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By H. G. Dwyer

CONSIDERATION for future expansion influenced the design of the new smelter at Chuquicamata. The section of the smelter now going into operation, while large, represents only little more than half of the ultimate plant. Equipment for the direct smelting of raw concentrates has been arranged in the conventional manner although needless details have been eliminated wherever possible while mechanical and electrical features not usually found in copper smelters have been incorporated. Access is afforded to all equipment but platforms and walkways have been held to a minimum. The poling and other platforms in the converter aisle, normally subjected to battering by ladles and other heavy equipment in transit, are of concrete construction with steel grid armor embedded in the exposed surfaces. Noticeably absent are the many platforms formerly required for the handling of converter flux since this material is charged to the converters through the side walls of the converter uptake hoods. Clear space with ample headroom is provided under each converter and converter puncher's platform to permit clean-up by small bulldozers. Remote controls have been introduced at many points. Electrical control rooms and control centers have been conveniently located. An electrical tunnel with cross tunnels and escapeways is situated under the entire length of the converter aisle for power feed lines. Skull breaking facilities are in a separate building near the smelter. A silica-slurry building with brick storage area is adjacent to the smelter building so that elevator service is available to both buildings. A smelter office, a change house and lunchroom are provided in convenient locations. Units of the smelter as presently constructed are shown in the plan and table on page 1194. The ultimate smelter will house 7 reverberatory furnaces, 14 waste-heat boilers, 9 Peirce Smith converters, 4 casting furnaces serving 2 casting wheels, and one holding furnace serving a straightline casting machine. Additional crane requirements in the converter aisle will necessitate the installation of special hoisting equipment in space already provided in a raised section of the roof over the casting end of the aisle. Thus a complete crane may be lifted high enough above the crane rails to permit the passing of another crane underneath thereby giving all cranes full use of repair facilities already installed. For the collection of gases, 2 dust chambers, each with separate stacks, will collect converter gases and 2 balloon flues with separate stacks will collect reverberatory furnace gases. Extensions necessary to complete the ultimate plant will be possible with little interruption to normal operations since the future converters and reverberatory furnaces will be added at the east end while additional casting facilities will be added at the west end of the present plant. The future units may be added singly or all at one time depending on requirements.

Jan 1, 1952

Truman H. Aldrich John H. Allen W. Lawrence Austin W. S. Ayres John Henry Banks Wm de L. Benedict Robert Bentley Andrew A Blair Walter M. Brodie

Jan 1, 1934

By L. H. de Wald, J. Walter Scott

The removal of lead by fire refining methods from copper of electrolytic quality is growing in importance. Particularly is this true of the refining of secondary copper and copper cathodes obtained from electrolytic cells where lead is used as the anode material, as in the electrolysis of ore-leaching solutions and foul electrolyte. When investigating the removal of small amounts of lead from molten copper in a reverberatory furnace during the refining of secondary copper of otherwise electrolytic copper quality,' it was found necessary first to know something about the distribution of the lead throughout the bath. The authors believe that it is not generally known how small amounts of lead as an impurity are distributed in a molten copper bath; that is, whether the lead is concentrated at the bottom or at the top, or whether it occurs more or less uniformly throughout the bath. The data contained in this report may help, therefore, in the development of a satisfactory fire refining method for the removal of this obstinate element from a molten copper bath. The first step was the development of a method for taking samples from various locations in the bath, at the surface and at various depths. The sampling device used, shown in Fig. 1, consisted of a 10-gram fireclay assay crucible fitted into a ring on the end of a steel rod handle of suitable length. A fireclay scorifier, 2 in. dia., with a ½-in. hole bored in the

