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By J. F. McLaughlin

The installation of the Chance flotation system for the preparation of anthracite demonstrated the need for some means by which the specific gravity of the fluid mass in the separating cone could be determined with a fair degree of accuracy. The means used (with the original cones) was a set of wooden balls attached to ropes or chains, each with a different specific gravity. In the Northern Anthracite field, where the cones are generally operated with the specific gravity at 1.65, these gravity balls are provided with a range of 1.6, 1.65, 1.7 and 1.75. The operator of the cone from time to time would place one or more of these balls in the cone and estimate its specific gravity by the action of the balls. The margin of error with this system was very great, because there was the possibility that the test ball might be held in suspension and carried around by material floating beneath the surface of the cone in what is known as "island " formation—a formation resulting from a low sand line in the separating cone. The inaccuracy of this system of determination of specific gravity was indicated also by periodic tests of the final refuse bank, which showed excessive losses of coal to the refuse. Further evidence of the unsatisfactory operation of the cone was the uneven appearance of the market product, which caused spasmodic condemnation. Maintenance of Even Preparation With the critical market condition of today, it is essential that the preparation units operate consistently to maintain an even preparation. If a plant wishes to maintain a 1.65 specific gravity preparation, which means that all material of a specific gravity greater than 1.65 is classed as "reject," if the specific gravity of the fluid mass runs to 1.70, the market product will not be satisfactory; or if the specific gravity of the fluid mass is 1.60, the market product will be entirely satisfactory but the losses at the slate ends will be excessive. The desired specific gravity of the fluid mass in the Chance cone is maintained by agitating the sand within the separating cone by means

Jan 1, 1931

By N. H. Mannakee

[Secretary's Note.—At the Spokane meeting of the Institute, in discussion of President Brunton's address on "Modern Progress in Mining and Metallurgy in the Western United States," and at the request of members present, Mr. Ayres gave an oral account of his new separator, which is here published as an independent paper, partly because of its inherent importance and partly because it describes an improvement which did not originate in the Western United States, and therefore does not fall, strictly speaking, under the title of President Urunton's address.—R. W. R.] A brief history of the growth of the anthracite-coal preparation will give a better view-point from which to judge the present problem of separating slate from coal. At the beginning of the commercial value of anthracite, 70 years ago, only the pure portions, or " splits," of the veins were mined and shipped to market, and without any preparation or screening other than the selection, while loading, of the glassy lumps, and the rejection of the fine material and the slate that had strayed accidentally into the coal. The next step was the crushing or breaking of the coal (hence the name "breaker" as applied to the preparation-building), and the sizing of it by means of bars or revolving screens. This stage of its development marked the advent of the " breaker-boy " as a slate-picker, with his ever-increasing capriciousness. As the richer veins or a splits " became exhausted and the market demanded a still greater output, the less-pure " splits " and the thinner veins were utilized to produce the coal. Carrying as they do a far greater percentage of impurities, particularly when removing the pillars, it became necessary to build new and better equipped preparation-plants. Finally, we are now at the highest stage of complicatioii yet known to the art of coal-preparation. We are dealing with varying specific gravities, frictional difference, hardness, structure, and form in the pieces of coal and slate coming from many different veins, and

Jan 1, 1910

By David Reger

Introduction THE exhaustion of oil and gas in the United States is proceeding at a rapid pace. This is especially true in fields where the light oils that furnish the most fuel for internal-combustion engines are found, leading to a late estimate by Arnold that the petroleum resources are 36 per cent. exhausted.1 The further remark by Dr. David T. Day at the New York meeting of the Institute, in February, 1916, that the operators of the country are now scarcely able to supply the ever-increasing demand for gasoline, leads the writer to believe that a determined effort will soon be made to secure deeper producing horizons in those regions that now furnish high-grade oil, and in which pump stations and pipe lines offer convenient means for the economical handling of new production. A study of Arnold's table shows further that the States that produce paraffin oil exclusively are 45 per cent. exhausted, this figure being much higher than the average of paraffin and asphalt States altogether. The most extensive of all the paraffin-oil-producing areas is the Appalachian Field, extending with the Appalachian Geo-Syncline from western New York southwestward through western Pennsylvania, western West Virginia, Ohio, Kentucky, and Tennessee, where oil of high gravity is always secured. Of these States, Pennsylvania and West Virginia rank highest in. the quality of their product. In West Virginia, the oil sands dip to a lower level than in any other locality in the Appalachian Geo-Syncline, the Pittsburgh Coal of the Pennsylvanian being only 50 ft. above sea level along the Nineveh Syncline at Wileyville, Wetzel County.2 The very significant fact that the sands of the Mississippian and Upper Devonian Measures furnished the richest oil pools of the State along this

