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By Ben Benally, Joseph M. Hans

INTRODUCTION The decontamination of the former Shiprock uranium millsite and the interim stabilization of its tailings piles began in March 1974, and ended in June 1978. Personnel monitoring for whole body gamma ray exposure was conducted for selected individuals working at the mill site throughout most of this period. The millsite and exposure conditions are briefly described to develop a basis for interpretation of the exposure data. Millsite The Shiprock uranium millsite is located in the community of Shiprock, New Mexico on a bench overlooking the San Juan River (Figure 1). Both Shiprock and the millsite are located on the Navajo Indian Reservation. About 1.5 million tons of tailings were produced on the millsite during its operation from 1954 to 1963. The tailings were pulped with waste process fluids and water, and discharged to form two tailings piles. The piles (ponds, dams) were formed with earth starter dikes and the dikes were then raised using the separated sand fraction of the tailings. The slime fraction of the tailings were discharged into the diked impoundment. The mill was shut down and partially dismantled before the lease expired in 1973, and the remainder of the mill and tailings reverted back to the Navajo Tribe which was the lessor. Portions of the millsite were then occupied by the Navajo Engineering and Construction Authority (NECA), a tribal enterprise. In addition to its regular activities, NECA established an on-site training school for heavy equipment operators. The trainees principally used the south tailings pile as a practice area and spread the tailings into a 10 hectare row of raffinate ponds located along the southwest side of the millsite. Overall, the training activities inceased the effective area of the tailings piles from 20 to 35 hectares. Another 56 hectares were contaminated with wind and water eroded tailings. Exposure Conditions Persons decontaminating a millsite may be subject to radiation exposures from several sources. Only whole body gamma exposures are discussed in this report. These exposures are arbitrarily divided into three levels: high, intermediate, and low. The levels are related to areas in which the decontamination work was being done and have no connotation concerning imminent hazards. High exposures would result when persons were working with raw tailings, and intermediate exposures would occur when persons were working in the ore storage and mill yards. Low exposures would result from working in areas where surface contamination exists from wind and water erosion. [ ] The millsite and its contaminated environs were divided into major areas to facilitate the decontamination work (Figure 2). Areas A, B, and C are defined as perimeter areas containing surface contamination. Work in these areas was expected to result in low exposures. Areas D, E, and F are the

Jan 1, 1981

By Robert W. Buechley, Charles R. Key, Jonathan M. Samet, Marion V. Morgan, Arnolfo A. Valdivia

