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By Egon Pohl, Friedrich Steinhäusler, Werner Hofmann

INTRODUCTION Anticipated increase of nuclear fuel production in the future coincides with growing concern about the occupational health risk of miners from inhaled radon decay products. As a consequence it has been suggested to even lower presently permitted exposure levels (NIOSH, 1980) and implement more stringent control on measurement programmes. Compliance with these regulations requires large investments in new or modified ventilation systems as well as in increased expenditure for staff and instrumentation for health physics operational monitoring. Since both costs are directly related to the overall ore production costs, this can have far reaching implications with regards to the economic feasibility of certain mining operations and consequently reduce estimates of low-cost minable ore reserves. In addition there is increasing evidence that the radon problem is not limited to the nuclear fuel cycle only, but can also represent a significant hazard to non-uranium miners. A key component in the cost-effective implementation of legislative control measures is the monitoring system employed. The choice of system is decisive for the total costs of installation and maintenance, manpower requirements and accuracy of nuclide determination. In the following operational specifications are defined for monitors in mining and milling environments. Different types of available instrumentation are discussed with regard to their suitability for practical radiation protection in underground mines, open cut mines and mills. ATMOSPHERIC CHARACTERISTICS Underground mines Radon in the air of underground mining operations is exhaled from surrounding rock surfaces, crushed material and, to a lesser extent, from water seepage. Whilst in uranium mines radon releasing ore bodies are generally localized in distinct areas, radon sources in non-uranium mines can be very dispersed throughout the system of mine tunnels. The ventilation scheme used influences the absolute atmospheric level of radon as well as the equilibrium conditions between radon and its decay products (factor F). In uranium mines mechanical forced-air ventilation is normally the only way to achieve and maintain legally required nuclide levels. This causes the F-factor to be rather low, e.g. in French CFA-mines F is of the order of 0.2 (Francois, Pradel, Zettwoog, 1975). At the same time the fraction of unattached radon decay products (fp) can increase due to the high air velocities employed. However, it is possible to find non-uranium mines with either natural draught ventilation only or assisted on demand by forced air ventilation during special operations or climatic conditions. Thereby the F-factor is more dependent on seasonal changes of temperature differences between outdoors and mine atmosphere and work routines. As a result F can cover a wide range from 0.02 to 0.95 (Steinhäusler, 1976). The use of filters or electro-precipitators in mine ventilation systems can modify the atmospheric characteristics twofold as it generally decreases the content of radon decay products, but at the same time increases the content of the unattached fraction fp . Average concentration levels of radon decay products are mostly lower in mechanically ventilated non-uranium mines than in equally ventilated uranium mines and are below 0.3 Working Level (UNSCEAR, 1977). However, some working places in non-uranium mines, specially with only natural draught ventilation can occassionally approach maximum permissible levels as defined for uranium mines (Strong, Laidlaw, O&apos;Riordan, 1975; Snihs, 1976; Sciocchetti, Scacco, Clemente, 1981). Open pit- and surface mines Radionuclide levels of radon decay products in the atmosphere of these mines are mostly too low to represent a significant inhalation hazard for miners, ranging typically from 0.03 to 0.1 Working Level (Steinhäusler, 1976). However, personnel using airpurifying respirators or working in cabins ventilated with filtered air can be exposed to a radon atmosphere with low value for the F-factor (F [<] 0.1) and high fp-value up to 80 % (Leach and Lokan, 1979). Mills Atmospheric radon concentration in crushing, grinding, drying and packing sections depends on the radium 226 content of the ore, ore storage methods and ventilation system employed. Providing adequate ventilation ([>] 2 air changes per hour) and control of dust production radon and its decay products represent only a minor problem (Saconney, 1979). MONITORING OF OCCUPATIONAL EXPOSURE Objectives Operational monitoring of the working places provides information on: - confirmation of appropriate control of the routine mining methods employed - indication of abnormal conditions. Although this type of monitoring enables the location

Jan 1, 1981

By J. A. Finch, Zhenghe Xu, Z. A. Zhou

A particle-bubble collision model was used to investigate the minimum recovery zone height, H(r,min) required for flotation columns. The criterion used to define H(r,min) is that height which gives unit collision probability. It is shown that H(r,min) is determined by particle and bubble size, gas hold-up and bubble-bubble interaction. Under typical operating conditions, for flotation of particles less that 10 micrometres in size H(r,min) is greater than 10 m, but it reduces to less than 1 m for particles greater than 100 micrometres

Aug 1, 1995

By M G. Hawtin

The Rocklands copper project is a multilode, iron oxide copper gold (IOCG) + cobalt mineralised system located in the Eastern Fold Belt of the Mt Isa Inlier. Rocklands orebodies cover a range of oxidation zones from oxide, to supergene and primary. Each zone contains different economic minerals of interest. The supergene zones contain significant amounts of native copper and the primary zones are dominated by chalcopyrite. The juxtaposition of these oxidation zones and minerals make it an interesting mineralisation system.Current mine production sampling collects a nominal 3 kg drill chip sample from a 1 m composite, RC drilled interval. A site specific sampling procedure was needed to quantify introduced sampling error when taking grade control samples. Due to oxidation and mineralogical variation in ore zones we needed to develop a practically achievable sampling methodology that ensures minimum sample mass and sample error requirements are achieved to ensure ‘representative’ samples for all ore types.These issues are planned to be resolved using an experimentally derived method using principles from the Theory of Sampling. This method uses an experimental calibration of parameters K and a for Gy’s fundamental sampling error formula using a modified 32-piece sampling tree experiment to compile sampling nomograms for the primary and supergene ores. This paper demonstrates the process undertaken to improve the quality and standard of existing mine production sampling procedures at the Rocklands operation.Results indicate a sample mass of 5 kg for RC drill chips is appropriate for both the primary and supergene ores to achieve a fundamental error below 0.01 or 10 per cent relative error, which is slightly higher than the existing sampling protocol of 3 kg. An additional experiment is recommended to compile sampling nomograms using the exact laboratory comminution and mass reduction stage information to assess the use of a 3 kg sample. The results also indicate no changes are required to existing sampling protocols when sampling RC drill chips in primary and supergene ore at Rocklands.Sampling nomogram results also provide a useful ‘rule of thumb’ reference guide to other mining and exploration companies to help understand minimum sample mass requirements and sampling error during exploration drilling, resource evaluation and mining production stages when targeting similar IOCG mineralised systems. A sample mass of 5 kg is recommended when RC drilling similarly IOCG mineralised prospects and deposits in the Cloncurry area. Future test work is recommended when additional open cuts become active as well as including resource drilling drill core to provide better three-dimensional sample representivity.CITATION:Hawtin, M G, 2017. Minimum sample mass from reverse circulation grade control drilling at Rocklands copper project, in Proceedings Tenth International Mining Geology Conference 2017, pp 337–346 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 20, 2017

