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By H. R., Collins

THE purpose of pulverizing coal before burning it is to make available every heat unit it contains. Machinery has been developed which will pulverize coal in one operation, delivering it to bins in front of the furnaces at an expenditure of about 17 hp.-hr. per ton, in a medium-sized plant. The cost of the operation depends upon the amount of moisture that must be expelled before pulverizing, the cost of labor, and the cost of coal delivered at the plant. At a small plant, requiring a pulverizer with a capacity of only ½ ton per hour, the cost per ton pulverized will naturally be greater than at a plant requiring the largest pulverizer, with a capacity of 7 tons per hour. The first step is to reduce large lumps to a size suitable for drying uniformly, before passing to the pulverizing mills; this is done in rolls, at a single pass. The second -step is the elimination of moisture, in order to facilitate pulverizing to great fineness, while also increasing the heating effect and the temperature attainable when the coal is burned. There are other mechanical advantages in the handling of, dried coal. Driers are now manufactured which are able to eliminate moisture without distilling any of the volatile combustible matter in the coal; they are fired by hand or with pulverized fuel. The heat first surrounds the shell of the drier, being confined within a chamber where complete combustion takes place; the heated gases are then led through a duct to the discharge end of the drier and enter the inside of the shell at a temperature not exceeding 300° F. This temperature is maintained by the operator and is indicated by a pyrometer. Volatile combustible matter is not likely to be distilled until the temperature rises above 400° F. On discharging from the drier, the coal is usually passed over a magnetic separator in order to prevent pieces of iron from going to the pulverizer. Two types are used: a magnetic pulley which automatically discharges its collection of iron, and a lifting type, from which the iron is removed by hand when convenient.

Jan 4, 1918

By Z. F. Danes

Combination of gravity data and density logs makes it possible to determine the structural relief. Under a wide class of geologic conditions, the solution is unique, or limited to a single parametric system of solutions. As a useful byproduct, the method yields a new way to construct derivative maps, continue the field up or down, subtract regional anomalies, and perform smoothing. In the past, the problem of interpreting gravity anomalies in terms of structural geology was solved along two main lines: 1) A body of a certain shape and density contrast was assumed and the corresponding gravity anomaly calculated, analytically for simple bodies, graphically and/or numerically for more complex ones; then the computed anomaly was compared with the observed one, and the assumed shape corrected according to the differences between the two. 2) The anomaly is assumed to come from a warped interface between two media of different density; the relief of the warping is assumed small compared to the depth of burial, so that the structure can be approximated by a flat horizontal surface of variable density distribution. The first method is a standard technique, described in every text book on applied geophysics, and is well applicable, for bodies with predominant vertical dimension. The second one, probably first used by Tsuboi and Fuchida,' is mathematically more elegant. It applies to the fashionable field of integral transforms (Bullard and Cooper, 2 Tomoda and Aki, 3 Dean 4) and is useful mainly in cases when the vertical relief of the structures is comparatively small. 30th methods, tacitly or openly, assume a density discontinuity across the boundary of the two media, while within each medium the density is constant. However, the recent invention of the gamma-gamma