INTRODUCTION The Uranium Epidemiology Study at the University of New Mexico was initiated in 1977 with the objective of assessing health effects of uranium mining in the Grants, New Mexico, area. The planned approach was a cohort study of mortality that would capitalize on the availability of health records at the Grants Clinic. During the first year, emphasis was placed on determining the feasibility of the proposed study: Were adequate records available? What was the quality of available information? Could a cohort be established? The first year's work confirmed the study's feasibility (Buechley, 1979). Subsequent efforts have focused on the health and exposure data base and comprehensive follow-up. A retrospective cohort study of early miners is in progress; a prospective cohort study of miners who began working under the current Federal standard is planned. This report will summarize the current status of the Uranium Epidemiology Study. A chronology of the study's activities follows (Table 1). TABLE 1. Uranium Epidemiology Study Timetable 12/77 Official start of study. 12/77-8/78 Abstracted Grants Clinic records of 21,780 persons examined for mining companies from 7/57 through 12/76. Checked company personnel records to establish who had worked underground 1 year or more through 12/76. 7/78- Obtained physical exam and mining history information on qualified miners from Grants Clinic and companies. 9/78- Obtained working level month information from companies and other sources. 12/78- Checked qualified miners list with New Mexico death tapes and New Mexico Tumor Registry files. 9/80- Submitted qualified miners list to Social Security Administration for search of its records. 12/80- Obtained copies or abstracts of early State Health Department and Mine Inspector's reports. 3/81-4/81 Conducted Pilot Respiratory Disease Survey. 3/81- Verified underground status of qualified miners via company personnel records. 3/81- Verified smoking and mining histories of qualified miners via Grants Clinic records. 3/81- Began establishing cohort with first underground experience after 1970. 4/81- Checked qualified miners list with Post Office, Voter Registration Office, Motor Vehicle Division, telephone directories, city directories, and other sources. BACKGROUND Uranium Mining in New Mexico The Grants, New Mexico, uranium belt measures some 110 by 25 miles in Cibola, McKinley, and Sandoval counties (Kelley, 1963). The three chief mining districts are, from east to west, Laguna, Grants, and Gallup. Average annual production was about 5,500 tons, or 45% of United States uranium, from 1966 through 1978. Despite a drop in this figure in 1979, to 40%, New Mexico still leads the nation in uranium reserves and production (Hatchell, 1981). In 1918 carnotite, a yellow uranium vanadate, was identified in the Shiprock area (Chenoweth and Learned, 1980; N.M. Energy and Minerals Department, 1981). It received little attention at the time because of its apparently limited volume. In 1950 an important uranium outcropping was discovered by chance in limestone at Haystack Butte, northwest of Grants. In 1951 deposits were discovered in sandstone at Poison Canyon, north of Grants, and at Jackpile, northeast of Laguna. In 1955 finding of radioactive cuttings at an abandoned oil test site near Ambrosia Lake led to drilling and discovery of the first underground deposits (Melancon, 1963). Ore bodies were also found near Church Rock in the 1950's (Rautman, 1977). A uranium rush followed. At various times during 1952-1962, 60 mines and 5 mills were in operation (Gay, 1963). In the early 1960's the price of uranium fell, and many miners left the area. Subsequent market fluctuations have produced shutdowns and consolidations, development and expansion. Many early uranium miners came to New Mexico with

Jan 1, 1981

By T J. Wilton, K A. Broadhurst

The United Kingdom, although a relatively small island, has a wealth of mineral deposits, coal, granite, limestone, gypsum, lead, florspar, etc. Coal is a major source of energy with 71 opencast coal sites, 52 using explosives. Many of these sites are in rural areas, the average area worked being 250 acres. The smallest is 35 acres, the largest 2,000 acres, but only 600 acres can be operated at any one time.

Jan 1, 1981

By D. C. Dahlin

The Bureau of Mines conducted laboratory batch beneficiation tests on samples from four potential platinum resources in southeastern Alaska to concentrate platinum and other precious metals. Thin Investigation was done as part of the mission of the Bureau to insure an adequate supply of minerals to meet the Nation's needs. Petrographic studies and beneficiation methods including gravity concentration, froth flotation, and magnetic separation were used to determine mineral association and to attempt to concentrate the platinum values with a primary mineral value such as chromite, copper sulfide, or magnetite. A high-graded sample from the Salt Chuck copper sulfide deposit yielded the platinum-group metal concentrate; the maximum grade attained was 0.04 oz Pt and 1.5 oz Pd per ton of high-grade copper sulfide concentrate. The other samples yielded low-grade primary mineral concentrates with low platinum-group values, and/or the platinum-group metals did not concentrate with the primary mineral value.