By R. Mohan Srivastava

"The goal of estimation is considered in the light of decision analysis. A simple example reveals the essential elements of decision-making in the face of uncertainty. A second example shows that these elements are also key features of many mining decisions. These two examples cause us to revive an old probabilistic notion that the goal of estimation should be to quantify what is unknown. The conditional probability distribution provides the complete statement of one&apos;s uncertainty. Once the conditional probability distribution has been established, one should retain as an estimate the value which minimizes some objective loss function. A third example shows how the ""best"" estimate depends on the chosen loss function. Current practice is almost exclusively based on the least squares method, which assumes a loss function which penalizes underestimation and overestimation equally. This is unfortunate not only because symmetry is an unrealistic assumption but also because mining offers a rare opportunity to state the loss function directly in terms of profitability. Both the theoretical framework and the practical tools exist to allow us to reconsider the conventional notion of the ""best"" estimate. IntroductionThis paper looks at estimation in the mining industry from the point-of-view of decision analysis. Virtually all decisions depend on estimates since few of the important parameters are ever known exactly. In this paper we will be referring mainly to the estimation of ore grades, though the approach discussed here is generally applicable to any decision in which relevant information is uncertain.In the first section we look at a very simple example of decision-making in the face of uncertainty. We will then tum to a mining example and see that the essential features of the simple example still pertain. From these two examples it will be clear that the decision analysis approach rests on two key concepts: the conditional probability distribution and the loss function. A third example demonstrates how different loss functions produce different ""best"" estimates from the same conditional probability distribution. In the final sections we willexamine the current practice of estimation in the mining industry. Though the tools for estimating the conditional probability distribution already exist, a change in the way we think about ""best"" estimates is needed. Very little attention is presently paid to the loss function even though it is readily available in many mining applications."

Jan 1, 1987

By D. H. Horswill

Red Dog, the world’s largest zinc mine, is located ninety miles north of the Arctic Circle in northwest Alaska on lands owned by the Inupiat Eskimo organization called NANA. NANA received a subsurface title over this resource as part of their settlement lands under the Alaska Native Claims Settlement Act and subsequently entered into a Lease Agreement with Cominco in 1982. Cominco, an international mining company focused on zinc, built and operated the mine. NANA’s principle objective in the development of the mine was to have it contribute to the development of sustainable communities within their traditional region. This meant creating job opportunities in a modern economy while preserving the traditional subsistence lifestyles. Preservation of their culture, language, and traditions are as important to the NANA people today as they were through ages past. The principles of sustainable development – matching economic, environmental, and social goals through continuous communications and high commitment to efficient and effective operations – have been core values for the Red Dog Mine from the beginning. This paper will describe how a large mining operation in a remote corner of the world has been able to contribute toward sustainable development in that region.

Jan 1, 2000

By Gillyeard J. Leathley

The Cassiar Asbestos mine is situated on McDame Mountain in northern B.C. at a mean elevation of 6000 ft (1830 m), some 6 miles (9.6 km) from the townsite /mill complex. Asbestos ore production averages 1,200,000 tons at 9% R.M.G. (Recoverable Mine Grade) and waste production is 6,000,000 cu. yd annually. Waste is loaded by P&H 1900 AL 11-cu.-yd shovels and transported by 75-ton and 85-ton end-dump trucks to the waste dumps. The ore is loaded by a 992 Caterpillar 10-cu-yd front-end loader into 50-ton and 75-ton trucks, which transport it to the crushing plant located at the mine. The ore is transported by a 15,370-ft (4685-m) tramline to the mill complex.

Jan 1, 1978

US 4,134,618-Method for restoring an underground reservoir subsequent to the solution mining of an ore body of uranium, copper, nickel, vanadium, molybdenum, silver, rhenium, or selenium. Clean water a cycled through the mined area, via the injection and production wells used during mining, at a flow rate sufficient to entrain residual dissolved solids. The solids are separated at the surface and disposed of, e g., in a deep disposal well, while clarified water is recycled C A. Kossack, assigned to Atlantic Richfield Co. Jan 16, 1979 4 pp. (299-5) Filed. 12-29-77 (865,646) US 4,155,982-In situ leaching of a subterranean uranium ore deposit The ore is leached with a dilute aqueous solution of ammonium bicarbonate containing a small amount of hydrogen peroxide in order first to convert tetravalent uranium in the ore to the hexavalent state and second to dissolve the hexavalent uranium preferentially to the other metals or elements commonly occurring In such ore The uranium- enriched leach liquor IS brought to the surface, and the uranium recovered by means of adsorption on a particulate ion exchange resin, elution with an acid, and treatment with ammonia to precipitate relatively pure ammonium diuranate. G G. Hunkin, T P. Fife, and J R Stano, assigned to Wyoming Mineral Corp. May 22, 1979. 13 pp (423-7). Filed: 10- 9-74 (51 3,445). Same: Canadian 1,027,849, dated Mar. 14, 1978. US 4,157,847-In the mining of sulfur by the Frasch process hot water accumulating above subterranean sulfur from previous mining operations is recycled downward to contact and heat additional solid subterranean sulfur. Advantageously, pressurized and superheated drive water from above ground is commingled with the underground water to raise its temperature before it is recycled downward R. L Williams, R L Latiolais, J. L Gabriel, and R. L Bouis, assigned to Freeport Minerals Co June 12, 1979. 13 pp (299- 6) Filed 7-28-77 (81 9,879). US 4,158,76%1n this method for determining the uranium content of a solution obtained in the solution mining of an underground uranium ore deposit, the intensity of gamma radiation from the Th-234 daughter of U-238 emitted by the solution is determined as it impinges on a scintillator after being passed successively through filter foils of a ytterbium- tin alloy, a lutecium-tin alloy, and a hafnium-zirconium alloy This scintillation is evaluated as a measure of the uranium content To assure that the uranium-thorium conversion has reached a stable state, the measurement should be made about three months after the uranium solution has been produced R C Smith, assigned to Westinghouse Electric Corp. June 19, 1979 11 pp. (250-255). Filed. 10-20-77 (843,770).

