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By John C. Franklin

INTRODUCTION Alpha-emitting radon daughter products are a recognized health hazard contributing to the development of lung cancer in persons exposed to excessive concentrations over an extended period. The primary control technique used by the mining industry is dilution with fresh air. Because the uranium is deposited in porous sandstone in most U.S. mines, the control of this hazard with dilution requires excessive volumes of air. As the mines become larger and deeper, the cost for ventilation becomes a major expenditure for the mining industry Most mines are maintaining levels of radon daughters lower than required by law but to maintain these present levels the mining companies are going to be drilling more vent holes or using other control techniques as the mine expands. These control techniques are being investigated by the U.S. Bureau of Mines to determine their effectiveness, cost, and safety for underground use. [Instrumentation] The Bureau's early research used batch sampling of both radon and radon daughters for determining the concentrations. In both cases it was determined that batch sampling did not give a complete understanding of what was happening to the concentrations. Therefore, continuous monitoring systems were developed for both. A continuous flow-through radon monitor was developed by the Bureau (Franklin, et al.., 1976) using a modified Lucas flask mounted on a photo-multiplier-tube (PMT). The air was filtered twice prior to entering the flask to remove dust, diesel smoke, moisture, and radon daughters. The signal from the PMT was then inverted, shaped, and driven to a data acquisition system (DAS) which accumulated the signals. The DAS has two comparators for setting the scan time and interval, a 24-hour clock with Julian date, a printer, and a perforated paper tape punch. Sampling time is determined by the type of experiment being conducted. Our normal sampling time is one 10-minute scan every 15 minutes, but could be set as low as one scan every minute or as long as a 16-minute scan every 167 minutes. There are two models of the DAS. One is a 10-channel system and the other a 30-channel system. On a 10channel system normally 6 channels are used for radon and radon daughters and the other four channels are used for temperature, absolute pressure, air velocity, and relative humidity as described by McVey (1977). Figure 1 shows a scan of five channels monitoring the radon concentration from intake to exhaust in a stope for a non-work day. If the radon concentration stayed this constant during work days, then continuous monitoring would not be necessary. [ ] Droullard (1977) developed a continuous alpha-beta monitor. This monitor has been modified so that it can be used with the DAS previously mentioned. Control Techniques The U.S. Bureau of Mines has conducted research in controlling airborne radon/radon daughter for 10 years. Major efforts have been conducted in sealing the rock to prevent the release of radon, bulkheading to seal nonproductive areas, backfilling with classified mill tailings, and positive pressure ventilation. Also, tests have been conducted to determine the mining cycles contribution to the radon problem. Surveys One stope with two intakes and one exhaust was continuously monitored to determine the effects of blasting, mucking, and drilling on the airborne radon concentrations (Franklin, et al., 1976). Two other detectors were located inside the stope which were close to the mining activity. In this stope no appreciable change in radon was observed during drilling holes for blasting and roofbolts. This study showed increases up to 40 percent in radon for slushing in the exhaust drift. The radon remained high and very eratic during slushing, as can

Jan 1, 1981

By V. A. Evans, D. L. Hellier

Introduction Preparing a report on fluorspar economics in the US could have been somewhat complicated at one time. This was because the number of producers, both large and small, derived production from different locations through-out the western US and the southern Illinois and western Kentucky fluorspar districts. Companies, such as Allied Chemical and Williamson Mining, operated in the Jamestown district of Colorado. In Walden and Salida, CO, the Colorado Fluorspar Co. operated mines and mills. In New Mexico, both Du Pont and Allied had mines and mills. In the Illinois-Kentucky district, US Steel, Rosiclare Lead and Fluorspar, Minerva Oil Co., and Marathon Oil all produced spar for the US market. Now all of the companies are out of the fluorspar business and, with the closing of the Seaforth Mineral Co.'s Inverness Mining Co., only Ozark-Mahoning Co. survives as a domestic producer. History Ozark-Mahoning Co., as it now exists, was formed in 1946 by the merger of Ozark Chemical Co. and Mahoning Mining Co. Ozark Chemical Co. was started in the 1920s to produce sulfuric acid for two refineries located in Tulsa, OK. Mahoning Mining was originally formed by the founders of Ozark Chemical Co. and the Youngstown Sheet and Tube Co. to produce fluorspar and associated minerals from properties in the southern Illinois and western Kentucky fluorspar district. In 1937, Mahoning Mining started an exploration program, employing contractor-operated churn drills to look for reserves in Hardin and Pope Counties in southern Illinois. Based on known reserves in the area and the results of the exploration program, work was started in the fall of 1938 to build a selective flotation plant at Rosiclare, IL. The mill was completed and put into operation in August of 1939. The company's first mining efforts were in the Cave-In-Rock district just north of Cave-In-Rock, IL. The deposits in this area were of the Manto or bedded type. They have been the mainstay of the company's operations in the Illinois-Kentucky fluorspar district. In 1950-51, the company expanded its operations by purchasing fluorspar properties from Colorado Fluorspar Co. at Walden, CO. This location was operated from 1952 through 1958 and from 1968 through 1973 when it was closed permanently. In 1974, Ozark-Mahoning Co. was purchased by Pennwalt Corp. Presently, it is a wholly-owned subsidiary of that corporation. As the only major domestic producer of fluorspar in the US today, the company's ability to survive under the flood of imported material is directly related to the four major segments of its operations. These are: exploration, mining productivity, maintenance, and marketing. All are done in-house. Exploration The Southern Illinois fluorspar district is located in the most complexly faulted area of the central US. The formations are sedimentary in nature and range in age from Devonian through Pennsylvanian. The deposits are small, they range in size from a few thousand tons up to 900 kt (1 million st), with the average minable deposit being around 544 kt (600,000 st). The ore occurs in both nearly vertical vein deposits and in Manto or flat lying bedded deposits. Due to the relatively small size of the deposits and the depth at which they occur, a good low-cost exploration program is essential to ensure that new finds of ore are equal to the mining extraction rate. Ozark maintains a leased land position in Pope and Hardin Counties that will fluctuate somewhere between a low of 40 km2 (10,000 acres) up to a high of 129 km2 (32,000 acres), depending on total size of tracts leased. Most all lands are obtained through a standard lease purchase option, with a small minimum royalty per acre during the normal 10-year option period. All leases are standard and without deviation to control land cost. All land leasing is done through Ozark's geological staff, who also

