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By S. S. Wang

Dialkyldithiophosphinates are a new class of flotation collectors for base metal sulfide ores. Structurally, they are quite different from the conventional collectors such as mercaptans, xanthates, thionocarbamates and dithiophosphates. Collectors Structures [Mercaptans R-SHS Xanthates R-OCS-M+ Thionocarbamates R-OCNH-R Dithiophosphates 2 R-CCNH-R' Dithiophosphinates (R) 2-PS-M+ ] In the above structural formula, R and R' stand for a hydrocarbon radical, and M+ is an alkali metal or ammonium cation.

Jan 1, 1982

By Joel K. Waterland

GENERAL DESCRIPTION Open cut-and-fill stoping for many years was prob¬ably the most widely used mining method in under¬ground metal mines. Then for a time this method was largely supplanted by the blasthole stope. It again be¬came popular as many mines reached depths or grades where methods requiring large open voids to remain open for extended periods of time became unsuccessful, often as a result of excessive dilution. The open cut-and-fill method is very flexible and is readily adaptable to almost any ore body. The standard application requires that a slice of ore usually 2.4 to 3 m (8 to 10 ft) thick be removed from the back of the stope, and as the ore is taken down, the back is dressed and rockbolted. After the back is secured, the broken rock is removed through rock passes to the level below. When the rock has been removed, the rock passes are extended upward the height of the ore removed, the stope is backfilled, and another cycle is mined. This method is best employed in plunging ore bod¬ies with considerable vertical extent, ore areas that re¬quire selective mining, ore areas where weak wall con¬ditions exist, and ore bodies that have an ore value that will carry this relatively expensive mining method. Blast¬hole stoping, shrinkage stoping, and other mining meth¬ods that do not employ rock passes in a stope are not efficient in plunging or flatly dipping ore bodies because the footwall makes ore removal quite difficult. Since mining is accomplished by taking down slices of the back, only small areas of the wall rock are exposed at any one time, and these only for short periods. Due to limited back height, areas of uneconomic rock may be left in place, or they may be mined and the material gobbed in the stope. Because the miners in the stope must work under freshly blasted areas, the amount of ground control is usually great. The volume of rock that is broken during one section of mining is relatively small and the amount of nonproductive work required is high. This results in limited productivity for the scope and, be¬cause of the nonproductive work that must be done on a regular basis, the production from the stope may be quite cyclical. SUITABLE ORE BODIES The open cut-and-fill method may be adapted to al¬most any type of ore body with a relatively high vertical extent. The ore body must be accessible at both top and bottom as well as at regular intervals throughout its vertical extent. Although adaptable to most ore bodies, the method is probably best employed where the ore has poor con¬tinuity and where most types of bulk mining would pro¬duce excessive dilution. In areas of poor ore continuity, the capability of continuous and extensive sampling dur¬ing the mining of each cycle makes this method very desirable. This capability also minimizes the amount of evaluation sampling that must be done before mining is started. Because of the extractive system used, the size and shape of the stope may be as readily changed as the sampling mandates. Probably the only ore characteristic demanded is that the ore has strength enough to be sup¬ported through the use of rockbolts or cable bolts dur¬ing the mining and backfilling cycles. Good planning, systematic sampling, and careful supervision will pro¬duce a product with less dilution than any other open stoping method. PLANNING Evaluation Planning Once it has been decided that the open cut-and-fill method would be the most efficient for mining a par¬ticular ore body, the next considerations would probably be the availability of an economical backfill material and the selection of an efficient transport system for this material. Although hydraulically transported mill tail¬ings are the most widely used product, this is not always practical due to mill location or the quality of the tailings. In such cases, backfilling may be used. The type of backfill and the type of equipment used will determine if a floor or capping on the backfill is required to minimize dilution during ore removal. The early selection of rock removal equipment is im¬portant since equipment usually determines the amount of development work required to bring a stope into pro¬duction and the size of the openings needed. The size and continuity of the ore body will usually determine the type of loading equipment. The use of slushers or load¬haul-dump (LHD) equipment captive in the stopes will minimize the amount of development. If the ore con¬tinuity is such that a ramp system for extraction can be used, the cost of development will be increased but the flexibility of continuous mining will minimize the cycli¬cal nature of the production. The height of the mining section usually is deter¬mined by the strength of the wall rock and the amount of back bolting required. Once this has been decided, the appropriate drilling equipment can be chosen. The number and sizes of the rock passes employed depends upon the type and size of the extractive equipment and the type of backfill that is to be used. Since the miners must enter and leave the stope each shift, the level inter¬val is usually maintained at approximately 45 m (150 ft). Access from the level above into the stope must be main¬tained at all times. The employees perform all the work in the stope and adequate ventilation must be provided. Stope Planning Due to the flexibility of the method and the vari¬ability of the ore zones, layout is usually done on a stope basis. In areas where continuity is a problem, the size of the stope is usually determined by the boundary of the ore (with all of the ore within that boundary being removed). In areas of good continuity where ramps are to be used, the length of the stope may be determined by the length of time each of the cycles (preparation, back¬fill, mining, and ore extraction) requires. The ramp work is then laid out so that access to the various parts