Jan 1, 1930

By F. L. Kadey

The petrographic microscope is a well known tool in nonmetallic materials research. Its utility in the study of perlite is, therefore, not surprising. Valuable information has been derived on samples smaller than those needed by most other methods of analysis. Relatively little has been written about the petrographic analysis of perlite for two chief reasons: perlite is a material of relatively new commercial interest; and the independent prospector or expander, responsible for most studies, attacks the problem from an immediately practical viewpoint. Hence, petrographic studies of perlite have been limited to academic aspects. Rapid development of commercial interest in perlite has changed the situation. Mining and expanding companies now have effective research tools; and such instruments as the petrographic microscope have added materially to practical knowledge about this mineral. Petrographic techniques employed in the search for suitable deposits screen worthless samples rapidly and inexpensively compared with large-scale expanding tests. Hence, only samples of potential quality are selected for further testing. Field appearance of perlite reveals little about its true identity and practically nothing about its commercial significance. Consequently, any method that can be used to separate the "wheat from the chaff" is valuable. In addition to the conventional thin-section and immersion methods used in evaluation of crude perlite, petrographic microscope techniques have contributed to the study of expanded perlite. When apparently similar materials behave as differently in use as perlites, the internal structure of the individual particles becomes a subject of great interest. The purpose of this paper is to discuss some of the petrographic techniques used to characterize this material. WHAT IS PERLITE AND WHERE IS IT FOUND? Perlite was at first described as a glassy rhyolite possessing a pearl-like luster and exhibiting numerous concentric cracks resembling an onion skin appearance.' Fig. 1 is an illustration of perlite in this classical sense. Recently, its definition has taken on a commercial connotation: an acid volcanic containing 2 to 5% of combined water that, when heated rapidly to the softening point of the glass, coincident with the volatilization of the combined water, expands or "pops". Fig. 2, although hardly fitting the classical description given by Johannsen, is an excellent perlite in the commercial sense. It may occur as a breccia in rhyolite, (Fig. 3) and sometimes it is associated with obsidian nodules, as illustrated in Fig. 4. The term perlite includes the expanded product also. In the United States, perlitic rocks are restricted to the western part of the country. 2 Elsewhere in the northern hemisphere, European occurrences follow a belt beginning in Iceland and extending in a southeasterly direction through Scotland, County Antrim in Northern Ireland, and through the Massif Central of France, Germany, and Hungary; thence, into Italy and Sardinia, and through some of the Aegean Islands. A general oriental belt extends from Japan through Australia and New Zealand. Rhyolitic areas elsewhere in the world undoubtedly contain perlite.

Jan 1, 1963

By P. Bardzil, J. Gurland

An experimental study of the variation of transverse-rupture strength with composition Anexperimentaland grain studysize has shown that the strength reaches a maximum for values of the mean free path between carbide particles of 0.3 to 0.6 microns. The fracture oforiginates in and proceeds through the carbide grains mainly. Impact strength and hardness also were recorded. MOST of the physical properties of sintered carbides vary linearly with composition. The transverse-rupture strength, however, shows a unique behavior. As the amount of binder metal is varied from 6 to 25 pct by weight, the strength at first increases; then, between 15 and 20 pct Co, it reaches a maximum of nearly 400,000 psi, and finally decreases with further additions of binder metal. This behavior of the transverse-rupture strength has been reported, among others, by Englel and Sandford and Trent.' The significance of these observations has not been discussed in the literature. That it may be of more than specific importance is indicated by the very similar variation of strength with composition encountered in other systems sintered in the presence of a liquid phase, such as Tic-Ni3 and Fe-Cu.' Experimental Details All compacts were prepared and sintered according to normal industrial practice. The average diameter of WC grains and the mean free path between grains were measured on metallographically prepared samples by a method of linear and planar analysis sing the relations where d is the average diameter of dispersed grains; Nl is the number of noncontiguous grains intersected on a metallographic plane by a line of unit length; N, is the number of noncontiguous grains per unit area; P is the mean free path between grains of dispersed phase; and f is the volume fraction of dispersed phase. Approximately 1000 grains were counted on each sample. For each composition d2 was plotted against P, the resulting straight lines serving as a check on the measurements. The distribution curves of the WC powders of different average diameters were homologous, and no attempt was made to influence the grain size by blending powders, adjusting the sintering conditions, or otherwise altering the particle size distribution. Densities were measured by differential weighings in air and water. The degree of densification did not vary consistently with either composition or grain size. The density is influenced by slight amounts of impurities, specifically by 0.1 to 0.2 pct Fe which enters the powders during ball milling. As an illus- -tration of the experimental variations, a number of measured densities are listed in Table I. They are expressed in percentage of theoretical density, as calculated from the published X-ray densities of WC and Co.' For the purpose of determining the transverse-rupture strength, rectangular test specimens (3/16x 3/8x3/4 in.) were broken by loading at the center of a 9/16 in. span. The average strength of five or more samples is reported for each composition and grain size. The compacts were ground on two parallel surfaces only. The load was applied at an average rate of 6,800 psi per sec. Since the elongation after fracture of the alloys is very small, it was assumed that the compacts deform elastically to failure and the fracture stress was calculated by the conventional beam formula. The data for one alloy are presented in Table II as an example of the results and scattering ranges encountered. The precision of the test, as measured by the average difference between duplicates, is of the order of 15,000 psi. The results are strictly comparable only to compacts prepared and tested in the manner described and of similar chemical composition and grain size characteristics. Unnotched Charpy specimens were used for impact testing. Each point represents the average of 3 to 20 samples, the larger number being used in an attempt to determine the influence of grain size. Considerable scatter was encountered, the range from lowest to highest value often amounting to 30 pct of the impact strength. The hardness is reported as a Ra reading, using a 60 kg load with diamond brale indenter. Compositions are given in weight percent.

