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By K. Osseo-Asare

The thermodynamics of solid-aqueous solution equilibria up to 200°C In the systems Me-NH3 H2 0 (Me=Cu, Ni, Co) are summarized with predominance area diagrams generated with the DIAGRAM computer program. The implications of the diagrams to ammonia hydrometallurgy are discussed.

Jan 1, 1981

By J. M. Johns

INTRODUCTION The Climax Mine is located on the continen¬tal divide, on Fremont Pass, about 145 Km south west of Denver, Colorado. Molybdenum, along with the by products, tungsten, tin, and pyrite are produced from both the open pit and under¬ground operations. The underground operation utilizes a panel caving mining system, and produces in excess of 27,000 mt per day. In the mid 1930's, the Climax mine began changing from a grizzly level layout to the slusher drift system which is still in use to¬ day. Over the years, the slusher drift layout has been refined to correct, improve, and sim¬plify minor operational problems. In common with most older hardrock mines, the development equipment used at Climax was the standard rail mounted, pneumatic mining machinery. This type of equipment remained in use, at least to some degree, until about four years ago. For production, 112 Kw electric slushers were used, while mine traffic was rail with trolley locomotives. Beginning about ten years ago, Climax began

Jan 1, 1981

By R. A. Pyper

Laboratory leaching experiments were conducted to ascertain the effectiveness of acidic solutions of thiourea, CS(NH2)2, for dissolving gold from finely disseminated gold ores. The ore tested was from the Chipmunk orebody of the Northumberland mining district in Nye County, Nevada, USA. Extractions of up to 90 percent were attained from the ore. Extraction was shown to be dependent on thiourea concentration, ferric sulfate oxidant concentration, acid concentration, and leaching time. Optimum conditions were an ambient temperature leach near pH 1.0 utilizing 100-200 kg/t thiourea and 50 kg/t ferric sulfate for 8 to 24 hours.

Jan 1, 1981

By Harold J. Schock

The cases of no auxiliary ventilation and ventilation with tubing were mathematically modeled, using a turbulent dispersion model. The scooptram that is used to load is modeled as a variable-speed, variable-strength source of pollutants moving in and out of the dead-ended drift. Model predictions are compared to experimental measurements. A parametric analysis was performed to determine the effect of exhaust pipe location and time spent out of the drift, as well as to determine if a relationship existed between CO2 and NO levels. It was concluded from this analysis that when modeling a load-haul-dump cycle, it is necessary to use a model that has the capability of predicting changes in source strength as well as source location. It was also determined that four relevant factors include effective heading height, apparent longitudinal diffusion coefficient, transverse concentration gradient at the face, and average dispersion width of the exhaust jet. As these factors change with heading geometry and ventilation configuration, they were determined experimentally, Carbon dioxide was used as a tracer to calibrate the model; since, knowing the fuel flow rate, it is possible to calculate CO2 production. The above four factors in the computer model were calculated using a regression analysis, comparing predicted local concentrations to measured local concentrations. The model was then verified after this calibration process using additional experimental results. With the calibrated model, it is possible to predict pollutant concentrations at any location in the heading using an engine map and the previously calculated flow field. A parametric analysis shows that the exhaust pipe location can have a significant impact on operator exposure to exhaust gases. It was also shown that NO concentration varies directly with C02 levels. Predictions using the dynamic model are contrasted with constant source strength (pollutant flow rate) results. An analysis of these predictions for computing time-weighted average NO shows the dynamic model predictions of operator exposure to this pollutant to be less than one-half of the value predicted using a constant source model.