Jan 1, 1982

By D. R. Babich, R. A. Allwes, L. V. Wade

The design of a longwall entry system is discussed in this Bureau of Mines paper. This is done primarily through the results of an in-mine case study made of two sets of longwall entries utilizing the stiff- yield pillar design concept. The rock mechanics instrumentation plan used in this case study is shown, including a description of the instrumentation layout and the types of instruments used. The analysis of the stress changes in the abutment and yield pillars is presented in detail as the longwall panels were mined. Conclusions concerning when the pillar first started to experience significant stress, pillar stress history as the longwall face approached and moved past the pillars, stress distribution within the pillars and relative stress changes between yield and abutment pillars are discussed. The paper concludes with a discussion of the redistribution of stress in the vicinity of the long- wall panel.

Jan 1, 1982

By Thomas L. Starks

Surface soil/spoil Samples (0-10 cm) from experimental test sites in western North Dakota were aseptically collected during August 1975 and each month (May through September) during 1076 and 1977. Selected samples were Cultured in a defined inorganic medium and after six weeks the algae were identified. Chlorophyll u content of the soil/spoil was used as a measure of the algal abundance. Soil physical-chemical properties, species variety and abundance were statistically analyzed. All test sites revealed distinct soil algal communities and demonstrated a successional trend as evidenced by significant increase, in species variety and bionmass over time. Some soil factors, calcium, copper, and phosphorus affected species variety and abundance at more than one situ. Others were important at specific sites; silicon and aluminum at the South Beulah test site; manganese sodium and zinc at the North Beulah test site No. 1 and boron, cadmium and magnesium at the North Beulah test site No. 2. Topsoiled areas had higher species varieties and abundances when compared with control areas, as did plots treated with amendments such as leonardite and fertilizer Generally.

Jan 1, 1982

Natural gas is a comparatively new fuel in Australia. It was first used in small quantities in Queensland in 1965, but it was not until 1969 that natural gas reached the capital cities Brisbane, Melbourne, and Adelaide in that order; then in 1971 it reached Perth. By the mid seventies, natural gas had secured a significant proportion of the energy markets in Queensland, Victoria, South Australia and Western Australia and in 1976 supply commenced in Sydney. In that year - 1976 - gas supplied about 12.5% of the total demand for primary energy in Australia excluding transport. In the current financial year - 1981/82 - natural gas is expected to supply nearly 19% of the market and forecasts for 1990 range up to 25%. From a starting point in the late 1960's when manufactured gas supplied a small but important proportion of the non-transport energy market - less than 3% - we have exper- ienced an average growth of nearly 24% per annum up to the present time.

