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By Steven J. Schatzel, James Addis, K. Vaibhav Raj, Berk Tulu

The National Institute for Occupational Safety and Health (NIOSH) is conducting research to enhance methane control and mine ventilation practices at coal mines, as a method to mitigate hazardous work conditions. NIOSH researchers conducted case studies to investigate the effectiveness of existing systems and improve performance to decrease the potential for unplanned or excessive methane emissions. At one mine, drainage was influenced by hydrofracking inefficiencies where two of the six wells showed good frack development. Methane content was variable, from 300 ft3/t to 600 ft3/t. This paper also summarizes overburden movement during undermining and the estimated quantities of gas transported to the mine using FLAC3D and CFD modeling. The findings from this research effort highlight considerations based on specific mine conditions that may also be applicable to most longwall mines utilizing methane control systems.

Feb 1, 2025

By Hongsheng He, Xiaofang Lv, Jianfeng Wang

This paper provided an overview of characteristics of limonite ores as well as their influence on sintering performance. Evaluation experiments focusing on basic characteristics, granulation characteristics, and high temperature sintering characteristics had been conducted by using two types of limonite ores, then the principles of high effective utilization technology of limonite ores were discussed, and the sintering pot tests were conducted. Limonite ores possessed low total iron content, high combined water content, high SiO2 content, large average particle size, low adhering particles proportion, high porosity, high water absorption rate, short water absorbing time, poor granulation ability, high melt formation ability, low melt fluidity ability, low melt penetration ability, high melt absorption ability, low bonding phase strength, and high forming ability of SFCA. By taking the technical measures of ore-proportioning optimization, increasing the proportion of burntlime and segmentation granulation, the usage of limonite ores could reach 50% in sintering.

Aug 27, 2021

By Milos Krofta, Donald B. Aulenbach, James P. Van Dyke

INTRODUCTION An innovative Ø.1-MGD (378.5-m3/day) biological and phyicochemical wastewater treatment system at Hancock, Massachusetts, USA, has been treating the domestic sewage generated from the four-season Jiminy Peak timeshare resort since 1985. The influent raw sewage is collected and equalized in an equalization tank, then biologically treated by two rotating biological contactors (RBC). A portion of the RBC effluent is discharged to a secondary sedimentation tank to simulate conventional waste treatment by a conventional RBC system. For parallel treatment the remaining portion of the RBC effluent is discharged to an innovative Sandfloat (SASF-8) clarifier to simulate a two-stage biological and physicochemical treatment by RBC-Sandfloat. Both secondary sedimentation effluent and Saandfloat effluent are later combined together and discharged to a tertiary sand filter and two UV disinfection'units for final polishing. The operation and performance data of Jiminy Peak plant in the period of May 1985 to March 1988 was fully documented in the Lenox Institute technical report No. LIR/Ø3-88/292 (391 pages, March 1988). This report presents only background information and complete plant data in the period of April 1988 to January 1989. Dr. Lawrence K. Wang and Mr. James P. Van Dyke have been the Consulting Engineer and Director of Special Projects, respectively, for the Jiminy Peak Project since 1985. (Key Words) Innovative and Alternative Technology, Rotating Biological Contactor, Sandfloat, Dissolved Air Flotation, Sedimentation, Tertiary Filtration, Ultraviolet, Disinfection. BACKGROUND INFORMATION Jimmy Peak, Inc., Hancock, Massachusetts, has designed and constructed a secondary wastewater treatment system consisting of preliminary/primary treatment facilities, two rotating biological contactors, a dissolved air flotation clarifier, a secondary sedimentation clarifier, a sand filter and two ultraviolet (UV) disinfection units. Total average wastewater flow is approximately Ø.1 ) (378.5 cubic meters per day). In accordance with the Consulting Engineer's preliminary design, the total wastewater stream is initially collected and equalized by an equalization tank, and then biologically treated by rotating biological contactors (RBC). Part of the RBC effluent is diverted to a Krofta Sandfloat (SASF-8) for physicochemical clarification in order to simulate an innovative RBC-Sandfloat two-stage waste treatment system. The remaining part of the RBC effluent is diverted to a sedimentation tank for secondary clarification in order to siumulate a conventional secondary RBC-sedimentation waste treatment system for the purpose of parallel comparison. Both Sandfloat effluent and sedimentation effluent are later combined together and discharged to the subsequent sand filter and UV disinfection unit for final polishing. Figure 1 shows the flow diagram of Jiminy Peak Plant. The Massachusetts Clean Waters Act (M.G.L.c.2l,s.26-53) was amended by Chapter 246 of the Acts of 1973 to authorize the Massachusetts Department of Environmental Quality Engineering (DEQE), Division of water Pollution Control, Boston, Massachusetts, to regulate discharges into all waters of the Commonwealth of Massachusetts, including groundwaters. The DEQE regulates discharges through the issuance of discharge permits which impose limitations on the amount of pollutants which may be discharged in the effluent, together with monitoring and reporting requirements and other conditions to insure adequate treatment of all liquid wastes prior to discharge. The DEQE reviewed the application submitted on behalf of Jiminy Peak, Inc. for a permit to discharge sanitary wastes from the Jiminy Peak secondary wastewater treatment system located off Brodie Mountain Road, Hancock, Massachusetts, and subsequently developed the conditions contained in the Groundwater Discharge Permit No. Ø-188. More specifically, Jiminy Peak, Inc. (i.e. the permittee) has been authorized to discharge into the ground from the Jimmy Peak secondary wastewater treatment facilities for which the Permit No. Ø-188 was issued, a treated effluent

