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By T. M. Allen, W. H. White

THE Greenwood-Grand Forks area of southern central British Columbia, known as the Boundary District, has a long history of mining exploration and production. At the turn of the century this was the premier copper mining camp in the British Empire, its total production amounting to some 20 million tons. Most of this ore came from the great Granby mines at Phoenix, but the Motherlode mine at Deadwood camp, 6 miles to the west, and several mines in Summit camp, 5 miles north of Phoenix, made important contributions. The large deposits were exhausted in 1918 and the district since has seen only desultory exploration and salvage operations. The orebodies are mineralized skarn zones in limestone members of a thick series of Upper Paleozoic sedimentary and volcanic strata. Chalcopyrite is the primary ore-mineral. Copper carbonates and silicates occur sparingly in outcrops, but the oxidized zone generally is very shallow. Much of the surface is mantled by glacial drift which in most places ranges in thickness from 2 to 15 ft. In some of the hanging valleys, however, the glacial drift may be as much as 100 ft thick and may assume drumlin-like forms. In 1951 an ambitious program aimed at the discovery of new orebodies and important extensions of abandoned deposits was launched by Attwood Copper Mines, Ltd. In this district so thoroughly searched by an earlier generation of prospectors, any orebody which had remained undiscovered must have little or no surface indication. Consequently, in addition to the basic detailed geological work, the program of exploration included magnetometer and self-potential surveys. Geological bets and geophysical anomalies were tested further, prior to diamond drilling, by a study of copper distribution in tree twigs and/or in the soil. The soil sampling and analytical methods used and some of the results seem of sufficient importance to warrant this paper. The authors had done some plant sampling in this and other districts, using the dithizone neutral-color-end-point method (Warren and Delavault, 1948, 1949; White, 1950),1-3 but they were unfamiliar with its soil application. Finally, after much experimenting in the field, they adopted the methods described here. These methods are not entirely original or defensible on theoretical grounds, but under field conditions of rapid sampling and analysis the results are reliable enough to be of use. Fig. 1, which shows the results of duplicate analyses of duplicate soil samples taken at 50-ft intervals across an anomalous zone, indicates the relative dependability both of the sampling and analytical methods. Sampling and Analytical Equipment A 2-ft piece of 1-in. solid drill steel, one end sharpened to a broad, conical point. The steel is marked at 1 ft from the point. A 2-ft piece of ½-in. black iron pipe, one end filed to a bevelled cutting edge. The pipe is marked at 1 ft 3 in. from the cutting end. A 3-lb hammer. A plastic or rubberized sheet about 18 in. square. Moisture-proof assay pulp envelopes. A 10-mesh seive made from window screen with the paint burnt off. A small assay spatula. A pan balance sensitive to 10 mg. Two ignition trays about 4 in. square, made of sheet iron turned up along the edges. A Coleman two-burner gasoline stove. An asbestos board about 5x8 in., used as a hot plate on the gasoline stove. A circular aluminum rack to hold 8 test tubes while refluxing (design of Almond and Morris). Pyrex Glassware Large refluxing test tubes, 25x200 mm, marked at 40 ml volume. Breakers, 20 ml. Pipettes, 1, 5, and 10-ml capacity. Graduate, 50 ml. Shaking cylinders, 100 ml, glass stoppers. Burette, 25 or 50-ml capacity, with holder. Chemical Supplies 1 N sulphuric acid. Hydroxylamine hydrochloride, solid crystals. Fisher Alkacid test paper. Copper standard solution. Dithizone standard solution 60 mg per liter. Water reasonably free of metals. Soil Sampling Method: The problem of how to take a soil sample is extremely crucial. The method outlined below, adopted after a number of tests, has the advantages of uniform pattern, uniform depth, and uniform size of sample. The area to be tested was marked off by chain and compass lines 100 ft apart, normal to the strike of possible ore deposits. Numbered stakes were set at 50-ft intervals along these lines and a soil sample was taken at each stake in the following manner. The drill steel was driven into the ground normal to the slope of the surface to the marked depth of 1 ft, moved slightly from side to side, then carefully withdrawn. The iron pipe was inserted to the bottom of this hole, tapped down to the marked depth of 1 ft 3 in. and withdrawn; the 3-in. soil plug in the