Jan 1, 1982

By R. L. Handy

The goal of this research was to adapt the highly successful rock borehole shear test for use in deep, small-diameter boreholes. Operating somewhat on the principle of a rock bolt, the borehole shear test measures both the expansion pressure and pulling force to give discrete measures of rock shearing strength as a function of the applied normal pressure. A new instrument was designed and built to fit into a 2-3/8 inch diameter hole. Unfortunately the design became so complicated, the operation is nut reliable. This is attributed to constraints of the small hole, and added requirements of an in-hole anchor to push against, and a rotating shear plate cleaner for repeated tests down-hole. A clue to the d: vice complexity is its length, which is over 10 feet. Furthermore it employs concentric cylinders and one hydraulic cylinder bored into the piston of another, or it would be even longer. Meanwhile drilling technology has advanced to greatly reduce costs of drilling 5- or 6-inch diameter holes. The downhole tester therefore should be redesigned and simplified for use in a larger diameter drill hole.

Jan 1, 1982

By L. C. Lang

INTRODUCTION The introduction of 165-mm (61/z-in.) holes to underground mining operations has made possible the application of Canadian Industries Ltd's (CIL) vertical crater retreat (VCR) mining method, illustrated in the accompanying sketches. This unique and revolutionary new application of spherical charge technology (see the Appendix), when applied to primary stopes and pillar recovery, eliminates raise boring, slot cutting, and dilu¬tion of ore by backfill; greatly improves fragmentation; reduces labor and time requirements; eliminates uphole drilling and blasting; and minimizes or completely elimi¬nates damages by blasts to the walls and retreating backs of the stope or pillar. If vertical large diameter holes are drilled on a designed pattern from a cut over a stope or pillar to bottom in the back of the undercut, and spherical charges of explosives are placed within these holes at the calculated optimum distance from the back of the undercut and detonated, a vertical thickness of ore will be blasted downwards into the previously mined area. Repeating this loading and blasting procedure, min¬ing of the stope or pillar retreats in the form of horizontal slices in a vertical upwards direction until the top sill is blasted and the mining of a stope or pillar is completed. The VCR method is also applicable to drop raises and has the potential to replace all other raising methods under most circumstances. PILLAR RECOVERY Levack Mine Inco Metals Co., Ontario Div., provided the first opportunity for the method in pillar recovery. The Levack mine's high grade ore pillar No. 4800 on the 975-m (3200-ft) level was used for the production¬scale experiment (Figs. 1 to 3). The pillar was about 49 m (160 ft) long, 6 m (20 ft) wide, and 20 to 26 m (65 to 85 ft) high. The mined area on both sides of the pillar was backfilled with a 1:30 cement:sand mixture. The pillar was removed in two phases. In phase 1, the standard uphole method was used to blast down the 18-m (60-ft) long section of waste from the bottom of the ore into the undercut. From the pillar's top sill, 165-mm (61/a-in.) holes were drilled downward into the pillar, and by measuring the depth of the holes, the results of the uphole blast were determined and roof line 1 was established. The bottom of each hole was plugged, then filled with sand to place the center of gravity of each spherical charge (loaded from the top sill) at a predetermined optimum distance from the horizontal free surface. The charges were then detonated. After detonation, both draw drifts were full of extremely well-broken material. The depth of each hole was measured again, and the plot of these depths re¬sulted in roof line 2. The same blasting procedure was repeated and the resultant new back elevation was marked by roof line 3. The poor results of the initial uphole blast at one location (notice the peak in area 1) appeared to influ¬ence the subsequent new backs. A third blast success¬fully evened the back, and resulted in roof line 4. An unblasted slab averaging 6.3 m (20.9 ft) thick remained below the pillar's top elevation as the final sill. In all three spherical charge blasts fragmentation of the blasted material was extremely fine. The backfill was fully exposed on both sides of the now-blasted pillar. The backfill remained undamaged and the ore was not diluted by sand. The remnants of all the 165-mm (61/2 -in.) holes remained clean and undamaged, and the holes had well-defined bottoms that could be easily measured and plugged. Each blast took down a 3.9 to 4.2-m (13 to 14-ft) thick horizontal slab of ore. Productivity was three times greater than that of the previously practiced cut-and-fill method. Since this was the first such experiment, blasting the remaining 6-m (20-ft) thick final slab was the sub¬ject of some deliberation. If the described method was repeated, we could have ended up with a 1.8-m (6-ft) thick sill unsafe to work on. It was therefore decided to blast the whole sill using two spherical charges prop¬erly placed in each hole, but with the application of