Jan 1, 1962

Agnico-Eagle Mines Ltd. comprises a Gold Division at Joutel, Quebec, and a Silver Division at Cobalt, Ontario. In 1979 the Silver Division milled 41,700 mt (46,000 st) of ore at a recovered grade of 139 g per mt (4.06 oz per st) for a production of 5,816 kg (187,000 oz) of silver. The 100 -year -old mining, camp of Cobalt has long been noted for its high-grade silver ores, and now seems to be on the threshold of being revived with several companies operating in the area. In 1979, the Gold Division milled at the rate of 27,200 mt per month (30,000 st) and processed 333,000 mt (367,000 st) at a grade of 6.58 g per mt (0.192 oz per st) with a mill recovery of 91.670 to produce 2,013 kg (64,700 oz) of gold. The average total cost per ton milled was $22.10 per mt ($20.05 per st) which was $3.66 per g ($1 13.90 per 0s). These figures make Agnico-Eagle one of the lowest cost producers in North America. Gold occurrence in the Joutel ore is fine-grained and contained within sulfide facies composed mainly of pyrite with minor pyrrhotite. Two underground mining methods are employed with long hole stoping accounting for about 75% of the ore mined and shrinkage stoping the remainder. Ore from underground is hoisted into a 545 mt (600 st) coarse ore bin from where it is conveyed to a 45 mt (50 st) surge bin ahead of the 10-51 Hydrocone crusher. Crushed ore is then screened to make a 13 mm (112 in.) product with the oversize being recrushed in a 7-60 Hydrocone. Mill feed is conveyed to two 900 mt (1,000 st) fine ore bins. The Hydrocones were installed as part of a new crushing plant in March of 1980 and have contributed substantially to an increase in mill capacity. Space has been left to install a third crusher in the future. The fine ore bins discharge to a 2.75 m by 3.66 m (9 ft by 12 ft) rod mill, followed by a 3. 7 m by 5.5 m (12 ft by 18 ft) ball mill in closed circuit with a Krebs 380 mm (15 in.) cyclone to produce a 7070 minus 37 micron (400 mesh) .flotation feed. This pulp is first treated in a 34 cu m (1, 200 cu ft) Maxwell flotation cell and then in an additional twelve 2.8 cu m (100 cu ft) cells. Combined flotation concentrates are reground to 9070 minus 37 microns (400 mesh) in a 1.8 m by 4.9 m (6 ft by 16 ft) ball mill in closed circuit with a single Krebs 254 mrn (10 in.) cyclone. Flotation tailings are sent to the tailings dam. Reground concentrates are then aerated with lime to a single Pachuca and thickened in a 16.75 m (55 ft) diameter thickener with the addition of diesel fuel to render any activated graphite inert to adsorption of

Jan 1, 1981

By Clarence Zener

IT has long been known that the density of a metal usually decreases with cold-work. Thus O'Neill1 observed as early as 1861 that cold hammering of commercial hot-rolled copper is accompanied by a decrease of density of 0.2 per cent, and that subsequent annealing restores the original density. The origin of this density change, however, has remained obscure. In particular, experiment has not yet decided whether it is due to the presence of small cavities and an increase in intergranular surface or to an actual expansion of the lattice itself. This uncertainty in the origin of the change in density is a reflection of our uncertainty as to the changes induced in a metal by cold-working. One school regards cold-working as a breaking up of the original grains into a large number of perfect crystallites somewhat randomly orientated with respect to one another.2 According to this viewpoint, the mere introduction of a large intercrystallite surface induces a net volume expansion. Another school regards cold-working primarily as a distortion of the grains. It is not obvious that this second viewpoint will lead to a net volume expansion. The purpose of the present paper is to show that this is indeed the case, and that this expansion is of the right order of magnitude to explain existing experimental data. The physical basis for the association of an expansion with distortion lies essentially in the variation of the elastic moduli with volume. This variation is considerable. Thus from Table I we see that a I per cent increase in volume for the case of copper will decrease the bulk modulus K by i0 per cent, the shearing modulus µ by 3.8 per cent. If we thus twist a rod through a given angle, we shall do less work upon it if it increases slightly in volume, thereby reducing the torsion modulus that involves as a factor the shearing modulus. A general principle' in physics states that the rod will so adjust itself as to have a minimum potential energy consistent with the end restraints. It will therefore expand. [ ] In order to obtain a quantitative comparison with experiment, we follow Maier4 in comparing the lattice expansion with the latent energy of cold-work. This latent energy has been measured directly by Taylor and Quinney5 on copper rods subjected to torsion, and may amount to as much as 15 per cent of the total work done upon the specimens. We shall assume that this

Jan 1, 1941

(Under this heading will be published notes sent to the Secretary of the Institute by members or other persons. ) Position in sales department is open for a mining engineer with experience in mine ventilation and possessing some knowledge of fan design; one competent to lay out ventilating systems to meet various conditions. No. 2. A large company manufacturing mining machinery has an opening in one of its Western agencies for a young mechanical engineer who has had some experience in designing along general machine shop lines, and who is familiar with cost systems. No. 3. A large university desires to secure a prominent metallurgist for a vacant professorship in metallurgy. No. 4. Opening at a mine in New Jersey for a technical graduate to begin work as chairman on the surveying squad. Salary $50 per month to start. No. 5.