Jan 11, 1986

By Francis S. Kendorski

INTRODUCTION Drift design problems in caving operations are a re¬sult of the geologic factors contributing to the overall success of the system, of the engineering factors dictated by economic and technical considerations, and of ore production practices. Combining these factors, a rational underground sup¬port system of rock reinforcement, light steel channel section or welded wire fabric, and shotcrete can be de¬signed based on rock fracturing, rock load, abutment loadings, ground movement, expected repair and desired flexibility. The design concept uses the effect achieved by restraining, reinforcing, and maintaining some of the intrinsic strength of the fractured rock mass composed of interlocked blocks of intact rock and rock fractures. Three different examples of drift support design in hypo¬thetical mines using the caving system are given. Caving is a system of underground mining where ore is extracted by means of gravity after the ore body is allowed to fail by removing support from underneath. The rock mass of the ore body fractures and flows ver¬tically downward to let gravity do as much work as pos¬sible. Caving differs from many other mining systems in that blasting is used only to initiate the rock mass failure by removing the rock supporting the ore but not to break the ore itself. The initial movement of the rock mass dur¬ing failure and the consequent crushing and grinding during the continued movement serve to reduce the ore to particles of a manageable size, with only limited sec¬ondary blasting necessary. The broken ore is extracted from the bottom of the failed rock mass through funnels of some sort pre-excavated in the rock. Ore extraction must continue or the swell of the broken rock will even¬tually fill the cavity and stop further rock mass failure and movement. The excellent general discussion on block caving in the SME Mining Engineering Handbook (Julin and Tobie, 1973) adequately covers the principles and application of this type of underground mining. Many rock mechanics aspects of block caving have been covered by others (McMahon and Kendrick, 1969; Swaisgood, et al., 1972; Mahtab and Dixon, 1975; King, 1946) and will not be reviewed further. Maintaining the stability of production drifts is one of the most troublesome problems plaguing the mine manager in a caving operation. Many factors contrib¬ute to drift support problems, and identifying the causes of instability and producing a reasonable support design are two steps toward achieving stability consistent with the mine plan. This chapter sets forth a technique for the design of support systems for production drifts in caving opera¬tions. The basic support system elements employed are rock reinforcement, welded wire fabric, and shotcrete. Recognized as contributing to the design are the factors of rock load, additional load from mining activity, rock fracture characteristics, repair expected, and flexibility. It must be emphasized that the drift must first be stabilized as for a tunnel, and the additional strengthen¬ing for mining-induced loads cannot contribute to the initial premining stabilization, or the reserve of strength is used up. MINE PLANNING The efficient mine planning engineer not only must satisfy the economic, human, and environmental aspects of his task but must also consider the mechanical con¬sequences of his plan. The problems created for the mine by placing parallel drifts too close, by crossing drifts on different levels with inadequate, if any, separa¬tion, and by installing connections and crossovers in the haulage plan without regard for the effective spans cre¬ated, are only a few of the problems a mine planning engineer can create for himself and the mine. The effect on immediately adjacent mine areas when an area is caved is important to drift design because the removal of vertical support from a rock mass causes the weight of that rock mass to be shifted elsewhere. The adjacent rock mass will carry this load and reach a new equilibrium with the applied stress. The advancing front of stress increase that results from caving (and many other mining systems) is generally called the abutment load and is the increase in stress over the gravity or tec¬tonic stress that already exists, as shown in Fig. 1. In general, the abutment load will be similar in nature to the stress change found around an opening in rock and will be taken as causing an increase in the vertical prin¬cipal stress, due to an approaching caving boundary, so that

Jan 1, 1982

By M. G. Ayat, B. C. Scott

J. Dasher Having an interest in coal slurry pipelines from a decade of arguments with Ed Wasp and crew at Bechtel about pumping thicker slurries slower, I immediately read this article and found nothing in it pertinant to the title. Ayat and Scott pumped five unidentified coals at 9% solids (50%-60% is the area of interest) in a 25-mm (1-in.) pipe around two square elbows through a 76-mm (3-in.) cyclone for up to 60 minutes and sized a large number of samples of unstated size on an unidentified wet screening device. They did not measure power consumption, discharge pressure, flow rate, or give pump tip speed or impeller diameter. They siad the cyclone had a "high" pressure drop (unmeasured) and did much of the degradation (measured) but more or less per unit of energy expended? Hargrove was not measured, so there are no data as to whether it would correlate with degradation. The authors conclude, with no attempt, correlation is "hard to establish." Please experiment before concluding. I am at a loss to know what "increasingly smaller size" means, much less what theory says such particles take "exponentially larger quantities of energy," which the authors neglected to measure. If such experiments without pertinent data or justified conclusions must be published, please attach a pertinant title. reply by M. G. Ayat The first and the last criticism of this paper is that the work is not pertinent to the title. Anybody who reads this article will immediately realize that the work describes the breakage of coal particles to finer sizes in coal slurry pipes and pumps. If this is so, why not title the work "Degradation process in coal slurry pipelines"? What could be more pertinent to this title than the investigation concerning the degradation phenomenon of coal particles in a pipe carrying a coal slurry? Mr. Dasher complains that he does not understand the meaning of the term "increasingly smaller size." The first sentence of this article defines the degradation process as "the breakage of coal particles to increasingly finer sizes." The term "increasingly finer sizes" here means successive breakage of a fine particle to finer and finer sizes. According to Hukki (Hukki, 1975), the probability of breakage is high for large particles and rapidly diminished for fine sizes. We apologize for being brief about some of these definitions. The degradation process, breakage of particles to increasingly finer sizes, is so widespread in the mineral industry that we did not feel it necessary to bore the reader with lengthy definitions of some simple terms. Mr. Dasher states that the coals examined and their original size consist are unidentified. Please look at Table I and Table 2 in the paper again where you will find the original size consist of the coals examined and their full specifications. We only named the coals A, B, C, D, and E to avoid identifying the coal seams that were more susceptible to degradation. Not identifying the name of the wet screening device used in this work has also been criticized. In our opinion, wet screening operation is such a routine and standard procedure that naming the device by which the screening is performed would serve no purpose but to promote a sales approach. This was not our intention. We did not conclude that the correlation between the degradation process and Hardgrove Grindability Index is hard to establish as Mr. Dasher writes in his letter. We stated, not concluded, that "It is reasonable to assume that some relationship between the extent of degradation and its physical properties, such as Hardgrove Grindability Index, does exist. However, any definite correlation is difficult to establish." This statement is based on other researchers work, which are clearly referenced in the article. The conclusions of this paper were based solely on the findings of the experimental work. As for the theory of comminution that Mr. Dasher asks, we would like to refer him to some basic comminution books and articles where various theories are clearly described. For example, Hukki (Hukki, R.T.,"The Principles of Comminution; An Analytical Summary," Eng. Min. 176, 106, 1975) suggests that the relationship between energy and particle size is a composite form of three laws (Bond's law, Kick's law, and Rittinger's law). A comprehensive analysis of coal breakage processes is also performed by Broadbent and Callcott (S.R. Broadbent, and T.G. Callcott, "Coal Breakage Processes, I. - A New Analysis of Coal Breakage Processes;" and "Coal Breakage Processes, II. - A Matrix Representation of Breakage," Journal of the Institute of Fuel, pp. 524-539, December 1956). These and many other relevant publications explain the relationship between the particle size and energy much better that what can be said in this short reply. We will, however, agree with Mr. Dasher on one particular point. We, too, believe that if the investigation concerning the degradation process in coal slurry pipeline were to be pursued further, one could choose to determine one or more of the variables available in the process, such as discharge pressure, power consumption, percent solid, pipe diameter, etc.