Jan 1, 1981

The Pinson Mining Co., which will start production in very early 1981, is the fourth Carbon-In-Pulp plant to be put into operation in the United States. The property is located in Humbolt County, Nevada, 38 km (24 miles) northeast of the town of Golconda. The famous Getchell Mine was also near Golconda and was the site of one of the early Carbon-In-Pulp plants in the 1940s. The deposit was discovered in 1971 by Reno based Cordilleran Explorations, which is founded by a consortium of Canadian mining companies. Cordilleran is a partnership consisting of John Livermore and Peter Galli. In 1972 a second but smaller deposit, called "Preble", was discovered at a location 15 miles south of Pinson. During 1980, a decision was made to place the Pinson deposit into production and Pinson Mining Company was formed. Ownership of Pinson is vested in Lacana Mining Inc. , Rayrock Mines, Inc. , Siscoe Metals, Inc., John S. Livermore, Peter E. Galli and Donald M. Duncan. Three distinct but connected ore zones occur at Pinson. Combined reserves of milling ore total 3, 200, 000 tons at an average (undiluted) grade of 4. 12 g per mt (0. 12 oz per st) Au. Initial mining will be from "A" zone which contains 1, 500, 000 tons with an average (undiluted) grade of 5. 8 g per mt (0. 17 oz per st) Au. An additional 5, 500, 000 tons of low grade ore will be mined from the Pinson deposit and processed by heap leaching. Mining reserves at Preble total at least 1, 500, 000 tons grading 2. 7 g per mt (0. 08 oz per st) Au. Production plans for the Preble deposit have not been formulated at the date of writing. Gold mineralization at both deposits is finely disseminated as sub micron sized particles. The ore at Pinson is structurally controlled and occurs in an Ordovician limestone-siltstone unit. Much of the ore has been silicified to a jasperoid. The ore zones have been entirely oxidized. Preble ore occurs in the Cambrian formation in a carbonaceous, limey shale. Only the upper part of the Preble is oxidized and amenable to conventional cyanide leaching. The Pinson ore body is adjacent to the mouth of Granite Creek in the Osgood Mountains at an elevation of 1, 615 m (5, 300 ft) with the mill being 1. 2 km (3/4 mile) from the mine at an elevation of 1, 495 m (4, 900 ft). Mine development work began in May of 1980 with ore being stockpiled for later processing. Pertinent factors affecting

Jan 1, 1981

By J. M. Smith, T. R. Horton, R. L. Blanchard, T. W. Fowler

INTRODUCTION Uranium mining operations release radioactive materials into both air and water and generate large quantities of solid wastes containing low levels of radioactive materials. Solid wastes produced by mining operations remain on the surface at many inactive mining sites in the Western United States. These mining effluents may present a potential health and environmental hazard. Therefore, Congress, in Section 114(c) of the Uranium Mill Tailings Radiation Control Act of 1978, instructed the Administrator of EPA to prepare a report identifying the location and potential health, safety, and environmental hazards of uranium mine wastes and to recommend a program to eliminate these hazards. Several facts and limitations helped shape the method and approach of the EPA study. Little information on uranium mines was available; measurement information that was available on uranium mine wastes was frequently influenced (biased) by nearby uranium mills; time and resources did not permit comprehensive field studies to provide additional data; and there are inherent variations between uranium mines and sites that complicate generic assessments of mine wastes. To accommodate these facts, the EPA developed conceptual models of uranium mines and made health and environmental projections from them. The models were based upon available data from the literature, supplemented with information from discussions with persons inside and outside the EPA, and by doing several short-term field studies in Texas, New Mexico, and Wyoming. When necessary, conservative (maximizing) assumptions were employed. This paper presents a brief account of a part of the EPA study dealing with the potential health and environmental effects caused by active surface and underground uranium mines. Airborne contaminants are emphasized, although solid and liquid effluents are also included. Due to the limited space, only the methods and parameters used and the results of the assessments will be presented here. Anyone interested in the source of the data used and the development of the parameters should refer to the EPA report (Blanchard et al., 1981). The occurrence and emissions of stable elements were included in the EPA report, however, due to space limitations and their apparent small impact, except for possibly at some specific mines, only radioactive sources will be included in this presentation. MODEL URANIUM MINES The model surface mine was located in the South Powder River Basin of Wyoming and the model underground mine was located in the Ambrosia Lake area of New Mexico. These are the prevalent type mines in those areas. The model mines were based on the average production parameters of the 63 open pit mines and the 256 underground mines that were operating in the United States in 1978 (Department of Energy, 1979) and on a report of an extensive study of open pit mines in Wyoming (Nielson et al., 1979). Information contained in environmental impact statements and in reports from federal and state agencies was also used. Parameters for the model mines are listed in Table 1. The surface mining scenario is that 7 pits are opened in the 17-year mine life with overburden from each successively mined pit used to backfill a previously completed pit, resulting in an equivalent of one pit of overburden (2.4 year production) stored on the surface. No backfilling is assumed at the underground mine. Overburden or waste rock, ore, and sub-ore are separated into separate piles that are either rectangular in shape with length twice the width or in the shape of a frustum of a regular cone. Both shapes have 45 degree sloping sides. To account for bulking, the volume of the material comprising the piles was considered to be 25% greater than the volume of material removed from the ground. It was assumed that dewatering was required at both mine sites. Wastewater discharge rates at the surface and underground mines were assumed to be 3.0 and 2.0 cu m per min, respectively. SOURCE TERMS The following radioactive contaminants at active uranium mines were assessed in the EPA report: 1. Radioactive particulates in a) wind suspended dust from waste rock (overburden) pile, sub-ore pile, ore stockpile, b) suspended dust from mining activities (rock breakage, loading and unloading ore and wastes), and c) vehicular dust, 2. Rn-222 emanation from waste rock (overburden) pile, sub-ore pile, ore stockpile, and mining activities, 3. Rn-222 emanation from mine surface areas, and 4. Radionuclides in wastewater discharged to land surface. Estimated average annual dust emissions (item 1 above) from the model mines are listed in Table 2. Emission factors and the assumptions used to estimate these dust emissions are described in detail in the EPA report. Radioactive source terms were computed for each of the sources; dust emissions were multiplied by the concentrations listed in Table 1