Jan 1, 1989

By A. Bhattacherjee

Industries in general and the mining industry in particular have become more safe than ever before. However, there is a need to increase the search for the causes of mining work-related injuries and to contribute to their reduction through job design, work procedures, and worker training. Mine accident and hazard control has to be practiced in a manner which is consistent with the overall objectives of the mine management, among which are the optimum allocation and effective utilization of the resources. The purposes of this paper are to (i) review the role risk analysis in safety management with references from mining and other industries; (ii) discuss the risk measures appropriate for injury analysis; (iii) present a set of analytical tools to injury experience data in under-ground coal mines; (i v) apply these tools on case study example; (v) discuss the results of the case study; and (vi) amplify the application limitations and the future research needs.

Jan 1, 1991

By Isa Nozari, Asghar Azizi

The extraction of copper from sulfate leaching liquor was investigated using Acorga M5640 commercial extractant in kerosene diluent. The effects of important factors such as the extractant concentration, contact time, liquor pH, and temperature were examined on the separation factor Cu/Fe and the extraction efficiency of copper. The results indicated that nearly 92.32% copper was extracted from the leach liquor at 10% (v/v) Acorga M5640 concentration, 1:1 aqueous/organic phase ratio, 20 min contact time, pH 2.5, and 25 °C temperature. Also, the further increase in the values determined had no positive impact on the extraction of copper. Under these conditions, the separation factor Cu/Fe was obtained about 29.64. The extraction mechanism was studied applying the plot of log(D) – 2 pH against log [extractant] and it proved the association of two moles of Acorga M5640 for the extraction of one mole of copper. McCabe–Thiele diagram was constructed and a single-step extraction process was found to obtain themaximumextraction. Additionally, thermodynamic datawere investigated and the extraction process was endothermic and non-spontaneous at 298 K (ΔH = 27.05 and ΔG = 15.52 kJ mol−1). Moreover, the loaded organic phase was stripped and 80.94% copper was recovered after 20 min using 10% sulfuric acid concentration.

Jul 31, 2020

By J. Hart, P. W. Conrad, S. Rosenthal, R. A. Hogan

The Mining Engineering Department at Montana Technological University (Montana Tech) has a new Mine Safety and Health Administration (MSHA) approved program for Part 46 and Part 48 mine health and safety training. The training program will utilize the facilities and faculty from Montana Tech’s Mining Engineering and Safety, Health, and Industrial Hygiene departments to provide superior mine health and safety training to the mining industry. The faculty of each department includes professionals with experience in safety training and the mining industry. The training program incorporates adult learning techniques and extensive hands-on activities in all training modules. The instruction will exceed MSHA requirements. The resources available on Montana Tech’s campus will be utilized to strengthen the training program providing the mining industry with a world class mine health and safety training program.