Jan 1, 1980

By Hugh Rose

JAY A. CARPENTER, Tonopah, Nev: (communication to the Secretary*).-This description of the Santa Gertrudis mill is of great interest to the operators of similar silver mills in Nevada. At the San Francisco meeting last September there were two excellent papers presented on Nevada silver mills; one by A. H. Jones on The Tonopah Plant of the Belmont Milling Co., and the other by E. E. Carpenter on Cyaniding Practice of the Churchill Milling Co., Wonder, Nev. At the same meeting, I presented a paper on-Slime Agitation and Solution Replacement Methods at the West End Mill, Tonopah, Nev., covering this .particular feature of our milling practice. A comparison of the data presented in these four papers is interesting in that, while the general design and practice is much the same in all cases, there is a wide difference in the design and practice in the various departments of &apos;the mills. It is to be regretted that Mr. Rose did not round out. this excellent article by following Mr. Jones&apos;s and Mr. Carpenter&apos;s&apos; lead in giving the average gold and silver content of the ore milled and the resulting tailing, with the consumption of chemicals and the? cost of milling. The practice of the Santa Gertrudis mill in the use of ?cyanide and zinc offers an interesting comparison with that of the Nevada mills, and a discussion of this point may bring out the reasons for the differences that exist.

Jan 12, 1916

By V. N. Cox, C. H. Grant

Both the underwater method and the rock cratering method contribute useful information in evaluating and comparing new explosive compositions. Results indicate that metallized explosive systems which contain no high explosives can be formulated to offer more energy and to do potentially more work than other slurry explosive systems. Commercial field trials are presently being conducted with a metallized explosive that does not contain any high explosive. Although there have been only a few commercial blasts to date, the results are encouraging. The Dow Chemical Co. has been engaged in explosives research since 1953. The incentive for this research was based on the desire to find additional markets and uses for magnesium metal of which DOW is a heavy producer. The thermodynamic calculations for explosives incorporating light metals indicate energies approximately twice that of nonmetallized explosives currently being used. Early in 1960, additional patent rights were acquired from independent inventors and a development station at Virginia, Minn., was set up to conduct a commercial testing and market evaluation program with this new product. The actual manufacturing of the explosive was to take place in the boreholes. Due to the complexities involved in controlling the manufacture of the explosive in the borehole, performance was sometimes in doubt. Furthermore, experience indicated that the metallized explosive systems could not be evaluated by the standard laboratory methods then being used, and it was necessary for us to look for more suitable research methods. Tests were conducted on several known methods for explosive evaluation. The most promising were the Navy underwater method for energy evaluation and a modified Livingston crater method intended to compare explosive performance in rock. After shooting over 2000 charges of 8 to 60 lb under water and in taconites and granites with many types of explosives, a metallized explosive was developed which is now being manufactured at a pilot plant in Virginia, Minn. and is being evaluated commercially. UNDERWATER METHOD An underwater method of evaluating explosives has a number of advantages. A permanent site can be established and operated indefinitely. A large number of charges may be fired in a short length of time. The medium in which the explosive is shot is exactly the same for successive shots. Data are recorded electronically minimizing human judgment. A disadvantage of the underwater method is the lack of confinement provided by the water medium. Many explosive compositions require a greater degree of confinement to develop good energy release. Although the relative significance of the two components of energy release that are usually measured, the shock energy and the bubble energy, and their relationship to breaking rock are not clearly understood, the underwater method is a valuable research tool for screening explosive compositions and gaining an insight into the manner in which explosives release their energy. Fig. 1 shows schematically the underwater apparatus. The piezoelectric gages are calibrated tourmaline crystals which convert the shock wave to an electrical impulse which is transmitted to an oscilloscope and recorded on Polaroid film. The shock traces are picked up by scopes 1 and 2. Scope 1 is timed to pick up the full trace whereas scope 2 is timed to cover the first increment of the shock trace expanded, and therefore allows full inspection of the propagation and peak pressures. Scope 3 records the bubble trace. When the explosion occurs, a bubble of gaseous detonation products is formed. The outward momentum of the water keeps expanding the bubble even after

Jan 1, 1963

By Douglas Ashmead

In his paper Mr. Braley makes no mention of the bacteriological aspects of the problem. It is now quite well established that certain bacteria play a major role in formation of acid mine waters, and it is a simple matter in the laboratory to show that under sterile conditions the rate of acid production from a pyrites suspension is only about one quarter of that obtained from a similar suspension inoculated with drainage from a mine producing an acidic pit water. Under sterile conditions the oxidation is due to direct chemical action and, from the evidence just given and from much other evidence, this increase under nonsterile conditions is due to certain bacteria. Experiments recently completed, and shortly to be published, have shown that this bacteriological oxidation can be prevented by the maintenance of pH conditions above 4. It was found that to raise this pH above 4 at the beginning of the experiments was not sufficient but that, due to the continuing chemical oxidation, alkali had to be added daily to maintain the pH conditions above 4. The amount of alkali added, however, over a fixed period, was only about one quarter of the alkaline equivalent of the acid produced when pH conditions were not controlled over an equal period. The opinion expressed by Mr. Braley that sodium hydroxide has little or no effect on the rate of oxidation of pyrites is not substantiated by the above experiments. The writer does not claim that these results show a practical solution to the problems, especially in abandoned workings, but feels that the application of an alkaline coating, such as lime wash, to exposed accessible workings might be well worth trying. S. A. Braley (author&apos;s reply)—In 1919 Powell and Parrl suggested that bacteria, or some catalytic agent, hastened the oxidation of pyritic or marcastic sulfur in coal. Carpenter and Herndon (1933)&apos; attributed the action of Thiobacillus thiooxidans. Colmer and Hinkle (1947)3 observed an organism similar to T. thiooxidans and another organism that oxidized iron. Leathen and Braley 9rst discovered this organism in 1947 in a sample of water from the overflow of the Bradenville mine (Westmoreland County, Pennsylvania). They characterized the organism in 1954" and gave it the name Ferrobacillus ferrooxidans. Although Temple and Colmer (1951)&apos; had suggested the name Thiobacillus ferrooxidans, since they claimed it oxidized both ferrous iron and thiosulfate, we have found that pure cultures of the organism do not oxidize thiosulfate, hence the name F. ferrooxidans. In 1955 Ashmead7 isolated an organism, similar to the one called Thiobacillus ferrooxidans by Temple and Hinkle, from acid mine water in Scotland. It is probable that this organism was F. ferrooxidans. In 1954 Bryner, Beck, Davis, and Wilsonh reported microorganisms in effluents from copper mine refuse. These organisms appeared to be similar but were not in pure culture. In view of this history of bacterial investigation of acid mine water and our own ten years of experience, we do not agree with Mr. Ashmead that bacteria play a major role in acid formation. We do not find that any of these bacteria will directly oxidize pyritic material. They do, however, augment the chemical formation of sulfuric acid by atmospheric oxidation. In two papers in 1953% eathen, Braley, and McIntyre discuss the role of bacteria in acid formation and postulate the mechanism through which they operate. Mr. Ashmead in his discussion of my paper has assumed that this work was carried on in the presence of acid mine water in which bacteria would be present. This was not the case. Strictly sterile conditions were not maintained, but the organisms present in mine drainages were definitely absent in these experiments. We believe that we have demonstrated that alkalis do not inhibit the chemical oxidation of pyritic material. This is also indicated by Mr. Ashmead&apos;s discussion in which he says that alkali must be added daily due to the continuing chemical oxidation. It is interesting to note that Mr. Ashmead finds that maintenance of pH above 4.00 decreases the activity of the bacteria. We have found also that a decrease in pH below 2.8 also inhibits its activity. Table XIII of published data&apos;" illustrates the decrease in activity with increased acidity, although pH values are not given. These values are in comparison with uninoculated controls and show the marked increase in acidity up to 22 weeks but a decline at 29 weeks, at which time the experiment was terminated. It is probable that after a longer period only chemical oxidation would have continued. From our studiesv we have postulated that the iron oxidizing bacterium (Ferrobacillus ferrooxidans) oxidizes the ferrous iron, resulting from chemical oxidation, to ferric iron. The ferric iron then aids the atmospheric oxidation of the sulfuritic material and is itself reduced to ferrous iron, which in turn acts as food for the autotrophic bacteria. Study of the physiologic properties of F. ferrooxidans shows that its preferred pH is about 3.00 and its activity decreases with variation in either direction. It is extremely inactive above pH 4.00 and below 2.5. This inactivity above 4.00 is indicated by Mr. Ashmead&apos;s observations. These properties of F. ferrooxidans then correlate perfectly with our hypothesis. Ferrous iron is oxidized very slowly by atmospheric oxygen in highly acid sohtion and since the bacteria become inactive, acid is formed only by atmospheric oxidation. At a pH of 4.00 or above iron is more readily oxidized by atmospheric oxygen, but the bacterial activity is decreased. However, with a pH above 4.00 the ferric iron is removed from the field of activity since its soluble sulfate hy-drolyzes and precipitates the iron as ferric hydroxide or a basic sulfate. As we have shown in the paper under discussion, the alkali does not inhibit the chemical oxidation, and thus the acid formation continues. This