Jan 1, 1981

The Tuscarora mining district, which has been highly productive of gold and silver ores in past years, is situated about 50 miles north- west of Elko, Nevada. The town of Tuscarora lies at an altitude of about 6,000 ft on the west side of the Independence Valley. Precipitation, particularly the winter snowfall, is considerably greater than in many other parts of Nevada, and temperatures are lower. From Tuscarora can be seen-across Independence Valley in the mountains to the east, a distance of about 12 miles, the Freeport Minerals Jerritt Canyon mine, which seems to be-destined to become one of the great gold mines in the western hemisphere. Tuscarora, a gold and silver camp, was active from the early 1870s to 1903 and produced about $10 million in those 30 years. Ten major mines and several smaller ones were operated in an area of about 1.45 square kilometers (one-half squaremile). The deepest shafts were only 230 m (750 ft) but most of the ore came from about the 105 m (350 ft) levels. Mineralization above the water table, at about 30 m (100 ft), was mostly native silver and cerargerite with free gold. Below the water table the ore was rich in sulfides including argentite, stephanite, proustite, pyrargyrite, enargite, arsenopyrite, bornite, chalcopyrite, pyrite, sphalerite and galena. A number of the properties contain sizeable dumps of what was con- sidered waste rock when it was mined. These dumps aggregate over 408,000 mt (450,000 st) with one dump, the Navajo, containing over one-third of the total. A few grab samples taken in the summer of 1978 from some of the dumps indicated interesting silver values with minor gold content. AgAu Exploration Ltd., was organized to explore and develop these dumps into a heap leaching operation. In the fall of 1978 a campaign was started to lease, with front money payments on a royalty basis, three groups of claims on which were located a few hundred thousand tons of dumps. In September and October of 1978, a systematic program of sampling was started on the Navajo dump which was most favorably situated in that it was near to an area which would furnish good sites for heap leaching pads. At the time it was fully recognized that sampling of dumps was a most difficult problem. Because of the topography on which they were situated and the manner in which they were formed, each dump required its own special sampling technique.

Jan 1, 1981

By Tim F. Dyson

Feasibility of employing automated processing and detection techniques in the mine disaster communications problem is demonstrated. Efficient processing methods are developed. These methods have been demonstrated both in a laboratory and in a field environment. Evalua¬tion of existing computer capacity is given along with recommendations for expansion.

