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By Gennaro Delisle

"Les Services TMG (Niobec) currently treat a niobium ore in the Chicoutimi area of Quebec. The mineral pyrochlore is present in a carbonatite deposit which assays 5% to 7% apatite.The ore is ground to minus 65 mesh and deslimed at 10 micrometres. Part of the carbonates are floated with a fatty acid emulsion and rejected to the tailings as a first step to ensure the proper conditions for the flotation of the pyrochlore concentrate. The carbonate concentrate is 65% minus 44 micrometres and is composed of dolomite, calcite, 22% apatite and a minor amount of silicates.A laboratory investigation was undertaken to determine the feasibility of extracting apatite from the carbonate concentrate. This led to the development of an economic flowsheet to produce high-grade apatite concentrate. The apatite is depressed with phosphoric acid, and the pH is adjusted at 4.8-5.0 with sulphuric acid. An additional amount of fatty acid is added to permit efficient flotation of the carbonates. 1n a second stage, the silicates left behind with the apatite concentrate can be removed by depressing the apatite with AIC3 while floating the silicates with a secondary amine.The results of the laboratory tests indicate the possibility of producing an apatite concentrate assaying 38.3% P2O5 (83.7% BPL) with a recovery of 66.2%. Pilot-plant tests undertaken at the concentrator yielded a concentrate of 34% P205 and recovery of 76%."

Jan 1, 1981

Robert C. Horton was sworn in as the new Director of the Bureau of Mines on October 13 in Washington. Horton, 54, was director of the Geology Division of the Bendix Field Engineering Corporation's operations in Grand Junction, Colorado, before being nominated by - it Reagan for this position. A geological engineer, he has been associated with the mining industry in the West for over 35 years. Legislation to expand the Defense Production Act for 1 year until September 30, 1982, was signed by the President on September 30. The Administration had requested, steel a 5-year extension, but the Congress chose to spend another year considering proposed revisions. The Act, among other things, authorizes the allocation of strategic materials for defense purposes and allows incentives including floor prices and Government financing to encourage domestic mineral projects.

Jan 1, 1981

By Paul R. Bluekamp

The Mather Mine property is composed of a 5.2 sq km (2 sq mile) area within the Cities of Ishpeming and Negaunee which are located in the Upper Peninsula of Michigan. Production in this mine started in 1943; it ran continuously until its closing in 1979, having produced slightly over 55 million tons of iron ore. This mine was a joint venture between several steel companies and The Cleveland- Cliffs Iron Company (CCI), with CCI being the fee holder and operator. For the first 17 years of operation, the ore was shipped in its natural state directly from the mine to the steel mill. By 1960, iron ore pellets were on the market and they proved so superior to the natural soft Mather ore that the latter became difficult to sell except for the coarser portions. It was decided to develop a pelletizing process for the Mather ore, and this was accomplished by 1965 when the first Mather pellets were produced. From this date, all but the coarse fraction of the Mather ore was shipped as pellets. The geological setting is that of a large east-west trending synclinorium which plunges to the west. The lowest member of this trough-like structure is, for the most part, a quartzite-like graywacke, the upper 20 m of which grades into a softer, fine grained slate. Lying conformably on this graywacke footwall member is an iron- formation member which is over 1,300 m in thickness. This iron-formation is cow posed of thin alternate bands of iron oxides and chert and is intruded by a number of diorite sills - some up to 122 rn (400 ft) in thickness. The north limb of this synclinorium dips at approximately 45' and bottoms out at about 1,000 m from surface in the central part of the mine. There are two sets of faults, many of which are intruded by diorite dikes, which trend east-west and southeast-northwest. Displacements are varied, reaching a maximum displacement of 243 m (800 ft). The ore is found lying directly on the slatey footwall and its position is largely controlled by the faults and dikes, with the bulk of the ore being on the upper side of these structures. The ore is composed of soft earthy hematite and martite with vertical thicknesses up to 122 m (400 ft) although the average thickness would be closer to 46 m (150 ft). The ore averaged 60% iron and 7% silica on a dried analysis. The Mather Mine is located on the north limb of the syncline and was worked from two shafts, the deepest of which. was 1,09 7 m. These two shafts are about 2 km apart and serviced a total of 8 working levels between them during the life of the mine. The level spacing was about 61 m (200 ft). The main haulageways were driven parallel to the ore/footwall contact in the hard cow petent graywacke wherever possible. On the lower levels, deeper into the footwall, naked development was common. This material graded into roof bolting ground towards the upper stratigraphic portion. As drifting progressed further into the upper stratigraphic portion of the footwall, progressively stronger steel sets had to be used. From the main haulageway, cross-cuts were turned into the orebody on 61 m (200 ft) centers and extended as far as needed to recover the ore available to that particular cross-cut. The main haulageways and cross-cuts were driven 3 m (10 ft) high, 3 m (10 ft) wide at the top and 4.6 m (15 ft) wide at the bottom. This configuration would accomodate a set composed of a 2.7 m (9 ft) cap on top of two 2.7 m (9 ft) legs angled out at 18'. A sill plate was used under each leg to prevent its sinking into the ore below. The type of steel used was dependent on the expected weight to be experienced. Sets were placed on 1.63 m (5 ft 4 in.) centers; however, in extremely heavy areas, it was sometimes necessary to install sets on 0.8 m (2 ft 8 in.) centers. On the top two levels (5th and 6th), the ore was considerably hard-er and was mined by sub-level stoping and long hole drilling. Very light steel sets were used in the main haulageways and cross-cuts and timber was used in the production drifts. Some con- crete production drifts were installed on 6th level, but proved to be uneconomical. However, as mining reached greater depths, the ore became softer and more massive, reaching its maximum vertical heights on 11th and 12th levels. On levels 7 through 10, yielding steel sets were used extensively in the slusher drifts. While they were satisfactory on 7th and 8th levels, their success diminished with depth and they were