Jan 11, 1913

By E. P. Mathewson

The experiments described in this paper were made at the Washoe Reduction Works, Anaconda, Mont., for the purpose of determining the relative speed of elimination of the iron, sulphur and arsenic during the process of bessemerizing copper-mattes. The samples were taken from a horizontal a barrel" converter, 8 ft. in diameter and 12 ft. 6 in. long, having an average charge of 9 tons of matte, and blown with a 16-lb. air-pressure. All samples were taken with the blast on, turning the converter from its normal blowing-position to nearly vertical, so that a conical mold, with a long handle attached to a horizontal rod, could be introduced from a platform above the converter into the mouth of the converter and a sample taken. The difficulties of taking a sample with the blast on are apparent, and the irregularities of' the lines in Figs. 1, 2 and 3 can be readily accounted for by the amount of slag in the samples. On the whole, the plotting of the results gives comparatively uniform lines, considering the conditions. The samples were taken by the regular sampling-force of the works at 10-min. intervals, as shown in Tables I., 11. and III., which also give the chemical analysis of the samples so taken. The test shown in Table I. was made with a new converter; that of Table 11. with a converter that had made three charges and had been cleaned out; that of Table 111. with a converter that had blown one charge to white metal, and had then been " washed " out and used for the test. In plotting these results, it was necessary to take, as a basis, the percentage of copper in the samples. This basis does not exactly represent the true condition, since the volume is constantly decreasing and a certain quantity of copper passes into the slag, but it is as near as can be obtained in practice. Take, for instance, in Table I., the sample No. 5, which shows: Cu, 59.9;

Jan 1, 1908

By E. Stansfield

THIS test was described in a privately circulated First Progress Report on Slacking Characteristics of Coal, by A. C. Fieldner and W. A. Selvig, May 1, 1928. The test was described in detail in U. S. Bureau of Mines Report of Investigation 3055, in November, 1930.1 The test, in brief, is conducted as follows: A sample of freshly mined coal consisting of 18 to 25 pieces approximating 1 ¼ cubes, and weighing 500 to 600 grams, is taken for the test. The sample is subjected to a "blank sieving" of carefully regulated degree, on an 8-in. circular sieve with 0.263-in. square openings; and the undersize and oversize weighed. The oversize is placed on a sheet-iron pan and dried for about 24 hr. in an A.S.T.M. air-drying oven maintained at 30° to 35° C. through which a current of warm air is drawn. The sample is then cooled to room temperature and immersed for one hour in water at the same temperature. The water is then drained off, the sample dried for 24 hr. as before, cooled, sieved as before, and the oversize and undersize weighed. The percentage weight of the undersize, is calculated and reported as the first cycle loss. This loss, corrected by subtracting the percentage loss in the blank sieving, is termed the slacking index. The sample may be subjected to further cycles of wetting, draining, air-drying, sieving and weighing. In a letter (August, 1932) W. A. Selvig approved making the blank sieving on a duplicate sample, rather than on the sample used for the test, where sufficient coal was available. This method is now employed in these laboratories. The accelerated weathering test has been employed in the labora-tories of the Research Council of Alberta, at Edmonton, since 1928, but not found entirely satisfactory as repeat determinations with samples taken from the same or adjacent mines frequently showed serious differ-ences. Differences must be expected as the coal may vary in the seam, and the selection of a typical set of lumps approximating 1 ¼ -in. cubes