Jan 1, 1981

By K. K. Kohli, S. S. Peng, R. E. Thill

Introduction Since the adoption of the Surface Mining Reclamation and Control Act of 1977, which mandates that surface subsidence be an intergral part of the underground coal mine design, there has been a sudden rise of demand on guidelines for minimizing surface subsidence and its associated structural damages. Unfortunately there is no sufficient amount of systematic subsidence data to fill such a gap, although considerable research has been devoted to the subsidence induced damages of surface structures in the past. However, it was said that many surface subsidence surveys have been performed and monitored by coal companies during the past twenty years, and that it will be a great source of information for subsidence engineering design if they can be collected and assembled together. An effort was therefore initiated in the Department of Mining Engineering, West Virginia University, under contract with the U. S. Bureau of Mines in December 1978 to contact each individual coal company in the Northern Appalachian Coal Field for any surface subsidence survey data. After 10 months' survey, the data collected include: 17 longwall panels and 2 room and pillar sections. The data in each panel and section cover the complete set of subsidence survey raw data, underground mine layout maps, surface topo maps, geological loggings showing the stratigraphic sequences and thickness of the overburden, and if available, locations and types of surface structures. On-site visits were followed to study the surface terraines and structures. Interviews were conducted with the owners of the surface structures in most cases about, if any, the history and characteristics of damages. Most subsidence surveys started with the first panel and ended with the second or sometimes, but rarely, the third panel. The first panel may be located in a virgin area but frequently it was adjacent to (on the tailentry side) a mined-out room and pillar section. It must be noted that all of the subsidence data is restricted only to the amount of vertical subsidence and that the horizontal displacement measurement was completely ignored. Primarily due to the fact that vertical subsidence is the most striking feature in a subsidence trough, which is mistakenly considered to be the only critical contributor to surface structural damages, and that a horizontal displacement survey is much more complicated and time-consuming.

Jan 1, 1981

By T. S. Mohibatsela, R. V. Ramani, C. V. Manula

The Mineral Engineering Department of The Pennsylvania State University has developed a suite of computer models for anthracite open pit mining (Fig. 1)1 Individual sub-systems have been developed to (a) interpolate geologic field data to obtain coal tonnages, waste volumes, and strip ratios; (b) simulate equipment performance on selected pit profiles to generate data required to choose coal and waste removal equipment; and (c) evaluate selected cost factors to determine either a rate of return for a specified selling price or vice-versa. This paper describes the application of two of these subsystems, namely, the Reserves Model (GEOL) and the Open Pit Materials Handling Simulator (OPMHS), and how these may be applied for the selection of equipment for a prime anthracite mining site.2

Jan 1, 1981

By Dennis H. Laubscher

INTRODUCTION The selection of mass mining methods is the situation facing not only planners of new operations, but also those involved with cur- rent operations, including the eventual change from open-pit to underground. Increasing ground control problems on many mines, plus a better understanding of dilution and ore loss problems have meant a review of current mining methods and support techniques. Experiments are only successful if conducted in the right mining environment and on the right scale. Mining personnel have been known for their conservatism and will tend to persevere with an established method rather than change; often, hidden costs will obscure the correct economic assessment. An established mining sequence is not necessarily correct in the long term. However, recommendations to change are usually met with opposition because of inconvenience, temporary production problems and the desire for a trouble-free guarantee. In any planning exercise, care must be taken that time and money are not wasted in trying to achieve the unattainable. During the decision-making period, production personnel must be involved, and, if possible, they should visit similar omrations to that proposed. Co-operative production personnel will help ensure success. Descriptions of mining methods are available in technical papers and books, and it is not intended to discuss this aspect, but, rather, the various factors which must be recognised before arriving at a final selection. In another paper presented at this Conference (Heslop and Laubscher, 1981) factors affecting ore recovery and dilution are discussed in greater detail for various cave mining methods. PRELIMINARY CONSIDERATIONS Comprehensive Geological Data. Geological information is available during the exploration stage, but, often, core from exploration drilling is not analysed with the selection of a mining method in mind. Drilling costs are high, therefore the maximum benefits must be derived from this operation. Holes which have not intersected ore may provide data on possible shaft sites, and 35 mm colour slides of core, sound logging and geomechanics classification, will provide a record which can be used with confidence at a later date. Intersections of the various rock types should be kept for future study; for example, is weathering likely to be a problem, bearing in mind that this can be accelerated in the warm, humid atmosphere of an underground mine. The accumulation of data required to select a mining method starts with initial geological work in the target area, whether this be surface mapping or a diamond drill hole. Publications. As technical papers, reports and books are used in the selection of mining methods, an appeal is made to authors to describe the geological environment so that readers can relate to their own conditions. Not enough emphasis is given to the problems experienced and the techniques employed to overcome them. Follow-up papers are seldom written and we should be prepared to discuss our failures. Diagrams are often misleading, as the ideal and not actual situations are depicted. The caved ore/waste interface is not a straight line, and whereas this may not be significant in a diagram of a 200-117 draw height, it certainly is in sub-level caving with draw zones only 1 - 2 m width but up to 20 m high. SUMMARY OF FACTORS AFFECTING THE SELECTION OF UNDERGROUND MASS MINING METHODS (a) Regional rock stresses and mining geometry. (b) The geomechanics rock mass classification of the orebody and surrounding rock mass. (c) Ratio of the surface area of the ore/unpay interface to the contained ore which determines ore losses and dilution. (d) The cavability and fragmentation data of the ore and hangingwall. (e) The mineral distribution in the orebody and its dilution zone.