Jan 1, 1957

By P. L. Allsman

A method of analyzing blasting action indicates that major cost savings are possible by revising practice and bringing the classical blasting formulas up to date; difficult problems such as taconite and throw-breakup can be attacked by engineering Denonation pressure — the peak pressure developed by the explosive reaction; "shatter blasting." Average detonation pressure — the average pressure over the time the reaction continues; "heave blasting." Radial presswe — the stress in a radially expanding sphere about the detonation. Corresponds to the diminishing peak pressure. Lateral stress - the stress in a circumferential direction, induced by outward movement of the rock away from the detonation. Limiting stress or Slim— the rock property at which failure will occur in the case at hand; may be tensile or compressive strength, or combinations. Recent developments in explosives technology, following the advent of nuclear explosions, have led to a rather complete understanding of their action. A survey was made for the purpose of uncovering any knowledge which might be applicable in the mining and quarrying industries, the principal users of explosives. Although no revolutionary techniques have become apparent, much basic data pertaining to blasting — fortunately not classified — have been developed by recent research, from which a good general understanding of the process can be derived. It is hoped this explanation of the scientific principles governing explosive action, together with a proposed analysis of blasting practice, will further the development of mining engineering. NATURE OF THE REACTION: The phenomenon of explosion, termed detonation, is in reality a rather complex chemical reaction, and as such is completely explainable by physical and chemical laws. Detonation is characterized by high temperature and pressure, extreme rapidity — often being complete within one microsecond — and most importantly by formation of a shock pulse which accompanies the reaction zone. As with any chemical reaction, there is a critical value of temperature and pressure below which detonation can not occur; this is termed the "critical point". Whereas an explosive substance will decompose slowly at ordinary temperature, with the formation of gases which readily dissipate; at elevated temperature the reaction is considerably speeded. If pressure is suddenly applied at some point in an explosive medium, adiabatic compression results, and the temperature is locally raised. When the temperature becomes high enough the decomposition will be so rapid that the gaseous products do not have time to dissipate, but contribute to a further build-up of pressure. This, in turn, furthers adiabatic temperature rise over a widening area, and with any heat generated by an exothermic reaction, causes a rapid increase in temperature. Both temperature and pressure are spontaneously built up in this manner until the critical point is reached, resulting in detonation. The values of pressure and temperature necessary to create detonation are established by the unusual nature of the process. If temperature alone is raised above the critical point, without sufficient increase in pressure, deflagration or "low-order detonation" occurs. In the true, or "high-order detonation", the shock pulse resulting from the violent decomposition becomes an integral part of the reaction, and itself aids in increasing the detonation

Jan 1, 1961

By R. Costantini

The author discusses the promising future of the use of pipelines for transportation of ore slurries over long distances, citing existing installations. Various criteria and factors affecting the use and size of pipelines are considered. Among the factors influencing the use of pipelines are particle size, specific gravity, abrasiveness, minimum velocities, and pumping considerations. Long-distance transportation of solids by hydraulic pipelines holds the promise of great economic benefits. Today within the confines of most mineral processing plants, the moving of solids about in slurry form by pipeline is commonplace, since the method offers great advantages when conditions are suitable. Yet, this method is surprisingly slow in being adopted on a broad scale for transportation of solids over long distances (arbitrarily the term will be applied to one over five miles long). This is especially surprising in view of the successful realization in this country within the last five years of two daring ventures in long-distance solids pipelines. A 72-mile long, 6-in. diam pipeline of the American Gilsonite Co. has been satisfactorily transporting 700 tpd of gilsonite from the mine at Bonanza, Utah, to the processing plant at Grand Junction, Colo., since April 1957. The 108-mile long 10-in. diam pipeline of the Pittsburgh Consolidation Coal Co. has been transporting about 4000 tpd of finely crushed coal from the mine at Cadiz, Ohio, to the Eastlake power plant of the Cleveland Electric Illuminating Co. on Lake Erie since January 1958. However, besides two projects in the operations of The Anaconda Co. there have been hardly any other significant solids pipeline developments in the minerals industry during the last three years. The two projects are the 14-mile long, 6-in. diam copper concentrate gravity pipeline at Anaconda&apos;s El Salvador project in Chile which has been transporting about 800 tpd since January 1959. The second is a 16-in. diam tailings pipeline at the Anaconda Reduction Works, Anaconda, Mont., designed to transport about 20,000 tpd of deslimed flotation tailings for dam building. This pipeline, which is currently undergoing tryout tests and which will be about seven miles in length, will have a maximum of three 20-in. rubber-lined centrifugal pumps in series at the starting point, each developing a maximum of 100 ft of head, to supplement gravity. It is the writer&apos;s opinion that the major reason for the hesitation in considering pipeline operations is confusion due to an inadequate understanding of the basic physical principles affecting the slurry flow process by many engineers in the industry. This in turn begets an attitude of fear and trepidation instead of confidence on the part of many management executives, which leads to a minimizing of the advantages and exaggeration of difficulties when considering the possibility of a long-distance solids pipeline. The promotion of better understanding and consequently the promotion of greater confidence in the potentials of long-distance solids pipelines when conditions are suitable is thus the main purpose of this article. To this end, after considering several examples of the possibilities for solids pipelines in the minerals field, the relative significance of the major factors affecting pulp flow, and some major aspects of slurry pumping will be discussed on a general qualitative level. Then a brief outline of the semi-empirical quantitative methods developed to date for estimating slurry flow behavior will be presented. SOME POSSIBILITIES To give an idea of the magnitude of the possible economic benefits, it should be interesting to mention briefly some of the projects currently under consideration and to suggest some likely future possibilities. At Anaconda&apos;s El Salvador Project in Chile (Fig. 1) it is eventually planned to extend the existing 14-mile pipeline by another 60 miles to bring the copper concentrates down to the coast of the Pacific Ocean. Preliminary tests indicate that the projected 6-in. pipeline will be able to transport approximately 2000 tpd of concentrates at a pulp density of 40 to