Jan 1, 1981

By Madan M. Singh

Mining under bodies of water such as oceans, lakes, rivers, streams, ponds, and other impoundments, is not a new technique. However, in the United States it has been fairly restricted to date. In order to recover the maximum amount of the nation's reserves, in the future, it is important that further mining under such bodies of water be planned. However, it is imperative that this is not done at the expense of danger to the nation's miners or appreciable property damage. If all the bodies of water are considered large, and it is assumed that these constitute a hazard when mining in their vicinity irrespective of their size and geometry, a significant percentage of the coal reserves are likely to be sterilized. How- ever, by adopting special mining plans and procedures, it may be possible to release reserves for exploitation that would otherwise be unmineable since the surface of bodies of water below which they lie are relatively small in size. In 1977, the Bureau of Mines issued Information Circular (IC) 8741, entitled "Results of Research to Develop Guidelines for Mining Near Surface and Underground Bodies of Water". This document was based on the work done by Skelly and Loy (under Contract No. H0252083) and K. Wardell and Partners (under Contract No. H0252021). How ever, these guidelines have limited application, because these did not (1) define the bodies of water of concern, and (2) take into account the nature of the strata being encountered. These guidelines assumed that all bodies of water which would be undermined were large. For the purposes of this discussion, water bodies may be divided into three categories, depending upon their potential to cause damage to the underlying mining activity: ? catastrophic ? major ? limited Bodies of water which pose the threat of catastrophic destruction in the mine, should there be an inrush, include oceans, large lakes, big inland reservoirs, and rivers that could flood the mine complete- ly in a short period. Those with major potential for damage would include lakes and streams which might present a danger to both life and property if not considered in the mining plans. Finite bodies of water which have a volume considerably smaller than the mine volume offer a limited potential for damage in the case of an inrush. These include flowing streams which have a relatively small rate of replenishment and which can be dealt with by pumping or isolation in the event of a breakthrough into the mine. Such bodies of water may be inconvenient and cause a temporary disruption of production, and could even cause some damage to equipment, but only in rare instances would personal injuries occur. The scope of this paper is limited to the exploitation of stratified mineral deposits with overlying -face bodies of water. Vein and dessiminate ore deposits are not discussed, nor are breakthroughs into adjacent mines in the same horizon. Overlying, water-logged strata also pose a serious threat to mining, but are not treated in this discussion. Pending further investigation these strata may be considered as bodies of water with catastrophic potential for damage, with the base of the strata being considered the bottom of the water body. When the body of water is large, so that it may cause catastrophic damage to the mine in the event of an inrush, the guidelines suggested in IC 8741 are applicable. In order to develop criteria and procedures for mining under smaller bodies of water, however, it is necessary that the nature of the strata disturbance due to underground mining be examined.

Jan 1, 1981

By D. Pearson

INTRODUCTION In 1975 278x10 3 tonnes of zinc was consumed in the United Kingdom, of which 69.3x10 3 tonnes was obtained from secondary sources. Of this secondary zinc 37.7x10 3 tonnes was used in brass making, the remainder was used in non metallic applications - eg oxides and salts. Many forms of zinc-bearing wastes arise, for example as galvaniser's drosses and residues, fumes from ferrous and non-ferrous metallurgical processes, and filter cakes from rayon manufacture and metal finishing. In the case of zinc-bearing dusts and fumes, the quantities available for recycle are at present restricted by the failure to collect them efficiently, hut environmental factors are leading to improvements in gas cleaning, with the result that quantities collected will increase and then exceed those required for the production of chemicals, so that some dusts may either have to be dumped or processed to metal. The recovery of zinc from steelmaking dusts is practised in W. Germany (1) and Japan (2) employing processes based on reduction in rotary kilns in order to produce sponge iron and a zinc- rich oxide suitable for feeding to the zinc-lead blast furnace. Although the zinc-lead blast furnace is the only zinc-making process used in the UK, no similar treatment of steelmaking dusts is carried out, a pilot scheme was tested by the British Steel Corporation, but development was abandoned for economic reasons. It appears to be cheaper to pelletise suitable dusts and export them for treatment elsewhere.

Jan 1, 1981

Provide reasonably priced, reliable fire protection for large walking draglines. Approach Fire hazards on large draglines were analyzed, then appropriate automatic fire sensing and suppression systems developed and in-mine tested. How It Works Five fire hazard areas identified for special attention are: electrical control room, high voltage transformer enclosures, circuit breaker rooms, collector ring/roller path area, and the automatic lubrication rooms. Automatic fire sensing and suppression systems were separately designed for each area. The design of each system was tailored to the hazards and equipment protected. Spot thermal sensing and Halon 1301 suppression were specified for the electrical

Jan 1, 1981

By Alan Peacock

Shield supports have been in use in the USA underground mines since 1975, and their development in the intervening time has been progressive and inovative. Specialized techniques are used to optimize their performance in a particular application and to design the component parts.

