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By L. BENEDICT

COMMENTING on the report of the latest Royal Commission for India, the September, 1926, issue of the National City Bank's monthly letter states, among other things, that "The decision of the Royal Commission is against the proposal for a gold coinage, but in favor of placing the silver coins and paper currency now redeemable in silver on a gold basis." Also, it quotes the following from the Royal Commission's report: The essence of the proposal * * * * * is that the ordinary medium of circulation in India should remain as at present the currency note and the silver rupee, and that the stability of the currency in terms of gold should be secured by making the currency directly convertible into gold for all purposes, but the gold should not circulate as money. It must not circulate at first, and it need not circulate ever. In this connection it may be stated that during the last one hundred and fifty years or so unsuccessful attempts have been made by the governments of different nations of the world to devise a plan or plans by which gold and silver could be used concurrently for money purposes and, particularly, as standard money; such, for example, as bimetallism, which, of course, is impracticable. But, since none of such plans has been adopted generally it is logical to conclude that their proponents did not shape them so as to conform to the basic principle of standard money, or else political issues were deemed to be dominant, or, perhaps, it was partly due to each.

Jan 1, 1926

By Charles Himrod

In presenting this paper it is not intended to enter into any discussion of the theory of the blast-furnace, but simply to give the results of a number of determinations of CO and CO 2 in furnace gases, and some conclusions arrived at, and the result of an experiment made in hopes of increasing the production of CO 2. In the first table, on page 198, is given the averages of a number of analyses of gases of different furnaces, and in the second, of two stacks of different sizes under varying circumstances. The conclusions arrived at as to what influences the production of CO2 are as follows: lst,. Kind of fuel. Bituminous coal not permitting the formation of as much CO, as coke. 2d. Kind of ores. Rolling-mill cinder being much less favorable to formation of CO2 than hematite ores, magnetites probably less than hematites, but more than cinder. 3d. Length of time of contact of gases with ore. Small furnaces and too rapid driving both waste fuel by riot allowing enough time for conversion of CO into CO2. 4th. The temperature of blast. The hotter the greater the quantity of CO2. 5th. Regularity of passage of stock through the furnace. The hanging and dropping of the stock, however slight, seems to pack it more solidly in the hearth, increases the resistance to the blast, and decreases the quantity of CO2. These are the most common causes noticed which influence the production of CO2, unless we except the one that might be considered the greatest of all, the one that has spoiled so many tons of iron, and the amiability of so many furnacemen, the one most dreaded, the great unknown. It would seem a self-evident proposition that the better the contact of the gases with the ore and the less with the fuel, the greater would be the amount of CO2 produced. To prolong the contact of gas with the ore furnaces are enlarged with advantage, and pains are taken to charge the stock in such a manner as will best promote it. Still, with furnaces of most approved construction and most skilful manipulation, the gases leave the furnace so rich in CO, that not half the heating power of the fuel is obtained, and the highest

By E. E. White

(Cleveland Meeting, October, 1912.) IN, August, 1911, I read a paper before the Lake Superior Mining Institute' on surveying and sampling diamond-drill holes. The present paper gives a more thorough description of these methods, together with a few changes and notes suggested by further experience. Since August, 1911, the methods, used constantly by the Cleveland-Cliffs Iron Co., have been developed to a higher degree of efficiency, and have given great satisfaction. Until a few years ago but little attention was paid to the determination of the actual course of drill-holes, to the accurate sampling and analysis of the material, or to the scientific location of the drill-holes. It has now been common practice for several years in many districts to test the inclination of drill-holes by etching glass tubes with hydrofluoric acid, but the direction of deviation from the vertical has not often been determined. Methods of determining the direction by means of a compass were first developed in connection with drilling in Victoria. The latest method used on the Rand in South Africa has been recently described by John I. Hoffman.2 In the same paper Mr. Hoffman describes an ingenious method of wedging-off a drill-bole so as to start new holes at several successive points. By this means a hole can be deflected in any desired direction, and the curvature can be controlled to a considerable extent, provided the rock is soft enough to permit the hole to be wedged off. Unfortunately, in the Lake Superior district, the jasper iron-formation is so hard that it is very difficult, if not impossible, to do this. I tried to deflect one hole so as to avoid drill-rods stuck in the hole, by using a 1 Proceedings of the Lake Superior Mining Institute, vol. xvi., pp. 100 to 120 (1911). 2 Recent Practice in Diamond-Drilling and Borehole Surveying, Bulletin No. 91, Institution of Mining and Metallurgy (Apr.. 11, 1912).