Jan 1, 1961

By A. W. Ruff

The seismic wave produced by an explosive is very important in blasting. A true understanding of the wave is only important when considering possible structural damage to buildings located near the blast, but it is also of primary importance when considering rock fragmentation. Because of the importance of the seismic wave to fragmentation, a finda-mental study of the seismic wave and its effect was made at the Cananea Consolidated Copper Co.&apos;s mine at Cananea, Sonora, Mexico. Strain gages were used to measure the strain in the rock when nearby explosive charges were detonated. The variance in the circuit voltage caused by changes in the resistance of the strain gage was measured with an oscilloscope and recorded with a Polaroid camera. CONFINED DRILLHOLE CHARGES To record the seismic wave from a confined drillhole explosive charge, a gage was cemented to the bottom of a 4-in diam drillhole. The horizontal gage hole and nine horizontal charge holes were drilled perpendicular to the center line of the 2-39 drift and 4 ft above the floor. The gage hole was 29 in. deep, and the charge holes were drilled so that each bottomed 1 in. deeper than the gage hole. The gage was installed with its axis horizontal and parallel to the 2-39 drift. The charge holes had the following horizontal distances from the gage: hole 3, 3 ft; hole 4, 4 ft; hole 5, 5 ft; hole 9, 9 ft; hole 10, 10 ft; hole 14, 14 ft; and hole 15, 15 ft. Figs. 1 and 2 show the relationships of the gage and the charge holes. Charges of 80 pct special gelatin, which weighed 50, 100, and 150 g were exploded in the charge holes and the seismic waves were recorded. All charges were stemmed with 200 g of mud and detonated with an electric blasting cap. Fig. 3 shows the recorded seismic waves. The reflected components of the waves were too weak to be objectional on these records. Table I shows the data taken from the wave forms. The approximate average velocity of a wave with a 9-ft travel distance was 12,900 fps. Since different charge weights were used, it was necessary to scale the data. The scaled maximum compress ion-strain data and the scaled maximum tens ion-strain data are shown in Figs. 4 and 5, respectively. Figs. 6 and 7 show the strain-distance data from Figs. 4 and 5 converted to a 1-lb charge. The results showed that the wave consisted of an initial compression pulse followed by a tension pulse. The amplitude of the tension pulse approached the amplitude of the compression pulse as the distance between the gage and the shot increased. The decay rate of the seismic wave is very important in the evaluation of the blasting characteristics of a rock. When high-amplitude stresses are produced, as in rock fragmentation, the rate of decay is much more rapid than when small amplitude stresses are produced. In underground mine blasting, the burdens are usually not greater than 6 ft and the decay is very rapid. The rapid seismic-wave decay rate in the fragmentation range means that in practice, if the minimum weight of powder is used to break a specified burden, only a very

Jan 1, 1961

By H. H. Waller, D. L. Monroe, D. P. R. Smyth

ORIGINALLY slusher drift development was conventional, advancing the drift full 10x13-ft size at 6 ft per round. This proved dangerous and costly because the weak fractured rock of the orebody cannot be left raw for any length of time without its reacting with air to cause spalling. With this method of drift development ground conditions were worsened by many blasts and the length of time it was necessary for the drift to remain open. The problem was practically solved by driving 5x7-ft pilot drifts, then enlarging to 10x13-ft raw size by slashing with 50 to 60-ft diamond drill holes. Calculated explosive charges are used to avoid wall and roof shattering. Still more recently jacklegs and sectional steel replaced the diamond drills, resulting in the same work for less than half the cost, see Figs. 1 and 2. Crosscuts at right angles to the main haulage drifts, which cross through the center of the ore-body, divide the ore into 200-ft blocks. Eight slusher drifts on 50-ft centers draw from one block. The slusher floor is 11 ft above the rail, inclined at 1 pct grade, and the top is 18 ft below the undercut. The finished slusher drift is 102.5 or 120 ft long, depending on whether it is north or south of the crosscut. Drawpoints are staggered on each side at 17.5-ft centers. The completed drift is concreted and has an inclined ramp and a vent at the back. When crosscuts and gangways have been excavated and concreted and slusher stations installed, slusher drift development is carried out in 10 clear-cut stages: 1—Pilot drift is driven full length, 5x7 ft. 2—Station is excavated at half-way point in pilot drift. 3—Pilot drift is enlarged to slusher drift size by drilling. 4—Lower part of slash is blasted and mucked for floor concrete. 5—Floor is poured. 6—Inside upper half of slusher drift is blasted, mucked, and concreted. 7—Outside upper half of slusher drift is blasted, mucked, and concreted. 8—Drawpoints are excavated. 9—Drawpoints are concreted. 10—Concrete ramp and sheave protection recess are constructed at back end of slusher drift. A 5x7-ft pilot drift is driven full length on the center line of the proposed slusher drift using a jackleg and detachable bits on 7/8-in. hexagon steel. A 10-hp air-driven slusher and 36-in. scraper pull muck directly into cars in the main haulage crosscut, see Fig. 2. Operations are on a three-shift basis. One man mucks the previous round, drills, and blasts his round in a shift. Average advance is 5 ft per shift. When the pilot drift is completed, 7 ft of drift in the center section is enlarged to full raw slusher drift size of 10x13 ft, for the set-up to drill long slash holes. The pilot drift is slashed to full raw size from this station with jacklegs and tungsten carbide chisel bits on 7/8-in. sectional steel. Crews line the holes without trouble, working from controls established by the engineers, and the pattern of holes for the slash is arranged to give the least shattering effect