Jan 1, 1982

By Lawrence K. Nordell

This paper reviews practices used In the selection and sizing of belt conveyors and stacker systems commonly used in crushing and grinding plant facilities. Historical and modern methods of sizing this equipment are presented. The new methods Include: 1) probability theory applications for establishing component reliability, 2) "Monte Carlo" simulation techniques used in evaluating alternative system designs and in establishing service factors and corresponding availability Indices of each system, 3) quantifying the system and component operating characteristics using computer assisted modeling tools simulating rigid body and elasto-mechanic behavior for starting and stopping large and/or complex belt conveyors.

Jan 1, 1982

By A. G. Journel

Few Earth Scientists realize fully the assumptions they are making, willing or not, when pooling into a same histogram data coming from different locations and considering such histogram as an estimate of a probabilistic distribution. First they need the hypothesis of stationarity in space; second the presence of spatial auto-correlations implies that data be unequally weighted according to their relative locations and the degree of spatial continuity of the underlying phenomenon. The novel technique of indicator kriging allows construction of local or global histogram which can be considered as unbiased estimators of precisely defined probabilistic distributions, full account being given to the underlying continuity. Such estimates of local distributions can be used in mine planning for estimation of local recoverable reserves.

Jan 1, 1982

By Lewis L. Oriard

INTRODUCTION In recent years, there has been a trend in the direction of larger drilling equipment and larger diameter blastholes. Although this change has improved the efficiencies and reduced the costs in many operations, it has increased the potential for damage to underground openings. In addition, in many instances one now finds more sophisticated delicate instruments, automated control facilities, and a large variety of structures in proximity to blasting activity. The combined effect of larger-scale blasting activity and its proximity to various features of interest is such that there is an increased need for a more refined analysis of blasting effects and their control. BLASTING EFFECTS ON ROCK SURFACES The Breakage Mechanism In order to develop techniques for controlled blasting, one must first understand the features of the mechanisms by which blasting causes rock breakage to occur. These features have not been easy to demonstrate, mostly due to the difficulty in making tests and observations at the high stress levels and short time durations involved. When an explosive charge is detonated, the material surrounding the charge is subjected to a nearly instantaneous, very high pressure [on the order of 1.4 to 13.8 GPa (0.2 to 2.0 X 106 psi), depending on the explosive]. If the charge is coupled to "average" rock, this pressure will pulverize the surrounding rock for a distance on the order of 1 to 3 charge radii in hard rock, and to a greater distance in softer rock (this is also dependent on the type of explosive). As the pressure wave passes into the rock, high tangential stresses cause radial cracks to appear, and the nearly discontinuous radial stress zones gen¬erated by the shock front may cause tangential cracks to appear. The extent of these cracks depends on the energy available in the explosive, how quickly the energy is transmitted to the rock, and the strength properties of the rock. The discontinuous shock front is quickly dis¬sipated, but the expanding gases generate a longer-acting pressure. A compressive pulse travels to the nearest face or internal rock boundary where it is reflected in tension. The tensile strengths of most rocks are roughly 40 to %o of their compressive strengths, so the rock may now fail in tension whereas it may have been able to support the diminished compressive phase without failure. The ten¬sile deflection typically produces a failure described as tensile slabbing or scabbing. Laboratory experiments and field experience have pretty well established that several mechanisms are involved. These include (1) the classical case of tensile parallel slabbing when the pressure pulse is reflected at a free surface; (2) failure under quasi-static compressive loading (the shape is normally irregular due to discontinuities in the rock); (3) radial cracking under the action of tangential stresses at the periphery of the expanding pressure pulse; (4) peripheral cracking at the discontinuous shock front which is quickly dissipated; and (5) additional mass shifting due to the venting of the explosive gases. The first three items have received much attention in the laboratory and the literature. The complex effects of gas venting are difficult to test in the laboratory because of the difficulty in reproducing the many weak planes and discontinuities typical of most field conditions, which play such a prominent role in determining the behavior of the rock mass subjected to blasting. Unfortunately, gas venting effects can be pro¬jected to significant distances under certain field conditions, and are sometimes difficult to control. It is not unusual for gas venting to be the overriding factor in determining the final geometric shape and physical condition of the finished excavation. Sources of Damage For the purposes of this discussion, damage includes not only the breaking and rupturing of rock beyond the desired limits of excavation but also an unwanted loosening, dislocation, and disturbance of the rock mass the integrity of which one wishes to preserve (such as mine pillars, underground openings, etc.). The sources of damage include, of course, all those physical features of the rock breakage mechanism. Each of these effects must be limited to the desired zone of breakage and excavation if the integrity of the remaining rock mass is to remain undiminished. The primary zone of rock breakage usually can be controlled in the normal process of field experimentation to determine proper charge sizes and location for primary excavation. However, it frequently happens that there is damage from sources which are more difficult to account for in the design process, which are often overlooked. These are (1) the overbreak due to poor drilling control, (2) dislocation of rock (mass shifting) due to venting of explosive gases, and (3) loosening or dislocation due to the influence of seismic waves (ground vibrations). CONTROL OF ROCK BREAKAGE Importance of Geometry In studying the rock mass and blasting design con¬siderations, it is important to keep in mind the geometric relationships among charge size, shape, and position, and the physical features of the rock mass to be preserved. The features of principal interest are the external shape and position of the rock mass relative to blasting, and the position and attitude of any weak planes in the rock mass. The Sequence of Blasting and Excavation Events Unfortunately, there are too many times when the task of preserving delicate rock is considered hopeless, and because of this attitude, no further effort is ex¬pended towards caution or control. In such cases there is often a failure to recognize the importance of the se¬quence of the procedures. Attention to this can greatly reduce unwanted effects at minimum cost. Perimeter Control The requirements for perimeter control are highly dependent on the special needs of each particular proj¬ect. The desirable degree of control is a highly variable