Jan 1, 1981

By B Mohanty

Several parameters such as, Energy, Strength, Brisance, Impulse, and Bubble Energy, are in common use in the explosives industry today to rate commercial explosives in terms of blasting performance. However, this multiplicity of terms, their frequent interchange and their purported one-to-one correlation with expected performance have given rise to a considerable amount of confusion, and have often resulted in unrealistic rating of explosives. In this paper, various theoretical and experimental methods of estimating energy and performance of explosives are discussed and their relative merits examined. These methods include, Computer Code Calculations, Cylinder Test, Trauzl Test, Ballistic Mortar Test, Underwater Test and Model-scale blasting. It is shown, in the light of actual blasting process, that such measures cannot be considered accurate indicators of expected blasting performance as the properties of rock to be blasted are either totally ignored or only partially accounted for in these methods. It is also shown that no single technique alone among those currently employed by the explosives industry can lead to realistic rating of explosives in terms of expected blast results. The need for the adoption of a two-tier rating system, one based on chemical potential "energy", and the other based on blasting "performance" is proposed, and specific approaches to achieve this objective are outlined.

Jan 1, 1981

By Roberto Pimentel de Souza, Helder Zenóbio

VRD started producing pellets in 1969 with one 2-million-t straight grate plant. In the beginning of April 1973, the second unit came on stream with an additional capacity of 3 million t/a of pellets. In January 1977, the third unit started, bringing the annual capacity to 8 million t/a. In the coming year (1978-79), three more units, 3 million t each, are expected to begin production. CVRD's experience with different binders and additives is described. Over the years several kinds of pellets have been produced to satisfy market requirements. General quality information is given for each.

Jan 1, 1981

By Philip C. Good

The Bureau of Mines investigated the feasibility of utilizing low-grade phosphatic shale as a feed for phosphoric acid production. Present practice in the phosphate mining operations of southeastern Idaho is to remove and either stockpile or waste low-grade phosphatic shales as the high-grade phosphate beds are selectively mined. Acid digestion studies were performed on uncalcined shale and on shale calcined to remove free carbon to determine phosphoric acid quality, yield, and amenability to filtration. The untreated shale, containing 21.4 pct P205, was digested with sulfuric acid in a single-tank reactor to yield phosporic acid and gypsum residue. Although 96 pct P205 recovery was attained in digesting uncalcined shale, only the calcined product gave satisfactory filtration rates. Calcined shale yielded phosphoric acid recovery of about 93 pct of the contained P205. It was higher in impurities than commercially prepared acid derived from high-grade western phosphate rock. An acid filtration rate of 745 lb P205/ft2/day was attained. This value is lower than the rate of 900 lb P205/ft that is generally considered to be acceptable in industrial production of phosphoric acid from high-grade phosphate rock.