Jan 1, 1934

By M. G. F. Söhnlein

In a Bolivian tin concentrator an appliance was needed to furnish a suitable product for fine jigging from a pulp of the following composition: Mesh Per Cent. + 20 8.0 40 36.5 + 60 9.0 + 80 10.5 + loo 5.5 4-150 1 4.5 -150 25.0 99.0 Loss 1.0 In the jigging process, particles of cassiterite as small as 0.1 mm. can be recovered, provided they are not present in excess, that is if the material treated on the jigs contains sufficient coarser sands to keep the interstices between the grains open. Removing the fine material from the pulp by screening, to prepare the jig feed, is impractical, because that means the use of an 80-mesh screen or finer, and is not as effective as hydraulic classification. A screen would eliminate mineral grains from the feed which can be recovered by jigging, and in addition would throw more gangue and low-grade middling on to the jig. Formerly a one-spigot Richards vortex classifier had been used to accomplish this separation, but the tangential openings through which the hydraulic water enters became clogged by fine vegetal matter with which the water was contaminated, and which it was impossible to remove from the water. Consequently the work of the classifier was imperfect and a good deal of slime was sent to the jig. When designing the plant referred to in this paper it was decided to use a hydraulic classifier as shown in Fig. 1. The essential features had been copied from the Anaconda classifier.' This remodeled classifier consists of a truncated pyramidal wooden hopper A with an attachment of cast iron

Jan 1, 1917

By David S. Butts

The high cost of energy has made many solution mining corporations turn from fuel consuming evaporators to solar energy and solar ponds. The generally held concept of a solar pond is an area enclosed by dikes into which brine is pumped, water evaporated by the sun, and the resulting salts scraped from the bottom of the pond. This paper will show that solar ponding is not simple. It is a science involving six main areas. These areas are (1) land conditions, (2) meteorology, (3) fresh water, (4) accessibility, (5) manpower and (6) mineral resources. Each area is discussed to identify important parameters that must be investigated before the viability of a solar pond complex can be established as a mineral recovery step.

Jan 1, 1985

By E. F. Johnson, F. H. Brinkman, H. J. Welge, A. L. Hicks

A method is presented for predicting the character of gas or water displacement in a radial system, which can be either horizontal or inclined. The latter case would comprise cone-shaped or dome-shaped symmetry. The method conriders the one dimension of radial distance, taken parallel to the bedding planes. Approximate allowance can be made for segregation of the two fluids perpendicular to the bedding planes. A complete saturation history can be obtained in about a day's time with the use of a desk calculator. Results of an example calculation agree well with the performance of simulated gas drive in a large model representing a section of a cone. This pie-shaped, sand-packed model was 8-ft long and 4-in. thick. The method of solution is useful in handling other partial differential equations of first order and first degree. INTRODUCTION A new analytical method has been developed for predicting the performance of gas or water drive in a reservoir having radial or cone-shaped symmetry (see Fig. 1). In addition, this new development offers a general method of handling partial differential equations of first order and first degree. This work extends those methods developed previously'- "or predicting the displacement of oil by gas or water in linear systems. The method of calculating oil recovery in a radial system involves two major parts. The first part is a calculation of the distribution of saturations at various times of depletion. The second part is the calculation of oil recovery and producing gas-oil or water-oil ratios through volumetric balance on the saturation distributions obtained in the first part. This second part differs only slightly in principle from similar calculations made with linear systems, for instance. The basic assumptions used in the calculations are as follows. 1. One-dimensional radial flow. An approach is explained later for approximating a second vertical dimension. (With- drawals of oil, and of the driving fluid after breakthrough, are assumed uniform around the periphery of the dome.) 2. Constant-volume fluid flow. Thus, in a gas drive there is no change in gas solubility, or in oil or gas volume due to viscous and gravity pressure drops. 3. Capillary forces are negligible. 4. Constant rate of injection. A program has been written for carrying out the calculations needed on a digital computer. However, a prediction can also be arrived at using a desk calculator; this latter method of work requires about a day's time for each prediction made. CALCULATION OF SATURATION DISTRIBUTION For definiteness, we describe the development of our method for the case of gas drive from the crest of the structure. The reader will note, however, that the method is just as applicable to an advancing radial water drive. The equation describing gas flow over an infinitesimal