Jan 1, 1955

By D. T. Van Zandt

IN the late 1920&apos;s rotary drills began to replace the churn drills in the petroleum industry, but until the middle 1940&apos;s the churn drill was the only widely accepted means of drilling large-diameter blastholes for quarry operations. The Calcite plant of the Michigan Limestone Div., U. S. Steel Corp., was one of the first to experiment with rotary drills for quarry blasthole drilling, and the first to employ compressed air on a fully rotary rig to cool the bit and raise the cuttings to the collar of the blasthole. The Calcite plant operates a limestone quarry near Rogers City, Mich., in the northern part of the lower Michigan peninsula. The formation quarried, a portion of the middle Devonian series, is the Dundee limestone, which is uniform, seldom massive, and characterized by definite bedding planes. The dip is southeast, 40 ft to the mile. Quarry faces vary from 20 to 116 ft in height. Vertical blastholes are used entirely, from three to five rows of holes being drilled parallel to the working face, spaced 18 ft apart with 18-ft burden and drilled 6 to 8 ft below shovel grade. Quarry operations coincide with the navigation season on the Great Lakes, as the bulk of the stone is transported by lake carrier. The normal operating season runs from April to December, the remaining time being devoted to stripping operations and plant and equipment maintenance. In the followirig discussion drilling rates mentioned refer to overall drilling time and include all operations such as moving from hole to hole, penetration and extraction of tools, and routine maintenance. Time consumed by such factors as power delays and major machine repair is not included in drilling time unless otherwise stated. Figures cover only operations at this one plant in the formation mentioned. Needless to say, a very different set of figures could be obtained in a different formation. However, the comparison of footage obtained with churn drills and rotary rigs in this particular formation has been used as an indication of what might be the expected performance of rotary rigs in other formations. Prior to 1950 the bulk of the blasthole drilling at the Calcite plant was done by electrically powered churn drills. Both crawler and wheel-mounted rigs were used. These machines, which mounted a 22-ft drill stem of 4½ in. diam and a spudding type of bit 2 to 4 ft long, drilled a hole of 5 ?-in. diam. Average drilling rate of these rigs in the Rogers City formation was 8 % ft per hr. In 1946 one of the first rotary blasthole drills offered to the quarry industry was put into use on an experimental basis. This machine, known as the Sullivan Model 56 blasthole drill, Fig. 1, was on 16-in. crawler pads and electrically powered at 440 v. The drill bit, a Hughes Tri-Cone roller bit of 5?-in. diam, Type OSC, was threaded into the end of the 4-in. square hollow drill rod or stem. These drill rods were 20 ft long with female threads on one end and male on the other to allow for addition of the desired number of rods for drilling holes of various depth. Rods were handled by a single drum hoist geared to the main drive motor and racked by a 30-ft derrick or mast when not in use. The cable from the hoist drum fed through a crown block on the top of the derrick back to the water swivel mounted in the top end of the drill stem in use. This cable remained attached during drilling operations and was used to hoist the tool string from the hole. Down pressure was applied to the tool string by means of a pair of 4-in. diam hydraulic cylinders acting on the drill chuck holding the drill rod. The first chuck consisted of flat jaws which gripped the flat sides of the stem. These jaws were controlled by set screws forcing them into contact with the drill stem. As these set screws had to be loosened and tightened by hand with each stroke of the hydraulic feed cylinders, there was great delay. For this reason the semi-automatic chuck was developed which automatically gripped the stem on the downward stroke but released for retraction of the hydraulic feed cylinders. Rotation was imparted to the tool string by a rotary table acting on the chuck and geared to the main drive motor through a separate gear train and clutch. A positive displacement water pump, mounted on the drill, fed water through a system of pipes and hose into the water swivel mounted on the top of the drill rod and through the rod and bit, washing the drill cuttings to the collar of the hole. Where water was scarce, provision was made to settle out the cuttings coming from the collar of the hole and re-use the water. Where water was abundant the stream coming from the hole was wasted. Drilling rate with this machine was about 20 ft per hr and bit life 1600 ft of hole. While this rate was more than twice that obtained with the churn drills employed, the problem of water supply and drill cuttings disposal rendered the machine impractical from an operating standpoint. Consequently it was used only in that part of the operation for which water was easily supplied, when the character of the formation made it least difficult to wash cuttings away from the collar of the hole. In October 1949 it was suggested that drill cuttings be removed by compressed air, long used for this purpose on pneumatic drills, and collected at the collar by suction. Thereafter, the water pump on the Sullivan 56 was replaced by a 500-cfm air compressor and a trial run made. Air pressure at

Jan 1, 1955

By James P. Dorian

Dramatic economic and political reform is under way in the transitional economies of the world. Transition extends beyond the restructuring of a nation&apos;s economy and political thinking. It also affects social, cultural and even religious aspects of society. Transition is a long-term process and generally unstable as the pace of reform is rarely the same across the many spheres of society. As nations modify their economies, industrial activities - including minerals exploration, extraction, processing and trade - are also being affected, according to a Financial Times of London management report. Mining policies are being revised to encourage greater foreign involvement, while legal regimes are being rewritten to protect foreign investments. From India to Peru, and South Africa to Russia, countries have been preparing, improving and expanding foreign investment laws, mining codes and economic stimulus measures to accelerate their minerals development. Privatization of mining is proceeding in many nations. Meanwhile, corporatization of former ministries and government agencies responsible for mining is also occurring. According to Minerals and Mining in the Transitional Economies: Poised For the Future, common objectives in the restructuring of a mining industry are to enhance overall economic and industrial growth, improve efficiencies in mineral recovery and use, acquire newer, advanced technologies, reduce environmental repurcussions of mining and improve safety conditions. The nature of mining is changing noticeably within the transitional economies. The industry is becoming more re¬sponsive to market forces and private sector concerns. The new report examines the changing characteristics of mining in the transitional economies. These include nations moving from centrally planned to market economies, as well as countries in the midst of economic and political reform, including the liberalization of industry. Information is presented to help business planners understand the impacts of transition on mining activities and re-evaluate their longterm strategies and market development policies. The document is divided into 10 chapters. Five of them are devoted to the evolving mining industries of Russia, South Africa, India, Peru and China. These countries were highlighted because their minerals sectors are undergoing changes. These include the opening up of their resources to foreign mining companies. According to the report&apos;s author, mining in the transitional economies during the next decade will be characterized by: ? Continued rewriting of national mining policies to create more favorable investment climates. ? Increased exploration for gold and selected base metals in newly opened regions. ? Innovative project financing schemes involving multiple partners and international lending institutions, such as the European Bank for Reconstruction and Development (EBRD). ? Increased gold and copper ore production capacity resulting from new investment and expansion of existing operations. ? Continued geographical shifting of metallurgical processing facilities from developed to other mineral-producing nations. ? Privatization of state-owned mining and processing assets requiring advanced technologies for improvement or expansion. ? Increased awareness of environmental concerns in business decisions related to mining and metallurgy.