Jan 1, 1981

By G. C. Thompson

A total system for the simultaneous management of various projects is presented. The objectives of the present organizational structure of the Coal Division in respect of projects are listed, as are the requirements for effective functioning of this organization. The processes involved in strategic planning and in the identification. evaluation, and development of projects are described. The monitoring and control of capital as well as project construction and production management. are discussed. Finally, the important role of project work in the development of management is stated, and the necessity for the continual assessment of the effectiveness of the project organization is stressed.

Jan 1, 1981

By Daniel E. Sullivan

This Bureau of Mines publication examines the potential impact on minerals of the tariff and nontariff agreements that resulted from the Tokyo Round of the Multilateral Trade Negotiations, which were conducted under the auspices of the General Agreement on Tariffs and Trade. These agreements will affect minerals by reducing tariff levels and by changing rules by which they are traded internationally.

Jan 1, 1981

By P. K. Hopke, A. Hubbard, K. H. Leong, J. J. Stukel, K. Nourmohammadi

INTRODUCTION In order to understand the transport and deposition of radon daughters in mine atmospheres, it is necessary to know the variation in the attachment of the daughter atoms to particles as a function of particle size, composition, number density, relative humidity, temperature, and radon concentration, the free gaseous diffusion coefficients of the daughters, and the variation in the mass transfer of the activity, both free and attached to particles, to mine surfaces as a function of particle size distribution, surface roughness of the mine walls, and the flow conditions. If all of these parameters are known in a model system, it should be possible to understand the transport and fate of the airborne radioactivity in real mines under certain well-defined flow conditions. There have been a number of recent investigations of the attachment of radon decay products to particles 1-4, but there are still a number of unanswered questions regarding the process. However, it is clear that for most real mine atmospheres, the vast majority of the activity is attached to particles. The size distributions for the activity-bearing airborne particles have been studied 5,6, and it has been found that most of the activity resides on particles with diameters in the range of 0.05 µm to 0.3 µm with an average mass median diameter between 0.1 and 0.2 µm. The behavior of the unattached radon daughter species has also been recently studied[ 7], and many of the previous problems regarding the value of the diffusion coefficient for Po-218 have been resolved. A major problem in the understanding of the airborne transport of radioactivity in mines is the lack of detailed knowledge of mass tranfer to and fluid flow over rough walls under fully developed turbulent flow conditions. This paper will report the progress on a project that is designed to obtained that information. MATHEMATICAL MODEL DEVELOPMENT Deposition of particles on smooth surfaces in turbulent flow has been extensively studied. A comprehensive review of these results has been prepared by Sehmel 8. There has not been such a comprehensive study of particle deposition on rough walls under such flow conditions. In recent years, only a single model has been proposed to explain such deposition 9,10 and in both of these papers the flow structure in the rough walled pipe was not taken fully into account. As part of the work being conducted on this project, a more complete model was outlined in a previous report [11]. The basic theory will be reviewed to provide a context for the flow measurements to be reported. The flux of particle deposited on the walls of a pipe in a turbulent flow is derived from the one dimensional form of Fick's law as given by [N = Dpdpp/dr (1) where N is the flux of particles deposited per unit area per unit time, D is the total eddy diffusivity of the particles, p is the airborne concentration of particles, and pr is the distance measured from the center of the pipe. The rate of deposition is best expressed by a deposition velocity Vp = NIP pb (2) where P b is the mean particle concentration in the sulk flow. The shear radius, V/ut and the shear velocity, u , are used to calculate a nondimensional distance, and velocity, respectively, where v is the kinematic viscosity of the fluid. The nondimensional form of equation 1 is given by Vd = DP dpp(3) V dr+ where Pp = Pp/ Ppb(4) By integrating equation 3 from the rough wall stopping distance, S , to the center of the pipe, the deposition velocity can be obtained. In order to make this calculation, it is necessary to have accurate descriptions for the particle eddy diffusivity, stopping distance, and shear velocity in order to insure that the influence of the flow structure has been properly accounted for. The shear velocity can be determined experimentally from the shear stress evaluated at the wall, Tw, and the fluid density, ut =VT w/p = ub V f/2 (5) where ub is the mean bulk axial velocity. The wall shear stress for a given pressure drop, dP/dL, and hydraulic diameter, Dh, is]