Feb 1, 2023

By A. A. Zatout, G. H. Sedahmed, T. M. Zewail, M. H. Abdel-Aziz, B. S. Ibrahim, E-S. Z. El-Ashtoukhy

In this work, the possibility of intensifying the cementation process in stirred vessels was studied by using the perforated baffles as turbulence promoters; the cementation reaction of Cu+2 on Zn was selected as a model. Zn was used in the form of a sheet lining the inner wall of an agitated vessel. Parameters considered are initial concentration of Cu+2, solution temperature, impeller rotation speed, and impeller geometry. Within the present range of parameters studied, the rate of cementation was found to increase with increasing initial concentration of Cu+2, solution temperature, and impeller rotation speed. At the same rotational speed, a flat turbine gave a higher rate of cementation than a 45° pitched turbine. The present suggested that baffles’ geometry gave superior performance in terms of mass transfer of copper cementation compared to the conventional baffle geometry. All parameters that affect the rate of cementation were correlated in the form of dimensionless equations suitable for the design and performance prediction of the agitated vessel. Potential applications of the obtained results in the field of hydrometallurgical processes and wastewater treatment were discussed.

Jan 12, 2021

By D. R. Spears

The U.S. Bureau of Mines has developed a Fast Air-injected hydrocyclone-Shallow Tank (FAST) flotation system. Conventional flotation cell kinetics are a compromise between two conflicting requirements, the need for an agitated zone for rapid bubble-particle collision, and a quiescent zone for bubble-pulp separation. The Bureau of Mines has designed a flotation system in which bubble-particle collision and bubble separation are carried out in two discreet unit operations. A bubble slurry was mixed with an ore slurry at a high rate in a modified air-injected hydrocyclone for rapid bubble-particle collision. The mixture was then fed into a unit specially designed for rapid separation of the bubbles from the pulp. A conventional flotation cell required 2.8 min to obtain 90 pet recovery at a 7-pct Cu rougher concentrate grade, whereas the FAST flotation system can obtain 85 pet recovery with a 13-pct Cu rougher concentrate grade in 0.28 min, and 92 pet recovery with a 7-pet Cu grade in 0.56 min. The Bureau's FAST flotation system is adaptable to a wider variety of flow sheets and could potentially allow for replacement of conventional flotation cells with smaller and faster units.

Jan 1, 1991

By Samuel A. Rasmussen

Lead anodes are used in commercial copper electrowinning operations, yet there is no method to calculate anode life. To be discussed in this paper are operating characteristics of rolled plate (Pb,Sn,Ca) alloy anodes at Morenci's electrowinning tankhouse along with fundamental operating conditions and how the anode reacts to these conditions. These include anode maintenance, electrolyte temperature and viscosity, cobalt concentration, and manganese concentration. Also to be discussed are anode substrate characteristics under normal operating conditions and how the corrosion rate is affected by changes in operating conditions.

