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Under the Town and Country Planning Act 1977 the Ministry of Works and Development has a responsibility for providing information to local and regional planning authorities on public works and matters of national development and importance, for the exercise of their responsibilities under this Act. One such matter which has recently become a source of concern is the development of South Island lignite. In 1979 the Ministry of' Energy produced a position statement outlining a programme of' research leading to the possible extraction of lignite in Southland or Central Otago for conversion into about 50 per cent of New Zealand's liquid fuel requirements by 1995. Understandably this possibility has caused considerable concern at the local level, expressed particularly by bodies such as Federated Farmers and the local authorities. In the absence of any factual knowledge short of mere conjecture been able to recommend to local authorities is that they limit their planning period to ten years. After this time more information regarding the size, nature, location and certainty of the proposal will permit a more accurate assessment of the implications of the York. While on holiday earlier this year I took the opportunity to visit opencast mining areas in Germany and England to gain an appreciation of the scale of impacts involved and discuss with those concerned the procedures used in planning for such operations. What follows is a brief description of these operations, a comparison with the likely development in this country and the extraction of what may be some relevant principles of' operation for the planning of an environmentally acceptable project. Please note that the views expressed in this paper are mine alone and are not necessarily those of the Ministry of Works and Development.

Jan 1, 1981

By Watson F. G

As an introductory paper of a series on the Melbourne (Australia) Underground Rail Loop extension to the extensive electrified suburban railway system the paper outlines the need for the Loop, legislative background, organizational establishment, engage- ment of consultants, design control, contracts, financial aspects, example of organizational flexibility and certification and transfer of the Loop to the Victorian Railways Board.

Jan 1, 1981

By Eugene F. Casey, John G. Ruggiero

INTRODUCTION Construction of tunnels in urban areas, especially reclaimed areas, pose many problems. The Red Hook Intercepting Sewer Tunnel, Contract 1A is a portion of the Red Hook Water Pollution Control Project located along the Brooklyn waterfront in New York City. This contract required the construction of a 2620 meter long (8600 L.F.), 3 meter (10-foot) diameter soft ground tunnel, eight Tunnel Manholes, nine Regulator structures, four Diversion and Tidegate structures and 765 meters (2500 L.F.) of branch intercepting sewers of diameters between .30 meters (12 inches) and 1.8 meters (72 inches). The construction contract bid price was $62 million with a duration of three years. The tunnel route crosses over five sets of operational twin cast iron subway tubes with clearances ranging from 4.3 meters (14 feet) to as small as 0.35 meters (14 inches). The route of the tunnel traverses historically interesting areas which have been reclaimed over the last 200 years. This paper is a case history describing the contractor's methods utilized to successfully advance the work through the above problem areas. LOCATION The Red Hook Intercepting Sewer Tunnel is located in the Borough of Brooklyn of New York City. The route is

Jan 1, 1981

By R. L. Loofbourow, Robert R. Miller

Deep, stable, unlined space promises secure, efficient, acceptable solutions to several current problems. With truly favorable conditions, it can be economic as well. The basic requirement is sufficient sound, tight rock at required depth. This and unusual surface requirements point to special testing. An objective is to describe conditions for planners, estimators and contracts. A first stage concept study needn't be prolonged but should overlook nothing significant. Although observations and tests on core are considered essential, time and expense can be cut by supplementing then with tests on walls of rotary or percussion holes. In some places initial drilling could be shallow.

Jan 1, 1981

By Konishi Y, Morioka Y, Okabe K

A part of the results of dissection survey of Chiba No.1 blast furnace, Kawasaki Steel Corporat- ion, is presented. Distributions of coke size and of slag and metal holdup are shown in the region of the dropping zone. A similar relation to the irri- gated packed bed is obtained. The distribution of coke size, coke temperature and liquid holdup are measured in and around the raceway. The shape of raceway in the horizontal and vertical sections indicates that each raceway worked individually. Liquid flow in the dropping zone and around the raceway on the basis of holdup measurements is Discussed.

Jan 1, 1981

By David K. Denton

Results of the mineral resource study of the Upper Priest Lake RARE II study area indicate a low potential for the discovery and economic recovery of silver and lead in the Prichard Formation near contacts with the Kaniksu Batholith in the vicinity of the Mountain Chief Mine (fig. 1). Silver and lead minerals locally occur disseminated and in irregular quartz fissure-fillings in the Prichard Formation. Between 18^6 and 1967, 105 lode raining claims were located in the study area. The area was designated for "further planning" in the Final Environmental statement but was later changed to "nonwilderness" by President Carter after work was scheduled for the area.