Jan 1, 1997

By Glenn C. Waterman

THE diamond drill is a very important tool in exploration and development testing and its use is increasing. In almost all cases results of diamond drilling are analyzed on the basis of grade and tons. A proper evaluation of core and sludge assays is important if drilling results are to be acceptable as a basis for geologic and engineering appraisal. The relatively wide variation in assay averages as calculated by various well-known combining methods indicates that the engineering choice of a method may affect the outcome of the drilling in terms of ore and waste. The problem of combining assay results from core and sludge samples has been discussed many times in conference and in the literature.&apos;-&apos; Most writers agree that the field of disagreement in methods is large and that the engineer on the job must consider features unique to his drilling, pick one of several combining methods, and depart from the rules when abnormal results come in. All the discussion to date can be summed up by the admission that as yet there is no fairly simple, generally acceptable combining method that is practicable over a wide range of drilling conditions, ground conditions, and ore occurrence. The combining problem is important in evaluating drilling results at Chuquicamata. Recently a reappraisal has been made of recovery variables and their effect on assays, with the result that a new combining method is offered which fits average drilling conditions and is mathematically reasonable. It is simple in application, fundamentally correct, and an improvement over most combining methods. At Chuquicamata diamond drillholes are used to outline the grade and position of blocks of normal and marginal grade oxide, mixed, and sulphide ore. Most holes penetrate all classes of material (and waste), and it is important for mining programs as well as ore reserves to know almost precisely the soluble and insoluble copper content of mineralized ground. At present three classes of ore are mined and treated differently. For an orderly sequence of mining operations which can provide regular daily tonnages of all three ore types and keep grade at certain levels with minimum variation, ore type and its grade must be predicted. Diamond drilling plus geologic mapping and bench sampling are tools for prediction. And drilling data are largely used to calculate grade of material more than a few meters away from bench faces. The orebody at Chuquicamata" is criss-crossed by millions of barren or mineralized weak to fairly strong slips and fault fissures. Mineralization is diverse and encompasses many quartz and oxide or sulphide-bearing copper veins, as well as seams and disseminated grains. Copper occurs in oxide or sulphide minerals or mixtures of these two mineral types. Rock conditions vary from intensely seri-citized (soft and porous) through clay-altered ground to almost fresh granodiorite. The result is an orebody which offers many obstacles to good and consistent core recovery in diamond drilling. Recovery varies considerably in the several alteration zones, the various types of oxide and sulphide ores, and the position and inclination of the drillhole within the complex fracture pattern. As core recovery drops sludge samples must be used with core samples to calculate grade. Many years of drilling at Chuquicamata indicate that in good grade oxide zones and in the sulphide areas core recovery is good and the ground uniformly mineralized. Moderate loss of core, therefore, does not markedly affect grade calculations based on core assays. An early core-sludge combining method used core assays at face value as indicating grade down to 50 pct core recovery, but below this recovery percentage sludge samples were used and weighted according to the standard Longyear chart. This method apparently did not introduce serious errors, but it abruptly used sludge assays with high weighting factors representing 100 pct return irrespective of actual percentage of sludge recovered. Recent drilling activities have been directed toward outlining the marginal ore areas. The non-uniform mineralization and generally poorer core recovery in such ground indicated that a more exact core-sludge combining method was required to equate wide differences between core and sludge assays and recovery. In fringe ore areas at Chuquicamata core recovery averages about 50 pct, from a minimum of 10 to 15 pct to a maximum of 100 pct. Sludge recovery is likewise variable and averages perhaps 80 pct even though holes are cemented as water return falls off. In a homogeneously mineralized area cut by many slips and faults, with hard and soft ribs, loss of core is loss of ground which has a grade similar to that of core recovered, and core assays approximate true grade. In this case sludge samples need not be used. However, it would be unusual to know beforehand that an area is uniformly mineralized, and in fact this condition is probably uncommon. Generally the distribution of valuable minerals in the ground does not exactly compare with their recoverability in core. Thus, in the usual case, loss of core decreases the

Jan 1, 1956

By Glenn C. Waterman

Richard Strong (Oliver Iron Mining Div., U. S. Steel Corp.)—Mr. Waterman states (p. 59, Trans., January 1954): "Core-sludge combining factors have been calculaied for any combination of core-sludge recovery on the basis of several logical assumptions: 1—The sludge sample loses reliability geometrically as sludge recovery drops. 2—The sludge sample should not be used if the percentage of sludge return is less than the percentage of core recovered. 3—A core sample should receive a weighting at least equal to the percentage of core recovery." Assumptions 1 and 2 seem justified at Chuquicamata and most geologists would agree to the validity of assumption 3. However, contrary to the statement quoted above, the core sample does not in general receive a weighting at least equal to the percentage of core recovery when the Chuquicamata methgd is used. Furthermore, the basic Chuquicamata method is identical to the Longyear or relative volume core-sludge assay combining method. Finally, the Chuquicamata method does not, in its present form, reflect true geometric loss in sludge sample reliability according to the relationship L = AR (N-1) The basic equations representing the two steps of the Chuquicamata method can be combined algebraically as follows (using the symbology of the original article except as noted): Vol A + B Assay 2 = ASA = Assay 1 Vol A VolB Core Assay [I] Vol A CA.F + ASA (100 —F) Combined grade =[21 100 Pct where F and (100 — F) are the core and adjusted sludge assay factors from Fig. 3. Substituting Eqs. 1 and 2 and collecting terms Combined grade = Vol B Vol A + B CAF (100—-F) + Assay 1 (100-F) Vol A Vol A 100 pct Eq. 3 is a unified and general statement of the Chuquicamata method. The true core factor includes not only the factor F from Fig. 3 but also the quantity VolB ------------- (100 — F), which is implicit in the value of Vol A the adjusted sludge assay. This quantity is real and positive for all values of F and core recovery less than 100. Therefore the true core factor is always less than F except when F is equal to 100. If sludge recovery is 100 pct, Eq. 3 can be further simplified as follows: Let F = percentage of core recovered. Let Vol A = Vdh (100 Pct — Pct CR) where Vdh is the total volume of the sampled interval and let Vol B = (Vdh — Vc.) Pct CR where V, is the core volume in the sampled interval at 100 pct core recovery. Then combined grade = Vo V. CA Pct CR- + Assay 1, 100 Pct — Pct CR Vdh Vdh [4] 100 Pct Eq. 4 shows that when sludge recovery is 100 pct the laboratory sludge assay receives significant weight in the Chuquicamata method, even at 100 pct core recovery. The maximum weight assigned to the laboratory sludge assay depends on drillhole size, as follows: Eq. 4, with suitable changes in symbology, is identical to Royce&apos;s formula 8,&apos; which is a statement of the Longyear or relative volume method of combining core and sludge assays. If Mr. Waterman&apos;s assumption 3 is of paramount concern to the geologist, the Royce method (for core recoveries of 60 pct and above) or a method employing the percentage of core recovered as a core weighting factor are the only recognized methods known to the writer which achieve the desired result. A modification of the latter method embodying Mr. Waterman&apos;s assumptions 1 and 2 would result from applying the factors from Fig. 3 directly to the laboratory core and sludge assays.