Jan 1, 1956

By Joseph A. Sutton, John D. Corrick

The continuing depletion of high-grade ore de- posits in this country has created a need to develop more effective methods for recovering valuable metals from low-grade ores. The use of microorganisms for their biochemical reactions is one possible way to solve certain phases of this problem. Bacteria have been used successfully by many diversified industries. Until recently, however, little attention has been focused on the use of bacteria in mining and metallurgical processes. This report summarizes one phase of the microbial studies being conducted at the U.S. Bureau of Mines, College Park Metallurgy Research Center. The two-fold objective of this study was, first, to determine if pure strains of the bacteria Ferrobacillus ferrooxidans, Thiobacillus concretivorus and Thiobacillus ferrooxidans could utilize the iron and sulfur occurring in sulfide minerals to produce appreciable quantities of ferric sulfate and sulfuric acid for dissolving copper and, second, to develop the chemistry involved in the microbial oxidation of sulfide minerals. These are of particular interest as they relate directly to the feasibility of employing microorganisms in leaching operations.

Jan 6, 1963

By Charles H. Doolittle, Royal P. Jarvis

The following notes are contributed, not with the idea of offering a complete history of the development of this very important process as applied to the Leadville district, but with the hope that the practice, as it was carried out in the Bi-Metallic Smelter, may still be of interest to others. The Bi-Metallic Smelter, at Leadville, Colo., was the first to make a commercial success of the process described in the following pages, which consisted essentially of smelting raw silver- and copper-bearing sulphide ores in blast-furnaces to form a matte, which was generally re-smelted with siliceous ores for a second high-grade matte. Robert Sticht, in his excellent monograph on the subject,' has outlined briefly the plant belonging to this company. He says, in effect: The Boulder, Mont., enterprise set the example for others with whom the patentee (W. L. Austin) had been negotiating, and the old La Plata lead-smelter at Leadville, Colo., was fashioned into a crude, temporary experimental plant for the Bi-Metallic Company, and the feasibility of the process demonstrated by a trial run. Subsequently, the plant was considerably improved, and, being supplied with very favorable ores, continued to run without interruption from August, 1892, to June, 1893. In September, 1803, smelting was resumed in a completely remodeled establishment, some $200,000 having been expended upon the new work. Thus, aside from the Montana experiments, which, under the direction of W. L. Austin, began as early as 1885 at Toston, and continued with indifferent success for six or seven years, the Bi-Metallic plant at Leadville may be said to have

Jan 1, 1911

By Eric Jette

The nickel-chromium alloys form the base f for many industrial heating alloys, so that this system is of considerable practical importance. The literature on these alloys, however, contains much conflicting evi-dence, not only regarding the solubility limits of the terminal solid solu-tions at different temperatures but even regarding, the number of phases existing in the system. Part of the difficulty undoubtedly arises from the fact that chromium of sufficiently high purity has been available only during the last few years. A second and possibly even more impor-tant difficulty arises from the ease with which chromium reacts with silicon and carbon compounds and the atmosphere to form silicides, carbides, nitrides and oxides. Pilling and kihlgren1 in 1929 reviewed the existing scanty information. They did not consider the diagram by which they summarized the various bits of information to be entirely satisfactory. Matsunaga2 investigated the system by thermal analysis, microscopic examination and other means. Electrolytic nickel of 99.9 per cent purity and thermit chromium containing 0.30 per cent al, 0.40 per cent fe, 0.25 per cent si and 0.04 per cent c were used. A tammann furnace with hydrogen flowing through it was used for melting. Sekito and matsunaga3 reported an x-ray, investigation presumably on these same