Nov 1, 1912

By T. L. Joseph

THE committee appointed to investigate the application of oxygenated air to the blast' furnace believes, after a careful study of the problem, that this enriched air will increase hearth temperature, tonnage, flexibility of operation, and quality of pig iron made, and that fuel consumption will be decreased. These improvements are supported by theoretical evidence; but a practical trial of the process is necessary to establish more definitely the magnitude of benefits realized. The commercial application, however, may profitably be preceded by a test on a smaller scale. A small furnace will, in general, respond to the enriched air in the same way as a commercial furnace and the magnitude of the benefits to be gained from varying additions of oxygen can be fairly well established at relatively little expense. A new experimental furnace, being built jointly by the Bureau of Mines and the University of Minnesota, at Minneapolis, will have a capacity of about 6 tons and will riot require more than 700 cu. ft. of natural air per minute. Its hearth is 36 in. in diameter and its height about 32 ft. Enough oxygen to enrich the blast to the extent recommended by the committee can be obtained from a small oxygen unit. It might even be cheaper to perform preliminary experiments with tanks of oxygen obtained from regular, dealers. A procedure in making the test that suggests itself is to start with operation on normal air; then to use blast, enriched in several steps. Systematic temperature measurements of the tuyeres, slag, and metal would reveal any changes in hearth temperature. When these changes become large enough, the burden may be gradually increased to just within the breaking point and the furnace brought back to normal, if possible, by the addition of oxygen. Such a test would show what might be expected in the way of a decrease in fuel consumption and the time taken to "straighten up" the furnace by means of oxygen. The new furnace at Minneapolis is so equipped that the furnace process can be observed from the stock line to the iron notch. We have been taking gas samples at various elevations in the furnace to study combustibility of the fuel and the reducing power of the gas. If either of these is affected by removing inert nitrogen, the change can easily be measured.

Jan 11, 1924

By V. J. Bishop, R. G. Beverly

Recent publicity has focused attention on the hazard of lung cancer among uranium miners. Many mining people feel the danger has been exaggerated. Whether or not this is the case, the industry must thoroughly evaluate this potential hazard and take corrective measures to minimize the threat of radiation after-effects and to maintain the level of radiation in mines within the presently-accepted limit. The measurement and control of radon gas and its solid radioactive decay products, Po-218, Pb-214, Bi- 214, and Po-214, should be the one facet of most concern to uranium miners in any radiation control program. To illustrate the radioactivity involved with small quantities of radon gas, the activity from the decay products of five milliliters of radon gas (imagine a 5-ml vial) would be adequate to raise one cubic mile of air to seven times the recommended maximum working level of 300 [ ] (micro micro curies) per liter. This would be enough radioactivity to raise to the maximum working level a 6 x 8-ft drift which extended around the world 160 times. Obviously, a little radon gas goes a long way.

Jan 8, 1961

By David H. Browne

(Pittsburg Meeting, March, 1910.) NICKEL has always been a fruitful mother of problems. Previous to the year 1906 nickel was regarded as an element replacing iron in copper-mattes, and it was believed that the same laws which governed the elimination of iron could be applied-to nickel. This conception has lately undergone a complete change, and nickel in all its stages is now regarded as an element replacing copper; in other words, copper-nickel is considered now as two elements, but as one metal. When we commenced our investigations on Monel metal we made a great many experiments on the conversion of mattes, both high copper and high nickel. We studied closely the composition of the mattes and the changes produced by the blow. We found in all cases certain agreements which we believe are due to some law. Briefly stated, our conclusions are as follows 1. Nickel is not an element replacing iron in matte. 2. Nickel-copper alloys act in the matte-blow like one metal. 3. Nickel-copper alloys follow, during the matte-blow, exactly the same laws as govern. the behavior of copper alone. This brings up the question : What are the laws governing the production of metallic copper, and what are the relations copper to sulphur and iron during the Bessemer blow ?