Jan 1, 1981

By Richard W. Hornung, Steven Samuels

A great amount of work in recent years has been devoted to refining older survivorship models, e.g. life tables, and developing newer ones, e.g. Cox regression. In this paper, these models are applied to data from an epidemiological study of uranium miners in four western states conducted by the U.S. Public Health Service. One of the principal problems addressed in this paper concerns the appropriateness of using cumulative dose as the measure of radiation exposure. An alternative is considered that incorporates dose rate into the model. This procedure does not implicity assume that equal cumulative doses produce the same lung cancer mortality, but instead considers the amount of time in which each miner's dose was accumulated. Results of the statistical analysis indicate that dose rate is a significant factor at higher cumulative dose levels, with lower dose rate apparently more damaging than higher dose rate. The shape of the dose response curve is curvilinear at higher dose levels. No interaction was found between smoking and radiation exposure on a multiplicative or relative scale.

Jan 1, 1981

Consolidated Gold Fields Limited published annually an excellent reference, GOLD 1978, GOLD 1979, giving the Supply and Demand for Gold, Recent Developments and Prospects, Soviet Gold Production, etc. This outstanding reference is authored by Christopher Glynn, Investment Department, London, Great Britain. The 1978 Report marked the tenth anniversary of Gold Field's review of world gold market developments; and from it and the 1979 report, the following information was derived. The recovery of gold is widespread in the Soviet Union. Originally, gold was mined principally in the Urals, but in the last few years several major deposits have been discovered in the southwestern areas of the Soviet Union. One of the most dramatic discoveries has been the Muruntau deposit in Western Uzbekestan. This deposit, discovered in 1958, is reported to be the largest and among the lowest cost producers in the Soviet Union. It is located in Armenia, approximately 750 km (465 miles) northeast of the Iran-Afghanistan border, lying in the center of the Kyzylkurn (Red Sands) desert at the foot of Muruntau Mountain. Free gold occurs in both fine and coarse size in quartz veins up to 20 m (66 ft) thick associated with a small percentage of sulfides, mainly arsenopyrite. The ore zone constitutes a large gold-enriched stock work of about 4 sq krn (1.5 sq miles). The development of Muruntau was a major undertaking with open -pit development and construction of the mill and refinery complex started early in 1967, and the first gold bar poured in mid-1969. At the end of the first construction stage in 1970, the mill building was an imposing structure about 38 m (125 ft) high and 330 (1,080 ft) in length housing, probably, 11 parallel milling circuits. The first construction stage also included six 50 m (165 ft) diameter thickeners and 30 Pachuca tanks. The original mill complex was tripled in size after large additional gold reserves were discovered at a depth of 600 m (1,970 ft); and the deposit is now being exploited by both open- pit and underground workings. Ore is ground in semi-autogenous mills with the free gold being recovered into a high-grade concentrate on shaking tables. Table tailings are purnped to duplex spiral classifiers in closed circuit with the grinding mills. Classifier overflow is then thickened before being transferred to the Pachuca building where cyanidation takes place. An ion exchange resin is then added in a countercurrent

Jan 1, 1981

"Eastern Canadian thermal coal mined at Minto, New Brunswick, and in western Cape Breton at St. Rose, Inverness County, Nova Scotia, contains 6-8 per cent sulphur. Both coals are high volatile bituminous. types. The Minto coal averages 18 per cent ash, 7 per cent sulphur and has a Btu rating of 11,900 per lb. Pyritic sulphur accounts for 75 per cent of the total sulphur and the remainder is largely organic; the amount of sulphate sulphur is negligible. This means that in a 7 per cent sulphur coal, 2 per cent is organic and 5 per cent pyritic sulphur. Pit-head St. Rose coal contains approximately 10 per cent ash, 7 per cent sulphur and 11 ,800 Btu/lb. The distribution of sulphur is different to that of Minto coal because about 12 per cent sulphate sulphur is present. Thus, a 7 per cent sulphur coal would contain 4 per cent pyritic, 2 per cent organic and 1 per cent sulphate sulphur.The iron sulphides, pyrite and, to a lesser extent, marcassite occur over a wide size range from large sulphur balls to crack fillings, lenses, and individual crystals and framboids. Coarse-grained pyrite and sulphur balls are easily separated during washing. It is the fine grain sizes, intimately mixed with the coal constituents, that are difficult to separate. Quantimet analyses of the fine grains in coal from both localities showed that about 50 per cent were smaller than 30 micron(1).No efforts have been made to wash the Inverness coal of Nova Scotia and in 1980 washing ceased in New Brunswick. Previously, the Minto coal was processed in an antiquated jig circuit which reduced ash from the strip operation run-of-mine value of more than 20 per cent to less than 18. More effective removal of ash and sulphur could clearly be obtained. Laboratory tests on both Minto and St. Rose coal have shown that ash rejection rates of 40 to 50 per cent at yields of 70-80 per cent were possible(2), Little information on sulphur removal was available."