Jan 1, 1982

By J. W. Howland

A demonstration of dozer blades, designed for recontouring area-mined spoil and shown effective at Kansas and East Texas surface mine sites, was conducted at a Northeast Arizona surface coal mine to determine their effectiveness in different climate and soil conditions. The specialized tools were: 48-ft angle blade (on side-by-side tractor), 24-ft grading bar, modified 24-ft "U" blade (on side-by-side tractor), and a 13-ft deep "U" blade. Interest in the increased dirt-handling capacity of a previously built 22-ft blade, resulted in the design of a new Centerflow Blade of good versatility, able to maintain increased flow in otherwise difficult-to-move spoil. The demonstration involved 263 acres and showed a reduction of 50% in reclamation costs and 20% in fuel requirements for this mine site over conventional equipment. This savings resulted from improved technology, trained operators, and rigorous application of supervisory and monitoring techniques. Procedures and results of the demonstration are presented and analyzed here, along with guidelines for application of findings.

Jan 1, 1982

By F. J. Arriagada

Where gold is present in a finely disseminated form within sulfide particles, the porosity and surface area of the calcined material strongly influence leaching since they control the penetration of solution inside the gold-bearing grains. On the basis of a stability diagram for the Fe-O-S in the temperature range of 400 to 750aC, the expected sequence chemical transformations associated with the oxidation of pyrite is discus-sed. By using analytical techniques which include scanning electron microscopy, X-ray diffraction and BET gas adsorption, the chemical and structural changes experienced by auriferous pyrite-arsenopyrite concentrates and pure pyrite samples are studied under an oxidizing roasting atmosphere of com-position 80% N + 20% O2 in the temperature range of 450 to 750°C. The results indicate that as the roasting temperature increases, the surface area of the calcines decreases sharply after a maximum value attained at about 450-500°C. SEM micrographs show that the needle-shaped tips which characterize the surface of the oxide product become wider and the size of the pores between them is decreased when the roasting temperature is in-creased from 550 to 750°C. This effect suggests that the formation of a more "closed" hematite product at higher temperatures is due to sintering associated with a faster growth and crystallization of the oxide. A strong correlation is shown between increasing roasting temperature and increase in gold lost in cyanidation tailings, an effect which may be partly attributed to the loss in surface area of the calcines.