Apr 1, 1910

By L. O. Howard

What was perhaps the first attempt at basic converting in the Southwest was made by the late Charles F. Shelby at Cananea early in 1907, when he removed the acid lining from one of the 8 by 12-ft. barrel-type converters then in use at the reduction works of the Cananea Consolidated Copper Co. and substituted a lining of one course of magnesite brick. In this shell, he blew 21 taps (about 180 tons) of matte to white metal before the brick gave out along the tuyhre line. In every case, the white metal so made was transferred and blown to copper in other converters. The cost of the lining was $700 and Mr. Shelby gave up the experiment as impracticable. No further serious effort to bessemerize copper mattes in a vessel having a more or less permanent lining of basic brick was made by any of the large copper smelters in the Southwest until some months after Messrs. Peirce and Smith had brought the question permanently before the public, and had proved it to be a metallurgical and financial success by their work at Baltimore and Garfield. Early in 1911, the Cananea Consolidated Copper Co. had in operation at its reduction works at Cananea, Mex., several 8 by 12-ft. barrel-type converters lined with magnesite brick which had previously been operating with acid linings. In the 18 months following, these converters produced 50,000 tons of copper. In June, 1912, this company installed the first of its new equipment of six stands of electrically operated 12-ft. Great Palls type converters, and shortly afterward discontinued the use of the smaller shells. In September, 1911, the Consolidated Kansas City Smelting & Refining Co. had installed at its El Paso plant two 10 by 26-ft. Peirce-Smith converters. Later, in March, 1913, one standard Great Falls type converter was added. About the time the Peirce-Smith converters were going in at El Paso, the Calumet & Arizona Mining Co. was experimenting with basic linings in 7-ft. by 10-ft. 6-in. barrel converters at its plant in Douglas, Ariz. These efforts were sufficiently successful to enable the plant to "handle the total output of copper in these small shells until the standard Great Falls converters were substituted in June, .1914.

Jan 1, 1917

By T. F. Bower, M. C. Flemings, H. D. Brody

A series of experiments are reported which show thai reasonable assumptions for analysis of solute redistribution in solidification of castings and ingols illclr(de: negligible undercooling before nuclcation of solid phases, negligible increase of solute in advance of the tips of growing dendrites, complete diffusion within the Liquid oer distances the order, of dentrite arm spacings, platelike dentrite t/o?$holog?.. Qzmli/ali,e and periment is obtained JOY Al-Cu alloys front measurements of amount of nonequilibrium second phase, minimum solute content at the centers dentrites arms (after solidification), and 3) change in minimum solute content in dentrite arms during solidification. In a previous paper' analyses were given describing solute redistribution in dendritic solidification. The analyses were undertaken primarily to permit prediction of "severity of microsegregation" in a casting or ingot after solidification and cooling to room temperature, but can also be used to predict "local fraction solid' as a function of temperature in the solidifying alloy. Numerical results were given for Al-Cu alloys. The aims of this paper are to examine the validity of some of the assumptions of that analysis, and to compare predictions of the theory with experiments. All experiments reported are on Al-Cu alloys, the bulk on Al-4.5 pct Cu alloy. EXPERIMENTS Experiments conducted were on small ingots, 41/2 in. in diam by about 10 in. high. solidified in a controlled thermal environment (i.e., in a resistance-heated, re-circulating air furnace, whose temperature could be controlled to 1°C). In one series of experiments, the ingots were solidified very slowly (solidification time was varied from 8 to 1000 hr) and at constant rate of heat extraction. In a second series of experiments the ingots were solidified ( 'unidirectionally". Fig. 1 is a sketch of the method of making the two different types of ingots. The unidirectionally solidified ingots were fully columnar, and the slowly solidified ingots had equiaxed structures, Fig. 2. In making the slowly solidified ingots, the mold employed was carbon steel, coated with inert refractory. Typically the mold and its environment were heated to 700°C. The molten alloy was then poured into the mold and the temperature of the system equilibrated above the liquidus. The environment of the mold was next cooled with a controller responding to a differential thermocouple that measured the difference between the temperature of the melt and the temperature of the mold environment. The differential thermocouple consisted of one chromel-alumel junction in the melt connected in opposition with a second junction in the furnace. The potential across the differential thermocouple was input to the controller which acted to keep this potential at a set. constant value; hence the temperature difference between the two junctions was maintained nearly constant. Major resistance to heat flow in the above apparatus was between the metal surface and the furnace atmosphere, not in the metal itself. By maintaining a constant temperature difference between the metal and the furnace atmosphere, the controller maintained an essentially constant rate of heat extraction during solidification, regardless of amount of heat released by the metal per unit temperature change. When the metal temperature dropped just below the eutectic temperature. the ingot was removed from the furnace and water-quenched. The unidirectionally solidified ingots were cast in a similar mold to the foregoing except that provision was made to extract heat through the bottom face of the mold by passing air or water through the mold base, Fig. 1. For these experiments, metal and mold temperature were again equilibrated above the liquidus. Temperature of the mold environment was then held