Jan 1, 1989

By L. Yuan, J. H. Rowland III

"In 1995 and 2003, the Mine Safety Health Administration (MSHA) conducted surveys to determine the number of atmospheric monitoring systems (AMS) being used in underground coal mines in the United States. The survey reports showed data for the different AMS manufacturers, the different types of equipment monitored, and the different type of gas sensors and their locations. Since the last survey in 2003, MSHA has changed the regulation requirements for early fire detection along the belt haulage entries. As of December 31, 2009, point-type heat sensors were prohibited for use for an early fire detection system; instead carbon monoxide (CO) sensors are now required. This report will examine the number of AMS currently used in underground mines in the United States to see if there is a change in their use. The locations and parameters monitored by AMS are also discussed. INTRODUCTION The last survey conducted by the Mine Safety and Health Administration (MSHA) on atmospheric monitoring systems (AMS) used in underground coal mines in the United States (US) was in 2003 [1]. The survey report states that since 1995 there had been a steady increase in the use of AMS in underground coal mines as the technology of CO sensors was improved by manufacturers. AMS had become more reliable and some CO sensors could be purchased with the capability to discriminate between the CO produced from a fire versus CO produced from diesel equipment. This increase in reliability had reduced the number of false alarms and the response time. The faster response time to a developing fire allowed for the miners to take the necessary action to put out the fire or exit the mine. It should also be noted that at the time of last MSHA survey, some mines still used point-type heat sensors to detect the presence of a fire on conveyor belts. This detection method works by measuring the temperature of the air around the sensor. As a fire starts to develop, smoke and hot gases are produced from the combustion process, and the air is heated up. Once the air and gases reach a designated temperature detected by the point-type heat sensor, a fire alarm signal is generated. Studies have shown that the response times for detecting the presence of a fire using CO sensors are much shorter than using point-type heat sensors [2]. Finally, since the last MSHA survey, regulations have changed on the type of automatic fire detection systems permitted to be used in conveyor belt entries. As of December 31, 2009, point-type heat sensors are not permitted to be used for fire detection in underground coal mines in the US; instead CO sensors are required as a replacement (30 CFR 75.1103-4). This study will examine how the changing of the above regulations has impacted the use of CO sensors in underground coal mines in the US compared to their use based on the last MSHA survey. Importantly, in the previous MSHA survey, the data was recorded for producing and non-producing coal mines. In this study, only the data from the producing coal mines was recorded. Data was collected from a total of 235 coal mines for this study with the help of the MSHA district offices."

Jan 1, 2017

By R. C. Jr. Babcock, V. S. Znamensky

The Pyke asbestos deposits were discovered by Kennecott Explorations (Australia) Pty. Ltd. in 1969 during the course of base metal exploration on the South Island of New Zealand. Extensive occurrences of chrysotile asbestos were found associated with the Red Mountain ultramafics in the headwaters of the Pyke River, and to the north and south along the ultramafic belt for a distance of 70 km. Surface evaluation and drilling by Kennecott in 1970 and 1971 identified numerous small deposits of good grade fiber in the Little Red Hills between the Pyke and Barrier Rivers. Additional drilling and three adits have been completed by other companies since that time. The property is presently held by Lime and Marble, Ltd., of Nelson, New Zealand, and Kennecott Explorations (Australia) Pty. Ltd. This chapter presents the regional and local setting of the asbestos deposits, including the geology and nature of mineralization. In addition, the exploration methods used by Kennecott in the discovery and evaluation of these deposits are outlined. Subsequent work by other companies is mentioned, but not de- scribed in detail. This subsequent work did not alter Kennecott's conclusions or produce any serious changes in the interpretation of geology and mineralization. The author acknowledges the contribution of the many employees of Kennecott Explorations (Australia) Pty. Ltd. who have been involved in the project. In particular, thanks go to Ramon Farmer, N.G. Corner, and A.W. McConie, who were responsible for completion of field work and reporting on the project. The responsibility for the conclusions and interpretations of this chapter are solely that of the author and result from his involvement in the discovery of the deposits and management of the exploration program. REGIONAL SETTING Ultramafic rocks on the South Island of New Zealand are thought to be Permian in age and occur between crystaIline and metamorphic continental rocks of Precambrian through Cretaceous age in the west, are thought to have once been part of the Antarctic subcontinent, and the extensive geosynclinal graywacke deposits of the Mesozoic New Zealand geosyncline in the east. They consist of the Dun Mountain ultramafics, a 130 km long segment of ultramafic rocks located at the north end of the South Island, and the Red Mountain ul- tramafics that occur in a zone 200 km long in the south half of the South Island as shown in Fig. 1. These two belts, once contiguous, have been offset a distance of 560 km by the still active Al- pine fault, which is analogous in character and offset to the San Andreas fault of the United States. The ultramafic rocks occur within a sequence of Permotriassic volcanic and sedimentary rocks that probably formed at the eastern margin of the continental plate. The ultramafics appear to be tectonically emplaced between Permian lithologies in the Dun Mountain belt (Coleman, 1966), and at the base of or between Permian volcanics and sediments in the Red Mountain belt. The Permotriassic ultramafic volcanic sequence of the Red Mountain ultramafic belt appears to be a typical ophiolite sequence. Ultramafic rocks, mixed mafic intrusive rocks and volcanics, volcanic rocks, and volcanoclastic sediments occur in apparent superposition from east to west in the area of Red Mountain. Farther south, ultramafic rocks are associated with basaltic volcanic rocks and some sediments, and still farther south are emplaced entirely in sediments. These structural compIications of the normal ophiolite sequence have resulted from the coincidence of Permian subduction, early Mesozoic extension and sedimentation, and Cretaceous subduction along the same plate margin (Landis and Bishop, 1972). The general geology in the area of the Red