Jan 1, 1982

By Larry D. Nichols

Poroplastic membranes containing ion exchange liquids were studied for separating cobalt from nickel-containing solutions similar to ammoniacal laterite leaches. The driving force for cobalt transport was differences in pH and composition between the leach and extraction streams. Using a diketone ion exchange liquid, a leach pH of 6.5 to 7.0, and a strip pH of 2.3, cobalt was extracted from a 750 ppm solution containing 2500 ppm of nickel at rates ranging from 1 to 10 µg/cm2 .min in single membrane cells, and from 10 to 100 in more closely spaced multiple membrane cells. Rates were variable, with evidence that cross -membrane flow pasterns were a key factor; a reproducible rate of 30 pg/cm .min would be economically viable. Loss of transport observed after several days of continuous operation was restored by flushing membranes with fresh ion exchange liquid. A major effort was devoted to preparation of thin, leak-free membranes; the resulting membranes endured operating conditions without mechanical failure.

Jan 1, 1982

By D. J. Barrett, M. A. Sochocky

INTRODUCTION Summary of Factors Influencing Comminution Circuit Design The selection of an appropriate comminution circuit for a specific ore is one of the most important decisions in the design of a processing plant. The importance is related to the fact that the capital and operating costs for crushing and grinding plant generally represent, as is well known, the major portion of the plant costs. This chapter will, therefore, identify as many as possible of the factors which influence the selection of a comminution circuit. The number of factors which will have to be considered for any comminution project will depend, among other things, on the type of project. If, for example, the project is an entirely new mining development, then most, if not all, the factors identified will have to be analyzed. If the project is an expansion of a developed mining operation or if the new orebody is adjacent to a viable operation, then the existing operation will influence the choice of factors or design constraints which need to be considered. The existing infrastructure or supporting systems may be quite adequate to accommodate the requirements of any expansion or new development. Hence, those factors may not have to be considered. To cover, however, the widest spectrum of projects, the factors which have been identified in this chapter are those which will have to be considered for the most complex of projects, that of developing an entirely new mining operation. The range of factors is as broad as the range of different ore types. Thus, each development must be treated with individualized care. The most important step in the development is the analysis and understanding of the ore characteristics. The hardness of the ore, its abrasiveness, friability, moisture content, grade, mineralization including gangue mineralization, liberation size, chemistry, and other characteristics should be the first to be analyzed in the investigation. The next should be the determination of factors such as plant size, throughput, location, climate, accessibility, and availability of water. The selection of samples for initial bench scale work and bulk sampling (when needed) for further

Jan 1, 1982

By S. V. Krishnan

In the upgrading of finely-disseminated iron ores, selective desliming is the critical step which must be controlled in order to achieve efficient flotation. A prerequisite for selective desliming is a properly-dispersed pulp; sodium silicate is commonly used as a dispersant. The mechanism by which sodium silicate acts as a dispersant in the presence of calcium ions was examined by streaming potential measurements, settling tests, abstraction density determinations, selective flocculation tests and scanning electron microscope observations. The study established that calcium silicate precipitates on the mineral surfaces dispersed the pulp. Although sodium silicate was a nonselective dispersant, the addition of corn starch aided the flocculation of goethite over quartz at low silicate additions, but became ineffective at higher silicate levels. A model is presented to explain the dispersant mechanism of sodium silicate in the presence of calcium ion for selective flocculation of iron ores.