Jan 1, 1981

By Kenric C. Owen

INTRODUCTION Situated 23 miles east of Pretoria the Premier Mine started diamond production in 1903. Two years later it produced the largest diamond yet discovered, the 3 106 carat Cullinan stone. For the period to 1932 when the mine was closed down due to economic circumstances, open cast mining was practised using inclined rope haulage traction and by this time had produced 48.5 million carats of predominantly industrial quality diamonds from 148 million tons of ore. In 1945 it was decided to re-open the mine using underground mining methods. A system using long hole benching was devised (Hodgson and Sewel 1960). This system was later modified but still forms the basis for approximately fifty percent of production. Following the successful introduction of block caving on associated diamond mines in the Kimberley area (Gallagher and Loftus 1960) this mining method was introduced to the Premier Mine in the late 60's and now accounts for nearly 50% of production. This paper will discuss block caving at Premier Mine in the light of our experience and discussion will be directed to the main geological and structural features of the orebody and host rocks and the related constraints imposed on the mining methods. Reference is made to the long hole benching mining method for comparative purposes. GENERAL DESCRIPTION OF OPERATIONS The ore from the block caving and bench mining areas gravitates via ore passes to a twin haulage, 500m below surface. Electric 13 tonne locomotives, each hauling a train of 10 Granby type cars, deliver the ore to two 42" x 48" (1.06m x 1.22m) jaw crushers. The ore is reduced in size to minus 0.15m before being hoisted to the treatment plant on surface. Hoisting is done with 12.5 tonne bottom discharge skips in a five compartment rectangular shaft using two 3 240 H.P. semi automatic Ward Leonard winders. A small single drum service winder operates in the other compartment. Men and materials travel in a separate shaft in a 5,4m x 2,8m cage operated by a Koepi winder. These two shafts are also the main intake airways to the mine. The two main extraction fans capable of 250 m3/s at 3,2 kPa are situated on surface and are connected to the underground workings by an incline and a network of air passes and return airways. GEOLOGY Although there are numerous occurrences of kimberlite bodies in the district, the Premier Mine kimberlite pipe is the only economic orebody. It is roughly oval in shape with surface dimensions of 900m on the long axis and 450monthe short axis. The surface area is 320 000 m2. The kimberlite has intruded a massive body of felsite and norite which is intersected by a number of faults. The contact between the kimberlite and host rock is clear cut and dips inward at an average angle of 80°. Cutting across the pipe and the country rock is a 75 metre thick gabbro sill, the top of which intersects the pipe below the 347m level and dips at 15' to the north west. The kimberlite has been metamorphosed a distance of some 20m both above and below the sill contacts. Age measurements on biotite from the gabbro date the sill as 1 115 million years, thereby providing a minimum age of the pipe. (Pre Cambrian). All other pipes in South Africa are of Cretaceous age (60 million years). A simplified geological plan and section of the Premier Mine orebody is shown in Figure 1. It is thought that the kimberlite intruded in at least three distinct phases. The kimberlites of these different phases can be distinguished most easily by their characteristic colour. BROWN KIMBERLITE. This is the oldest kimberlite and forms a crescent shape in plan on the south eastern side of the pipe. In depth it increases in relative area. Although the brown kimberlite carries the highest diamond grade in terms of carats per ton all the various kimberlites in the Premier pipe are economic to mine.