Jan 1, 1995

By Y. Lahaye, O. Martinsson, B. I. Palsson

"Four main and three minor rare-earth-element (REE)-bearing minerals were identified and quantified in the Kiirunavaara apatite iron ore tailings using optical microscopy, an electron probe microanalyzer (EPMA) and a mineral liberation analyzer, and their chemical compositions were analyzed by the EPMA and laser ablation inductively coupled plasmamass spectrometry. REEs are shown to be contained in the minerals apatite, monazite, allanite, titanite, zircon, thorite and synchysite. In zircon, thorite and synchysite, they occurred in only trace amounts and contributed limited amounts to the total REE budget, and these are consequently of minor importance. Monazite occurred as inclusions in apatite and as free particles, 90 percent liberated. Allanite occurred to some degree in mixed grains with magnetite but also as free particles. Monazite mainly reported to the apatite concentrate, while allanite and titanite largely went to the tailings, the latter preferably to those fractions smaller than 38 µm. The amount of titanite in the finest tailings fraction was 2.3 weight percent, containing close to 1 percent REEs, with heavy rare earth elements (HREEs) making up 28 percent of the total REEs. However, a texturally distinct group of titanite grains showed an HREE/REE ratio of up to 67 percent. Furthermore, titanum dioxide analyses indicate that titanite is preferentially released into the tailings from the secondary magnetic separation step in the concentrator. Our data therefore suggest that titanite, occasionally enriched in HREEs, can be extracted from the processing stream and might thus be considered a new source for REEs at Kiirunavaara and similar deposits.IntroductionDeposits of apatite iron ores are the major source for iron in the European Union and are exploited by Luossavaara-Kiirunavaara AB (LKAB) in several mines in the Kiruna-Malmberget area in northern Sweden. Kiirunavaara is the largest of the apatite iron orebodies in Sweden, comprising more than 2 Gt of iron ore with 60 to 68 percent iron (Fe). At the end of 2015, the total accumulated production of crude ore from openpit and underground mining since the start of mining was 1,566 Mt, with ore reserves of 644 Mt grading 46.3 percent Fe (LKAB, 2016). The deposit defines a steeply dipping tabular body, approximately 6 km long and up to 180 m thick, extending at least 1,500 m below the surface. It is situated in the middle part of the Palaeoproterozoic Kiirunavaara Group of volcanic rocks and follows the contact between a thick pile of trachyandesitic lava and overlying pyroclastics (Martinsson, 2004)."

Jan 1, 2017

By A. L. Hinde

The recovery of heavy minerals by gravity concentration during grinding is practised widely in the minerals industry. Studies have shown that the grade and recovery of the concentrate depends not only on the performance of the concentrator but also on the degree of liberation achieved in the mills and the scavenging action of the hydrocyclones closing the grinding circuit. In the current study, consideration is given to the development of mathematical models to quantify and unify the operations of comminution, liberation, classification and concentration. The models are based on test work done at production scale with gold ore.

Jan 1, 1985

By Ali Ekrem Aritan

Exposure to hand-arm vibration (HAV) among the labor force in the mining industry can result in permanent vibration diseases such as behavioral disorders, visual impairment, osteoporosis, pain (mostly in the hands, wrists, and shoulders), circulatory system disorders, and malnutrition. In this study, the vibration exposure of excavator operators working regularly in 25 different natural stone quarries was measured. In addition, the relationship between vibration exposure and rock properties was investigated. Gloved and ungloved vibration measurements were performed and revealed that antivibration gloves reduce excavator operators’ vibration exposure by 7.10–25.95%. Samples from each quarry were characterized with rock mechanics tests. The correlations between the hand-arm vibration measurements and the natural stone physical–mechanical properties (uniaxial compressive strength, Shore hardness, Knoop hardness, apparent density and total porosity) were determined to be over 70%. One finding of the study was that the health risk in excavators cannot be reduced sufficiently by the sole use of antivibration gloves. Furthermore, the excavation of any rock with a Knoop hardness above 121 is associated with an 81% risk of HAV. To increase awareness of the danger of vibration exposure for natural stone miners, training programs should be provided, and technical and operational changes should be implemented.

By Mohammed Laraba, Rodrigo Alvarez Garcia, Juan Maria Menendez-Aguado, Roula Abdelmalek

The aim of this research is the study of the WHIMS parameters on the improving process of a local (DD: Djebel Debbagh, East of Algeria) kaolin. The obtained results of performed XRF, SEM, and XRD characterization showed that this local material is an alumino-silicate containing kaolinite and halloysite with MnO and Fe2O3 impurities (respectively 39.30% and 34.67% alumina and 38.18% and 39.70% silica in both samples (DD2 and DD3). After homogenization, crushing, and milling, a multiparameters factorial experimentation (grain sizes, intensity, and dimension of matrix balls) showed that WHIMS is an effective process to improve the quality of this kaolin, mainly: removing their impurities. Dimensions of matrix balls (Db) = 2 cm, intensity of the applied magnetic field (I) = 1.5 T, and grain sizes (Dp) = > 100 < 250 μm) are the optimal values to reach 91.06% for manganese and 93.45% for iron as removing yield. The statistical analysis of these results sorted the investigation parameters and showed that Db has the highest and most significant influence on the recovery of iron and manganese particles. The magnetic field intensity (I) has a less important influence while the grain size has an average influence (positive on the iron recovery but a negative one on the recovery of manganese).