Jan 1, 1981

By James A. Browning

Following an unsuccessful attempt earlier in the program to develop and market a drill channelling machine, we have designed and tested a new continuous channelling method which reduces operating sound levels to below 100 db. This new method relies on a small diameter oxygen-enriched burner to produce narrow channels. Sound intensities are reduced substantially compared to conventional channel widths. They are further reduced by a water spray directed immediately above the flame. Baffles placed along the top of the channel will reduce sound levels to below 90 db. Cutting rates are greater than those of existing channelling burners with costs per unit of channel area substantially reduced for those cases where oxygen costs are reasonable.

Jan 1, 1981

By Griffiths J. B

From its mining lease at the northern end of Cohens Reef at Walhalla in the North Gippsland Mountains, the Long Tunnel Extended Gold Mining Company produced 13,695 kg of gold, which at 1980 prices represents a value significantly in excess of $200 million. In the period 1885 to 1908 the mine was one of Australia's principal reef gold producers, being top Victorian producer in six of these twenty-four years. Prior and subsequent occupants of the lease produced an additional 160 kg of gold. The Extended and the other mines on Cohens Reef whilst contributing importantly to the financial resource of Victoria, were the sole important support for the substantial town of Walhalla that had grown up around them. When the mines closed, Walhalla gradually reduced from a bustling developed town to the present attractive mountain village so favoured by tuurriSts. To allow the visitor to obtain a better ;ri.icrstanding of Walhalla's mining past, important parts of the Long Tunnel Extended Gold Mine workings and site are being