Jan 9, 1916

ONE of the interesting-though not unnatural-features of the whole Porphyry Copper development is the way in which the history of each property dovetails with that of one or more of the others. The same names keep cropping up as one digs into the accounts of events from 1900 to 1915. J. Parke Channing made examinations of Nevada, Ray, and Inspiration, as well as of Miami. Henry Krumb did important work in the early stages of the exploitation of Utah, Nevada, Ray, Braden, and Inspiration. Pope Yeatman was favorably impressed with the Braden property before he had ever visited it, when he saw a batch of Braden ore sent to the United States for testing. It reminded him immediately of Nevada Consolidated ore: indeed, he was unwilling at first to believe that it came from South America. And it was while working on the examination at Utah Copper that Philip Wiseman became enthused with the possibilities of the Ray mines. In 1905, Mr. Krumb, under the general direction of A. Chester Beatty, was strenuously engaged at Bingham, Utah, in arriving at an answer to this question: Should or should not the Guggenheim interests back Jackling to the tune of several million dollars to build the big concentrators at Gafield? Seeley W. Mudd was the Guggenheim engineer on the Pacific coast and in that capacity was indirectly responsible for the work at Bingham. At his suggestion Wiseman came up from Los Angeles as one of Krumb's assistants-at $5 per day, incidentally. A little later Mr. Mudd, in Utah on one of his visits, went down to Bingham. That evening in the old Hills Hotel in Markham Gulch he was going over reports with Mr. Krumb. Wiseman and some of the other engineers on Krumb's staff were present. During a breathing spell, Mudd expressed

Jan 1, 1933

By H. H. Otto

Fourteen years ago, J. B. Warriner presented before the Institute a paper on anthracite stripping,' describing the progress of stripping in the Anthracite Region from its beginning with an old quarry stripping at Summit Hill, which was operated more than a century ago, through various types of strippings used in 1917. This paper will briefly review the status of anthracite stripping methods from 1917 to the present day and describe in detail the stripping operation of The Hudson Coal CO. at Clinton collicry. It is not intended as a review of the work in the district—that would require a much longer paper. History and Statistics Mr. Warriner said, in part: "The sizes of shovels have gradually increased until now 70 and 80-ton shovels are the rule and larger equipment is coming in." He also referred to a large electrically operated dragline which weighed about 225 tons and had an operating radius of 125 ft. In 1917, a number of strippings used shovels of various sizes, ranging up to the 60 Marion and Bucyrus 70-C railroad type. Material was hauled away in narrow-gage side dumpers, usually of 4 or 5 cu. yd. capacities. Star-type churn drills with 4-in. bits were generally used for drilling purposes. Since 1917 wages in the anthracite field; have increased, but the higher operating costs of strippings are now being successfully combated by continuous improvements in stripping equipment, such as the development of the caterpillar or crawler shovels, the use of larger cars, better drilling equipment, the introduction of shovels of increased capacity and the adoption, in a number of instances, of standard-gage equipment for transportation. In 1918 another large dragline was introduced in the southern fields, which had been in service in the Middle West and is in use today, dipping vertically about 180 ft. to raise coal out of a stripping. About 1918, small shovels with caterpillar traction were introduced. In 1922 the 6-in. well drills began replacing the 4-in. drills, with an attendant saving to the operators. About the same time A. E. Dick placed the first pair