Jan 1, 1956

By C. B. E. Douglas

The following analysis was stimulated by a previous article on evaluation of the water problem at Eureka, Nev., which describes a method using formulas especially devised to calculate flow potential of extensive aquifers characterized by relatively even porosity and permeability throughout. The present discussion submits that the method was unsuitable for solving the kind of problem occurring at Eureka, where the amount of water available, rather than the flow potential, may have been the vital factor. IN an interesting article on evaluation of the water problem at Eureka, Nev.,1 W. T. Stuart describes how a difficult water problem, or one phase of it, may be evaluated by means of a small scale test. Test data are plotted by a method rendering, under certain conditions, a straight-line graph that can be projected to show how much the water table will be lowered by pumping at any specified rate for a given time. A formula is then used to determine the size of opening, or extent of workings, necessary to provide sufficient inflow to enable pumping to be maintained at that rate. At first glance this might seem the answer to a miner&apos;s prayer, but a word of caution is in order. It may not be the whole answer. Moreover, results obtained by the method described are reliable only for conditions approximating those assumed. Even where conditions do not meet this requirement, however, it may be possible to draw helpful inferences from the results, perhaps enough to facilitate another approach to evaluation of a problem. The two formulas Mr. Stuart used, the Theis formula and the one developed from it by Cooper and Jacob, were given field checks a number of years ago in valley alluvials by the Water Supply Div. of the U. S. Geological Survey and found to be reliable when the aquifer is very large in horizontal extent and sufficiently isotropic for the test well and observation wells to be in material of the same average permeability as the saturated part of the aquifer as a whole." Extensive valley alluvials, sands, and gravels can be evaluated in this way, and there are even cases in which the method could apply to porous limestones, such as flat beds of very large areal extent that have been submerged below the water table after extensive weathering. These are sometimes prolific sources of water for towns and industries. It is necessary for them to have been above the water table for some geologically long period of time in a fairly humid climate before submergence because the necessary high porosity and permeability, and large reservoir capacity, are the result of weathering, that is, of solution by the carbonic acid (H,CO3) in rainwater formed by the absorption of CO, from the air by raindrops, and this dissolving action must cease when all the H2CO3 has been consumed by re- action with the carbonate to form the more soluble bicarbonate. Consequently this weathering process is largely restricted to a zone that does not extend much below the water level, and submergence is necessary after the weathering to provide large reservoir capacity and good hydraulic continuity. On the other hand, water courses tend to form along faults and fractures in limestone, and to become enlarged by solution, well below water level when, as often happens, fresh meteoric water is circulated rapidly through them to considerable depth by hydrostatic pressure, as through an inverted syphon. Although the reservoir capacity of such water courses is relatively small they may extend far enough to tap more prolific sources. Cavities, and sometimes caves of considerable size, are found in limestones where the acid formed by the oxidation of sulphides has attacked them. This action can take place as deep below water level as surface water is carried by syphonic or artesian circulation, because the oxygen it carries in solution will not be consumed until it reacts with some reducing agent, such as a sulphide. Moreover, the formation of acid and solution of limestone in this way is not confined to the immediate vicinity of the sulphide. Oxidation of pyrite, for example, results in formation of acid in several successive stages, each taking place as more oxygen becomes available, as by the accession of fresh water into the circulation at some place beyond the sulphides. When the acid thus formed attacks the limestone, CO, is liberated and the ultimate effect of the complete oxidation of one unit of pyrite will be the removal of six times its volume of limestone as the sulphate and bicarbonate, both of which are relatively soluble. The reaction may be continued or renewed along a water course far from the site of the sulphides, where the small electric potential produced by contact with the limestone helped to start the reaction. Mr. Stuart refers2 0 caves in the old mining area in the block of Eldorado limestone southwest of the Ruby Hill fault at Eureka, Nev., and to the cavities encountered in drillholes in the downthrown block on the other side of the fault. Although he interprets these cavities as evidence that this formation was sufficiently isotropic (evenly porous and permeable) to give reliable results by the method he describes, they may, in fact, be entirely local conditions. There is reason to think they were probably formed

Jan 1, 1956

By Eugene McAuliffe

The mining press, as well as certain federal and state bulletins, refer from time to time to the relative hazards that attach to loading bituminous coal by hand when compared with the so-called "mechanical loading" process. Much of the coal coming under the head of hand-loaded coal is won in part by the use of machinery; the coal-cutting machine is in general use, the electric power drill employed to a lesser extent. Therefore it may be said that the principal difference in the process employed, mechanical loading versus hand loading, lies in the fact that some type of loading machine is employed in the mechanical loading process to lift the coal from the face of the room, place or entry, after it is mined, placing it in the pit cars for transport to the tipple. It is a common practice to compute accident records on the basis of tons mined per fatal and per nonfatal accident, disregarding the fact that it is possible to produce by the aid of coal-loading machines the same daily, monthly and annual tonnage with a much smaller force than is required where the coal is all shoveled by hand. Inasmuch as the men who are released by the substitution of machinery for hand labor must find employment elsewhere, suffering some measure of hazard in their new occupation, we are of the opinion that the true relative hazard that attaches to the processes of mechanical loading and hand loading can be determined only by using as the basis of comparison "man shifts worked" or "hours of exposure" experienced per accident. The Union Pacific Coal-Co., beginning with the year 1929, undertook to maintain a true record of man shifts worked per compensable accident, giving equal weight to fatal and nonfatal accidents. Accidents lacking sufficient severity to come under the compensable terms of the Workmen&apos;s Compensation Act were not included in the compilation, the Act providing that "no compensation, except the expense of medical attention, shall be allowed for the first seven (7) days of disability, unless the incapacity runs beyond the period of twenty-one (21) days, in which case the compensation shall run from the time of the injury." With the further thought that all accidents suffered by employees engaged in any service, in or about the place where the loading process

Jan 1, 1931