantitative data relative to ventilation, headgate dust controls, cutting drum sprays, external machine sprays and cutting sequence. Attempts were made to correlate dust control parameters in place at the time of the visit to dust concentrations measured at the tail, on the tail shearer operator and to the average concentration of dust determined from averaging all the measurements from the occupations sampled. The parameters in place to control dust were also compared to the gamut of technology known to be available to control dust. This paper presents a discussion on the effectiveness of the parameters in place to control dust and on the availability of technology for the control of dust on longwall mining operations.

Jan 1, 1990

By Rudolph K. Hoagberg

Silica sands, primarily in coarse size ranges, are used to prop open fissures or voids created in the bedrock by hydraulic fracturing. This procedure increases the flow of gas or oil or other fluids through the bedrock. The sands used must meet various specifications relating to their physical and chemical characteristics; these characteristics are in turn related to the geologic provenance of the sand deposit. The Jordan Sandstone, a Cambrian formation which is exposed in Minnesota and Wisconsin, meets the specifications of a proppant sand by virtue of its depositional history and its sedimentary properties.

Jan 1, 1980

By Yu. A. Polozov, V. A. Lagunov, O. Yu. Lushinkova, Yu. I. Svirskiy, Eh. Ya. Kipko, Roy A. Williams

The integrated grouting method, as developed in the USSR, can be applied to multi-purpose operations that in¬clude the construction of shafts, drifts, and tunnels. In the USSR it has been used extensively for primary and final grouting of underground excavations in mining engineering and civil engineering projects. It has been applied in projects where the primary objective was waste contain¬ment. It has also been used to control subsidence beneath buildings located over mined out openings and to eliminate the flow of ground water beneath and around dams. This chapter considers several industrial applications in the form of case histories. The word integrated means that grouting activities were integrated with other activities. 9.1 CASE HISTORIES OF GROUTING FRACTURED ROCK WHEN SINKING VERTICAL SHAFTS The integrated grouting method has been applied most widely in the sinking of vertical shafts in fractured, satu¬rated rock. As explained previously herein, grouting oper¬ations preferably are carried out from the ground surface where they are integrated into the schedule of the setting up of the shaft excavation and construction equipment. Such integration reduces significantly the length of time required for preparation of shaft construction and the time required to sink the shaft. The increased efficiency is achieved by the elimination of cementing operations from the working face upon the penetration of each aquifer. Concomitantly, labor and energy consumption during shaft sinking are minimized because the more complicated and labor-intensive work is carried out at the surface section. Table 9.1 presents the essential parameters for the pri¬mary grouting of saturated fractured rock carried out by the integrated method when sinking vertical shafts. 9.1.1 NAGOL'CHANSK MINE NO. 1-2, VENTILATION SHAFT NO. 1 The integrated method of grouting saturated fractured rock through holes drilled from the surface was employed for the first time during the sinking of vertical shaft No. 1 at the Nagol'chansk mine No. 1-2 in the Don Basin. The shaft had an inside diameter of 6 m. The grouting operations commenced by using cement grout with a density of 1.7 to 1.8 g/cm3. Clay-based grout with cement and other addi¬tives was used only for grouting the aquifers where cement grout proved to be ineffective. By the time of the grouting of ventilation shaft No. 1, hydrodynamic analytical methods were well in hand and a preliminary method for designing isolation curtains had been developed. A method for designing and drilling ori¬ented directional drillholes whose natural curvature re¬flected the characteristics of the fracturing had been devel¬oped. The majority of the necessary equipment had been developed. Clay-based grouts were considered ready for the first industrial application. The hydrogeologic environment at the Nagol'chansk mine No. 1-2 contained many prolific aquifers. According to monitored drillhole data, the total expected inflow into ventilation shaft No. 1 without grouting was 425 m3/hr. This large ground water inflow rate was supported by in¬formation obtained during the sinking of the main shaft and from the auxiliary shaft of the "Nagol'chansk" mine No. 1-2. The construction contractor at the site began sinking ventilation shaft No. 1 to a depth of 217 m by cement grouting from the working face of the shaft. As a result of the cement grouting operations, the inflow of water was reduced from 105.4 to 40 m3/hr over this 217 m interval. However the residual inflow rate caused the shaft to become functional at the depth of 217 m. At this point all grouting operations were transferred to SPETSTAMPONAZHGEOLOGIA (STG) in an effort to improve the operation. In order to carry out preliminary grouting from the sur¬face, seven drillholes were installed and designed for re¬ceiving grout. Drilling was implemented by the ZIF-1200A rig to the design depth of the shaft. The geometry of the principal fracture system that would be encountered by the shaft was estimated while breaking ground using informa¬tion from other shafts. The drillholes were arranged so that they cut through the aquifers uniformly around the shaft. Some preference was given to intercepting the aquifers that were located along the rock dip above the shaft bottom because it was established that the ground water in the aqui¬fers flowed downward along the dip into the cone of de¬pression of the shaft. Investigations in the drillholes using the DAU-3M flowmeter described in Chapter 3 showed that all of the nine aquifers were penetrated by the time a depth of 690 m had been reached. The flowmetric information showed that aquifer numbers [1 and 2], [4, 5, and 6], and [7 and 8] had practically identical hydrogeologic characteristics; so it was decided to connect them into combined stopes and inject the grout using five stopes instead of nine. Table 9.2 shows the results of the grout injection into the aquifers. The volume of grout pumped into the separate intervals through single holes varied from 204 to 15 m3. The trend of grout consumption showed a reduction of grout consump

Jan 1, 1993

By David Burgess

The paper describes experiments at the Somerset Mine, Colorado, under a cooperative agreement between U.S. Steel and the U.S. Bureau of Mines. A test was made of the tube bundle method of mine air sampling which was developed for European longwall mines; some 300,000 running feet of tubing provided continuous monitoring at 39 locations. Carbon monoxide concentrations were obtained with standard deviations of less than 1 ppm and oxygen concentrations with standard deviations of less than 0.20 percent. A base level of CO index (ppm CO evolved/pct 02 removed)) was determined to be 0-20 through most of the passageways of the mine. However, in one working section where the coal was initially hot (about 180°F within a pillar) the CO index was demonstrably higher, >20 trending upward to about 60 when smoke appeared and the section was sealed. The data are discussed in terms of laboratory experiments on CO evolution from coals.

Jan 1, 1975