Jan 1, 1931

By S. J. Hardison

IN the fall of 1912, the appearance of water in the oil of the Nevada Petroleum Co., Coalinga, Cal., made necessary the installation of a dehydrating plant to reduce the water below the 3 per cent. limit prescribed by the agency. Unlike the mining industry, technical literature of the oil industry is limited and extremely unsatisfactory. Until the recent efforts of the Petroleum and Gas Committee of the American Institute of Mining Engineers, no concerted movement has been instituted to secure publication of papers dealing with the problems of the oil business, and because of the additional fact that dehydrating of oil has not been practiced in California to any extent up to comparatively recent time, it was found necessary to experiment in order to determine the most satisfactory plan for this purpose. It is easy to write regarding successful enterprises, but, while not so pleasant, it is equally desirable to write of failures so that others may be saved the loss incident to such investments. The water in this oil occurs both free and as an emulsion. Free water easily settles out, but the emulsion requires treatment. So far as can be determined, the emulsion consists of globules of water surrounded and enveloped by a film of, oil. Starting with this hypothesis, the theory has been evolved that in order to break up this emulsion the water must be heated at least to boiling point, when an explosion takes place destroying the globule. Unfortunately, this condition was not recognized in the installation of the first plant, which was planned on the following lines: A heater was arranged so that all the oil from the wells could be heated before reaching the shipping tanks, these being already fitted with coils for steam heating. Tanks were filled with about 4 ft. of water, live steam was turned into the coils and the water brought nearly to the boiling point, when the heated oil was run in. This for a time was partly successful, but the length of time necessary to apply this heat was too great for practical operating.

Jan 3, 1915

By Richard McCaffery

THERE has been some discussion among blast-furnace operators regarding the relative merits of tuyeres of the converging type and tuyeres designed in an attempt to produce a diverging jet of air. An article1 on Progress in German Blast Furnace Practice indicates the interest in this subject abroad. This article illustrates a diverging type of tuyere which is shown here in Fig. 1A. The design appears to be rather poor as regards known properties of air flow. The adoption of this. tuyere, the article states, resulted in a decrease of blast pressure of about 4 lb. per sq. in. and an increase of the output of the furnace. If the improvements mentioned are due entirely to the change in tuyere design, the subject of tuyere design seemed worthy of investigation and this paper is a report of such work. Because of the gain in efficiency brought about by the use of a diverging tuyere in the German plant, it was thought desirable to investigate the shape of the air streams issuing from various-types of tuyeres to ascertain whether the gain was due to the tuyere itself or to the distribution of the air in the furnace. Since it is difficult to do experimental work on a full-sized furnace, or even on full-sized tuyeres, and because of the equipment available for the work, it was decided to carry out the work on scale models of tuyeres. Therefore, in this work, single tuyeres were used and the blast discharged into the atmosphere. The shape of the air stream was determined by means of a Pitot tube suitably supported in the air stream at various positions. The air was. furnished by a Roots blower driven by a 10-hp. direct-current motor having field control, and capable of discharging 250 cu. ft. per min. with a pressure on the discharge of 4 in. of mercury. A modern blast furnace with 21-ft. hearth requires about 50,000 cu. ft. per min. of free air at from 16 to 23 lb. per sq. in. pressure, depending upon the type of burden. This pressure is the total pressure required to force the air through the burden and through the tuyeres. Upon computation the pressure drop through the tuyeres is about 1 to 2 lb. per sq. in. By making the tuyeres to a one-fifth scale, and by using a pressure of 1.5 in. of mercury, the shape of the air streams should be

Jan 1, 1931

By Donald E. Cummings

Certain dusts are dangerous when inhaled, but most hazardous of all dusts are quartz or other forms of pure crystalline silica. The inhalation of dusts containing silica in combination with other elements gives rise to changes in the lung which differ from those produced by pure silica. As a generality, silicates are less dangerous than pure crystalline silica and vary greatly among themselves in harmfulness. Repeated analyses have shown that approximately 70 per cent of the total number of particles involved in any dust-producing operation are below 1 to 2 microns in diameter. (A micron is approximately 1/25,000 in.) Similar studies of dusts recovered from silicotic lungs post mortem have demonstrated essentially the same size frequency distribution, except that the upper limit of size retained in the air sacs is approximately 10 microns. Experiments carried on by Gardner and the author have further demonstrated that greater changes were produced by a fixed weight of very fine quartz particles (2 to 3 microns and smaller) than by the same weight of slightly larger particles (8 to 10 microns). Any attempt to control a dust hazard should therefore be directed toward the removal from the working atmosphere of particles less than 10 microns in size, and particularly those below 3 microns. All industrial hygienists agree on this matter, but there remains considerable dispute as to the lower limit of particle size capable of causing injury. It has been generally conceded, however, that particles of harmful dusts smaller than 1/4 to 1/2, micron are not essentially concerned in the development of pulmonary fibrosis. This opinion is partly substantiated by the existence of large numbers of these small particles in practically all atmospheres that are not associated with any known hazard. Such particles are so small that they have very little mass, fail to settle even in quiet air, and probably are not easily retained in the lung. Until further research proves that they are involved in the dangers provoked by the inhalation of certain dusts, engineers will generally agree that it is better to disregard them.