Jan 1, 1981

The Rio Tinto - Cerro Colorado operation was the first major cyanide plant built out of South Africa since the Carlin and Corte mills came on-stream in 1965 and 1969. Its flowsheet was presented in the original monograph, "Gold and Silver Cyanidation Plant Practice, " published in 1975, with a brief description of the plant but without the operating costs. The authors have recently received these data and it was believed worthwhile to present them in this volume along with the aforementioned flowsheet. The Cerro Colorado ore deposit is located at Minas de Rio Tinto in the province of Huelva, Spain, approximately 88 km (55 miles) north- west of Seville. The previously mined copper ore bodies lie on either side of the deposit which is mainly copper ore. The crest of the deposit is a capping of "gossan" containing hematite, limonite, gold, and silver. Separate facilities for primary crushing, concentration of copper ores and cyanidation of gold-silver ores are provided. Fine crushing is common to both ores. The run-of-mine open-pit ore is fed from a stockpile via a surge bin and apron feeder to a jaw crusher. After crushing to minus 150 mm (6 in. ), the ore is washed in a rotary scrubber to remove the clay material ahead of fine crushing. Minus 6 mm (1/4 in. ) screen undersize is pumped directly to the grinding circuit with the screen oversize being crushed to minus 19 mm (3/4 in. ) for feed to the primary ball mill; this is followed by two subsequent stages of ball mill grinding to 80% minus 43 microns (325 mesh). Grinding is with water only; no cyanide is used. After thickening, lime is added to the first two of seven Pachuca tanks for conditioning ahead of cyanide addition in each of the five remaining Pachuca tanks. This flowsheet indicates rapid cyanide consumption and the need to maintain a high strength cyanide solution for silver extraction. Agitation for dissolution of the gold and silver is followed by standard CCD practice employing four 67 m (220 ft) washing thickeners in series, with repulping between each thickener. A departure from American and Canadian practice is the use of Stellar filters developed in South Africa for pregnant solution clarification and recovery of zinc-gold-silver precipitates. The precipitate is treated with acid for zinc removal, the sludge filtered in presses, dried, calcined and smelted to produce dore bullion of 900 fineness containing 12% gold and 78% silver.

Jan 1, 1981

By T. D. Brown

"A review of the current status of gasification technology shows that many engineering systems are immediately available for use in the gasification of coal. Widespread international experience can be brought to bear on the atmospheric pressure gasification of all Canadian coals.Unless a high ash content dictates the use of a fluid-bed gasifier, it appears that bituminous coals can be used in gravitating (fixed] bed systems and that the strongly caking coals are most suited to the high-pressure entrained-bed systems . Existing and developing technologies in fluid-bed gasification are attractive for reactive lignites and subbituminous coals.The Fischer-Tropsch based process is the only commercial embodiment of coal liquefaction; the use of Canadian coals in this process depends only on their gasification characteristics. Current developments in hydroliquefaction are directed toward carboniferous coals of medium or high rank. It is not clear that cretaceous coals of the same rank will give similar process yields and economics.IntroductionThe twin coal conversion processes of gasification and liquefaction have a history which dates back beyond the current energy shortages of the western world. Indeed, liquids and gas from coal were produced before this century.Coal gas produced from a coal carbonization process was originated before 1800, and the development of gasification of non-coking coals was patented by Siemens in 1860.This patent represented the first use of air and steam together in the gasmaking process. The pioneering development of coal gasification Using oxygen and steam was demonstrated by Lurgi in the late 1920s. This historical sequence was all carried out at atmospheric pressure, and it was not until the immediate pre-war (1939/45) years that operation of the coal gasification process at elevated pressure became a reality."

Jan 1, 1981

By Robert M. Mann

The Ag/S specific ion electrode has been evaluated for analysis of both free and total cyanide. Metal cyanide solutions were prepared from reagent grade compounds and by stoichiometric fixture of the metal with a potassium cyanide solution. Solutions containing Cu+1, Ni+2, Fe+2, Fe+3 and Zn+2 complexes were analyzed prior to and following EDTA pre treatment. This effort provides an understanding of the effect of each metal cyanide complex on the SIE analysis of free and "total" cyanide.