Jan 1, 1934

By C. M. Young

This is an attempt to bring together some of the knowledge of the coal-forming conditions obtaining during the Pennsylvanian period in the Western Interior Coal Province, to sketch briefly the present extent and character of the coals produced, and to show the beginning of correlation of beds over wide areas. Coals of later age than the Pennsylvanian are not included and, for reasons that will appear later, little attempt is made to cover the districts of eastern Oklahoma and western Arkansas, and northern Texas is omitted entirely. Geological Conditions The conditions that made possible the formation of large coal beds in the United States were most favorable in the Pennsylvanian period, formerly known as the Upper Carboniferous. The geological conditions can be determined from the records, but not the reasons for them. Preceding the Pennsylvanian was the Mississippian, formerly known as the lower Carboniferous and still later as the sub-Carboniferous. During this time, large bodies of limestone were deposited and at the close of the Mississippian they were warped upward and for a long period were eroded. Then there began a period of subsidence, more or less interrupted at times, during which the Pennsylvanian strata were deposited (see table on next page). These consist of sandstones, limestones and shales, with occasional coal beds. At the beginning of the Pennsylvanian there was subsidence, which allowed the sea to extend over the eroded Mississippian surface to the north and east. It passed over eastern and central Texas, a large part of Oklahoma, with a projection into Arkansas, over Kansas and part of Nebraska, all the northern and western part of Missouri, a large part of Iowa except a strip in the north, most of Illinois, practically all of Kentucky, northern Tennessee, most of Indiana, all of the eastern part of Ohio, practically all of West Virginia, western Maryland, the west end of Virginia, and a large part of Pennsylvania. The Ozarks were already up and separated the main body of this sea, which covered the northern

Jan 1, 1936

By D. B. Kibler

The Florida hard-rock field extends from Suwanee and Columbia Counties in northwest Florida to south of Croom, Florida, in Hernando County. This area is approximately 100 miles long and varies from 2 to 30 miles in width. Florida hard-rock phosphate was discovered at Hale Springs, about one mile northwest of the town of Dunnellon, Marion County, in the year 1888, by Alber-tus Voight, who lived within a few hundred yards of the spring. Soon after this discovery there was a great influx of land speculators and wildcat operators into the vicinity of Dunnellon. A few years later phosphate was found in Alachua, Levy, Citrus and Hernando Counties, and a great number of small crude plants were built from Newberry, Alachua County, all the way to Croom, Hernando County. Phosphate was found in many places as "outcrops." Looking for outcrops was the first method used by the early prospectors in searching for deposits. Several years passed before any systematic method of prospecting was used. Early Mining The early method of mining hard-rock phosphate was with picks and shovels, wheelbarrows, mules and wagons. The pick-and-shovel method was used altogether until about 1900, when some of the larger operators began to use railroad-type steam shovels of from ½ to I-yd. capacity Picks and shovels, with negro labor loading the phosphate ore (called "matrix") into wooden side-dump skip-cars, was continued until about 1907. From the discovery of hard-rock phosphate until approximately 1905, overburden was removed with mules and scrapers. At some places, the overburden was loaded into side-dump skip-cars by negro laborers and the cars were then pulled up an incline with a steam hoisting engine. The maximum depth of overburden that could be economically removed by hand labor, mules and scrapers was approximately 15 feet. About 1905 one of the companies erected a producer-gas power- I house near Dunnellon and generated its own electricity and about the same time the larger companies began using hydraulic methods for the removal of overburden. From the year 1888 through the year 1939 approximately 75 companies were operating at one time or another in the hard-rock territory, on approximately 500 locations. The hard-rock deposits are more irregular and have higher mining costs than the pebble-rock deposits, therefore 1 they have not been competitive in the domestic market, but because of the higher grade of the hard rock it is in demand for I the export trade. Discovery of phosphate outside of the United States has made it increasingly difficult in recent years for the hard-rock companies to survive, so today i only three companies are operating. These