Jan 1, 1967

By H. H. Klepfer, M. E. Snyder

UNTIL very recently, the commercial availability of the rare earths as metals has been very limited. Fabrication of mill products from these metals has not been studied in most cases. This note reports the results of the development of fabrication techniques for thulium. Thulium has a melting point of about 1550°C and a hexagonal close-packed crystal structure. It oxidizes in air to give a black oxide (TmzO,). The procedures for producing thin thulium foil were developed on one ingot weighing about 220 g and were subsequently applied in processing about a pound of metal to foil. Several alternates for the various fabrication steps were investigated and will be discussed. The metal fabricated was in the form of commercial chill-cast ingots 1 in. in diam and 2 in. long weighing approximately 220 g. Impurities in the ingots were reported by the vendor to be 4000 ppni tantalum, 2000 ppm calcium, 200 ppm nickel, 100 to 200 ppm iron, 100 ppm europium, and less than 100 ppm copper, lutetium, and ytterbium. In addition to these impurities, several salt-like inclusions as large as l/8 in. in diam were revealed along the center line of the one ingot sectioned. Preliminary tests indicated that small wafers cut from the as-cast ingot would not fabricate readily by rolling. Forging of copper-jacketed wafers was therefore attempted. At 1550"F forging was satisfactory but an apparent reaction of copper with thulium demanded investigation of lower temperatures. Therefore, the remainder of the test ingot was forged at 1450°F—with only minor cracking. All ingots forged were inserted into copper tubes of 1 in. 1D and 0.125-in. wall thickness. The jackets were sealed by flattening the ends of the tubes and welding under helium. Heating time at 1450°F was 30 min. Press forgings of 1/8 in. per pass were used, followed by 10 min reheats. When the ingots had been squared and reduced to 0.250 in. in thickness, the original copper jacket was stripped off and replaced by a new jacket in preparation for hot rolling. Hot rolling at 1450" F without edge cracking was readily accomplished after forging. Excellent results were obtained with 10 pct reductions of thickness followed by 5 to 10 min reheats. After reduction of thickness from 0.250 to 0.100 in. the copper jacket was removed. It was found, in fact, that hot rolling in air was possible. A tenacious black oxide similar to that seen on zirconium was formed during 3 min reheats at 1450°F. Reduction in air to 0.010 in. foil was possible taking 10 pct reductions per pass. The oxide coat formed during hot rolling in air could best be removed by sand blasting and pickling. Common pickling solutions containing polar solvents were found to attack the metal too rapidly and a concentrated nitric-hydrofluoric acid mixture attacked neither the oxide nor the metal. The most satisfactory pickling solution was 52 vol pct concentrated nitric acid-48 vol pct glacial acetic acid. After forging to 0.250 in., vacuum annealing and cold rolling was found to be another satisfactory alternate to hot rolling in copper jackets. After forging. a hardness of Rockwell B76 was found. Annealing in vacuum (2 X l0-5 mm of Hg) for 1 hr at 1200"F did not alter this value. Annealing at 1470"F for 1 hr brought the hardness down to R;]63. With cold rolling (5 pct per pass) the hardness returned to about R,1:76 after 20 pct reduction and edge cracking became noticeable. However, cold rolling to a total of 40 pct reduction in thickness (RI,,83) was possible before edge crack propagation became serious. Good surface finish was obtained, and the metal loss due to oxidation was minimized by cold rolling and vacuum annealing. Using this procedure the yield of 1.25 in. wide by 0.010 in. thick foil from a 1-in. diam ingot was about 40 pct.