Jan 1, 1981

By W. R. Bush

Radiation protection in uranium mines and nuclear reactors is compared, with the objective of determining if any radiation protection practices which have worked well in one area can be adapted to the other area, especially from reactors to mines since radiation protection is more highly developed in reactors. Several areas are identified where improvements can be made in radiation protection in mines, the most important being improvement in ventilation at the workplace, greater attention to special monitoring during upset conditions, and a need for a substantial increase in radiation protection staff. It is concluded that, although considerable room exists for improvement, the general approach to radiation protection presently being followed seems appropriate for the circumstances encountered in mining.

Jan 1, 1981

By Raymond E. Kary, John W. Mcklveen

The four unit Cholla coal fired power station at Joseph City, Arizona, consumes coal from the McKinley mine located in western New Mexico. Since coals may contain relatively high uranium or thorium concentrations and may release quantities of detrimental progeny of these elements to the environs, there is concern that releases may pose an adverse environmental impact or even a health hazard to the workers. Thus, an effort was made to quantify the radionuclide content of the coal and compare the results with radionuclide concentrations already present in the vicinity of the plant. Additionally an attempt was made to assess any potential for significant enhancement of the levels of radioactivity in the environment or to plant personnel. Background gamma radiation measurements at and near the plant ranged from 8 to 12 [µ]R/hr (70 to 105 mR/yr) with a few localized anomolies measuring to 18 [u]R/hr (160 mR/yr). Levels on top of the coal storage piles decreased to ~ 6.5 [u]R/hr (60 mR/yr). Apparently the direct radiation emitted from the coal is less than the background levels in the vicinity of the plant. Measurements performed on top of several coal seams at the mine confirmed low radiation levels emanating from coal, but readings more than doubled when the overburden was monitored. Laboratory gamma spectrometry confirmed the lower amounts of natural radioactivity in the coal when compared against soil and water samples obtained in the vicinity of the plant and the overburden at the mine. In an effort to see if there is biological enhancement, animal feces was sampled. The radiation variations were found to be influenced primarily by naturally occurring potassium-40 which varied from ~ 1 to 18 pCi/gm rather than fluctuations in the uranium and thorium series. The potassium-40 concentrated in the ash were similar to the concentrations in the environs around the disposal ponds while Radium-226 concentrated in the fly ash was ~ 3 times that found in the immediate surroundings. Radon-222 gas concentrations in buildings at or near the plant ranged from 0.5 to 1.8 pCi/l and were generally lower than indoor concentrations measured in the Phoenix area. The burning of coal at the Cholla station does not appear to increase ambient levels of radioactivity found in the surrounding environs or to the workers. However, the use of byproduct fly ash for construction purposes may need to be examined.