Jun 11, 2022

By A. Nieto

This article is a review of previous discussions about underground mining method selection and classification with the purpose to clearly define and classify the two groups of variables (indicators) involved in the mining selection process: key deposit indicators (KDIs) and key mining method indicators (KMIs). This paper provides a simple and clear approach for the classification and selection of a feasible underground mining method based on a predetermined set of favorable deposit characteristics, which are defined as key deposit indictors (KDIs). This article also gives a summary of the key advantages and disadvantages associated with every underground mining method based on their key mining indicators (KMIs). The selection method given in this paper is based on defining field KDIs and comparing them to the KDIs that are favorable to every considered mining method. By cross validating a matrix of favorable KDIs, the paper presents a simple approach to rank several underground mining methods using a scale from favorable to less favorable accordingly to the ore deposit characteristics (KDIs). KMIs are used to further complement KDI rankings by analyzing every method’s KMI performance based on the expected productivity of the mining operation being considered. To further assist the reader through the selection process, this paper gives basic sketches representing each of the primary underground mining methods discussed in this paper. Each method is given using the same 3D isometric view as spatial reference to assist the reader with visual interpretation and comparison between methods.

Jan 1, 2010

By Per-Martin Sandgren

LKAB&apos;s ores have specific mineralogical properties that make them especially suitable for the production of super-grade concentrates. Conditions are particularly good for this purpose at Malmberget, where a concentrate for high-purity sponge iron production has been produced for many years. Until 1978 the approximate assay was 71.2 % Fe and 0.45 % SiO2. Since then the assays have been improved by process modifications to approximately 71.5 % Fe and 0.20 % SiO2. Scope exists for still further improvement. In recent years the production of fractioned concentrates has been started. The material lies within the size range 74-125 µm and the approximate chemical analysis is 71.6 % Fe and 0.30 % SiO2. Other special products are under discussion. The Malmberget ores are also suitable for the production of high-purity direct reduction pellets with a SiO2 content below 1 %.

Jan 1, 1981

Geostatistical literature has been dealing with recoverable reserve estimation techniques for the last 30 years. There have been well established support correction techniques (Le. affine correction, indirect log-normal correction, etc.) to deal with volume -variance relationship globally in determining what will be the reserves above a cutoff grade during mining. Kriging has become the accepted technique in the mining industry for ore reserve estimation. Many large companies are using this method for gold and copper deposit modeling. In general, the estimates coming from kriging are smooth. The variances of the block grades estimated by kriging are often lower than the material that will be mined selectively. As a result, kriging tends to under estimate reserves at high cutoff grades and overestimates them in low cutoff grades. It is necessary to perform additional steps after kriging if the goal is to determine recoverable reserves at some cutoff and not the in situ resources. In the mining industry, it is becoming more and more common for the low grade heap leach projects to consider whether the higher grade portion of a deposit may be processed by a mill at higher profits. Similarly, a concentrator -refinery operations are implementing heap leach -SX-EW option for the lower grade portion of their deposit. These options are making it very critical to estimate the low and high end of the grade distributions more accurately. Ore reserve estimation is normally done by using widely spaced exploration data which has very small size support (Le. core samples). Selective mining is carried out on ore control that is based on closely spaced blasthole data normally averaged into larger polygons based on some cutoff grade. Estimated grades from exploration data need to reflect the variability of the grades of the selectively mined units if they are used without further adjustments to predict recoverable reserves.

Jan 1, 1997

By Gary Peach, Khalid Saif Al-Khayareen, Vigil Fernandez, Furqan Hameed

The Musaimeer pump station and outfall project (MPSO) is in Doha Qatar. This technical paper will discuss and explain the three major stages of geotechnical investigation employed. Stage one the tender geotechnical data and how this was used to design the TBM and identify various areas of risk. Stage two was the selection of supplementary geotechnical investigation methodologies. Stage three was the geotechnical investigation by specialist equipment installed on the TBM and providing further data on a real time basis as the TBM was advancing. Analysis of all the data will be provided along with the lessons learned.