Jan 1, 1942

By Rich Akerman

as 1866, I wish to assure both these gentlemen that I had not seen the section when I designed mine, :md even if I had, I should not have then dared to put it forth as a standard for English rail makers; tind the soundness of my opinion has a simple but conclusive proof in the statement made concerning the extra price which was at that time required for their production. I need not say that I highly approve of the original Welch section, for it is in principle similar to the design which I made ufteen yeam subsequently; and from a professional point of view I would fully concecle to Mr. Welch the credit of first designing the modern steel-rail section. From com- mercial, matter of tict, and practical point of view, he was, however, before his time by many years; or to put it in another way, rail making was ten years behind its time when the new metal steel was first introduced for rails. If his design has been incorrectly identified with my name (as the sections are similar in principle, though they differ in thickness of flange and web), I beg to say (1st) that I much regret the misun- Understanding which has arisen through no fault of niine, aud (2d) that no confusion has been intentionally created by me. It is a pity that the section designed so long ago has not been made public at an earlier date, as it would make a really goal stand- and pattern, anil even now be quite equal to present requirements. In conclusion, I wish to say that I have never desired to claim any invention or novelty by the publication of my sections; they Were put forth as being such as English railmakers could and would produce at the ordinary market. price, and as being the best and most suitable to be so obtained. Had Mr. Welch or Mr. Holley put forward theirs ten or fifteen years ago, I should have been the first to recommend theru. In fact I should have been glad to see their narnes associated with rail patterns in their own country, in the same way as mine has been. PROFESSOR RICH. AKERMAN, STOCKHOLM, SWEDEN: i AGREE with your (Mr. Sandberg's) views that it is too early yet to stipulate only one definite chemical composition in the rails. Such a stipulation, I think, might he justified, if it were not for the manganese; but this metal changes the properties of the iron aud steel in tho same direc- tion ae carbon, silicon. and phosphorus, by increasing its tensile strength, stiffness, hardness, and brittleaess. If the iron and steel does not only contain carbon and manganese but also phosphorus,

Jan 1, 1881

By Dix, E. H.

THE percentage of manganese used in commercial aluminum alloys is small, and yet this element is an important addition to some very valuable alloys. When used alone with commercial aluminum containing iron and silicon as impurities, the usual percentage in wrought materials is from 1 to 1.5 per cent., whereas castings may contain as much as 2 per cent. An aluminum-manganese alloy containing nominally 1.25 per cent. manganese is much used in sheet form for the manufacture of cooking utensils and other articles requiring a little greater strength and hardness than could be obtained with commercially pure aluminum. It is also used for various purposes in the form of rods, tubes and pipes. The aluminum-manganese alloys, both wrought 'and cast, offer excellent resistance to many kinds of corrosion attack. The U. S. Naval Gun Factory uses an aluminum alloy containing from 1 to 2 per cent. each of copper and manganese, for both castings and forgings. Manganese is also used, up to about 1 per cent., as a minor alloying .element in several of the strong aluminum alloys of the duralumin type. It is felt that the role of this element when thus used has not been fully understood, but it is believed that the results obtained in this investigation will throw additional light on the subject. The earlier work on the system aluminum-manganese was general in nature, and no detailed investigations of the relations existent in the useful aluminum-rich alloys have been reported. Guilletl1in 1902 surveyed the system, and Hendrichs2 in 1908 published an equilibrium diagram which, although based on .few data, is often used in the literature. Fragmentary work has also been done by other investigators."