Jan 1, 1937

By Edward Keller

In order to conduct intelligently the sampling of copper of various forms and grades, solid or in molten furnace-charges, a knowledge of this subject is essential. Yet figures and tests have been persistently withheld from publication, and the books at our command give us little or no information in this regard. This gap in our copper-literature has led to a series of experiments, the results of which are given below. Nearly all gold- and silver-bearing copper, destined for shipment, is cast into the form of bars or pigs. For the sampler, the question arises, Does copper, so cast, remain homogeneous ? If it does, sampling should present no difficulties; but, from experience, we know that this is not the case, and we are thus led to study the behavior of the precious metals arid impurities in copper when the latter changes from the molten to the solid state. For the purpose of the experiments here described, all available grades of copper were selected, and all of them mere subjected to the same conditions, an important requisite for proper comparison. The molten copper obtained directly from reverberatory- and blast-furnaces was cast in a cast-iron mould, the sides and bottom of which were 1 inch thick, and would, therefore, give the copper a fair length of time to chill. The inside dimensions of the mould, and, consequently, of the block of copper obtained, were: length, 10 inches; width, 9 inches; and depth, 5 inches. Fig. 1 illustrates the method employed to obtain samples from these blocks. A was cut away, and the face so produced BALTIMORE COPPER WORKS,

Jan 1, 1898

By Charles Pack

Die castings may be defined as metal castings made by forcing molten metal, under pressure, into a metallic mold or die. It is necessary to keep this definition in mind to avoid confusing this process with other permanent-mold casting processes. The fundamental principles of the process have been known and practised many years. The simplest application is embodied in the modern linotype machine in which molten metal (usually tin-lead alloy) is forced under pressure into a metallic mold. The pressure is derived from a piston and cylinder immersed in the molten metal. Progress in the art of die casting may conveniently be divided into three groups: Machine for imparting pressure to the metal, material for the die or mold, casting alloys. The problem of delivering molten metal under pressure into a die is comparatively simple, when dealing with low-fusing-point alloys, as the alloys of lead and tin, but it is much more complicated when dealing with metals of higher fusing points, such as the alloys of zinc, aluminum, and copper. Although the art of die casting is comparatively new and, to a large extent, unknown, the records of the patent office are replete with patents on the subject. Fig. 1 shows the Underwood machine patented in 1902; this is probably one of the first machines designed for the production of commercial die castings. The relation of this machine to the linotype casting machine is clearly apparent. A cylinder and piston are immersed in the molten metal, the application of power to this piston forcing the molten metal, under pressure, into the mold or die. The Doehler machine, Fig. 2, patented in 1907, is based on the same general principles. This machine is used to a large extent at the present time, throughout the United States, for the production of zinc, tin, and lead alloy die castings. In the machine shown in Fig. 3, patented by Doehler in 1910, compressed air is used for forcing the metal into the die. In Fig. 4 is shown another of this type of machine. Here compressed air is applied to the surface of the molten metal to force it into the die. In a machine patented by Chandler in 1914, shown in Fig. 5, the