By W. A. Roper

Since 1950, several papers have been published which have described various methods for studying mobility ratio effects. The methods which have been described for studying mobility ratio effects include X-ray shadowgraph techniques, stepwise potentiometric model procedures, stepwise numerical procedures, gelatin model studies and fluid mapper model studies. The purpose of this paper is to describe a procedure for using a resistance network for studying mobility effects. DESCRIPTION OF APPARATUS The basic unit of the experimental model consisted of a network of 840 resistors fastened to a sheet of black Lucite which was mounted on a wooden frame, Fig. 1. The dimensions of the Lucite sheet were 36 X 36 in. The resistors were fastened to the Lucite sheet at mesh points by means of snaps. The mesh points were arranged in a square pattern, 30 X 30 in., spaced at intervals of l1/2 in. The snap fasteners permitted several resistors to be "stacked" in parallel. By adding additional resistors to the basic unit, the specific resistance on each side of the flood front could be varied. The square network of resistors represented one of four symmetrical elements of a five-spot well pattern. The current input was analogous to an injection well and the current output was analogous to a production well." The power supply was essentially a full wave transformer-rectifier system for converting alternating-current into direct-current. The DC voltage could be varied from 41.43 to 290 v in steps of 41.43 v by selecting the proper outlet jacks. The major components of the power supply were: a Triro power transformer, R 11-A; a dual 5Y3 vacuum tube; two gas filled OD-3 tubes; and a Triplett voltmeter, model 227-1, range 0 to 500 DC v. The voltages at the mesh points were determined with a Keithley electrometer, model 210, No. 185, manufactured by Keithley Instruments, Cleveland, Ohio. A Heathkit vacuum tube volt-

By D. Malhotra, R. M. Hoover, F. N. Bender

The paper discusses the effect of aeration and agitation on grade and recovery of molybdenite. A series of factorially designed flotation tests were carried out in a "Leeds Automatic Laboratory Flotation Cell." Analysts of flotation results indicate that molybdenite has fast floating and slow floating components. The rates of fast floating components, which do not create a problem inflotation, could be varied by changing agitation or airflow rates. However, the rate of slow floating components was stable and was only a function of residence time. Aeration has a significant effect on the grade of MoS2 and only a slight effect on the recovery of MoS2. Agitation has a slight effect on the grade, but does not effect the recovery of molybdenite. The implications of the results on the design of flotation cells and control and optimization of the existing circuits are also considered.

Jan 1, 1981

By William Kent

ABOUT a year ago, the writer had occasion to assist Mr. John L. Gill, Jr., of the Pittsburgh. Car-wheel Works, in making a trial of his new testing machine. A piece of Bessemer steel, of about .34 carbon, was used as a test-piece, turned to a sectional area of one square inch for a length of five inches between shoulders, and pro¬vided with large round heads, which were fitted to what is called the " Wade grip." Measurements of elongation, accurate to , 1/10000 of an inch, were made by a special apparatus containing two micrometer screws, with electric contact attachment. The machine was operated by hand-power. Readings of elongation were taken at each increment of 2000 pounds stress per square inch, up to 44,000 pounds, the elongation corresponding to the latter figure being .0088 inch. The elongations up to this point remained almost exactly proportional to the stresses, the strain diagram, as shown on the Plate, from the origin to the point A, being a straight line (the deviation of no observation from the straight line being greater than .0002 inch). There being no indication of an approach to the elastic limit, the attendant at the hand-wheel and scale-beam was told to proceed as usual and move the poise forward 2000 pounds, or to 46,000, and turn the wheel until the beam was balanced. Noticing that the beam did not rise in the usual time, and that the elongation was increasing rapidly, the writer called to the attendant to stop turning the hand-wheel and to push the poise backwards until the scale-beam balanced. Instead of balancing at between 44,000 and 46,000 pounds, as was expected, it balanced at 40,000 pounds, and the elongation had increased to .0343 inch (or more than four times that recorded at 44,000 pounds), as plotted at the point B, the dotted line between A and B showing the supposed course of the strain diagram between these two points. Immediately upon recording the elongation the writer said to Mr. Gill, who was standing alongside: "This phenomena is without a parallel in my experience, and I know of but one way to account for it. The piece must actually be broken in its interior, through a flaw or weak portion of the metal. The test-piece is now equivalent to a tube, with