Jan 1, 1986

Conveyor belting-Dunlop Belting Division has published a manual on its Starflex plied conveyor belting. The design section of the manual contains advice on the calculation of tensile strength and horsepower needs while the section on belt selection offers helpful recommendations. Circle 200 on reader service card Hydrostroke feeders-A pamphlet from Kone Corp. highlights the uses and operating principles of its hydrostroke feeders. Circle 201 on reader service card Electric cylinders-A 24-page catalog from Raco International Inc. describes applications for its electric linear actuators in addition to the electronic options for computer - controlled operation. Circle 202 on reader service card High torque drives-T. B. Wood's Sons Co. offers a 56-page booklet providing features and specifications on its high torque drives. Information includes a step-by-step drive selection proce¬dure. (HTD) Circle 203 on reader service card Sludge depth meter-The model 600 sludge-depth meter that locates the sludge bed in clarifiers and settling tanks is described in a four-page bro¬chure from Markland Specialty Engineering Ltd. (600-84) Circle 204 on reader service card Cavity pumps-An eight-page bulletin is available from Robbins & Myers Inc. It features the application of Moyno progressing cavity pumps in handling composite slurry fuels. (400) Circle 205 on reader service card Roller chain-A roller chain catalog shows heavy duty drive chains and other specialty conveyor chains. It is available from Peer Chain Co. (PC200) Circle 206 on reader service card Belt filter-Phoenix Process Equipment Co. has available a pamphlet detailing its belt filter press. The unit is designed to dewater refuse and clean coal, yielding easily handled dry filter cakes. Circle 207 on reader service card Capabilities - Literature from International Mineral Services Ltd. highlights its services and capabilities to the mining industry. Circle 208 on reader service card Motor analysis - How to select the proper electric motor by comparing life cycle costs, power costs, rate of return, and other factors, is described in a brochure from Westinghouse Electric Corp. (SA-11376) Circle 209 on reader service card Cavity pumps - A product application data sheet is available from Robbins & Myers Inc. It details the use of Moyno positive-displacement, progressing cavity pumps in handling ground limestone slurry. (PC-21) Circle 210 on reader service card Dust collectors - "Dust Collector Selection Guide," from American Air Filter, describes dry mechanical collectors, wet collectors, fabric collectors, and electrostatic precipitators. (CAD-1-901G) Circle 211 on reader service card Wet scrubber - A 12-page bulletin from The Ceilcote Co. provides a comprehensive description of its ionizing, wet-scrubber system. (12-19) Circle 212 on reader service card Metric o-rings-Simrit Corp. has published a 16-page brochure detailing its full line of standard metric o-rings. Information includes graphics and dimensional charts, and specific data on materials and application ranges. Circle 213 on reader service card Hearing protection - A 16-page catalog from Cabot Corp., EAR Division, provides information on its hearing protection devices and noise control products. Circle 214 on reader service card Toxic gas detection - Sensidyne Inc. is offering a guide for toxic gas monitoring. A description of the electrochemical sensors, as well as ranges, complete specifications, and interference charts are included. Circle 215 on reader service card Hydraulic bolting systems - Ingersoll-Rand Co. is offering a brochure on its line of hydraulic bolting systems. These systems, hydraulic wrench and power console, are designed for heavy duty bolting applications. Circle 216 on reader service card Temperature monitoring - A brochure describing the Ramsey Engineering Co.'s micromonitor temperature monitoring system is available. Three types of switches are available for monitoring bearing temperatures. (80.300) Circle 217 on reader service card Product catalog - Shadbolt & Boyd Co. has published a product catalog. Among items described are hoist slings and cranes; compressors; hydraulics; wire rope, chains, and fittings; and materials handling and shop equipment. Circle 218 on reader service card Cylinder controls - A 12-page booklet presenting Hanna Corp.'s line of electrical controls for cylinders is available. It features proximity and limit switches for hydraulic and pneumatic cylinders, and standard and 3-amp reed switches for pneumatic cylinders only. (550) Circle 219 on reader service card Slurry pump - Pettibone Corp. has published a 24-page booklet covering its heavy duty pumps made with 'diamond alloy' materials for handling slurries of abrasive materials. Circle 220 on reader service card

Jan 9, 1985

By Stanley Dempsey

INTRODUCTION The mining industry has been under- going a significant restructuring for the past several years. Some firms have been liquidated. Some have been merged with other firms. A number of mining divisions of larger firms have been spun off, in some cases in leveraged buyouts by management. Recently, there have been a number of spin outs of gold subsidiaries. There are also a number of start-up situations where new firms are emerging as significant gold producers through mine development or acquisition. This restructuring has created an attractive market for specialized mining finance firms, and a number of firms have recently entered that field in the United States. These firms provide financial advisory and valuation services to mining firms and project developers, and access to venture capital. In some cases, they invest for clients or for their own account. I have been asked to describe how our firm works, and what kind of information we expect to receive when mine developers seek our assistance in financing a project. HOW WE OPERATE Denver Mining Finance works for both mine developers and investors. We also provide expert advice on mining issues to governments and to firms who are in litigation over mining issues. A lot of our work involves valuing mineral interests for various purposes. Our principal expertise is in the technical, commercial, and financial area. We help put mining deals into shape commercially so that they can be financed, and we help investors get into mining deals that are commercially sound. We work on projects, but we also advise mining firms on the structure of their business and how to make the most out of one or several mines that they might own. In addition to our work for clients, we are also prepared to invest in mining projects for our own account, and to recommend gold properties for acquisition by our parent company, Royal Gold, Inc. We are, to some extent, a venture capital firm, in that we sometimes work with projects that are not "bankablerV1 and are prepared to invest in attractive early-stage situations in mining. However, I want to emphasize that our principal focus is on development of commercially attractive mining projects that can be brought into production in the best possible financial terms. If we can make a project qualify for a project loan from a commercial bank, we will not advise the developer to use high cost venture capital. For the purpose of this paper, I will focus primarily upon our functions with respect to development of precious metal projects. In a typical case, a firm with an advanced-staged project approaches us for help with financing the final feasibility study, and construction and operation of a mine. In some cases, the developer has already developed a financing plan and wants only our access to investors. If we agree that the project is sound and that the financing scheme will be attractive to investors, we will help the developer reach investors on a commission basis. In such cases, we are particularly insistent that the financing plan and our participation meet all regulatory requirements, particularly compliance with state and federal securities laws. In most cases, however, the developer asks us for advise both on how to structure the financing, and how it can be implemented. Here we work for fees based on hourly rates for advisory work,