Jan 1, 1919

By James S. Foster, Paul G. Shewmon, Frank V. Nolfi

Analytical expressions are developed for the kinetics of dissolution or growth of a spheroidized, solute rich stoichiometric Precipitate in the surrounding matrix. There are two limiting cases; in one the rate of dissolution or growth is limited by long range solute diffusion through the matrix and in the other by the transfer of atoms across the mrdrix-precipitate interface. In the former case local interfacial equilibrium exists while in the latter it does not and the rate of dissolution or growth is controlled by an interfacial reaction. The analysis treats both limiting cases as well as mixed control in a single development, where it is found that the thermodynamic state of the interface is completely specified by a single parameter, o, defined by where K is the reaction rate constant, D is the matrix solute diffusivity, and R, is the precipitate radiz~s. If K is infinite then a is zero and local equilibrium prevails; if K deviates infinitesimally from zero, then a is very near unity and an interfacial reaction is rate controlling. Mixed control occurs for intermediate values of o. From the analytical results, numbers are computed which are useful in determining the duration of interfacial reaction or mixed control; following this time period, long range diffusion is rate controlling. THE rate controlling process or processes governing phase boundary migration has been the subject of much speculation and investigation. Traditionally, the most widely assumed boundary condition in quantitative treatments of this problem has been that of local interfacial equilibrium. This condition implies that long range diffusion of one or more elements controls the migration and is referred to as "diffusion controlled migration" (herein DCM). In contrast to the above view is the proposition that local interfacial equilibrium does not exist and that transport of atoms across the interface is rate controlling. Consequently, this mechanism is referred to as "interface controlled migration" (herein ICM). In processes of the latter type, the activity, a, of one or more elements is discontinuous at the phase bound- ary and the flux, J, across the boundary of a given element is some function of its activity discontinuity, a. One of the most comprehensive contributions to a mathematical description of phase boundary migration is due to am.' His analyses are applicable to precipitation in binary systems in which a) the precipitates are spherical* solute rich compounds of constant

Jan 1, 1970

By A. H. Kline, E. B. Zimmer, Zay Jeffries

It is well known that many properties of a given metal vary with the size of grain or cell. For most industrial purposes, where high ultimate strength and high elastic limit are desired, the manufacturer tries to produce a fine-grained structure. For some purposes, for example in transformer iron, a coarse-grained structure is desirable. The terms fine-grained and coarse-grained are used only in a relative sense. For example, a fine-grained cast copper might have grains 100 times larger than coarse-grained high-speed steel. Similarly, a finegrained steel rail might have grains or cells 100 times larger than coarsegrained wire, which was made of steel from the same heat. In some metals, the change in grain size is more appreciable khan the changes in any of the properties determined by the tensile test. Furthermore, in metals which have been annealed or subjected to high temperatures, grain size may prove to be a better indication of the life of the metal in use than the tensile test. One of the authors was able to supplant the tensile test with grain-size determinations in the control of metal for a specific use, the latter determination being more indicative of the life of the metal part than the tensile test. A grain-size determination with a tensile stest is, of course, more valuable than either singly. At present there is a decided dearth of data on actual grain-size determinations in metals, and consequently a corresponding lack of ability to interpret the determinations after they are made. It would be desirable to have correlation made between grain size and properties of metals, comparing thousands of determinations. In this way, grain-size determinations might be interpreted in much the same manner that tensile tests are now. Several Methods of Measuring Grain Size One of the chief reasons why more grain-size measurements have not been made is the excessive work involved in making a determination.

Jan 1, 1916

By Harry W. Brown

This paper presents results of static capillary pressure measprements made both by the restored-state and by the mercury-injection methods, and of dynamic capillary pressure measurements made by the Hassler technique. The results indicate a variation from sample to sample of the factor used to convert capillary pressure obtained by mercury injection to those applicable to a water-gas system. Static capillary pressures measured by the restored-state method were found to be in close agreement with dynamic capillary pressures, indicating that data obtained from static measurements may be utilized with confidenre in the solution of dynamic problems of fluid flow. Experimental data are presented on the use of "J-curves" for the correlation of capillary pressure-saturation relations for a number of cores from a particular geologic formation. It is shown that the correlation was improved in some cases by reqtricting it to rertain lithologic classifications. INTRODUCTION The Meaning and Importance of Capillary Pressure The coexistence of two or more immiscible fluids within the voids of a porous medium, such as reservoir rock, gives rise to capillary forces. Because of the interfacial tension existing at the boundary between two immiscible fluids in a pore space, the interface is curved, and there is a pressure difference across the interface. This pressure difference is termed the capillary pressure. The magnitude of the capillary pressure between two immiscible fluids which together fill a porous medium depends upon several factors: the textural properties of the medium, the wettability of the medium, the interfacial tension between the fluids, the respective saturations of the fluids, and the manner in which these saturations are attained. For a given pair of immiscible fluids in a particular sample of reservoir rock, the capillary pressure is a unique function of the fluid saturation, provided that the saturation of the fluid which wets the rock has previously been decreased unidirectionally from an initially complete saturation. Similarly, for a given pair of fluids at a particular saturation which has been achieved by a