Jan 1, 1981

By Peter J. Rafferty, Friedrich Steinhäusler

INTRODUCTION The Nuclear Energy Agency plays an active role in promoting international cooperation among its member countries in the field of nuclear energy. In addition to various other functions, it plays a major role in encouraging harmonisation of government regulatory policies and practices, promoting exchange of information, and coordination of research and development in the field of radiological health and safety associated with nuclear fuel cycle activities. The work of the Agency is carried out through a number of specialised standing committees. In particular, the Committee on Radiation Protection and Public Health (CRPPH) is responsible for the Agency's activities concerned with radiological protection and related environmental problems. Its functions include review and discussion of national radiation protection policies and practices, review of developments in radiological protection, interpretation of ICRP recommendations and the study of the means of their translation into practical applications, including the establishment of radiological protection standards. Its functions also include the preparation of technical studies and reviews on specific problems requiring attention, and coordination of further research and development at the international level. Major attention is presently being given to the NEA programme of work on problems associated with radiation protection and environmental impact of nuclear fuel cycle activities, with particular attention to the front-end (uranium mining and milling) and the back-end (waste management) of the fuel cycle. In this context, the increasing attention that has been given in several countries to the problems associated with the exposure of man to radon, thoron and their daughters, and with their dosimetry and measurement, were readily appreciated by NEA, which began an active programme of work in this field in 1976. Because of the detrimental health effects,as demonstrated by epidemiological studies,caused by prolonged exposure to excessive levels of shortlived daughters of radon, particularly in poorly ventilated underground mines, the bulk of attention and needed effort has been focussed on radon and radon daughters in uranium mining. In certain countries some concern has been expressed also about the significant levels of exposure experienced by workers in non-uranium mines,and members of the public who live in particular areas or in dwellings built with particular materials which produce higher than average levels of radon and radon daughters. However, at the time of the first NEA involvement in this field it was considered that one of the most urgent problems to be solved was that of ensuring adequate personal dosimetry for uranium miners. Consequently, the NEA was urged to organise a specialist meeting on personal dosimetry and area monitoring for radon and radon daughters to provide an international forum for exchanging information and reviewing problems in this field. The meeting was held in Elliot Lake, Canada, in October 1976. A second specialist meeting, on the same subject, was held in Paris, in November 1978, to review further developments in this area. These meetings demonstrated that the overall problem associated with exposure to radon and radon daughters had many facets, each of which in recent years has been the subject of considerable attention in many countries for different reasons. It emerged from the meetings that further work was required on a variety of issues in two main areas: 1) dosimetry 2) metrology and monitoring. Conclusions and recommendations which emerged from the two NEA specialist meetings were discussed by the CRPPH. As a consequence the Committee approved, in September 1979, a detailed programme of work in the area of dosimetry and monitoring of radon, thoron and their daughters, and approved the setting up of an international Group of Experts on Radon Dosimetry and Monitoring to undertake the work. The work has been divided into two phases - phase I, on dosimetric aspects, and phase II, on metrology and monitoring aspects. The terms of reference of the work are given in Appendix 1. A list of national representatives on the Group of Experts on Radon Dosimetry and Monitoring is given in Appendix 2; many of these persons are participating here in the Conference. The Group of Experts met for the first time in April 1980 and met again in September 1981. The work of phase I is nearing completion and a report is expected to be submitted soon to the CRPPH for its consideration. Three technical papers providing interim information on the study appear elsewhere in the proceedings of this Conference. The papers cover the three principle areas examined in the study so far: 1) dosimetric aspects 2) review of the "working level" 3) review of objectives and requirements for measurement and monitoring of radon, thoron and daughters. The authors, their affiliations, and the titles of these papers are listed in Appendix 3. A brief overview follows giving the principal results.

Jan 1, 1981

By Erland A. Tillman

The U.S. National Committee on Tunneling Technology under the aegis of the National Academy of Sciences has long recognized that productivity in major underground construction projects could be increased through better contracting practices and better management. Between 1974 and 1978, this Committee has prepared three reports on these subjects. This paper reviews these reports and their recommendations. It covers the extensive actions by the U.S. National Committee on Tunneling Technology to bring the recommendations to the attention of leaders in the underground construction industry and thus further their adoption and implementation. These activities are continuing. Visibility in this session will give these reports and recommendations increased exposure.

Jan 1, 1981

By Charles H. Arnold

This topic addresses historic and current techniques of excavating, supporting and lining tunnels of 4 m (13.1 ft) and smaller. Special considerations relate to the economics of working in close quarters and the sizes of tunnel which achieve the required facility in the optimum economic manner are discussed. The excavation discussion includes TBM, conventional and special techniques for hard rock, soft ground, subaqueous, free and compressed air tunnels. Internal and external support methods discusses timbering, steel ribs, bolting, strapping, shotcreting, grouting, and freezing. Lining methods include cast-in-place concrete, concrete pipe, segmented linings, shotcrete, and others.

Jan 1, 1981

By Bradley C. Bockrath

Some thirty coals were evaluated for liquefaction activity under a mixture of CO and H2 at 425° C. The viscosity of the liquid products was taken as an indicator of the degree of liquefaction. It was found that the direct liquefaction of a particular coal may be beneficially influenced by the presence of pyrite. Active pyrites are often found in the native mineral matter of the coal itself. For example, subbituminous and bituminous coals and lignites of relatively high pyrite content were found to liquefy more readily than similar coals or lignites which were poor in pyrite. The reactivity of inactive coals could be improved by using blends with active, pyrite-rich coals. In addition, pyrite was effective when added separately as pyrite concentrates obtained from coal beneficiation operations. Various other iron- and sulfur-bearing additives were also effective when used in combination as added slurry-phase catalysts, including iron oxides and mineral residues derived from liquefaction of active coals. Under liquefaction conditions, pyrrhotite ( Fe I- x5) is formed from pyrite or from the iron oxides and many sulfur containing species. Pyrrhotite was shown to be a hydrogenation/dehydrogenation catalyst for tetralin but is not active for initiating a variety of free radical reactions that do occur when pyrite is heated with tetralin. The role of pyrite in coal liquefaction is complex and still not well understood.