Jun 23, 2024

By J. R. Dyni

Colorado nahcolite deposits, estimated at 29 billion metric tons, formed in an asp metric permanent-lake basin whose waters were rich in algae and sulfate-reducing bacteria. Nahcolite (NaHCOg) occurs in oil shale of the Green River Formation (Eocene) as aggregates (62 percent), disseminated crystals (24 percent), and impure beds (14 percent). Because much of the nahcolite is nonbedded, it will be mined with oil shale and can be concentrated to marketable grade by crushing and screening Impurities in nahcolite are kerogen, quartz, pyrite, and carbonate minerals. A promising new market for nahcolite is in the process of removal of sulfur dioxide from industrial stack gases. When heated, nahcolite reacts with sulfur dioxide to form sodium sulfate. Differences in thermally induced microporosity may explain why nahcolite is more reactive than trona, a possible competitor with nahcolite for SO2 removal. Limited production of nahcolite at one or possibly two localities is expected in the next several years, but large-scale production is probably at least 10 years away.

Jan 1, 1980

By Ingemar Marklund

INTRODUCTION Today at LKAB, the primary mining method used is sublevel caving, independent of ore-body size, shape, dip, and geographical location. The mining of small ore bodies with relatively flat dip involves a lot of develop¬ment work, a large initial investment, big dilution, and large ore losses. In 1975 serious investigations into other possible methods, especially for those small ore bodies which do not lie in direct contact with other ore bodies, began. Different methods were studied and a number of in¬teresting ideas came from Russian mines. An analysis was made of one of the proposed methods, sublevel benching, in which mining is carried out from raises, and which uses the throw from the blast to place the broken ore. However, there were reservations about working in big open rooms at great depth, and doubts about the possibility of getting acceptable drilling results in an acceptable environment from raised platforms. Testing was therefore delayed for a time. However the investigations continued and in the fall of 1976, it was decided to try vein mining in a small ore body (Indian) (see Figs. 1 and 2). MINING RESULTS The results are now available from the mining of Indian and are as follows: Ore plan area 1200 m2 Level height 50 m Ore volume 60 000 m3 Ore tonnage 270 000 t Iron content 62% Dip 0.69 rad (40°) Ore thickness 14 m Ore compressive strength 26 MPa Ore C-factor (fracture density) 0.21 Hanging wall compressive strength 28 MPa Hanging wall C-factor 0.27 About 40 000 m of the holes were drilled by ring drills with an average productivity of 85 m/shift; loading of a round (2 rows, 750 m of hole) was usually accomplished in one shift. The total ore recovery was about 270 000 t with an average iron content of 59%, and three to four shifts were required per round (4700 t) to load. The principal of "throwing" the rock during blast¬ing worked well, and no ore remains on the footwall. ROCK CONTROL Before mining commenced at Indian, it was dis¬covered that the rock strength in its vicinity was rela¬tively poor. Since the ore is flatly dipping, and mining was to take place with open rooms, the strength of the hanging wall was of great importance. A supplementary program of exploration drilling was done during the fall of 1976 in order to assess fracture spacing and rock strength from the drill cores. Results showed that the strength of the hanging wall was low because of the many fractures. A decision was made to try cable bolting of the hanging wall by extending some of the ordinary blast¬holes. Cables (20 mm diam) were grouted in these larger holes and the rest of the hole was then utilized for ordinary loading (see Fig. 2). About 100 bolts altogether were installed and, to date, only small falls have occurred and the hanging wall is holding. Its exposed area is 2100 m2 DEVELOPMENT WORK On the 350-m level, four crosscuts (two per mining room), one longitudinal drift, and two elongated drifts (one per raise) for the raise climber were developed as shown in Fig. 1. Of the 230 m of drift, 60 m were in ore and the rest in waste. For mining room A alone, a total of 115 m of drift were needed with 30 m in ore. The drift dimensions were about 4.5 X 5.5 m. In the drifting on the 300-m level, two connecting drifts were driven from one existing transport drift forward to the respective raises. The total development was 93 m, all in waste (Fig. 1). The two raises having a length of 80 m each were driven at an angle of 0.68 rad (39°) from the hori¬zontal. They were driven along the footwall contact with a separation of 30 m and had an area of 10 m2. LONG-HOLE DRILLING The drilling of long holes with the ring drilling unit began in the fall of 1977 in raise A. The equipment consisted of an Alimak raise climber, equipped with a lower platform from which the charging was done.