Jan 1, 1987

By J. J. Bean

The development drilling program of the Copper Cities mine indicated that the molybdenum content of the ore would be about 0.011% MoS2' The molybdenum content of the Miami mine at the same time was about 0.025% MOS2' The Miami mill had a very successful molybdenum operation which used bleeder turbine steam from the Miami power plant for the steaming operation. The much lower grade of the Copper Cities ore, the absence of a source of low cost steam and the use of sodium aerofloat as the copper collector which Miami experience indicated did not respono readily to the steaming technique made very doubtful the success of a similar operation at Copper Cities. However, the relatively few pounds of copper that could be recovered from the low grade Copper Cities heads made very desirable the recovery of the molybdenwa,as a source of additional revenue. Accordingly, shortly after the Copper Cities mill began operation, a research program was instituted to develop a suitable method of molybdenum recovery. This program explored first, the successful steaming process then in use at Miami and as expected, found it unable to decompose the sodium aerofloat copper collector. The known chemical depressants were then tried without much encouragement. Finally the copper concentrate was dried, not roasted, to 2-3% mOisture and this along with sodium ferro-cyanide and cyanide resulted in sufficient copper depression that a final molybdenum concentrate containing 1.0% CU could be made. The bench scale testing was followed by pilot plant testing. Sixteen

Jan 1, 1972

By Williams R. Reynolds

Studies of the northeastern Mississippi bauxites were initiated primarily for the purpose of ascertaining the quality and quantity of the ore. Subsequently, this ore was found to have a rather dubious economic value as was previously suspected but never established in earlier economic reports (1, 2, 3). However, the quality of ore was found to be higher than previously thought. Even though the ore bodies were determined during the early stages of this study to be too thin, small and widespread for commercial production a continuation of the study was not abandoned. On the contrary, the present investigation was intensified concentrating less on the economic application and more on stratigraphy and genesis. Considering the ever increasing demand for aluminum products the northeastern Mississippi bauxite district which includes portions of Tippah, Benton, Pontotoc, and Union Counties may soon become, in spite of inferior quantities of ore, an attractive area for bauxite production. Presently, however, the meager quantity of ore discourages exploitation. On the other hand, these northeastern district bauxites are facies of more extensive units of lignite bearing high aluminum clays, and considering the advances in aluminum extraction from clay these clays may in the future beckon mining interest. The bauxite of the northeastern Mississippi district, mapped as a part of the Betheden Formation (4). has an outcrop trend along the county lines of Tippah and Benton Counties and the extreme western portion of Union and Pontotoc Counties. Studies of exposures and cores stratigraphically positions the outcrop between the middle Midway Porters Creek Clay and the lower Eocene Wilcox sands and

Jan 1, 1977

By Joseph R. Hatch

Previous studies of the Thompson and Temperly ore bodies show alteration aureoles in the Middle Ordovician host-rock limestones, dolomites and shales extending laterally as much as 60m from the ore. Thickness, texture, carbonate and clay mineralogy, major-, minor-, and trace-element contents; and carbonate carbon and carbonate oxygen Isotope compositions of the host rocks all have been significantly altered by hydrothermal processes. Organic geochemical analyses of samples from the basal 8-cm-thick carbonaceous shale bed and immediately overlying limestone bed of the Quimbys Mill Member of the Platteville Formation, collected from within the ore, the alteration aureole, and the unaltered country rock show: 1) solution of carbonates and residual concentration of organic matter, 2) decreases in the volatile (Si) and pyrolytic (S2) hydrocarbon yields In samples of the basal shale from within the ore, 3) decreases in the relative amounts of the lower molecular weight (<nC25) saturated hydrocarbons in samples from the ore, and 4) no changes in either the organic carbon isotope composition or the Rock-Eval or biomarker thermal maturation Indicators. In light of the 125-170°C depositional temperatures documented for the early marmatitic sphalerltes in these deposits by fluid Inclusion studies, the lack of thermal alteration of the organic matter suggests a relatively short mineralization episode, a conclusion supported by the relatively simple sphalerite band stratigraphy. The decreases in the volatile and pyrolytic hydrocarbon yields for shale samples from the ore supports a model for mineral deposition which obtains sulfide through reaction of brine sulfate and indigenous organic matter.

Jan 1, 2013

By J. W. Stevenson

The need for a separate ventilation department within the management team was recognized by the management of U. S. Pipe and Foundry before the company was purchased by Jim Walter Corporation in January, 1970. Because the coal beds of the deep underground coal properties owned by Jim Walter Resources lay 1300 to 2200 feet below the surface, efficient mine ventilation practices are essential through shaft openings varying from 16 to 22 feet in diameter. The extremely high methane gas quantity indicated to be in the coal bed because of the depth of cover and documented by various U. S. Bureau of Mines tests and demonstrated by early mining proved the need to implement this early management decision. At U. S. Pipe and Foundry mine ventilation was a function of the Mining Engineering Department with ventilation personnel assigned to the mine receiving their day to day work assignments from the mine manager or mine foreman. At Jim Walter Resources, the General Manager - Ventilation reports to the Senior Vice President, Mining Engineering, who reports directly to the President, Mining Division, Figure 1 (Company Organization Chart). Position descriptions are static; however positions are dynamic and consequently the degree of emphasis of specific duties and responsibilities change depending not only upon the aptitude of the individual performing the job but also upon his ability to meet the needs of his supervisors; and ultimately upon his ability to react to the economic forces directing our society. The following describes the duties and responsibilities of the General Manager - Ventilation, Figure 2.