Jan 1, 1880

By H. K. van Poollen, H. C. Bixel

A treatment is given of the transient pressure behavior of a well located at the center of a circular region surrounded by a radial discontinuity. On either side of the discontinuity, the values of permeability, viscosity, compressibility and porosity are uniform but may be different from those on the other side of the discontinuity. The results are obtained by solving a pair of finite difference equations. The numerical solution to these equations is obtained using a digital computer. The results are compared with previously published, approximated analytical solutions to the same problem. Solutions, presented graphically, show pressure decline for constant rate fluid production. The range of variables studied include dimensionless time from '0.001 to 100 and storage capacity ratio from 0.001 to 1,000. The well behaves as if it were in an infinite reservoir for dimensionless times less than 0.25. Reservoir properties near the well can be estimated in the usual manner. An overlay technique is used to match an experimental curve with one of the theoretical curves. It is possible to estimate the distance to a discontinuity by substituting the actual time t and the corresponding dimensionless time tD at which a match occurs into the equation a = (0.000264 tk1/tD , µ1 cI)I/2 where kI/ I µI cI is the diffusivity near the well, and may be estimated from data taken at early time. Several buildup curves are computed. These curves show that for early shut-in times, correct values for transmissibility are obtained from conventional analysis. However, erroneous values of static reservoir Pressure are obtained unless data at large shut-in times are taken. INTRODUCTION A mathematical treatment of the transient pressure response of a well located at the center of a region bounded by a circular discontinuity is given. Within a region (Fig. 20) the properties of both the rock and the fluid are considered to be constant on either side of the discontinuity. While these properties are considered constant, they may be different on opposite sides of the discontinuity. A discontinuity of this type could be a fluid-fluid contact or a sudden change in rock characteristics such as thickness, porosity or permeability. Analytical solutions to this problem are available. 1-3 However, they are so involved that they are of little practical use. An approximate solutionl is available but the range of times over which it is valid has not been specified. A numerical solution to a pair of finite difference equations is used to obtain the solution given in this paper. PREVIOUS WORK One of the first solutions to the problem was published by William Hurst.1 Hurst considered unsteady-state flow of fluids through two sands in series with different mobilities in each sand. He used Laplace transforms to obtain a solution for a single well located at the center of the circle enclosing the first of the two sands. In this case, the solution for the pressure change in Region I (Fig. 20) is2

In an earlier day, a moderate-sized 600 tpd mine could easily require 60 working faces and might use five supervisors. The miner had a tonnage rate as an incentive and was largely his own boss. With the advent of conventional mining machine crews, the typical pattern of resource used changed radically. The number of work places could be cut by a factor of ten, but supervision only by a factor of somewhere around three, or at most, five. Obviously, supervision was increased. Capital use rose at the same time by a factor of approximately eight. Materials used, such as powder and bits, probably in- creased in both quality and unit cost, too. It certainly required a better-trained man to run the machines, but fewer miners were needed. It is clear that each input factor such as land, labor, capital and management, has changed significantly and it would be foolish to attribute all the gain to improved machines. However, concentration of work was a vital consideration in the change, sketched above, and such focusing of work in fewer areas could not have occurred without high capacity mining machines.