Jan 1, 1955

By Henry Sturgis Drinker, Rossiter W. Raymond

Jan 1, 1914

By W. R. Hibbard

THE classical work on the electrical conductivity of alloys was carried out by Matthiessen and his coworkers1 in the early 1860's. He attempted to correlate the electrical conductivity of alloys with their constitution diagrams, but the information regarding the latter was too meager for success. Guertler2 reworked Matthiessen's and other conductivity data in 1906 on the basis of volume composition (an application of Le Chatelier's principle with implications as to temperature and pressure effects), and obtained the following relationships between specific conductivity and phase diagrams (plotted as volume compositions) : 1—For two-phase regions, electrical conductivity can be considered as a linear function of volume composition, following the law of mixtures. 2—For solid solutions, except intermetallic compounds, the electrical conductivity is lowered by solute additions first very extensively and later more gradually, such that a minimum occurs in systems with complete solid solubility. This minimum forms from a catenary type of curve. Intermetallic compound formation with variable compound composition results in a maximum conductivity at the stoi-chiometric composition. Landauer" has recently considered the resistivity of binary metallic two-phase mixtures on the basis of randomly distributed spherical-shaped regions of two phases having different conductivities. His derivation predicts deviations from the law of mixtures which fit measurements on alloys of 6 systems out of 13 considered. Volency (Ionic Charge) Perhaps the first comprehensive discussion of the electrical resistivity of dilute solid-solution alloys was presented by Norbury' in 1921. He collected sufficient data to show that the change in resistance caused by 1 atomic pct binary solute additions is periodic* in character. The difference between the period and/or the group of the solvent and solute elements could be correlated with the increase in resistance. Linde5-7 determined the electrical resistivity (p) of solid solutions containing up to about 4 atomic pct of various solutes in copper, silver, and gold at several temperatures. He reported that the extrapolated"" increase in resistance per atomic percent addition is a function of the square of the difference in group number of the solute and solvent as follows: ?p= a + K(N-Ng)2 where a and K are empirical constants and N and Ng are group numbers of the constituents. This empirical relation was subsequently rationalized theoretically by Mott,8 who showed that the scattering of conduction electrons is proportional to the square of the scattering charge at lattice sites. Thus, the change in resistance of dilute alloys is propor-t,ional to the square of the difference between the ionic charge (or valence) of the solvent and solute when other factors are neglected. Mott's difficulty in evaluating the volume of the lattice near each atom site where the valency electrons tend to segre-gate: limited his calculations to proportionality relations. Recently, Robinson and Dorn" reconfirmed this relationship for dilute aluminum solid-solution alloys at 20°C, using an effective charge of 2.5 for aluminum. In terms of valence, Linde's equation becomes ?P= {K2 + K1 (Z8 -Za)2} A where K1 and K2 are coefficients, A is atomic percent solute, Z, is valence of solvent, and Zß, is valence of solute. Plots of these data for copper, silver, gold, and aluminum alloys are shown in Fig. 1. The values of K1 and K2 are constant for a given chemical period (P), but vary from period to period. The value of K, increases irregularly with increasing difference between the period of the solvent and solute element (AP), being zero when AP is zero. The value of K, appears to have no obvious periodic relationship. All factors other than valence that affect resistivity are gathered in these coefficients. Because of the nature of the coefficients, Eq. 1 is of limited use in estimating the effects of solute additions on resistivity unless a large amount of experimental data are already available on the systems involved. It is the purpose of the first part of this report to investigate the factors that may be included in the coefficients of Linde's equation. On this basis, it is hoped that the relative effects of solute additions on resistivity can be better estimated from basic data, leading to a more convenient alloy design procedure. It is well 10,11 that phenomena that decrease the perfection of the periodic field in an atomic lattice, such as the introduction of a solute atom or strain due to deformation, will also increase the electrical resistivity. Thus, in an effort to relate changes in electrical resistivity to alloy composition, it appears appropriate to consider the atomic characteristics related to solution and strain hardening