Jan 1, 1980

By Donald D. Haas

Contract mining has a definite place in the mining industry, both in the United States and around the world. A mine owner should evaluate the utilization of contract mining in the devel­opment and production of mining properties. This evaluation of contract mining should be performed at various times in the life of a mining prope1iy from the development and initial operation of a new mining project, periodically throughout the mine's life, and at mine closure. Many factors must be considered in ana­lyzing and evaluating the potential of contract mining and those factors are both quantitative and qualitative. Quantitative factors are those factors for which costs or measures can be determined or estimated objectively; and qualitative factors are those factors for which costs or measures must be determined subjectively and do not lend themselves to objective determination or estimation. Some of the qualitative aspects may provide only yes, maybe, or no answers while others will provide a tendency or partiality toward one answer or the other without a quantified cost. The major quantitative aspects include operating and capital costs, monitoring costs, capital preservation, acquisition and disposal of new and used equipment, direct payroll and administrative costs, recruiting and training costs, and employment and termi­nation of staff and hourly personnel. The major qualitative areas include corporate philosophy, risks associated with capital and operating costs, financing the project, operating control, operat­ing flexibility, recruiting and training of personnel, employment and termination, administrative concerns, and owner/contractor relationship.

Jan 1, 1998

By Gregg Sherry, Bob Gordan

INTRODUCTION This paper report presents the general case history for the construction of the North Outfall Replacement Sewer (NORS) project. A 1982 City of Los Angeles Wastewater Facilities Plan recognized that a 19.3 km (12¬mile) section of the North Outfall Sewer (NOS), from Culver City to the Hyperion Treatment Plant, was deteriorating and had insufficient capacity. Built in the 1920s of unreinforced concrete and lined with vitreous clay tiles, the originally unventilated NOS suffered considerable corrosion by H2S gas. This caused roof tile and concrete to fall to the bottom of the pipe, which, over the years, reduced the 11.9 M3/S (420 cubic feet per second (cfs)) design capacity of the NOS to a current level of approximately 8.5 M3/S (300 cfs). This worsening condition, thereby contributing to the polluting of Santa Monica Bay. This situation culminated in the State of California Regional Water Quality Control Board, Los Angeles Region, issuing a Cease and Desist Order on January 27, 1986, which required the City of Los Angeles to eliminate such dry weather overflows through a number of prescribed actions. One of these was to construct the North Outfall Replacement Sewer (NORS) by April 1, 1993. PROJECT DESCRIPTION The NORS Project involved the construction of over eight miles of tunnel and related structures. It consisted of a main tunnel conduit to convey sewage to the Hyperion Treatment Plant and four diversion lines connecting the main tunnel to major existing

Jan 1, 1993

By G. G. O. O. Uwadiale, M. A. U. Nwoke

The recovery of iron from low-grade (31.6% Fe) Muro iron ore was investigated using reverse anionic flotation of quartz. Tests were performed to investigate the effect of pH, the effects of using distilled vs. tap water as the flotation medium and the effects of using cassava starch vs. dextrin for the depression of iron minerals. The optimum results were obtained around pH 10 and 11, and cassava starch was superior to dextrin in depressing iron oxides. Investigations at pH 11 using cassava starch and distilled water resulted in a concentrate grade of 69.4% Fe with a recovery of 84.4% Fe. Under the same conditions, the use o f dextrin produced a concentrate grade q f 65.9% Fe with a recovery of 78.6% Fe. Investigations at pH 10 using cassava starch and distilled water resulted in a concentrate grade of 60.2% Fe with a recovery of 83.9% Fe, and under these conditions the use of dextrin produced a concentrate grade of 56.1 % Fe with a recovery of 81.5% Fe. Flotation studies conducted at pH 11 using tap water and cassava starch resulted in a concentrate grade of 70.4% Fe with a recovery of 92.7% Fe. This demonstrated the superiority of tap water over distilled water in the reverse anionic flotation process. Investigating the effect of calcium ions on the activation and flotation of quartz with oleic acid revealed that a calcium ion concentration of 50 to 100 ppm was optimum.

Jan 1, 1996

By R. C. Gilmore, A. Juganda, Jr. Bogin, S. Saki, S. A. Lolon, J. F. Brune, Jr. Zipf, J. W. Grubb

"Methane-air mixtures form explosive gas zones (EGZs) in underground longwall coal mine gob ventilation systems. Such EGZs have caused a number of fatal mine explosions, including the 2010 disaster at the Upper Big Branch mine, where 29 miners lost their lives. Researchers at the Colorado School of Mines use Computational Fluid Dynamics (CFD) modeling to predict EGZ locations. In this paper, a model of a bleeder-ventilated gob system includes two adjacent panels, an active longwall and a mined-out gob. Boundary conditions for the simulations match statutory ventilation conditions for methane concentrations and air flow rates at common measurement points throughout the mine. The modeling results predict the persistence of EGZs in the gob that may compromise the effectiveness of the bleeder ventilation system. INTRODUCTION Mine operators must maintain statutory air flow and air quality requirements at various locations throughout the mine ventilation network. U.S. regulations generally require mine operators to establish bleeder ventilation systems surrounding longwall gob areas (30 CFR §75.334.) Gilmore et al. (2015) documented that explosive gas zones (EGZs) exist within bleeder-ventilated gobs and may extend into the bleeder entries that must be examined weekly, presenting an immediate fire and explosion hazard to mine examiners. In a project funded by the National Institute for Occupational Safety and Health (NIOSH) under contract 200-2009-31409, researchers at Colorado School of Mines have developed a tool for modeling gob air quality using Computational Fluid Dynamics (CFD). This CFD model is capable of examining a wide range of longwall mine ventilation systems. Marts et al. (2014a, 2015) studied progressively sealed gobs with U-type ventilation and using a back return. Saki et al. (2015) examined the impact of gob ventilation boreholes and face air quantity effects on tailgate methane concentrations. Gilmore et al. (2014) modeled bleeder-ventilated gobs and Gilmore et al. (2015) studied the effects of bleeder regulator settings on EGZ volume and locations in the gob. CFD studies for progressively sealed gobs were generally carried out using 5000 ft (1500 m) panels because the gob atmosphere remains largely unchanged from about 500 m (1,500 ft) inby the face. A modular meshing methodology (Gilmore, 2015) permitted researchers to extend bleeder models to 10,000 ft (3050 m) to include the full extent of bleeder entries. This paper further expands the bleeder CFD model to study the gob atmospheres of two adjacent longwall panels, each 10,200 ft (3100 m) in length, with the active face located 6,700 ft (2040 m) from the startup room."