Jan 1, 1961

By D. S. Choi, R. Stefanko

There are a number of complex factors that influence the effectiveness of anchorage to maintain tension in rock bolts. However, a plastic analysis of the anchorage site employing certain simplifying assumptions with application of the Mohr-Coulomb criterion appears to explain the observed phenomena. Such an analysis has been made and a correlation sought with field and laboratory tests. Field tests were made in an anthracite mine in eastern Pennsylvania and included pull tests and long-term tests of a variety of anchorage devices in two basic lengths, 30 and 42 in. in two widely differing seams. Performance is reviewed for wedge, expansion shell, and resin anchorage. Laboratory tests duplicated many of the field conditions but in addition compared the performance of shells with normal and reversed serrations. This performance was compared with the predicted results from the plastic analysis. One of the major problems in conducting long-term underground tests is the selection of suitable instrumentation. All installed bolts were fitted with spherical and hardened washers to insure the best possible torque wrench readings. In addition, commercially available load cells were used. Finally, the performance of a specially developed strain-gage-equipped ring cell is viewed. Rock bolting as a method of support continues to increase with applications in many other industries in addition to mining. Nevertheless, with nearly 55,000,000 roof bolts installed in coal mines alone last year, this remains as the single greatest use. While bolts have frequently supported ground where conventional timbering could not, there are relatively few design criteria; and trial-and-error procedures prevail. Furthermore, there has been a lag in development of suitable instrumentation that is simple to install and read out, sensitive, durable, reliable, safe, and economical in evaluating the effectiveness of a bolt over long periods of time. Therefore, the pull test continues to be the most popular method of evaluating the applicability of a certain type of roof bolt under specific installation conditions. At The Pennsylvania State University in the Dept. of Mining, research has been conducted for a number of years to measure bleed off in carefully controlled laboratory experiments as well as in underground investigations."-' Unfortunately, most of the instrumentation developed has been primarily suitable only for research purposes, not possessing all of the aforementioned characteristics desirable for routine underground use. Other groups also have met with restricted success. Therefore, while relatively crude, the torque wrench continues to remain as the most widely used load measuring device. While both field and laboratory tests continue to be con- ducted, analytical analyses are attempted to discover the more important design parameters in order that more efficient anchorage might be devised. Bolts are being used for a greater variety of purposes in mining. Suspending wire sideframe belt conveyors from roof bolts is a common application. The suspension of a monorail transportation system presents yet another. One such system has just been installed in a recently reopened anthracite mine and is presently being evaluated under production conditions. Preliminary studies revealed that a considerable cost reduction was possible by suspending the monorail on bolts anchored in the top. The monorail was to be installed under two widely differing conditions—a competent sandstone above the Buck Mountain seam and a softer shale top above the Skidmore. The type of anchorage device, length of bolt, and long-term performance, consistent with economy and safety, had to be established for the installation once the decision was made to suspend the system on rock bolts. This paper describes some of the testing procedures leading to a final selection. Theoretical Analysis of Expansion Shell Anchorage A detailed look at an expansion shell assembly might shed some light on the factors involved in the design of a suitable shell, Fig. 1. When a bolt is rotated, the tapered plug is forced downward, expanding the leaves laterally to grip the sides of the hole. Two friction surfaces are present: (1) the interface of the plug and leaf and (2) the interface between leaf and rock. The relationships of these friction planes, geometry of expansion shell, and properties of the rock are important in the design of an expansion shell. Therefore, an analysis assuming the rock to behave as a rigid plastic material with its yield governed by the Mohr-Coulomb criterion was made." Furthermore, the effect of friction between the leaf and rock produced by serrations was analyzed.

Jan 1, 1971

By J. M. Smith, G. P. Willhite, J. S. Dranoff

Heat transfer rates were measured in sandstones with flow of gases perpendicular to the direction of energy transfer. Effective thermal conductivities ker ranged from 0.7 to 1.7 Btu/(br)(ft)(°F). The contribution of the solid phase appeared to be the most important in these consolidated materials, although the thermal conductivity of the gas had some effect. The velocity of the gas through the pores of the sandstones had no influence upon ker up to values of 168 lb/(br) (sq ft) in agreement with data obtained for unconsolidated beds of glass beads. The present results indicated that gas mixing, and hence heat transfer by convection in the pores, is less for perpendicular transfer of energy than when fluid flow and energy transfer are in the same direction. HEAT TRANSFER PERPENDICULAR TO FLUID FLOW IN POROUS ROCKS Heat transfer in porous media with pore sizes in the micron range depends upon the fluid in the pores and the geometry of the solid phase. The best characterized system is a bed of solid spherical particles. Heat transfer in this system has been studied extensive1y (Ref. 8 summarizes the literature up to 1959) when the pores contain stagnant fluid. When the fluid is in motion the directions of flow and energy transfer have an effect on the heat transfer rates, as demonstrated by comparing the work of Willhite, et al l3 for perpendicular flow and that of Kunii and Smith9 for parallel flow of energy and fluid. For perpendicular flow, no increase in effective thermal conductivity ke was noted up to mass velocities G of 77 lb/(hr) (sq ft). In contrast, for parallel flow ke, increased with G in the same range of flow rates. These higher values of ke in the direction of flow also have been observed3,4,10 in beds of larger particles, 0.1- to 0.5-in. diameter. For beds of consolidated materials, such as porous rocks, data are not available for these compar- isons, although Adivarahan1 reported results for the parallel case. Hence the primary objective of this work was to measure ke values for perpendicular flow of fluid and energy in porous rocks. Of interest also was the variation in effective conductivity with fluid velocity. APPARATUS AND PROCEDURE In the experimental method, a constant heat flux was applied to the inner surface of an annular section of the porous rock. By cooling the outer wall, a temperature gradient through the annular sample was established and measured with thermocouples placed within the sample at various radial positions, and at three elevations (A,B.C). The location of the 2-in. O.D., 3.75-in. long sample in the apparatus is shown in Fig. 1. Fluid entered the bottom (1) of the 3-in. I.D. (approximate) steel shell, flowed upwards through the sample and out at the top (3). pressure taps (2,4) were used to check the permeability of the sample. The energy flowed radially from the centrally-located electric heater through the sample and was absorbed in the water-cooled jacket. The samples studied were naturally occurring sandstones from different locations with the properties given in Table 1. These materials are identical with those used by Adivarahan 1 for the parallel flow of energy and fluid. Prior to use they were refluxed with toluene to remove hydrocarbons and leached with distilled water to remove soluble salts. Each sample was visually examined and discarded if large nonhomogeneities, such as cracks or stone particles, were noted. Eight copper-constantan thermocouples were inserted in holes drilled radially into the sample with