Jan 1, 1981

By Edward Burris, Terry Gorsuch, Joseph M. Hans

INTRODUCTION In any large earth-moving operation, good planning and management are necessary to complete the operational tasks promptly and successfully. When an earth moving operation is complicated by radioactive contaminants, normal earth moving techniques and procedures must be modified. Any planning and management, therefore, must include the radiological aspects of the operation. It was found that the radiological aspects dominated most of the planning and management activities and were extended to all facets of the decontamination work at the former Shiprock uranium millsite. These planning aspects are discussed and their use to develop a work plan is described. The management aspects are discussed and their use to establish a management structure are also presented. PLANNING Some method of procedure, formulated beforehand, was necessary to govern the decontamination work at the former Shiprock uranium millsite. This procedure was expressed in the form of a work plan which served several listed purposes. 1. It defined the work to be done and the sequence it would follow. 2. It was used as a yardstick to measure progress. 3. It was used to assign organizational responsibilities. Several factors were considered to aid in the development of the plan. These factors are discussed below: Goals It was established that radiation exposure was occurring to persons working at the millsite, and in an around the community of Shiprock, from airborne radioactive mill wastes and radon-222 exhaling from the tailings piles. The goal set for the decontamination work was to reduce on-site exposures to levels acceptable for the millsite occupants. The attainment of this goal would also have a substantial impact in reducing off-site exposures. The objectives necessary to achieve the goal were consolidation and containment of the wastes. The former objective implies decontamination of the millsite and environs, and the later implies stabilization of the wastes. In practice, a total and complete decontamination of the millsite and contaminated environs would be very difficult and costly. The costs for decontaminating them could be high enough that an alternative method might be more cost-effective for reducing human exposures (i.e, move the affected people away from the source). The interim guide "Radiological Criteria for the Decontamination of Inactive Uranium Millsites" was used for the decontamination criteria (EPA 74). Briefly, the criteria state that off-pile decontamination should be effective enough to reduce the net above ground exposure rate to less than 10 [u]R/hr for unrestricted use of the affected area. When decontamination cannot readily be achieved, the exposure rate levels could be relaxed to 40 [u]R/hr; however, the affected area has to be restricted. The second objective, waste containment, means isolating the wastes from the biosphere. Since no method of containment was available at the beginning of the millsite decontamination effort, temporary containment (interim stabilzation) became the objective. The tailings pile and decontamination wastes would be covered with clean fill. The interim stabilization should last from 5 to 10 years until the final disposition of the wastes will occur. The goal, therefore, would be achieved by decontaminating the off-pile areas to less than 10 uR/hr where practical. The decontamination wastes would be used to plate the surface of the tailings pile and would be covered with clean fill. Radiological Survey The radiologial survey is the key factor for planning a decontamination activity. The survey should delineate the spread and depth of the contaminants relative to the decontamination criteria. Surface wastes, in general, can be evaluated for spread and depth with reasonable radiation survey equipment. Subsurface wastes on the other hand can be missed entirely, as happened during the radiation survey at the Shiprock site, although numerous exploration holes were bored and dug. The survey results can be used to define areas that may not be amenable to decontamination because of complications or safety reasons. For example, no decontamination of the bluff base was to be attempted because of the possibility the bluff might collapse on the personnel and equipment. Contaminated bottoms of decant ponds on the flood plain were not removed because they would be slurried by ground water. Slurry removal was deemed inefficient because the contaminants would be scattered and no equipment was available for its transport. In summary, the radiological survey defines the boundaries of the decontamination work and provides

Jan 1, 1981

By David M. Lake

Conversion to trolley assist is a means of minimizing diesel fuel consumption, which in many mining situations offers a rapid payback of capital investment plus the inherent flexibility and maneuverability of truck shovel operation. When diesel electric truck shovel mining operation with trolley assist is compared to alternative mining methods such as conveyor systems, the trolley assist system can be easily justified in terms of return on investment, payback and increased productivity. A review of present technology, performance requirements and past experience in the mining industry has led to the development of a trolley assist system for diesel electric haulage trucks which, is proven in the mining environment, offers significant reduction in fuel consumption and increased productivity.

Jan 1, 1981

By M. Kanefsky

Two channel polarity coincidence and polarity difference statistics are analyzed. The signal, common to both channels, consists of sinusoidal bursts where the exact frequency of the signal is nearly known, but other parameters such as amplitude, phase, and pulse starting time are unknown. The noise inputs are dependent, narrow band, Markov processes. It is shown that the performance depends not only on the signal uncertainties, but on the precise shape of the cross-correlation functions between the noise in-puts. By using two polairty difference statistics in addition to the polarity coincidence statistic, it is shown that the decrease in performance, as well as the cost of hard limiting due to correlated inputs, can be made small.

Jan 1, 1981

By E. J. Peterson

Oil shale retorting induces mineral and chemical reactions to occur on the macroscopic and microscopic levels in the kerogen-bearing marlstone. The nature and extent of the reactions is dependent upon process variables such as maximum temperature, time at temperature, atmosphere, and raw shale composition. This report describes the investigation of the mineral, chemical, and trace element release properties of spent shales retrieved from an experimental in situ retort at Occidental Oil Shale, Inc. 's Logan Wash site in Garfield County, Colorado. Correlation between mineralogy of the spent materials and the mobility of major, minor, and trace elements are indicated, and relationships with important process parameters are discussed. The progress .of carbonate decomposition reactions and silicat ion reactions is indicative of the processing conditions experienced by the shale materials and influences the mobility of major, minor, and trace elements when the solids are contacted by water. Shale minerals that are exposed to the extreme conditions reached in underground retorting form high temperature product phases including akermanite-gehlenite and diopside-augite solid solutions, kalsilite, monticellite, and forsterite. The persistence of relatively thermally stable phases, such as quartz, orthoclase, and albite provide insight into the extremes of processing conditions experienced by the spent shales. Leachate compositions suggest that several trace elements, including vanadium, boron, fluoride, and arsenic are not rendered immobile by the formation of the high-temperature silicate produce phase akermanite-gehlenite.