Jan 1, 1927

By N. J. Grant, A. W. Mullendore, Y. Ishida

Creep tests were performed at 500°F on polished specimens of A1-3 pct Cu at stresses of 2000, 4000, and 6000 psi. The effects of the size and distnbution of the second phase were studied in connection with the mechanisms of deformation, especially grain boundary sliding. Although the amount of grain boundary sliding varied considerably for the different heat treatments, yestriction of grain boundary sliding is not due primarily to precipitate on the grain boundary. NUMEROUS prior investigations of grain boundary sliding, notably those of McLean, 1 Rachinger, 2 Rhines, Bond, and Kissel, 3 and Brunner and Grant,4 have established the importance of this phenomenon in the deformation of pure aluminum and its solid-solution alloys at elevated temperatures. An extension of this kind of study to the case of a simple two-phase structure is an important step towards understanding the deformation behavior of more complex alloys. A two-phase A1-3 pct Cu alloy was selected for this study. The creep behavior of alloys in this system has been studied by Sully and Hardy,5 by Underwood, Marsh, and Manning,' and by Pelloux, Chaudhuri, and Grant.7 In these studies, the prior heat treatments were such that some precipitation took place during the tests in most cases. Pelloux, Chaudhuri, and Grant7 studied the creep of A1-2 and 3 pct Cu alloys in the overaged condition. Their results indicated that the structure most effective in promoting high-temperature strength consisted of a fine, closely spaced grain precipitate and large, widely spaced particles on the grain boundary. The present work consists of quantitative meas- urements of grain boundary sliding during creep of an A1-3 pct CU alloy with four different precipitate distributions. The aims were to quantitatively define the role of the second phase on grain boundary sliding, and to obtain information regarding the mechanism of grain boundary sliding. MATERIALS AND PROCEDURE A high-purity A1-3 pct Cu alloy, kindly supplied by Alcoa, was used in this investigation. The composition of the alloy in weight percent was as shown in Table I. This composition is such that A12Cu exists as a second phase at temperatures below 850°F. A homogeneous grain size of 0.6 to 0.8 mm was obtained by annealing the machined test bars for 5 min at 1000°F, furnace cooling to 930°F, and homogenizing for 5 hr. Following the recrystalli-zation and homogenizing anneal, four different heat treatments shown in Fig. 1 were utilized. The F7 and F5 treatments utilize a furnace cool to the aging temperature while Q7 and Q5 have a water quench from the homogenization temperature before aging. Figs. 2 to 5, and Table 11, show the size and shape of precipitate dispersions obtained with the above heat treatments. The heat treatments produce variations in the dispersion both in the grains and along the grain boundaries. Precipitate particles along the grain boundaries are rounded and large (about 1.5 p diameter) in the F5 and F7 structures (Figs. 3 and 5). The grain boundary precipitate was somewhat smaller (0.9 p diameter) in the Q7 structure and no grain boundary precipitate was detected in the Q5 structure using light microscopy. The pre-

Jan 1, 1964

By Dr. O’Neil Thomas J., Donald W. Gentry

There are mines that make us happy, There are mines that make us blue, There are mines that steal away the tear-drops As the sunbeams steal away the dew. There are mines that have lost the ore chute faulted, When the ore's forever lost to view, But the mines that fill my heart with sunshine, Are the mines that I sold to you. -Anonymous INTRODUCTION The preceding chapters have discussed, in various degrees of detail, the primary considerations necessary to perform a cash flow analysis of a mining-related investment opportunity. Although the previous discussions regarding the manipulations, calculations, and interrelationships of the various components of the pro forma income statement were illustrated in the text, many individuals are un- familiar with the format and appearance of a final spread sheet for the evaluation of a mining property. As a result, it seems appropriate to conclude this book with an abbreviated case study of an actual mine property (currently in production) showing the cash flow analysis and some back-up data. Obviously all of the supporting data and calculations for the case study presented cannot be provided; however, reference is made to the major assumptions utilized and to the appropriate procedures employed in the calculations, as described elsewhere in the text. The problem associated with using any case study for illustrative purposes is that one tends to be overwhelmed by detail. For instance, specific costs, prices, expenditure timings, and other variable values may be challenged because they either do not conform to one's expectations or are not consistent with specific operating experiences. When this occurs the instructional value of the problem becomes obscured. The purpose of most case studies, and particularly the one presented in this chapter, is to expose the student to the methodology, procedure, and format associated with performing a cash flow analysis of a mining property. It is not to laboriously wade through the specific details associated with operating and capital cost development, mine plans, or equipment schedules. It should be noted, how- ever, that the values represented in this case study result from specific mine plans, operating procedures, equipment spreads, manpower requirements, and other en-

Jan 1, 1984

By D. Jenkins

The Franklin Laboratory was designed mainly for the analysis of the products from the two concentrating mills situated at Franklin and Sterling Hill, the most important determinations being the zinc, iron, manganese, lime, moisture, and silica contents of the ores. Within the last 4 years, however, it has been found advisable to examine most of our supplies, and therefore additional space and equipment was allotted for the examination of such substances as oils, greases, soaps, alloys, fuels, paints, explosives, and water. The building as shown in elevation (Fig. 1) and plan (Fig. 2) consists of a main or routine laboratory, an experimental laboratory, a water analysis room, and the other accessories such as stock room, library, sampling and drying rooms, and office. The main laboratory is further divided into a balance room and zinc analysis room. As the greater part of our determinations are volumetric, good light is a very important factor and with this in view the interior malls of the building were constructed of white enameled brick laid in Keene cement. These also have the additional advantage that they give a clean and attractive appearance to the room. The windows are large, being 76 by 44 in. and even on very dark days the light is good throughout the whole laboratory. Our zinc titration requires a constant light and as, on many days, especially in winter, the light is variable, we installed two Artificial