Jan 1, 1981

By W. K. Hartley

INTRODUCTION AND HISTORY The Kimberley Division of De Beers Consolidated Mines Limited includes four underground mines in Kimberley viz. Dutoitspan, Bultfontein, De Beers and Wesselton Mine and two outside mines, viz. Finsch, an opencast operation and Koffie- fontein, an underground operation. Figure 1 is a map of the Kimberley district showing the location of these mines. Finsch is 150km west of Kimberley and Koffiefontein is 80km south of Kimberley. The head office of the De Beers Company is in Kimberley. The diamond sorting and trading companies for diamonds produced in South Africa are also in Kimberley. Diamonds were first discovered in the Kimberley area at Hopetown in 1866 on the banks of the Orange River. This, and later finds, triggered the start of the great "Diamond Rush". In 1870 the Jagersfontein (now closed) and Dutoitspan pipes were discovered, followed by those at Bultfontein, De Beers and Kimberley Mine (now closed) in 1871, Koffiefontein in 1880 and Wesselton in 1891. Finsch Mine was a more recent discovery in 1961. Finsch, due to go underground in 1986, and Koffiefontein, are fully self-contained with their own treatment plants, housing, services etc. The four Kimberley mines deliver ore by conveyor to a central treatment plant and use joint ser- vices. ,+lining is presently carried out mainly by Block Caving which has replaced the earlier Chambering method. De Beers Mine, on a limited scale, and Wesselton Mine have recently started to use Sub-Level Caving, and both Bultfontein and Du- toitspan are scheduled to use the same method at a later date. The combined ore tonnage produced by the Kimberley Mines is 17 400 tons per day, Koffiefontein 12 000 tons per day and Finsch Mine 17 000 tons per day. THE EVOLUTION OF MINING METHODS Initially, the open-cast mining was carried out to as great a depth as safety would permit. In the early stages of open-cast mining, the Kimberlite was removed in horizontal slices from levels 13m to 20m apart. The rock broken in the workings was raised to the surface by means of endless rope haulage inclines. As lower levels were attained these rope haulages were replaced by aerial hoists, inclined shaft haulages or inclined skip-ways. These methods were dependent upon the strength of the unsupported pipe walls, which became increasingly unstable with depth and a system was adopted whereby the ground mined in open excavations was tipped down passes to tunnels connecting with vertical hoisting shafts sunk in the country rock. As work in the open pit became more and more impracticable due to extensive falls of ground, this system was replaced by a combination of open-cast and underground mining, and eventually by wholly underground methods. Chambering With the advent of underground mining, a system known as "Chambering" (or "Inclined Chambering") was evolved. It was essentially a combination of shrinkage stoping and caving, whereby advantage was taken of the high pressure exerted by the loose rock that had accumulated in the open excavations due to the collapse of the pipe walls. This mining method was introduced around 1890 and was successfully employed, with little change, up to the early nineteen seventies. In this system each mine was served by a main hoisting shaft situated about 300m in the country rock. Main haulage cross cuts were driven from the shaft to the pipe at 200m (600ft) vertical intervals. A sub-vertical shaft, situated close to the pipe provided access to the working levels. Chamber levels were established from the sub-vertical shaft at 13m (40ft) intervals. See figure 2. On each of these levels successive 6,8m wide “cuts" were mined out across the pipe, from one rock contact to the other, and extended vertically to the footwall of the level above. Every 6,8m cut was made up of a chamber and a "back-pillart1, each approximately 3,401 wide. The chamber was mined by overhand shrinkage stoping. The levels were carried "en echelon", with each chamber on a level approximately three cuts ahead of the chamber on the level

Jan 1, 1981