Jan 11, 1968

By J. F. Myers, R. J. Tower

"There is nothing new under the sun." All over the world, mineral-dressing engineers are working at their problems, no two of which are alike. Each encounters equipment and process problems. Many devise simple ways and means of overcoming their difficulties. Many of these problems are quite common in the industry, yet, since no one human being seems capable of thinking of everything, these problems at some plants often are looked upon as necessary evils. The purpose of this symposium is to gather together, from competent engineers, devices and practices that have been useful to them. Most of the ideas contributed are not new, yet there is no plant operator, old timer or youngster, who will not find some idea herein that will give him a suggestion for improving his operation. Space does not permit a complete elaboration on all of the contributions offered. Each item is summarized, but this will be ample for those skilled in the art of mineral dressing. A list of contributors appears at the end of the paper, and the superior number at the end of each item identifies its source. General Equipment A signal light to show that a steel bin, containing relatively line, moist ore, is full, can be made by suspending a copper rod in the bin to a sufficient depth to ensure ample contact with the ore. One wire of a 32-volt circuit should be attached to the rod and one to the side of the bin. Sufficient current will flow to make the signal light burn. A 110-volt circuit could be used, but 32 volts ensures safety.10 Fig I shows a circular bin for crushed ore. It is a concrete structure footed on the ground, using no supporting columns. Tailing from the sink-float was used for aggregate. Cross conveyors for discharging ore are not required, nor is a tripper for filling. Segregation is minimized by drawing from different gates. An exhaust fan on top of the bin eliminates condensation due to temperature of the ore.21 A beam with an adjustable point of suspension near the center permits rigging rolls for removal in a very few minutes. This makes unnecessary the lacing of the cable through the hub bolts on one side, around the shaft on the other, and the trouble of balancing the load.8 The ratchet handle on a Symons crusher may be replaced by a drum with a cable wound on it. When the cable is pulled by a crane or hoist it causes the drum to revolve and thus screws or unscrews the bowl. This reduces handwork and materially shortens the time required for the job.' TO Spread ore feed from an apron feeder over a belt conveyor, place a heavy steel bar about midway between the top

Jan 1, 1949

By J. F. Bryson

A FULL description of past mining conditions in eastern Kentucky coal mines with reference to the cause of coal bumps and the preventive methods was given by the writer in a previous paper.1 At that time the method had been tried for a period of 13 months, successfully. The method has been enlarged add carried on for over two years, and is still successful. Recapitulating briefly from the previous paper: The No. 3 mine of the Mary Helen Coal Corporation, Coalgood, Ky., operates in the Harlan seam, and is worked by a room-and-pillar method. Cover varies from 1200 to 2000 ft. and the immediate roof over the coal is a hard, massive, sand rock, estimated to be from 70 to 90 ft. thick. The average height of coal throughout the mine is 46 inches. The reason for adopting the method were that in this mine five men had been killed by coal bumps between 1931 and 1934; many other men had been injured and coal-cutting machines had been thrown from under the cut. The morale of the workmen was at a very low ebb. Many remedies had been tried in other mines troubled by coal bumps, one method being to leave chain and room pillars in every 500 ft., to serve a bumping block; the idea being that if a bump took place it would act on the pillars left in the gob. Another method adopted was to board up a pillar, leaving a small opening at the bottom sufficient for shoveling coal. In this system the idea was that if a bump did occur on the pillar, the boards would stop flying coal so that no one would be injured. Many other methods were tried and it is safe to say that each mine troubled by coal bumps tried various remedies, none of which proved effective. On considering all that had been attempted in an effort to eliminate coal bumps, and the information gained by experience of coal bumps in various mines, it was thought that a long face with gob support would eliminate the trouble. Accordingly, on Jan. 15, 1935, a long face was started with wood cribs on 20-ft. centers. When the former paper was written on this method, the long face had been brought back a distance