Jan 1, 1981

The discussion in this paper centres on examples of mining that are drawn mainly from the coal mining industry. The intention is to review the opportunities that exist for incorporating environmental planning into mining industry practice in New Zealand. The coal mining examples serve to make some general points which have some application for many types of mining.

Jan 1, 1981

By Bennet A. G

An overall review of the selection of rock strength and deformation characteris- tics for the design of the Melbourne Under- qround Rail Loop (MIJRL) rock tunnels is presented. Particular emphasis is placed on aspects of in-situ testing, including large diameter plate bearing tests. Material on the interpretation of test results for use in the design of the tunnel linings and openings is included, with a brief discus- sion of desiqn methods.

Jan 1, 1981

By Milton E. Wadsworth, J. Brent Hiskey

INTRODUCTION Fifteen years ago a distinguished electrochemist referred to electrochemistry as an "underdeveloped science " In the article in question, Professor Bockris (1) stated that many areas of.natura1 phenomena and applied technology are not perceived as being electrochemically based At that time, this was especially true for much of non-ferrous extractive metallurgy and was a particularly accurate description of hydrometallurgy During the last decade, there has been considerable advancement in applying the theory of electrochemistry to our know1 edge of hydrometallurgical processes and in understanding the mechanisms of underlying electrode reactions. Impressive as these advances are, there is still a growing need for electrochemical research of mineral systems and electrochemical engineering to apply this work in practical situations In a broad sense electrochemistry is defined as the physical chemistry of ions in solution and of phenomena which occur at the electrified interface. Bockris and Reddy (2) refer to electrochemistry as being divided into two fundamental domains, ionics and electrodics ionics is restricted to the behavior of ions in solutions, examples being many familiar redox processes Electrodics represents the main part of electrochemistry, and is the study of charged interfaces and the conditions governing charge transfer reactions across them. Considering hydrometallurgical applications, electrodics suggests the following relationship for a conducting solid (bulk or surface)

Jan 1, 1981

By John E. Callahan

"A stream sediment geochemical survey covering an area of 13.000 sq. km near Churchill Falls. Labrador. was conducted for Brinex (British Newfoundland Explorations Ltd.) prior to partial flooding by the Churchill Falls power project. Stream sediments (minus 80-mesh) were analyzed for Cu, Pb, Zn, Ni, Ag, Mo. U and Mn, and the heavy mineral, nonmagnetic fraction (minus 60-mesh) for Cu, Pb, Zn, Ni, Ag, Mo, Ba and Mn. A binocular examination of the concentrates for sulphides and pyrope garnet was done on the heavy mineral concentrates. The concentrates were also examined for fluorescent minerals.Two distinct areas with high trace-element values indicate the need for follow-up studies. The first is in an area of gabbro erratics containing massive and disseminated sulphides (pyrrhotite, pyrite and minor chalcopyrite) on the western edge of the survey area along the Esker Road. The second, largest area (several hundred square kilometres) is anomalous in U and Mo and associated with felsic intrusives (Red Wine Complex?), syenites. granite gneiss and sedimentary rocks (Seal Lake Group?).The binocular study showed that almost 15% of the heavy mineral concentrates contained pyrite and no other sulphides. This indicates that this technique may be worthwhile in determining the presence of sulphide minerals in stream sediments from glaciated environments. The only fluorescent mineral observed in the study was zircon."

Jan 1, 1981

By B. W. Brooks

"After an extensive period of University research and field trials, large scale reclamation of the Waite Amulet tailings discharge areas commenced in the spring of 1978. The Noranda program consists principally of sampling the complete dam to determine the vegetation potential of the various areas, addition of the required amelioration materials, and the use of good agricultural practices. This paper describes the development work that led to the establishment of the Noranda-Guelph Contract and the subsequent findings concerning the revegetation of tailings dams. Current revegetation procedures are presented. INTRODUCTIONThe Amulet Mine Ltd. was staked in 1924 and by 1930 had constructed a small 300 ton per day flotation plant. During the late twenties and early thirties, Noranda Mines Ltd. increased its equity in the company to gain a sixty percent share of this small mine. In 1933 the mine was amalgamated with another small producer, the Waite Ackerman Montgomery Mine Ltd., which was more than eighty per cent controlled by Noranda to form the Waite Amulet Mines Ltd.An interesting historical side point to the story of the Waite Ackerman Montgomery property is that a concentrator was never constructed. In fact, the first ore to go through the Horne mill of Noranda Mines Ltd., built in 1928, was not from the famous upper H and L ore bodies but custom shipments from the Waite."