Jan 1, 1955

By E. O. Wagner, V. F. Parry, W. S. Landers

Development work has shown that the yield of primary tar from coal is proportional to the heat in the volatile matter of the coal and that the yield of tar from noncoking coals may vary from 10 to 45 gal per ton on the ash-free and moisture-free basis. Pilot plant operations have proved that 100 to 135 pct of the bench-scale carbonization-assay yield of tar can be obtained. FOLLOWING investigations by the Denver Bureau of Mines on drying fine coal in the entrained state,'. ' Texas Power & Light Co. employed the fluidized technique to upgrade Texas lignite for use in power plants. Because of the rapidly rising cost of natural gas in Texas the company agreed to cooperate with the Bureau of Mines in studying further the upgrading of lignite by low-temperature carbonization, since potential value of the tar would help offset the cost of fuel. Considerable interest was aroused, therefore, when Aluminum Co. of America decided to use dried lignite in its 240,000-kw power plant at Rockdale, Texas, operated by Texas Power & Light. Carbonization of coal before burning in a power plant is not new. Plants in Germany have been operated for many years on carbonized brown coal and lignite briquets produced in large integrated plants. Large-scale experiments in carbonizing pulverized coal with hot flue gases were made 30 years ago at Milwaukee," and many inventors have proposed other processes. In the U. S., technical and economic problems have prevented successful operation of large plants on carbonized coal. but new techniques of handling solids in fluidized beds may overcome these difficulties, particularly if the higher volatile noncoking coals are used. Any bituminous material will decompose when heated to about 900°F to yield primary tar, gas, and char. The quantity of tar obtained from various coals depends on their rank and volatile content and to some extent on rate and method of heating. The liquid products derived from thermal decomposition of peat, brown coal. lignites, oil shale, and noncoking bituminous coals can be easily liberated by rapidly heating these materials in a fluidized bed at about 500°C (932°F). Since it takes only a few minutes to heat small particles of fuel to carbonizing temperature in a fluidized bed, processing plants can be built to handle up to 50 tph at relatively low cost. The pioneering development work being conducted at Rockdale, Texas, by Alcoa and Texas Power & Light will solve many engineering problems of such operations. The Bureau of Mines at Denver has studied the carbonizing properties of several hundred coals by small scale assay at 500°C, deriving a simple correlation between the potential yield of tar and the proximate analysis. This correlation, shown in Fig. 1, compares the heat in the volatile matter of the coal with the potential yield of primary tar. The yield of primary tar varies from 12 to 50 gal per ton of pure coal." About the same amount of heat is needed to * MAF equals moisture-and-ash-free basis. distil any dry coal at 500°C, and the time required to carbonize a given particle size is about the same for any coal. It is therefore obvious that coals rich in volatile matter will show the greatest profits when processed. Texas lignites are rich in volatile matter, which contributes some 5700 Btu per lb of MAF coal, and the potential primary tar and oil yield from this lignite averages 12 pct by weight or 27 gal per ton from the pure lignite. Texas lignites are considerably richer in potential tar than lignites of North Dakota and Montana. During 1950 the Bureau of Mines conducted many experiments on carbonizing fine dry coal in the entrained state, publishing a report4 n major features of pilot and commercial-size units required to carry out the process. Additional experimental work has been done at Denver and the pilot plant has been improved, particularly the facilities for studying removal of fine dusts from tar vapors and gases at high temperatures by electrostatic precipitation. A number of materials have been carbonized, including sawdust, peat, brown coal, Brazilian oil shale, lignites, and the lower rank bituminous coals having noncoking properties. Straight coking coals cannot