Jan 1, 1981

By Stephen Utter

One function of mining is to supply oil shale of the required grade and characteristics to surface retorts. The mining methods must he safe, efficient, and environmentally acceptable. Although the Bureau of Mines no longer conducts research on methods of mining fossil fuels, including oil shale, its many years of activity in that field resulted in a significant body of findings. This article summarizes some of these findings, with respect to oil shale mining. The large areal extent of Green River formation oil shales and the diversity of geologic, hydrologic, topographic, and other conditions, preclude the use of any one single mining and waste management technique. Instead, in most instances, methods and equipment will be chosen to meet site-specific conditions. Such methods include conventional surface mining (open-pit and strip), underground mining (room-and-pillar, sublevel stoping, caving) and in situ extraction (true in situ from surface drill holes and modified in situ from underground retorts containing rubblized oil shale). Combination or hybrid systems may prove suitable. Both surface and underground backfilling of spent shale waste may be practiced. Tremendous amounts of oil shale must be mined, processed, and the waste shale properly disposed of in order to meet projected industry production goals. For example, about 1.4 tons of oil shale averaging 117 L of shale oil per ton are required to produce one bbl of oil. A plant with a daily production of 8 dam' (50 000 bbl) of oil would thus require some 68 kt of raw shale daily. This is more than the average daily production from any underground mining operation.

Jan 1, 1981

By Won C. Park

Experiments showed changes in the mineral content of dolomitic oil shale at temperatures and exposure times which bracket conditions encountered during in situ retorting. Laboratory results correlate with cores from a spent in situ retort at Occidental's Logan Wash test site at Garfield County, Colorado. Dolomite, CaMg(CO3)2, decomposes rapidly to form periclase, MgO, calcite, CaCO3, and CO . Both calcite and periclase are intermediate products that can be stabilized by CO atmosphere at lower temperatures, while steam accelerates the formation 2 of silicates. The decarbonation of calcite appears to involve a direct reaction with both quartz and periclase in forming these silicates. Decarbonation preceeds the formation of akermanite, Ca2 MgSi 207, and diopside, CaMgSi 206, within several hours of heating and is followed by the nucleation of forsterite Mg2SiO4, and monticellite, CaMgSiO4. Under in situ retorting conditions, akermanite and diposide are the major minerals above 800°C. Equilibrium is not attained within 200-300 hours. In situ retorting temperatures above 800°C for extended periods of time should reduce potential ground water leaching by the formation of silicate minerals. These silicates are less soluble than free lime, magnesium or alkali oxides. There is a corresponding decrease in the natural pH of these silicates from approximately 12 down to 7.5 to 9:

Jan 1, 1981

The planning, design and construction of a service tunnel, under a major arterial road between two hospitals are described. The multiple uses of the tunnel combine to provide a favourable cost-benefit ratio for the tunnel as part of a rationalisation of services between hospitals. The driven tunnel provides economic, program and approval advantages over that of a cut and cover construction. The cross sectional shape of the tunnel was determined during feasibility studies by the degree of weathering of the Silurian mud- stones, the shallow cover and the proposed excavation equipment as well as the functional requirements. The site investigations covering geotechnical, building, noise and level surveys provided input to the design documentation and construction phases. Monitoring of settlement levels and primary support stresses during early construction stages confirmed the adequacy of the proposed tunnelling methods and support to be used adjacent to the major building foundations. The 140 metre long tunnel was success- fully completed in April 1980 at a cost of $650,000.

Jan 1, 1981