Jan 1, 1981

By Stein Wennberg, Snorre Tessem

INTRODUCTION The mine is situated in the middle of Norway 150 km north of the town Trondheim. The iron ore deposits were discovered in 1906 and have been continuously mined since. The first owner was A/S Nordiske Grubekompagni. In 1912 Fosdalens Bergverks-Aktieselskab was established; since World War II it has been a wholly state-owned company. The location is advantageous, with mining facilities, ore dressing plant and stockpile situated near the northern head of the Trondheimsf jord. The iron ore always occurs in Ordovician lavas and sediments, now altered to greenschists, keratophyres and limestone. The ore bearing zone is 150 km long. The deposits consist of two parallel linear ore bodies lying one above the other in the same stratigraphic horizon which dips 70' to the south. The upper body is 50-100 m high, the lower one 250 m high, while the distance between the two bodies in the East mine varies between 20 and 60 m. The ore width varies between 5 and 8 m. The ore bodies strike WSW and plunge between 10' and 15" in this direction. To date continuous ore has either been worked or opened up along a total strike length of 10 km. Until 1969 all ore was mined in the western part of the deposit called the Western mine. In the spring of 1969 production started in a new mine, the Eastern mine. The ore bodies in this new mine are geologically the same as in the old mine, but they have been displaced 700 m downwards along a major fault in the Fosdalen valley.

Jan 1, 1981

By Larry J. Tucker

INTRODUCTION The orebody at Pea Ridge strikes NE-SW and dips at about 800 to the SE. The ore and surrounding rocks are Pre-Cambrian in age. These rocks are usually very competent. Locally strong jointing has caused problems and toward the hangingwall dikes have caused particular problems during or after the open stoping phase of production. About 396 m of sediments lie over the ore. The shape and mineral content of the orebody has been fairly consistent so far with production areas lying from the 1375 to the 2475 levels. The deeper limits of the ore- body haven't been found yet. The ore is high grade magnetite with a strike length of over 600 m and a maximum width of over 180 m on some levels at the center of the orebody. Production at Pea Ridge for the first three years of operation came from two types of open stopes. The original stopes were recovered with horizontal production rings being drilled and blasted from vertical raises. Later a more typical sublevel stoping method using vertical production rings from one or more drilling sub- levels was used. There were a number of reasons for the selection of open stoping as the initial production method at Pea Ridge. First, it was judged that the ore was not of a block cave quality. After fifteen years of production, this assumption can be verified. There was some concern during the planning days for the potential for high flow rates of water into the mine from aquifers in the sediments lying over the orebody. Dewatering wells from surface and underground were put in as a precaution and nothing came of this potential problem. Open stoping and its resulting pillar structure put off this problem by supporting the sediments a few years longer. For a moderate investment in level development some production from open stoping was quickly realized. The first open stopes were brought in on what was then the bottom mine production level (2275), near the shaft pillar and situated on the footwall. This production practice of stoping on the heels of development from shaft pillar westward and from footwall to hangingwall continued through the open stope phases of production at Pea Ridge for many years. From the standpoint of mine efficiency and pillar stability, this would not seem to be a recommended mining practice. 74 570 w electric slushers, using 1.37 m dippers, pulled the open stope ore out into 18 ton bottom dump cars. Ordinarily five of these cars made up a train, which was pulled to the dump area by a trolley powered electric locomotive. Open stoping had the disadvantage that 100% of the recoverable portions of the orebody had to be developed in order to recover 33% of the ore. 67% of the ore was left behind in a large pillar system. For many years we had no planned method for recovery of these pillars. Mining methods are of course, limited by the existing technology. Advancing technology, coupled with the limited production capabilities of open stoping, low efficiencies and high relative cost early in the history of the operation recommended that other mining methods be looked at. In Sweden, considerable practical experience, coupled with a large amount of theoretical work had become available on sublevel caving. To Pea Ridge it offered the apparent benefits of: 1) The flexibility of diesel powered trackless haulage. 2) The efficiency of two machines to one operator on longhole drilling. 3) A multilevel production picture with a large number of drawpoints for ore recovery. 4) 100% recovery of the mineable portion of the orebody in one production pass, with no pillar recovery problems left behind. This paper outlines the experiments and experiences which Pea Ridge has gone through with sublevel caving from 1967 to this date.