Jan 1, 1957

By N. G. W. Cook

The conventional cyclic system of deep-level mining by drilling and blasting gives rise to an inadequate degree of stope sorting when mining thin reefs. This results in poor utilization of the capital facilities of a mine in the form of shafts, haulages, airways, and associated equipment. Continuous and controlled removal of the thin gold-bearing portion of the reefs would permit better stope sorting and hence greater utilization of capital facilities. Results of experiments to develop hard-rock cutting machines for mining are reported and the benefits which could be derived from their use are discussed. Mining, from exploration through refining, is essentially a process of sorting in which the payable mineral, or metal, is progressively separated from the other constituents of the earth's crust with which it was originally associated. This takes place more distinctly at each step of the operation. What is it that determines the degree to which sorting should be carried out at each of the several steps comprising a whole mining operation? The formal answer is that degree of sorting at each step which results in the lowest overall cost for the complete separation. In practice, individual steps are chosen from those available in current technology, each of which effects a degree of sorting such that the quantity of material which must be sorted in the succeeding step is economically acceptable. It follows that any new technological development has repercussions throughout the whole mining operation and, more important, that the solution to excessive costs in any one step of the operation may lie not in improving the costly operation itself so much as in increasing the degree of sorting preceeding that operation. This concept, particularly in relation to deep-level mining of thin, tabular gold-bearing reefs in South Africa, is discussed here, and the most recent results achieved in the development of hard-rock cutting machines for stoping more selectively than is possible with explosives are presented. Deep-Level Mining Deep-level mining involves operations which are either not encountered, or are of only trivial importance, in near-surface mining. Near-surface, the major operations are those of rock breaking, transport, and milling. In deep-level mining, hoisting, environmental control, and strata control assume major importance. Some idea of the relative magnitude of these operations may be gained by comparing the separate amounts of energy which are required, or which must be controlled, to effect the various operations when, say, mining a tabular deposit 40 in. thick at 8000 ft below surface, Table 1. It is true that the costs of handling a given quantity of energy are not the same for each operation. Nevertheless, Table 1 does emphasize the fact that the operations of hoisting, strata control, and environmental control are of unique and major significance in deep-level mining. In particular, hoisting and environmental control place a heavy load on the reticulation system of the mine—the shafts, haulages, and airways. Typically, a new, deep gold mine with an annual revenue of about $35 million requires a total capital expenditure of about $140 million of which some $100 million is invested in developing and equipping this reticulation system. The ratio between annual turnover and capital invested of about one-quarter is exceptionally low, and it typifies the poor utilization of capital by the current technology of mining hard rock at depth. The average thickness of the reefs in the new South African goldfields varies from 10 to 30 in.,l and even in the thicker reefs the gold is often confined within a small fraction of the nominal thickness. Nevertheless, it is universal practice to mine these reefs at a stope width of about 40 in. or more, so that the quantity of rock broken in the stopes and hoisted out of the mine is between two and ten times the quantity of rock actually carrying a significant amount of gold. The reason for the adoption of such excessive stope widths is to be found in the method of rock breaking by drilling and blasting. The only free surface to which a blast hole can break is the stope face. It follows2 that each hole cannot have a burden in excess of the height of the free face if it is to break satisfactorily. To

Jan 1, 1971

By A. M. Gaudin

THE large increase in the use of selective flotation as contrasted with the collective flotation of a few years ago has focussed attention on the desirability of achieving accurate control of the pH of mill circuits. Although it is well-known that alkaline circuits result in dropping pyrite and that acid circuits conversely float it, but drop the gangue, there is little that is known, beyond question, ass to the action of the acidity or alkalinity of the circuits. A complete answer is not available at present but experiments -are now in progress at the University of Utah, whose aim is to find out what the effect is on recovery of a change in hydrogen-ion concentration and how this change is brought about. Broadly speaking, a change in hydrogen-ion concen¬tration seems to affect either the mineral surface, as was suggested by A. W. Fahrenwald in his paper "Surface Reactions in Flotation"1 (TRANS. 70, 647) or the attachment of the collecting agent on the mineral surface.

Jan 1, 1929

By C. E. MACQUICC

WITHIN a period considerably less than two decades, the engineering view of alloy steels has greatly changed-both as to their composition, and applications. Inasmuch as the elements used in manufacturing alloy steels are substantially the same as those which have been used for the past two or three generations, a statement like the foregoing may reasonably call for some explanation. The first commercial use of alloy steels in this country dates back to the sixties of the last century, when one of the earliest alloy steel patents, covering the use of chromium, was taken out by Bauer. About ten years later, German and French toolmakers began to introduce tungsten into tool steels. Developments along this line were more or less static for 25 years or more, the tool steel, compositions progressing along lines worked out by Mushet and one or two others in Europe. The next step in advance, made by Taylor and White inn the late nineties, consisted of heat-treating the alloy steels to enable much higher cutting speeds, thus introducing what soon became known as "rapid" or "high-speed" steels. All these developments being aimed toward the improvement of cutting tools, the compositions made were essentially of a hard, brittle nature. Following a different line of development started in the seventies, naval ordnance material began to embrace the nickel steels. These were rapidly popularized, became, and still are, with the addition of one or more elements, as chromium, molybdenum, vanadium, standard for heat-treated armor plate. During that period, however, no notable uses were made of alloy steels for engine parts or other high-stress parts where they are now considered indispensable; moreover, the present-day use of metals in the chemical industries was probably not even imagined. After the turn of the century, the necessity for making heavy-duty. machine parts, and especially the develop¬ment of the automobile, caused alloy steels to rapidly supplant plain carbon steels for engineering purposes.