Jan 1, 1951

By S. S. Rumsey

Introduction IN order to realize the economies which would result in operating the mine pumps at the Chapin mine with electricity. generated by water power, the Oliver Iron Mining Co. recently installed a hydro-electric plant at the upper Quinnesee, falls on the Menominee river, and constructed a transmission line to the mine. On account of the large flow of water in the Chapin mine, and the great depth from which it has to be handled, it was necessary to make a thorough investigation of the merits of both the motor-driven centrifugal and motor-driven plunger pumps, in order that the most economical results as regards power, cost of installation, maintenance, and operation, be obtained. This investigation resulted in the selection of motor-driven centrifugal pumps, as the cost of the pumps and motors, installation, and maintenance was considerably lower, and the space required in the underground workings very much smaller, than for the plunger pumps. The overall efficiencies guaranteed by several of the centrifugal-pump manufacturers were so nearly those guaranteed by plunger-pump manufacturers that the saving in power which would result from the installation of plunger pumps did not justify their increased first cost and the increased cost of installation under the conditions prevailing at the Chapin nine. In order to substantiate the representations of the centrifugal-pump builders it was deemed necessary to insist upon a factory test of the completed units before shipment, the test to be conducted under conditions of head and capacity approximating as nearly as possible those under which the pumps would operate in service. Furthermore, it was considered necessary to demonstrate, by test, that there were no deficiencies in design or capacity of the pumps or motors before their final installation in the mine, as it would be extremely difficult to make alterations in the pumps or motors after they were installed.

Jan 10, 1914

By Cleland N. Conwell

Intermittent coal mining has been conducted in Alaska for over a century. The first report of coal in Alaska was by the Veechy expedition of 1826 and 1827. Whaling ships used coal from Corwin Bluff near Cape Beaufort during the early and latter parts of the 19th century. The first coal mine in Alaska-and probably the oldest in the Pacific coast of North America-was opened by the Russians at Port Graham on the southwest tip of the Kenai Peninsula. The mine was operated from 1855 until 1867 when the United States took possession of Alaska. From 1880 to 1915 coal mining was carried on intermittently at several localities including Unga Island, Herendeen Bay, Chignik Bay, Kachemak Bay, and several points along the Yukon River. Small coal mines along the Yukon River provided fuel for the river steamers traveling from St. Michaels to Whitehorse. The area of sustained coal production began with the completion of the Alaskan Railroad from Anchorage to the Matanuska coal field in 1916 and to the Nenana field in 1918. Coal production continued to increase until 1953, at which time the oil and gas fields in Cook Inlet were developed. Oil and gas eventually replaced the Matanuska coal as a means of providing heat and power. The military in the Anchorage area shifted from coal to oil in 1967 causing a decrease in production. The increased demand for electric power in the Fairbanks area offset the decline and coal production started to increase. In 1969-1970 the Golden Valley Electric Assn. (GVEA) constructed a mine-mouth power plant 70 miles southwest of Fairbanks near Healy and a power line to Fairbanks. Location of Coals in Alaska Coals are found in all geographical subdivisions of Alaska. Coal in Alaska ranges in age from Carboniferous to Tertiary, and grade from lignite through anthracite. Fig. 1 shows the locations of Alaskan coal fields. The only fields that have been developed extensively are the Matanuska and Nenana. Fig. 2 illustrates the annual production of coal in Alaska and the ratio of underground to strip mining.

Jan 1, 1973

By Frederic Benard

PRIoR to 1936, the Ontario Refining Co. received all incoming blister copper from The International Nickel Company's smelter in the usual form of 460-lb. cakes, or slabs. These were received in open cars in lots of 250 pieces. At the refinery the cakes were unloaded onto narrow- • gauge cars, for convenient handling and pickup by charging cranes. In March 1936, an investigation was started on the feasibility of transferring converter copper in the molten state from the smelter to the refinery anode furnaces, a distance of 1 1/4 miles. Although common in steel practice, the transportation of blister copper in this form had not previoisly been attempted and a preliminary study of the various factors involved was therefore necessary. For this purpose, a series of experiments was made in which molten copper, removed at different stages during the furnace cycle, was held in a small portable holding furnace. This vessel was essentially a refractory-lined steel cylinder provided with ports for filling and emptying, heated by an oil burner. Dimensions of the steel shell were 3 ft. 10 in. diameter by 7 ft. 3 in. long, and thickness of refractories depended on the design and combination used for the particular experiment. The vessel was filled with approximately 2 1/2 tons of copper at a predetermined temperature, the lining having been preheated to a definite point by means of an oil burner. All openings were well luted and temperature readings were made every 15 min. by means of an immersion-type thermocouple. In this manner the rate of heat loss resulting with different refractory design could be developed. The ratio of exterior surface to weight of metal was taken into consideration in comparing the experimental results with what might be expected for a full-sized unit, and observations were also made on type and degree of slag accretion. It was not possible to get as much information on this latter point as desirable, but all evidence showed that the problem would be control of slag build-up rather than attack on refractories. From the information derived, and in consultation with the engineering staff of the supplying company, an order was placed for the first