Jan 1, 1937

By Bruce A. Kennedy

In recent months, public hearings held by the Senate Subcommittee on Minerals, Materials and Fuels on the Stillwater copper-nickel complex in Montana have been reviewed at great length in the press. Although the meetings were well managed by the Subcommittee and questioning was kept in line with the aim of the hearings-namely, how can re- sources be developed in a manner compatible with the preservation of the environment-it was alarming to read the apparently uninformed and questionable arguments presented by some witnesses who were attempting to prohibit mining in the area altogether. One spokesman even went so far as to urge that mining of the area be rejected be- cause, even if the mining operation did not pollute the areas, the effluent and waste produced by the people themselves could constitute a hazard to the environment and to themselves. This, of course, would be true to a point. However, if this or similar types of reasoning were used to reject mining of a large number of orebodies in the future, how is the United States going to meet its future mineral demands and what effect could this have on the economic strength of the country? Fall of the Roman Empire An interesting article in a recent issue of Time magazine came up with a new theory to explain the fall of the Roman Empire. Claire Patterson of CalTec, during a recent study of air pollution, investigated the historical worldwide distribution of lead. As lead was a by-product obtained from silver mining during historical times, he turned his attention to studying the production of silver during those eras. He discovered a very interesting fact-that, I unless the stock of metal is continually being replenished from the mines, accidental loss diminishes a country's stockpile of silver at a rapid rate, as much as 2 per cent per year. Patterson, a geochemist, theorized that with an accidental loss rate of 2 per cent per year, if the country's silver supply dried up, the entire existing stockpile could disappear in about a century.

Jan 1, 1971

By C. J. Delisio

The basic concept of tunnel boring has not changed since the late 1800's. R. Stanley of Great Britain obtained a Canadian patent as early as December 1891. Mr. Stanley's machine was a device that consisted of a rotating wheel equipped with drag-type teeth; it had a means of thrusting the cutterhead against the face or heading; it had a means of clamping or gripping the machine in its bore; it had a means of collecting and depositing the cuttings into a screw-type conveyor located near the invert of the tunnel. It even had a means of varying the speed of the forward motion of the machine to suit different hardnesses of rock encountered. This description essentially fits the latest tunnel-boring machines being built today. In short, there has been no significant advance in the basic concept in 70 to 80 years. It may well be that no significant change can be expected for some time into the future. Today's machines, while still basic in concept, offer many advances aimed at faster tunneling. Among these are 1 ) improved cutters, 2) guidance and steering devices, 3) dust-control devices, and 4) mechanisms to aid in installing tunnel supports. CUTTERS The most important part of a hard-rock tunnel borer is the cutter. If the cutter can cut the rock, it is a boring machine. If the cutter cannot cut the rock, no other part of the machine can. When Jarva, Inc. decided to build a machine, the first problem was to find a good cutter. In our search, we looked at most of the cutters on the market. We selected Reed Drilling Tools' cutter because the design of their cutter was aimed at hard-rock tunneling. We feel that they build the best cutter available and provide the best possible service to their customers. As tunnel-boring machine design advanced, the transition from drag bits to roller bits was natural because harder formations required a roller bit, and it is the goal of all mole users to bore harder formations. Roller bits

Jan 1, 1970

In a letter to the Secretary, Sir Robert A. Hadfield encloses a statement of the terms and conditions under which his prize is to be awarded by the Institution of Mechanical Engineers, and also a table comparing the hardness of a large variety of steels, examined both by Brinell and by scleroscope methods, together with the tensile and compressive strength of the same specimens. The conditions for the award of the prize are as follows: SIR ROBERT HADFIELD PRIZE For a new Method of Determining the Hardness of Metals 1. Sir Robert A. Hadfield, D. Sc., D. Met., F. R. S., Vice-President, has placed in the hands of the Institution of Mechanical Engineers the suns of £200, which, with any income therefrom, may he awarded at the discretion of the Council of the Institution as a prize or as prizes, for the description of a new and accurate method of determining the hardness of metals, especially of metals of a high degree of hardness. 2. The ordinary tests of hardness, such as are described in the report of the Hard-ness Tests Research Committee (Proceedings of the Institution of Mechanical Engi=neers, 1916, pages 677 to 778), which should he consulted by competitors, fail to some extent when the hardness of the material exceeds about 600 to 800 Brinell. What is

Jan 3, 1918