Jan 1, 1981

By S. Bassignani, D. Ippolito

In Italy at the moment there is a program which involves for the future the exploitation of two uranium mines which are now in the research stage. About one hundred and fifty workers are at present employed in those facilities. An health physics surveillance program is under way and has been designed to fulfill the requirements due to the future exploitation of the two mines. A complete evaluation of the total dose (external + internal) received by the workers will be given together with a detailed analyses of the technical methods employed for such evaluations. The external doses are evaluated on the basis of individual TLD dosimeters, while the evaluation of the internal doses is mainly due to an enviromental survillance program of the radon and radon daughter air concentration inside the mines. The criteria established for the assessment of the sampling procedures, sampling intervals and dose evaluation related to the radon doughter exposure are reported. The external dose received by the workers are lower than 0.2 m Sv~month and the internal exposure less than 6.4 IO-3 J.h.m-3 /year or using the old units 2 WLM/year.

Jan 1, 1981

By Jay S. Silhanek, Vernon E. Andrews

[Introduction] Section 122 of the Clean Air Act Amendments of 1977, Public Law 9595, directed the Administrator of the Environmental Protection Agency, to review all relevant information and determine whether emissions of radioactive pollutants into ambient air will cause or contribute to air pollution which may reasonably be anticipated to endanger public health. In August 1979, EPA published a preliminary report on assessments of public health impact resulting from emissions of radionuclides into air from a broad spectrum of major source categories (1). A section of this document presented a theoretical analysis of the radioactive airborne emissions from several non-uranium mines including iron, copper, zinc, clay, limestone, fluorspar, and phosphate. Since 1978 EPA's Las Vegas Laboratory has been gathering field data on actual radionuclide emissions from these mines to support the earlier theoretical analysis. The purpose of this paper is to present the results of those field measurements in comparison with the assumed values for the theoretical analysis. Data were presented in EPA's preliminary report on both gaseous radon-222 and particulate uranium-238 emissions from non-uranium mines. Most of the emissions data were rough estimates based upon the uranium content of the ore. Preliminary data from a few field studies were also used. Although both the theoretical and field studies by EPA were concerned with radon-222 and radioactive particulate emissions, this paper will concentrate only on the radon effluents. Radon from mines seems to have the greatest potential for possible health effects due to radiation. Data on particulate emissions are presented in other technical reports (2-5). [Theoretical Study] The theoretical emission data are presented in Table I. The estimates for the iron, zinc, and clay mines are based upon preliminary measurements taken by the Las Vegas laboratory. The estimates for the other mines were made assuming that 20% of the total radon-222 in the ore was released during the mining process and the radon release rate from the exposed ore surface was a factor of two greater than that resulting from mining. [ ] Dose assessments for radon from these mines were carried out using model facilities located on generic sites (1). The AIRDOS-II computer code (for estimating dispersion of airborne pollutants) with some minor modifications was used to determine working level exposures to a maximally exposed individual and a regional population. A conversion factor for risk/WL-year(O.021 fatal cancers/person-WL-y) was developed primarily from information on uranium miner exposures (6). This computer code and conversion factor were then applied to the assumed emission rate of radon-222 from each type of mine to develop estimates of working levels, lifetime risks and number of fatal cancers. A summary of this data is presented in table II. [ ] The highest estimated number of cancer deaths per year of operation is 0.05 cancer deaths or 1.0 death due to radon during the 24 year life of a phosphorous mine. For comparison, EPA estimates that underground uranium mines would cause .03 cancer deaths per year of operation in a regional population(1). [Field Study] As a followup to the theoretical study, EPA's Las Vegas laboratory began making measurements in 1978 of actual emissions of radon-222 and radioactive particulates from mine vents at several facilities for each of the industries listed in Table I and II. They were intended to verify general information and radon emissions from EPA's theoretical estimates.

Jan 1, 1981