Jan 1, 1918

By S. F. Shaw

IN 1913, the Tiro General mine, at Charcas, S.L.P., Mexico, which had been making from 125 to 150 gal. of water per min., was allowed to become flooded, after all the pumps had been removed, and in 1918 the problem of unwatering it came up for solution. There are two vertical shafts, the Tiro General, 1390 ft., and the San Fernando, 1225 ft. in depth, with five connecting levels between them. The San Fernando shaft had been partly retimbered, while the Tiro General had been newly timbered from the collar to, 1090 ft. in depth. It had been possible to place guides in only one compartment of the Tiro General, from the collar to the seventh or 850-ft. level, to which it was necessary to lower the water before station pumps could be installed. Soundings indicated that compartments in the Tiro General were open to 1082 and 428 ft., and in the San Fernando shaft to 471, 488, and 586 ft. The water had risen to a point 140 ft. below the collar of the Tiro General shaft. PUMPING EQUIPMENT The steam plant at the Tiro General shaft includes one 300-hp., one 150-hp. and one 100-hp. Erie City water-tube boiler, and two 160-hp. Heine water-tube boilers. All but one of these boilers were supposed to be in good shape, but it was soon discovered that six years of idleness in a climate that is foggy for several months of the year had left them in poor condition. The compressed-air plant comprises one Nordberg cross-compound compressor with piston displacement of approximately 2500 cu. ft. of free air compressed to 96 lb. per sq. in.; and another Nordberg cross-compound with piston displacement of 1235 cu. ft. of free air, to the same pressure.

Jan 2, 1920

By R. B. Roof, J. J. Lombardo

DURING the course of investigations on U-Cb alloys by X-ray diffraction techniques, an impurity or an additional phase was observed in severa1 alloys of various compositions. The X-ray dif-

Jan 1, 1959

By W. N. Poundstone

The mining and ventilating system used in development work in the Pittsburgh Seam in northern West Virginia is discussed. The seam conditions and the nature of the accompanying methane gas are described. The type of equipment and the mining cycle will be discussed, showing how they are well suited for very gaseous development work. Face ventilation in development work is possibly the fastest growing problem of the industry. The coal mines of the future will be faced with the prospect of mining from under increasing depths of cover. Consequently, larger and larger amounts of methane gas probably will be found. The Pittsburgh Seam, in northern West Virginia, is an example of an area already faced with this problem. At the present time, most of the development work being done in this seam lies beneath 500 to 1200 ft of cover. The Pittsburgh Seam in this area has always been very gassy—even near the outcrop—and the recent development work has been accompanied with extremely large volumes of gas. In many cases, a single development section has liberated in excess of 1,000,000 cfm in 24 hr. This problem of heavy gas liberation was the chief concern, several years ago, when continuous mining equipment was first considered at Christopher Coal Co. All of us were apprehensive about the liberation that would accompany rapid extraction in a single working place. However, the experience during the past few years has shown that this ability to mine only one place at a time, is actually the key to working this type of coal. With all of the mining or advancement concentrated in one place, the ventilation can also be concentrated. By this it is meant that continuous mining permits the active working place to be ventilated with a maximum amount of fresh air, taken directly from the intake source without first passing another working place. Continuous mining (and a good ventilating system) also permits a much greater concentration of attention or vigilance to the actual working place. There are two things that are very important to the mining of coal having high rates of liberation. First, adequate volumes of air are necessary. Second, and perhaps more important, a mining and ventilating system must be used that will provide an uninterrupted flow of air to every portion of the working section. Liberations of this magnitude take only a few seconds of interruption for a dangerous accumulation of gas to occur. With adequate volumes of air available, the ability to concentrate ventilation more than offsets the concentration of gas emission that is inherent with continuous mining. The mining system used with continuous mining equipment at Humphrey Mine is similar to the system used at many mines in the area for development work. This system is designed to favor ventilation, realizing that other efficiencies are meaningless if the equipment must stop because of ventilation difficulties. This plan is especially well suited to minimizing ventilation interruptions. Basically, the overall plan of mining is to develop headings into virgin coal and encircle or block out large areas. These blocks are generally at least 2000 ft sq. The purpose of this blocking out is to bleed gas from the area before pillaring. Experience has shown this method to be quite effective, even in the most gassy areas. The gas in this field seems to migrate or flow readily from the solid coal into the outside return headings of the development work. The numerous clay veins and slips that are found in the area are extremely good avenues for gas flow. A block of coal, surrounded with development headings, usually bleeds off readily; and since it is cut off from the virgin coal, it is not subject to gas migration, through the seam, from this source. However, the outside return of the encircling development work, adjacent to the virgin coal, may liberate gas for years. This liberation from the outside ribs in the virgin coal is the reason split ventilation is used in development work. If split ventilation were not used, there would, in many cases, be a serious build-up of gas in the intake before it could reach the working face. Fig. 1 shows a typical development section having seven headings. The two outside places on each side are returns, and the three center headings serve as intakes. This section is equipped with a ripper-type continuous mining machine. An off-track loading machine is used to load from a surge pile on the