By W. F. Carew, F. A. Crossley

IT has been reported that the Ti-8 pct A1 alloy is ductile as water quenched from 800°C but brittle as annealed at 650 °C." The present, somewhat limited, investigation was undertaken to discover the cause of the embrittlement. In order to evaluate to some extent the continuity of the embrittling phenomenon, alloys of 6, 8, and 10 pct A1 were studied. Subsequent to the completion of most of this work, Ence and Margolin reported that a Ti,A1 phase exists and that two or more unidentified compounds may exist in the Ti-A1 system.' Materials, Equipment, and Techniques The alloys were prepared using Kroll process sponge of 124 Bhn hardness. In addition, some 8 pct A1 material was made with iodide titanium. The sponge-base alloys were prepared as melts of 4 lb or more by double arc melting. The iodide Ti-8 pct A1 alloy was prepared as 200-g buttons. The alloys were forged at temperatures of 980" to 1150°C. Results of chemical analyses are given in Table I. A Baldwin-Southwark 60,000-lb capacity tensile machine, equipped with an autographic stress-strain recorder, was used for the tensile tests. Standard tensile test specimens, 0.252 in. diam, were used. Standard Charwy V-notch impact specimens were tested using a -kiehle machine. specimens were sealed in evacuated Vycor bulbs for heat treatment. Powder samples for Debye-Sherrer anal- ysis were prepared by filing and screening through 325 mesh. The —325 mesh fraction was taken in an effort to concentrate a microstructural constituent assumed to be less ductile than the a phase. The powder samples were stress relieve2 by annealing

Jan 1, 1958

By C. W. Dunham

DESIGN of the structures for the Morenci Reduction Works involved many interesting problems. Naturally, the chief purpose of these structures is to house and support the equipment and other things necessary for or incidental to the conduct of production. hence most of the features of the design were developed to suit these particular conditions; however, several are so generally applicable as to warrant description.

Jan 1, 1942

By James R. Jones

The concern for air pollution goes back centuries as will be seen from this quotation: "Strife and coal, it seems, have a hand-in-hand historical relationship. It was thought by some . . . in the Middle Ages that burning coal filled the air with dangerous poisons. So great was the concern that in 1306 King Edward I of England passed the first and unquestionably the harshest, air pollution legislation edict that history will record. His proclamation declared the use of coal an act punishable by death." As related to coal mining the air pollution problem is primarily that of its customer when using the fuel. Thus the major impact relates to markets for coal rather than the actual mining operation itself. In the vicinity of the coal mine the problem is principally related to fugitive dust from roads, transfer facilities, loading facilities, crushers. etc., and particulate matter (dust or fly ash) from thermal dryers and air cleaning tables. FUGITIVE DUST Most state air pollution regulatory agencies now have rules and regulations requiring control of fugitive dust. This refers to dust generated from normal processing operations. A typical air pollution control agency regulation reads as follows: "Fugitive Dust" "No person shall cause, suffer, allow, or permit any material to be handled, transported, or stored; or a building, its appurtenances, or a road to be used, constructed, altered, repaired, or demolished without taking reasonable precautions to prevent particulate matter from becom-