Jan 1, 1930

By E. F. Reed

CONSIDERABLE deep hole prospect drilling has been done in the last few years in the Globe-Miami mining district about 70 miles east of Phoenix, Arizona, and in the San Manuel-Tiger area about 50 miles south of the Globe-Miami region. More than 205,000 ft of churn drilling have been completed by the San Manuel Copper Corp. at their property in the Old Hat Mining District in southern Pinal County. The deepest hole on this property is 2850 ft; there are 49 holes deeper than 2000 ft. At the adjoining Houghton property of the Anaconda Copper Mining Co., where only one hole reached 2000-ft depth, there were 27,472 ft of churn drilling and 3436 ft of diamond drilling. Three churn drill holes were deepened by diamond drilling methods. Near Miami in the Globe-Miami district the Amico Mining Corp. drilled four holes by combined churn and rotary drilling methods, the total amounting to 13,879 ft, of which 2256 ft were drilled with a portable rotary rig. In the same district, besides doing a large amount of shallow prospect drilling, the Miami Copper Co. drilled two holes of 2560 and 3787 ft, respectively, which were completed by churn drilling methods. The rocks encountered in drilling at San Manuel and Tiger are described by Steele and Rubly in their paper on the San Manuel Prospect' and by Chapman in a report on the San Manuel Copper Deposit? The rocks are well-consolidated Gila conglomerate, quartz , monzonite, and monzonite porphyry. In some places these formations stand very well while being drilled, and three holes were drilled without casing, the deepest of which was 2200 ft. In other holes faulted and fractured ground made drilling difficult. In the Globe-Miami district the deep drilling was done in the down-faulted block of Gila conglomerate east of the Miami fault and in the underlying Pinal schist. The geology of this area is described by Ransome.3 In the Amico holes the conglomerate varied from material consisting entirely of granite boulders and fragments to a rock made up of schist fragments in a sandy matrix; in the Miami Copper Co. holes there were more granite boulders and the material was poorly consolidated. Drilling was much more difficult and expensive in the Miami area than in the San Manuel district, mainly because of the depth of the holes and the formations drilled. All the deep hole prospecting described in this paper was done with portable rigs. The churn drill rigs were of several types, of which the Bucyrus-Erie were the most popular. Bucyrus-Erie 28L, 29W, and 36L rigs were used on some of the deeper holes on the San Manuel property. A few Fort Worth spudder types were tried, and the deepest hole at San Manuel was drilled with a Fort Worth Jumbo H. The spudder type is considerably larger than most other rigs used on this work and required a larger location site. The spudders were belt-driven machines with separate power units, and time required for setting up and moving was much longer than with the more portable drills. All the churn drilling was done by contractors or with machinery leased from them. A few of the contractors had complete equipment, including most of the necessary fishing tools. Unusual and special, fishing tools were obtainable from the supply companies in the oil fields of New Mexico or in the Los Angeles area. Most of the contractors used equipment with standard API tool joints, so that much of it was interchangeable. Failure of tool joints is one of the principal causes of fishing jobs. It can be minimized if the joints are kept to the API specifications and the proper sized joints are used in the various holes. The minimum sizes that should be used with various bits are as follows: 12-in. and larger bits, 4x5-in. tool joints; 10-in. bits, 3 1/4x4 1/4-in. tool joints; 8-in. bits, 2 3/4x 3 3/4-in. tool joints; 6-in. bits, 2 1/4 x3 l/4 -in. tool joints; 4-in. bits, 1 5/8 x2 5/8-in. tool joints. Two rotary drill rigs were tried at San Manuel on the same hole, and a portable rotary drill rig was used on the Amico drilling for test coring the formation and for drilling in holes 3 and 4. Rotary drilling differs from churn drilling or cable tool drilling in that the bit is revolved by a string of drill pipe and the cuttings are removed from the hole by a thin solution of mud pumped through the drill pipe. The principal parts of a rotary rig are the power unit, a rotating table to revolve the drill pipe, hoists to raise and lower the pipe and to handle casing, and a pumping system to circulate the drilling liquid. The rig used on the Amico property at Miami was mounted on a truck. The larger rig used on the San Manuel property was hauled by several trucks and had separate turntable and pumping units. Diamond drill coring equipment was used successfully with the rotary rig in the holes on the Amico property, To allow for 2 3/8-in. drill pipe with tool joints, 3 1/2-in. core barrels and bits were used. With the standard 3 1/2-in. core barrel there was considerable difficulty in maintaining circulation with mud, so a barrel was designed with a smaller inner tube and a broad-faced bit. This allowed coarser material to circulate between the barrels. Rock bits of 5 5/8 to 3 7/8 in. were used with the rotary rig for drilling between core runs. Diamond drill equipment is much lighter than churn drill tools, so that fishing tools can usually be obtained from supply houses by air express when needed. Three churn drill holes on the Houghton property at Tiger were deepened by diamond drilling with Longyear UG Straitline gasoline-driven-machines. The open churn drill hole was cased with 2 1/2-in. black pipe. In deep hole churn drilling, casing is one of the most important items, especially in drilling in unconsolidated material like the formations drilled by