Jan 1, 1961

By H. A. Dierks

PENNSYLVANIA's anthracite region lies in the heart of the richest and most densely populated area of the U. S. Nearly 70 million people live within a radius of 500 miles, in which 130,000 manufacturing plants employ more than 8 million workers. The region comprises some 3000 sq miles and contains 484 sq miles of coal measures, which folding and erosion have divided into four separate and distinct fields, known as the Northern, Eastern Middle, Western Middle, and Southern. One million people live in the anthracite region proper, and their prosperity depends to a great extent on the anthracite industry, which produces more than $200 million in new wealth for the U. S. each year. Nearly 80 pct of the tonnage shipped out of the region is handled by nine railroads, which depend on this traffic for a large part of their revenue. It follows that the economic health and the very existence 0: this industry is of real concern not only to the region itself but also to the state of Pennsylvania and the country as a whole. Anthracite mining is one of the oldest mineral industries in America. About 1760 burning rock or stone coal was discovered on the hillsides of northeastern Pennsylvania. Within a few years a systematic search for coal bed outcrops was under way, and hand mining was begun. At first production was very limited, since anthracite was difficult to burn and was not readily accepted as a fuel. Pennsylvania anthracite is one of the densest and hardest of all coals, and its content of volatile matter is low. Consequently it does not ignite easily and it burns without smoke. About 1808 Judge Fell of Wilkes-Barre burned anthracite in an open grate fireplace and demonstrated proper handling of an anthracite fire. From

Jan 1, 1958

By Charles M. Rolker

The Allouez Mine is situated in section 31, town 57, north of range 32 west, Michigan. The mine is being worked in a conglomerate bed, which conglomerate is generally conceded to be the continuation of the Albany and Boston conglomerate of Honghton County. Besides the well-known Calumet and Hecla with the adjoining Osceola, both working on the same bed, and the new location, the Seneca, now being tried on the Kearsarge conglomerate, the Allouez Mine is the only one working on a conglomerate bed. The Albany and Boston, in Houghton County, of which, as said before, the Allouez is supposed to be the continuation, shares with all the remaining conglomerate beds, near and around Portage Lake, the property of being non-quartziferous, while the conglomerates of Keweenaw County are all highly quartziferons in character. This change takes place, according to Professor Pumpelly, about six miles northeast of Portage Lake. The Allouez conglomerate, with a dip of 38" 15' and a general course of about 36' east of north, is overlaid by a trap, and has as its footwall an amygdaloid. The trap is followed by a series of amygdaloidal melaphyres and true amygdaloids, further on by the greenstone and ashbed; towards the lake shore alternating traps and conglomerates are found. At the footwall of the Allouez conglomerate, an amygdaloid lies, the thickness of which is not known, which is also the case with the other formations, accurate surveys and geologiwl researches of this district having, as I was informed, never been made. About 198 meters (650 feet), horizontally measured, from the Allouez conglomerate another amygdaloid bed can be traced, on which the present rockhouse stands, further on a melaphyre is exposed, and then follow the diverse conglomerates, the Calumet, the Kearsarge, and Kingston, which, according to the last geological report, are at about the re-