Jan 1, 1973

By W. F. Mullen, C. W. Stickler

WITH few exceptions, unit operations in the Pennsylvania slate industry in 1950 did not differ appreciably from production methods described by Behrel and Bowles2-4 several decades ago. Many traditionally picturesque but relatively inefficient hand operations continued to contribute to high operating costs in an industry in which the margin of profit, for various reasons, was admittedly low. As part of a general program in which the Pennsylvania State College has been assisting the slate industry to solve some of its problems, time and method studies were used in an effort to determine the bottlenecks of production flow. The results are as applicable to other stone industries as they are to slate. A study of the basic elements in the production of roofing slate, structural slate, and slate blackboards among selected slate producers revealed comparatively few refinements. Any marked difference in operating method was characterized largely by the available equipment, a noticeable difference in the working properties of the rock quarried by a particular company, and by either the habit thinking or ingenuity of the individual operators. With the hope that significant economies in manufacture might result, all the unit operations were broken down into elements for study and analysis. The greater part of the study encompassed those operations from the quarry rim to the finished product; however, to obtain a reasonable synthesis of time allotments per operation, several analyses of quarrying from the finish of the wire saw cut are included for reference. The complete study of quarrying made by Bowles several years ago resulted in the introduction of the wire saw into the operation; there appears to be a need for as fundamental a change in the processing of finished slate as the wire saw was in the quarrying process. Fig. 1 illustrates the flowsheet of quarried rock from the parent bed to the quarry rim with the basic elements which contribute to the difficulty of removing slate from the ground. In Pennsylvania roofing slate is produced by one of three procedures, many operations of which are quite similar. In the traditional or classical method the quarried rock is reduced by sculping and auxiliary manual treatment to workable size and then is split and dressed (trimmed to size) in the conventional shanty. Archaic as this method might appear, there are several elements that result in substantial savings over more mechanized operations. In a modified version of this method, the block is reduced by sawing and gouging preparatory to splitting and dressing in indoor stations. This method is particularly adaptable to rock which fractures unevenly by conventional sculping methods. In the third and more modern method, reduced block received from the block maker is diamond-sawed to length and finished on a production line. Fig. 2 illustrates the flow pattern in each case. In the modified and more modern plans, dressing is done on mechanically-operated trimming machines, which appreciably reduce the fatigue factor of the operator. However, it should be noted that it is advantageous for the operator to be able to control the speed of the knife to prevent breakage of certain classes of stock. In the classical plan the splitter and dresser normally act as blockmakers also and carry their eighter slabs (equivalent in thickness to eight shingles) into the shanties themselves, all of which adds to fatigue and reduced efficiency. The extremely low investment and operating expense of the classical method has undoubtedly been of paramount importance in its continuation. Production of Mill Stock The production of mill stock is accomplished by less picturesque methods than enter into the production of roofing slate but it is, in the main, relatively more efficient because of the increased use of machinery for handling and for finishing. Mill stock can be classified into two categories as far as difference in production method is concerned: structural slate and blackboards. The significant difference in operation is caused largely by the nature of the rock used for each product with the best and easiest cleaved rock being reserved for blackboards to permit the splitting of large slabs to as little as 1/2 in. in thickness. Structural stock, on the other hand, is split into panels no thinner than 11/2 in. and

Jan 1, 1952

By William H. Case

The Bertha zinc-mines of the Bertha Zinc and Mineral Company are in that part of the State of Virginia popularly known as Southwest Virginia, and are included in that division of the " Great Valley " named in recent years the " New River-Cripple Creek Mineral-Region of Virginia." The mines lie wholly in the southeast part of Wythe county, two miles from Pulaski county on the east, and three miles from Carroll county on the south. The only other zinc-mines that are opened in the region are those of Manning & Squier, which adjoin the Bertha zinc-mines on the northeast, and the mines of the Wythe Lend and Zinc Company, at Austinville, eight miles to the southwest. A little farther southwest of Austinville, near Ivanhoe, a new company is preparing to mine zinc-ores. The mining properties named are included within a distance of ten miles along a direct line, bearing north 55' east, on the south side of New river, which here flows northeastward in a tortuous course. The greatest distance of any of the mine-openings from the river is a little less than two miles. It is not the magnitude of the operations in zinc-mining in this particular region, as compared with that of other well-known districts, that merits especial notice. Aside from the exceptional purity of the spelter made directly from the Bertha ores, some peculiar features of the zinc-ore deposih, as shown at the Bertha mines, and