Jan 1, 1944

By H. J. Rahilly

Mine fires, in the Butte district, have been a source of trouble and expense for the past thirty years, for while the actual fire area in most of the mines has been comparatively small, the handling of the gas has generally been a much larger problem than the actual putting out of the fire. The chief causes of these fires have been: Defective electrical equipment, incendiarism or carelessness, and spontaneous combustion. While fires from the first two causes have been numerous, they have generally been discovered and extinguished before they became very serious. Spontaneous combustion has been the cause of most of the large fires. Sufficient heat has been generated by ground movement, high rock temperatures, and decomposition to set fire to inflammable materials (tarred ropes, canvas, dry timber, manure, and hay) left in the gobs some years ago. Also, the heat resulting from the oxidation of the fine broken sulfide ore is sometimes sufficient to set fire to the timbers. In 1906, a fire was discovered on the 1100-ft. level of the Minnie Healy mine about 400 ft. (122 m.) west of the Minnie Healy shaft, which presumably originated, in one of the upper levels, from spontaneous combustion. The fire was walled off by concrete bulkheads and a pillar of ground 100 ft. (30 m.) thick was left below and mining resumed underneath. During subsequent operations, this pillar became cracked and broken and the fire passed into the new workings. It travelled downward and spread rapidly until, in 1917, it had become necessary to wall off portions of the Tramway, Leonard, and West Colusa mines on the 1300, 1400, 1500, 1600, 1800, 1900, and 2000-ft. levels, and the tops of the 2200-ft. raises were sealed. The fire area now consists roughly of a block 1000 by 1200 ft. and from the 600-ft. to the 2200-ft. level. This block is composed largely of a series of parallel northwest-southeast nearly vertical veins and stringers that form a large orebody, with no well-defined walls but badly fractured and faulted, and known locally as "horsetail ore." These veins, or orebodies, are mined by the square-set or the rill method.

Jan 1, 1923

By Donald C. Branford

The place of geophysics in the curriculum of a college or an engineering school has been much discussed. There is uncertainty as to whether the graduate may be called a "geological geophysicist" or a "geophysical geologist." Some have used the term "geophysical engineer." The question of the specialized undergraduate curriculum versus a general undergraduate training in the parent sciences with the addition of cultural electives has been and is still being argued. The literature during the past Io years shows that there are two rather definite schools of thought on this subject. The prospective employer, engaged in active geophysical exploration, visualizes the graduate in terms of the routine work he will do in the field. The school administrator, grounded in many years of tradition in the cultural type of education, visualizes the graduate in terms of his position in a political and cultural society as well as in his work as a professional geophysicist. Both are right and both wrong. A third school of thought, that of the student himself, has been almost totally ignored. It would seem best, for the purpose of this paper, to treat these various points and allied material separately, and then to integrate the conclusions as a set of conditions that should be fulfilled as far as possible in the ideal curriculum. At the end the relationship of our findings to the curriculum that has been tentatively adopted at the University of Pittsburgh will be considered. Divisions in Definition or Field of Geophysics Geophysics is that field of science in which geological and earth problems are solved with the use of physical and chemical techniques. Altogether too much attention has been paid to the artificial separation in this science of the commercial applications from the so-called "pure" or academic geophysics. The only real difference in the two types of work is one of scale of operations. In the commercial field the scale is large in terms of fmancial backing, immediate economic importance, and speed of operations. In the academic field the scale is large in terms of breadth of interests, in eventual economic importance, and length of time allowable for theoretical investigations. In exploration work the scale is small in terms of dimensions, since the methods are applicable only to the outer 20,000 ft. or SO of the earth's crust. In academic work, the student is not pressed for results of immediate economic importance, and therefore can afford to spend more time on problems that seem unimportant at the time. In the sense of background knowledge of the parent sciences necessary to the geophysicist, no essential difference can be found. Physics, chemistry, geology, and mathematics together, and in their proper proportionate amounts, form the matrix from which geophysical ideas and methods may spring.