Jan 1, 1981

By Bernard L. Cohen

INTRODUCTION The most serious occupational radiation episode in U.S. history was the exposure of uranium miners to radon daughters in the 1945-1968 time period. Although the maximum permissible dose at that time was 12 WLM (working level months) per year, average exposures were several times that figure, and ten times higher exposures were not uncommon. Starting in the early 1960s, mounting epidemiological evidence appeared for an excess incidence of lung cancer among these miners; up to 1974, there was an excess of 134 cases (159 observed vs 25 expected) among the 4000 in the group under study (NAS-1980). When this excess became apparent, the Federal Government took jurisdiction, greatly tightened enforcement, and in 1969 lowered the maximum permissible exposure to 4 WLM per year. These standards were met largely by a great improvement in ventilation. As a result of these measures, average miner exposures were reduced 5-fold from 1965 to 1968 and by another factor of 3 by 1970. In 1978, the 4 WLM/year maximum permissible exposure was exceeded by only 37 of 7500 underground miners, for only 16 wars it exceeded by more than 25%, and for only 8 was it exceeded by 502 (AIF-1980). The average exposures to the various groups are listed in Table 1. We see that the average exposure for all miners was 1.03 WLM, and for those who worked essentially full time underground it was 1.45 WLM. For purpose of later discussion, it will be convenient to choose a single value for average exposures. In view of the fact that employment situations and job categories are bound to vary over a lifetime of work, we take this to be 1.3 WLM/year. The purpose of this paper is not to dwell on history, but rather to address the question of whether or not this present situation is satisfactory. This question was considered recently by a study group under the auspices of National Institue of Occupational Safety and Health (NIOSH-1980) and it answered the question in the negative, concluding that the maximum permissible exposure should be substantially reduced. However, their arguments were incomplete and were lacking in perspective. The principal thrust of this paper is to provide some of the material overlooked. RISK TO URANIUM MINERS FROM RADON EXPOSURE In order to quantify the risk to uranium miners under present working conditions, it is first necessary to estimate the risk of lung cancer per WLM of exposure. The most straightforward way of doing this is to use the data on the group of uranium miners under study. These are listed in Table 2 for 8 exposure ranges and for the total group (NAS-1980). If all data are given equal weight, the risk is seen to be 3.5 x 10-6 per year per WLM. Much can be said in favor of using this value for the risk as dose independent; it is within one standard deviation (SD) of the observations in 5 of the 8 dose categories (as expected from the definition of SD), and within 2 SD for all 8, and for the three cases differing from the mean by more than one SD, two are above and one is below. However, the uranium miner exposures in the present situation will be far below those covered in Table 2, so it is worth considering the possibility that the risk differs from this at low doses. From the last column of Table 2 we see that the experience from all exposures below 600 WLM indicates a risk higher than the mean for the entire group by more than two standard deviations (the value 600 WLM in this comparison was deliberately chosen to maximize this deviation; for exposures below 360 WLM and 840 WLM, the excess is only 1.0 and 1.6 standard deviations respectively). In view of this tendency for the risk to be higher at lower dose, we will take the risk to be 5.0 x 10-6 per year per WLM in what we will refer to as model A. This risk estimate is based entirely on exposures much [higher] than those that will be experienced by

Jan 1, 1981

By A. V. Petty

In accordance with its objective to maximize minerals recovery from secondary domestic resources, Bureau of Mines conducted research on recycling chrome refractory wastes. Since 20 pct of the U.S. demand for imported chromite is used in the production of refractories, primarily for the steel, copper, and glass industries, samples of used chrome-containing refractories from copper converter and reverberatory furnaces were investigated to determine their potential for recycling. The samples were beneficiated using a combination of wet magnetic separation, flotation, and leaching techniques, and the concentrates were reformed into briquets for refractory evaluation. Small refractory test specimens produced from the beneficiated concentrates gave modulus of rupture values at 1,350° C comparable to those of commercial mag-chrome brick of similar composition; however, grinding the waste refractories to at least minus 65-mesh was required in order to liberate the metallic copper. Most of the copper was easily recovered using conventional beneficiation techniques, but the finely ground chrome concentrate would require briquetting, calcining, and crushing to produce a coarse grain suitable for recycling into secondary mag-chrome refractories.

Jan 1, 1981

By James Lee Hibbard, Doraiswami Ramkrishna

A population balance model accounting for the transient variations in drop size distributions in a stirred liquid-liquid system is analyzed for its effect on chemical reaction conversion in a stirred, batch reactor. The analysis is applied specifically for phase transfer catalytic reactions, which indicates that transient effects of evolution of drop size distributions may be important in such systems.

Jan 1, 1981

By Bennet A. G

As part of the construction of the Melbourne Underground Rail Loop it was necessary to construct two 7 metre diameter tunnels, each 100 metres in length, by shield driven excavation through an alluvial valley at Elizabeth Street. The tunnels had a primary support of bolted precast reinforced concrete liner blocks and were excavated with the aid of compressed air. The work was undertaken by the Tunnelling Group of the Melbourne and Metropolitan Board of Works as nominated sub-contractor to John Holland (Constructions) Pty. Ltd. The paper describes some of the problems associated with the construction of the above tunnels and the methods used to overcome them. In particular

Jan 1, 1981

By V. Coffin, T. Cienski

"(Figure 1) Reagentized pulp enters into the column at about 1/4 of the distance from the top. Inside the column it meets with a downward flowing stream of water and with an upward flowing mass of air bubbles. Flotation takes place in this counter current system while at the same time fine gangue particles are washed off the froth before it reaches the overflow lip at the top of the column. The main operating features of the column flotation are as follows:a) No moving parts. Solids are kept in suspention not by an agitating action of a propeller but by settling down against the rising air bubbles.b) Wash water. Spray pipes are installed a few inches below the froth surface and deliver required volume of water. The purpose of this water is twofold:(i) it keeps in balance the flow of material through the column and(ii) it washes rising froth in the froth layer as well as removes unwanted fines attached to the bubbles at the lower flotation section. As a result the concentrate produced is of higher purity than in the conventional flotation. Column is characterized by very high up-grading. However, even temporary disruption of wash water supply results in a drastic drop of concentrate grate.c) Flow balance. (Figure 2) For the correct operation volume of the underflow (tail) is larger than that of the feed. The difference is made up by the wash water. This difference called bias is predetermined and ranges from some 4 to 15 USGPM depending on the size of the unit and process requirements."

Jan 1, 1981

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