Jan 1, 1894

By Arthur Notman

SacRamento Hill is a mass of granite porphyry intruded along a fault between Paleozoic sediments and pre-Cambrian schists in the Bisbee district, Cochise County, Arizona. The intrusion invaded both walls of the fault to a large extent, and the accompanying extensive mineralization of the Paleozoic limestones is the locus of the great high-grade copper orebodies of the Bisbee mines. Prior to 1909, although the presence of ore in the masa of the hill itself had long been suspected, systematic exploration had not been attempted. A few shallow pits on the surface here and there had disclosed some chal-cocite with associated carbonates as stringers in a mass of strongly brccciated, silicified, and iron-stained porphyry. The workings in the surrounding limestones had penetrated the porphyry in a dozen or SO places for some distance from the contact but had disclosed no ore. However, the rock had been largely altered to quartz and sericite and very thoroughly impregnated with pyrite and traces of the copper sulphides. A close geological study of the conditions, both surface and underground, strengthened these suspicions, and plans for prospecting the hill were made in 1909. Reference to the accompanying illustrations, Figs. 1 and 2, will give an idea of the general appearance of the surface. The mineralization is more intense in two main zones, one at the west end and the other at the east end of the hill. The outcrop of the hill is elliptical and approximately 3,000 ft. long by 1,400 ft. wide. The Sacramento shaft, which is the main hoisting shaft of the mine and situated at the east end of the hill, as shown by the accompanying map (Fig. 3), offered a convenient base for prospecting that end of the hill, while the workings from the Holbrook shaft offered a similar opportunity at the west end. The choice of level for the first work was determined by analogy with the most favorable depth for enrichment below surface obscrved in the other disseminated deposits of the Southwest.

Jan 1, 1916

By F. E. Turton

THE town of Rancagua, at the junction of the state and Braden Copper Co. railroads, is located 82 km south of Santiago, the capital of Chile. From Rancagua 70 km to the east, situated in the Andes at an elevation of 7000 ft, is the town of Sewell, the location of a millsite. Two and one half km east of Sewell by underground railroad at an elevation of 7500 to 10,000 ft is the Teniente orbody. The history of El Teniente dates from the eighteenth century, when according to legend a Spanish lieutenant, or teniente, a fugitive from justice, escaped to the Andes and camped at what afterwards became known as the Mineral El Teniente. This man reputedly discovered a vein of high grade copper ore. The first official records date from 1819, when some exploitation was begun, and continue until 1870. The orebody was large and of relatively low grade and was beyond the capabilities of the small groups attempting to operate it. No effort was made to erect a concentrator until William Braden became interested in the property. In 1901 Mr. Braden promoted the necessary capital for the first practical exploitation of the orebody. Barton Sewell and E. W. Nash, both of the

Jan 1, 1952

By Guy Bjorge

IN recent years the Homestake mine has been served by three shafts, the B. &. M., the B. & M. No; 2 and the Ellison, supplemented by an inside shaft, the Milliken, extending from the 2000-ft. level to the 3200-ft. level. In 1932, when a new operating shaft was under consideration, the bottom of the mine was at the 3200-ft. level. The B: & M. and the B. & M. No. 2 shafts were down to the 1550-ft. level and the Ellison was being deepened from the 2600 to the 3200-ft. level. All ore was hoisted at the B. & M. and-Ellison shafts. Levels below the 2600 level, which were under development only, were served by the Milliken shaft. Below the 1550-ft. level, the mine was served only by the Ellison shaft from the surface. Within a few years, when ore reserves above the 1550-ft. level would be depleted, operation would be limited to the Ellison shaft and the auxiliary underground shaft. The Ellison shaft alone was not adequate for hoisting the entire output from the deeper levels and also handling all men and supplies. Further-more, assurance of safety and uninterrupted operation required- two fully equipped shafts to the deeper levels. Therefore a new operating shaft was authorized in November, 1932, and was named for A. J. M. Ross, the mine superintendent.

Jan 1, 1935

By F. B. Plummer

PERHAPS the greatest progress made in the technical methods of oil production during the last year has been in handling gas from the new fields that yield light distillate fractions. At least sixteen recycling plants have been perfected and installed in South Texas fields alone during 1939 having a daily capacity of more than one billion -2 cubic feet of gas and capable of producing more than a million gallons of distillate daily. The process consists of running the gas under pres- sure through especially designed plants, separating out the condensate, then compressing the gas to 3000 lb. or more, depending on the subsurface pressure in the formation, and injecting gas back into the reservoir to be produced again later, along with more of the liquid hydrocarbons, and again recycled. Hence production in these distillate fields in sense has changed from that of producing a constantly diminishing and much-wasted product to a more or less continuous manufacturing process, in which practically all the gas is conserved-a real triumph in engineering practice.

Jan 1, 1940