Jan 1, 1982

By Rafael Moure

On April 21, 1980 the Oil, Chemical and Atomic Workers Union (OCAW) submitted a petition urging the revision of the Mine Safety and Health Administration (MSHA) permissible exposure limit to radon daughters in underground mines. This petition asks for: 1. Annual worker exposure to be lowered from the current 4 WLM/year to 0.7 WLM/year; 2. Full economic protection for workers whose exposure exceeds the proposed 0.7 WLM/ year; and 3. The primary means of lowering workers&apos; radiation exposures shall be environmental controls. I will present the scientific rationale for the lower standard requested in our petition to the Department of Labor. [1. Radiation Exposures to Uranium Miners] Two types of radiation exposures are present in the environment of uranium mines: A. [External Exposure] to gamma sources which until 1978 was considered as negligible compared with internal alpha radiation. R. E. Utting (1), in an investigation of three Canadian uranium mines, found levels of exposures adding to an annual dose equivalent of 3 rems (30 mSv)/year. He concluded that gamma dosages may be a major fraction of the maximum permissible dose for many individual miners. The U.S. experience has shown that there is significant gamma exposure in U.S. uranium mines and, therefore, it could contribute to the total radiation exposure to the lungs in combination with internal alpha radiation. Before ventilation improvements in uranium mines, the contribution of external radiation to total lung dosage was relatively small. After improvements in ventilation, the reduction in internal radiation exposure appears to be in 1980 as low as 1/10 of the 1950 dosages. However, ventilation improvements do not affect external gamma radiation. Gamma radiation remains at the same 1950&apos;s level, and it represents today a higher proportion of the total radiation burden to the miner&apos;s lungs. MSHA regulations (57.5-47) state that if gamma exposure is below 2.0 milliroentgens per hour, then workers are not required to wear gamma dosimeters and cummulative gamma radiation records need not be kept. Assuming an exposure of 1.9 milliroentgens per hour per 173 hours/month for 12 months,a miner would accumulate a dose from external radiation of 3.9 rems/ year. This dosage would, of course, go unrecorded. MSHA, therefore, should consider a downward revision of the external radiation permissible exposure limit of 2.0 milliroentgens/hour. MSHA should make clear that this exposure level, when added to the internal exposure, does not make for sound health physics -even though the measured current levels do not approach 2.0 milliroentgens/hour. B. [Internal Exposure] to alpha radiation is still the larger contributor to lung irradiation in uranium miners. The target organ is still the lung. The parameters to be considered in determining the biological radiation load appear in Table I for radon daughters present in the upper respiratory tract (URT). TABLE I RADIATION DOSE PARAMETERS IN THE URT OF URANIUM MINERS Parameter Uranium Underground Mining [1. Quantity of the Absorbed Radon daughters (Po-218, Nuclide Pb-214, Bi-214) deposit in bronchial and lung tissue. (Measured by air concentration, Working Levels) 2. Type of Radiation Emitted Alpha Particles 3. Energy of Radiation High-Linear Energy Transfer 4. Half-life of the Nuclide Ra 222, 3.8 days in the Organ Po-218 (RaA), T1/2 = 3.1 minutes Pb-214 (RaB), T1/2 = 26.8 minutes Bi-264 (RaC), Vi = 19.7 minutes 5. Selection Factor of the Direct inhalation with mine Organ dust to be deposited in the tipper Respiratory Tract 6. Mass and Form of the Organ Especially cancer sensitive, tracheobronchial region and bronchi* 7. Distribution of the Activity Dependent on deposited dust in the Organ distribution in the Upper Respiratory Tract 8. Chemical Nature of All compounds (RaA, RaB, RaC) Incorporated Substance biologically insoluble. (Only localized, irradiation) *See Reference (2)] The parameters of Table I depend, in turn, on the demography of the exposed population, especially: 1. Age 2. Sex 3. Health Status 4. Genetic Constitution 5. Lifestyle (Smoking, Diet) Some of these demographic characteristics of uran-

Jan 1, 1981