Jan 1, 2016

Western Energy Co.&apos;s Rosebud mine in Colstrip, MT, has implemented innovative mining techniques in the past,* and following in that tradition, new techniques for overburden removal were recently introduced at the Montana coal mine. In April 1979, falling coal production left Western Energy in a bind. At that time the company was one of the first to implement a dozer/scraper stripping system for additional stripping capacity. This time around, however, the challenge was to eliminate spoil rehandle with a dragline while digging down 33 in. "Rehandle will increase 20% by going down an additional 3 m past the 24-m level," according to Bob Cope, mine supervisor for Western Energy. To cut this additional expense by eliminating the need to rehandle, Western now uses scrapers to bench down for the dragline. This system utilizes the flexibility of scrapers as earthmoving tools. Although scrapers were originally purchased to fill an immediate need for productive, low-cost overburden stripping, they are now benching and continue to optimize the overall effectiveness of existing draglines. According to the company, the draglines couldn&apos;t keep up with production requirements if they had to rehandle spoil material. Benching with the scrapers eliminates spoil rehandle. Western Energy expects to produce 10.6 Mt of coal this year from three active pits in the Rosebud mine. Western estimates that there are enough coal reserves at Rosebud to support a 40-50 year operation. The sub-biturninous coal has an energy value of 9.2-9.4 Mj, 0.87% sulfur, 8% ash, and 20% moisture. Pit A, approximately 610 m long and 45-55 m wide, is the mine&apos;s greatest producer. A Marion 8050 dragline with a 46-ms bucket digs 30-33 m to uncover the 7.3-m Rosebud coal seam. Here, the scrapers were put to work benching for the dragline. Long Construction Co., which operates the mine equipment for Western Energy, uses four Caterpillar 637 tandem-powered scrapers, two Cat 631s and two Terex TS-24s. The number and combination of scrapers actually used for benching periodically changes, depending on which ones are needed for topsoil removal and reclamation. Scrapers benched down about 9 m and removed nearly 2.3 Mm of over-burden at pit A. About 1 Mm was placed in a 9-m lift on top of the high-wall at the head of the pit. The scrapers built up one side of the lift first and then the other on successive passes to

Jan 2, 1981

By Rudolf E. Greuer

Introduction The Institution of Mining and Metallurgy and the Institution of Mining Engineers organized the Third International Mine Ventilation Congress held in Harrogate, England. Sixty-one papers were presented and about 300 participants registered. About 30% of the participants came from Great Britain, 20% from South Africa, and about 5% each came from Australia, Canada, China, France, US, and West Germany. The remainder of the participants came from 18 other countries. Eleven of the papers presented dealt with methane, three with diesel exhausts, four with dust, four with radioactivity, 18 with heat, and 15 dealt with main and auxiliary ventilation. Six of the papers dealt with mine fires, which is a boundary region between main ventilation and gas concerns. Ventilation Congress proceedings are available. See ME, December 1984, page 1687, New Books page. With CH4 and radioactivity topics, the Australians dominated since they are becoming large coal and uranium producers. With diesel exhausts, the most important problem in highly mechanized mines, the North Americans were prominent. The South Africans, working in 90% SiO2 in their gold deposits, were preeminent in presentations on dust. They also led in topics related to heat, joined by the West Germans whose mines are getting deeper. In main and auxiliary ventilation, Great Britain and West Germany provided the majority of contributions. Methane The large number of methods for the precalculation of CH4 liberation in longwall mining all contain three elements: gas content of the coal, gas emitting zone (or influence zone of face), and degree of gas emission. Determination of the gas content, commonly accomplished by taking coal samples, does not pose much difficulty. But this is not the case with the other two factors. Most existing approaches rely on rock mechanics observations. Some only rely on intuition or speculation. The West German coal mines conducted a large research project between 1977 and 1982, in which gas pressures around longwall faces were measured. Gas pressures and gas contents can be related. Therefore, influence zone and degree of gas emission can be determined simultaneously. The precalculation of CH4 liberation in room-and-pillar mining is simpler. Since foot and hanging walls remain essentially intact, gas pressure distribution and gas flow can be calculated using hydraulics equations. A key to these calculations is permeability. A group of Australian researchers reported that the permeabilities of rock under three dimensional stress differs from rock under destressed conditions. This fact was known, but no systematic observations existed. These were provided. Methane drainage has been practiced for more than 40 years. It is used with great success in longwall operations. Some of the West Germans think that methane drainage from the footwall is neglected. Under certain geological conditions, they claim this is as important as drainage from the hanging wall. A research project was presented in support of this claim. Considerable efforts have existed for more than 10 years to use methane drainage in room-and-pillar mining. A paper described the accomplishments of Consolidation Coal Co. Since coal and rock are not fractured as much as in longwall mining, gas transport takes place by the slow diffusion from micropores into the cleats. Then, it is transported by laminar flow along the cleats until a drainage borehole or the mine workings are reached. Boreholes about 300 m (1000 ft) long can reduce the gas content of a band of coal 100-m (330-ft) wide by 50% in one year. A paper from India described methane drainage from gob areas. A French paper reported that especially high methane concentrations can be drained from gob areas if the face ventilation is descensional. Methane buoyancy and ventilating pressures compensate each other, and little air dilution by leakage takes place. Another French paper described a newly developed gas and air velocity monitoring system making use of a microcomputer. Four papers on methane dealt with gas outbursts. Two Australian contributions described the problems encountered with CH,-CO2 mixtures and the attempts to solve them with drainage boreholes having 10 to 20-kPa (1.5 to 3-psi) suction pressures. A Hungarian paper described problems and attempts to reduce the risks of gas outbursts in Hungary. Prediction of gas outbursts in coal seams was the focus of an English paper. Observations show that they are frequent when coal, through its geological past, has become soft and brittle with a resulting increased desorption rate. An instrument for gravimetric de-