Jan 1, 1946

By John W. Bell

IN the introduction to an excellent pamphlet, John Gross1 makes the following statements: Although marked progress has been made along mechanical lines, the theory and conception of underlying principles have not advanced so rapidly. This lack of advance in theory may be attributed to the unfortunate situation resulting from the controversy as to whether the Rittinger or Kick law is applicable to crushing. Many pages of the technical literature are burdened with theoretical discussions in favor of one or the other of these laws, which have tended to cloud rather than to clear the atmosphere. Realizing the situation, the Milling Committee of the American Institute of Mining and Metallurgical Engineers, through its secretary, E. A. Hersam, studied the milling industry and, because of the apparently hopeless lack of agreement in the Kick versus Rittinger dispute (5, 13, 14, 16, 18, 19, 21, 22, 23, 24, 30. 34), concluded that a foundation for the theory of crushing was of primary importance. The numbers in the paragraph quoted refer to papers listed by Mr. Gross in his Bibliography on Crushing and Grinding.1 One is obliged to infer that these papers show "hopeless lack of agreement in the Kick versus Rittinger dispute." No. 22 is a paper by Arthur O. Gates2 and No. 23 is one by the writer.3 Employing the subject matter contained in papers Nos. 22 and 2323 the writer will show that their authors formulated identical conclusions, based solely on experimental evidence. Exactly the same defect in the Stadler hypothesis was described in each of the two papers, although entirely different experimental methods were employed in uncovering it. Both of the authors, independently, formulated the conclusion that Rittinger's was a. true hypothesis and Stadler's false. The evidence in a third paper4 will be employed to prove that Stadler's hypothesis, when examined experimentally, is worthless. STADLER'S HYPOTHESIS WORTHLESS It is the writer's opinion that the evidence available in 1915 and the corroborating evidence in 1916 was amply sufficient to eliminate the need for any further consideration of Stadler's hypothesis. Certainly the evidence produced in 1918 dealt this hypothesis its coup de grace. It follows, therefore, that in the year 1918 there was only one credible hypothesis left-the Rittinger hypothesis. Gates' Opinion What information is available in Mr. Gates' paper?2 His conclusions are not consecutively summarized, none the less they are perfectly clear. On page 876, Gates says, and it is a very significant utterance: But my results show, among other things, that many more of Stadler's energy units are obtained per foot-pound-applied in coarse than in fine crushing. This suggests a doubt as to the correctness of his unit, since one would expect a foot-pound-applied to produce the same number of "energy units," whether the feed and resulting product were coarse or fine. When Rittinger's theory is applied, the number of mesh-tons (my unit of surface produced) is

Jan 1, 1942

By I. L. Greninger

THE property of the Andes Copper Mining Co. is in the central part of the Republic of Chile, Province of Atacama, 147 km. (91.5 miles) eastward by rail from the port of Chanaral. At this point the Andean system is made up of two principal ranges. The mine is on the westerly slope of the outer, or western, range, at a mean elevation of 3250 m. (10,500 ft.). A meter-gage railroad extends from the company's port at Barquito, near Chañaral, to the reduction plant at Potrerillos. This road has a total length of 151 km. (94.0 miles), of which 93 km. (57.79 miles) are owned by the Andes Copper Mining Co. The remainder of the road is owned by the state. The company operates its trains over the portion of the road owned by the state, under an agreement with the Chilean Government. The mine is connected with the reduction plant by a meter-gage railroad, 10 km. (6.214 miles) long, which extends from the coarse-ore bins at the metallurgical plant to the underground storage bins at the mine. Tunnels make up 60 per cent. of the total length of this railroad. As is characteristic of the Andean region in Chile, the topography of the locality is extremely rough, and a number of points near the mine have elevations of from 15,000 to 16,000 feet. History of Potrerillos Mine The ore occurrence has been known for many years, and many of the claims under individual ownership were worked in a small way before the property was acquired by the present owner—the Andes Copper Mining Company. The highest grade oxide ores near the surface were mined by primitive methods and, after sorting, the product, assaying around 15 per cent, copper, was hauled by mule carts a distance of about 80 km. (49.71 miles)

Jan 1, 1929