Jan 1, 1982

To develop and design a bit that will rapidly drill smooth and precise instrumentation holes in soft rock. Approach Use a one-piece bit made by combining a drag bit and a diamond bit. Because a smooth-walled borehole with a precise diameter is needed for borehole pressure cells (BPC) and other borehole instruments, a diamond bit is usually used. However, diamond bit drilling is slow, and the bit will tend to bind when squeezing ground is encountered. On the other hand, the drag bit will cut faster, but the walls of the drill hole will be rough and the diameter will not be as precise. The combination will provide a bit which penetrates as rapidly as a drag bit, but retains the smooth-wall and precise borehole diameter of a diamond bit. How It Works A three-foot-length of EW-size drill rod is centered in a diamond bit by welding it to three equally-positioned spacers at each end of the bit. A drag bit is attached to the EW rod with a standard EW-to-drag bit adapter as shown in the diagram. Three equally-spaced diamond inserts are set to the desired hole size and welded to the diamond bit barrel behind the crown. These inserts increase the smoothness and precision of the hole by cut-ting away any hole closure. Three equally-spaced shims are set approximately 0.020-inch under the overall insert diameter size and welded to the rear of the diamond bit barrel. The combination bit should be used on a diamond drill with either water, mist or air as the coolant.

Jan 1, 1982

By Thomas D. Boyce

Minewater from many western U.S. uranium mines often contains radioisotopes in concentrations higher than permissible for discharge to unrestricted areas. .Radium 226 has the lowest permissible concentrations of all radioisotopes and usually greatly exceeds this limit in untreated effluents. This radionuclide is derived from the natural radioactive decay of uranium: [ ] The treating of minewater effluents has been studied previously (1) and the classical removal method for radium has been adsorption on barytes or freshly precipitated barium sulfate. This paper describes the removal of radium from minewater by utilizing a selective complexer based on a crosslinked organic polymer matrix.

Jan 1, 1982

By Walter E. Dean

Deep-sea manganese nodules are considered to be potential ores of manganese, nickel, cobalt, and copper. Considerable time, effort, and funds have been applied to the study of Se distribution of nodules on the sea floor and the location of economically important areas, but little effort has been spent in understanding the origin of nodules. Most theories that have been proposed to explain the origin of manganese nodules have been based on inorganic physicochemical processes. It is not by coincidence, however, that the area considered to be most economically feasible is in the zone of high biological productivity in the eastern equatorial Pacific Ocean and underlain by siliceous biogenic sediment. The genesis of manganese nodules and the concentration of many trace elements are intimately associated with biological processes.

Jan 1, 1982

By Anil Mahyera

Towards the goal of achieving uniform-high-tensions in roofbolts to provide superior roof support[1-7] for extended periods, Terra Tek installed mechanical-anchor bolts in a mine in eight different ways -- under torque/thrust control with and without hardened washers, belleville washers and tension equalization; in a conventional manner with and without a hardened washer. The various bolt groups were analyzed in terms of their tension uniformity and mean tension levels. Installing the bolts accurately with ? a hardened washer under the bolthead, and ? proper torque/thrust control resulted in the mast uniformly tensioned bolts -- with a coefficient of variation of only 14%, an improvement of 3 times over a conventional installation. Also, these bolts maintained a high tension level -- 80% of the installation tension, compared to 50% achieved conventionally.

Jan 1, 1982

By Paul R. Thomas

The Bureau of Mines evaluated the impact of the depletion allowance upon the economic availability of copper from domestic deposits. The potential long-run percentage depletion deduction, cumulated over the life of the properties, was estimated at approximately $2.5 billion for producing properties and up to $11.7 billion for nonproducers. The income transfer, in the form of higher Federal income tax payments if the allowance were repealed, was estimated at $1.5 billion for producing properties and up to $8 billion for nonproducers. The percentage depletion allowance, through its role in determining Federal income tax liability, was found to be very significant in affecting profitability. Repeal of the allowance could result in major domestic copper availability reductions at all price levels, now and in the future, as a result of reduced profit margins for current producers, and postponement of development decisions for nonproducers. It was demonstrated that a 50-percent reduction in the percentage depletion rate applicable to copper would also result in a significant decrease in profitability for nonproducing properties. For producing properties, the percentage depletion rate on copper was shown to be less significant than the income limitation that is placed upon the total amount of the depletion deduction. Repeal of the minimum Federal tax provision was shown to be significant in enhancing profitability for both producing and nonproducing deposits. The report concludes that the percentage depletion allowance operates as an important method of capital recovery and production incentive.

Jan 1, 1982