Jan 1, 1985

By Aladar B. Dory

INTRODUCTION Practical experience in mines with known presence of radon daughters in the mine atmosphere in Canada and elsewhere shows that a very high concentration builds up in an unventilated dead end heading. As Holaday et al1 observed, even a minimal air movement results in a drastic reduction in radon daughter concentration. It is therefore obvious that the main objective of radon daughter control in the working environment is to design the ventilation system providing an optimized flow of fresh air into the workplace, resulting in acceptable climatic conditions and achieving radon daughter concentrations resulting in exposures as low as reasonably achievable. BASIC OBJECTIVES Large mining companies, having extensive material resources and professional expertise, have utilized elaborate electrical modelling in the design of mine ventilation systems as early as 1950 (coal mining industry in Europe) and with the advance of computer modelling techniques, their utilization in ventilation systems design is on the increase. Unfortunately, these methods are usually not available to small mining companies and even the large companies might not achieve the fullest benefit from utilizing them, if proper limiting factors are not considered in the modelling. When an evaluation of a ventilation system of a mine is undertaken in literature, a measure of the amount of air supplied underground per one ton of ore mined is used as an indicator of the efficiency of the ventilation system. Yet, even the greatest amount of air forced into the mine might not result in an acceptable working environment if a proper distribution of this air into individual working places is not achieved. The volume and the age of the air are probably the two most important factors in achieving acceptable radon daughter concentrations in the workplace, but other factors also have to be considered. DIRECTOR MINE - ALCAN, NEWFOUNDLAND FLUORSPAR WORKS ST. LAWRENCE, NEWFOUNDLAND, CANADA Ventilation To illustrate the effects of the design of the ventilation system on the control of radon daughter concentration, let us review the gradual development of the ventilation system of this mine from the earlier years of its development up until its final years of operation. This mine, located near the community of St. Lawrence on the south coast of Burin Peninsula was developed in the late thirties and reached full production by 1942. Unfortunately as was customary at that time, the only source of ventilation was a natural draft. The mine was extremely wet, and no significant attention was initially given to possible health effects of dust. It was not until the mid-fifties, when a number of cases of silicosis had surfaced, that de Villiers and Windish2 observed a significant increase of lung cancer incidence among the miners in comparison to its incidence among the general population of Newfoundland. Suspicions regarding radiation as a cause of the lung cancer were expressed, but it was only in surveys taken in late 1959 and early 1960 that Windish3 and Little4 established the presence of radon daughters in the mine atmosphere in very high concentrations. Windish, de Villiers and Hurley suggested that the most likely source of the radon in the mine was the mine water which dissolved radon during its passage through the granitic country rock in the surrounding geological area. This conclusion was confirmed by analyses of water from various areas of the mine by the Atomic Energy Canada Limited laboratories. The radon values in the samples varied from 4,240 to 12,850 pCi/L5. Following the discovery of the presence of radon daughters in the mine, the company took speedy action to install mechanical ventilation for the mine. The system was not designed as a total unit, but fans were installed rather on a trial and error basis. The basic system installation began in March 1960 and was completed by 1962. It remained basically unchanged with only minor modifications until August 1973 when a wholly new, redesigned ventilation system was implemented. A schematic section of the mine and its ventilation system for the period prior to March 1960 is given in Figure "A", for the period 1960-1973 in Figure "B", and for the period after August 1973 in Figure "C". The ventilation system prior to 1960 is not known. All workings of the mine were ventilated only by natural ventilation. If any measurements of airflows at different or any times of the year ever existed, no records have been preserved. The very minimal natural ventilation was augmented by "blowing" air from compressed air supply lines and exhaust air from drills. It is known that the compressor capacities of the mine were limited and therefore no significant air movement was probably created by the "blowing".

Jan 1, 1981

By R. E. Utting

INTRODUCTION Until recently, gamma doses had been largely ignored in Ontario uranium mines. This has been due to the assumption that these doses are small and have been more or less unchanged with time and hence their effects have been included automatically in the epidemiological studies that led to the establishment of radon daughter exposure limits. This assumption had to be challenged for two basic reasons. The first was that radon daughter exposures to miners have been progressively reduced over the years due to improved ventilation and ever more stringent regulations, while gamma exposures have presumably remained relatively unchanged. Therefore it must be assumed that the ratio of gamma to radon daughter exposure has gone up. The second reason is more philosophical. It is clearly inappropriate to make judgements on the significance of a potential industrial hazard when the magnitude of that hazard has not been fully assessed. Having decided that some sort of assessment of gamma exposures to uranium miners must be made, it was than necessary to determine how this should be done. Several options were available, for instance: (i) Wholesale personal gamma dosimetry for all mine and mill workers, (ii) Personal gamma dosimetry only for those workers suspected of receiving the higher doses, coupled with area monitoring to estimate the exposures of other workers, (iii) Area monitoring coupled with dose rate times time calculations for all. This would correspond to the generally prevalent method of assessing radon daughter exposures. It was argued that since radon daughter exposures are the major radiological hazard in uranium mines, to invest resources for assessing a lesser hazard to a greater degree of precision was not cost effective. (iv) Since gamma dose rate is related to ore grade, individual doses could be assigned from knowledge of work location and ore grade. Before deciding which of these options would be most appropriate, it was necessary to have some idea of the magnitude of the problem. Very few data were available in the literature and with the exception of a few spot dose rate measurements, and the results of a few gamma dosimeters issued to selected individuals by some of the mining companies, nothing was available. A rule of thumb of obscure origin is often quoted within the industry indicating that gamma dose rates underground will be about 0.25 mR/h per lb/ton or 5 mR/h per % U. This had been used by some to justify neglecting gamma radiation at least for ore grades of the order of 0.1% or 2 lb/ton, on the grounds that gamma dose rates would be of the order of 0.5 mR/h and therefore give rise to annual doses of only about 10 mSv (lrem). That is, it was assumed that gamma radiation was of limited concern compared to the hazard associated with the inhalation of radon daughters. We were thus faced with the situation of just assuming that no regulatory limits were being breached. This situation could not be allowed to continue. A program was initiated to investigate the gamma doses absorbed by uranium miners in three mines in Ontario, and extensive gamma surveys were conducted in the Quirke 2 mine of Rio Algom Ltd, Elliot Lake; Denison Mine, Elliot Lake; and Agnew Lake Mine, Espanola. Negative reaction was received from several mine company officials to the possibility of all miners being required to wear personal gamma dosimeters due to the logistical difficulties involved, and therefore part of the project was aimed at determining if a reliable correlation between gamma dose rate and ore grade in the work location could be deduced, in order that dose rate times time calculations might be used for gamma dose assessments. The results of these programs provided evidence that the gamma dose for some employees in the three mines investigated may be a significant fraction of the current maximum permissible annual dose of 50mSv (5 rem). When combined with radon daughter exposures in the manner recommended by the ICRP at their 1980 Brighton meeting (ICRP 80) the results indicated that some individuals will come close to the resulting limit and may even exceed it. The results also indicate that is probably not feasible to develop a reliable formula for

Jan 1, 1981