Jan 1, 1952

By E. R. Borcherdt

IT is notable that the first large-scale mine operation equipped entirely with detachable bits was the Badger State mine of the Anaconda Copper Mining Co. in Butte, Montana, just 30 years ago. This mine in 1922 was producing approximately 1200 tons of ore per day. Much of the data presented in C. L. Berrien's article' describing the development and installation of the Hawkesworth detachable drill bit were obtained from these operations. As in any pioneering effort, no precedent existed and many difficult problems required solution, so that the changeover to detachable bits at all Butte hill mines was not completed for 6 years. There was widespread disbelief as to the probable efficiency of the new installation. Some attempts were made in 1931 by the owners of the Hawkesworth patents to interest Ontario gold mine operators in the bit. These efforts were not successful, but they undoubtedly stimulated thinking which resulted in the invention and patenting of several well-known Canadian detachable bits, one of which is now a widely used throwaway bit. The success of the Butte installation also led to the development of the threaded type of bit connections by several well-known manufacturers, and in 1935 these bits were introduced to the mining industry on a national scale. The original Hawkesworth bit was not provided with a water hole but, depended upon water passing through the clearance opening between the tongue in the bit and the groove in the rod to flush cuttings from the drill hole, see Fig. 1. In December 1935 it was found that this method of introducing drilling water to the bit face resulted in high dust counts. To correct this a water hole was drilled on the central axis of the bit, passing through the tongue. Unfortunately, quenching water would rise through the small water hole, spot-hardening the tongue to cause breakage, never completely eliminated. In the fall of 1936 large-scale tests indicated that savings would be effected by use of a threaded type of bit, which was therefore adopted as standard for all Butte mines. This type of bit was used until 1947, when it was superseded by a one-use slip-on type. Since the first use of the Hawkesworth bit every detachable bit of importance has been investigated, and where advantages which might reduce costs or increase efficiency were indicated, substantial tests of the bit were carried on in the Butte mines. When tests demonstrated the advisability of changing from one kind of detachable bit to another the change was made at one level or in one area each day until the new rod and bit equipment was used throughout the mine. This involved a minimum of cost and disruption of drilling. Intelligent selection of a detachable bit to obtain optimum results requires careful consideration to achieve a balance between the three principal types of equipment used in the drilling process: 1—drill bits, 2—drill steel, and 3—drilling machines. Optimum results imply maximum output and minimum cost per unit of output. Since every rock type differs in drillability and it is generally impractical to provide equipment for more than one or two types of rock which may occur in one operation, selection of equipment must encompass average drilling conditions. However, on exceptional occasions several widely differing conditions may make it mandatory to provide equipment best suited to each condition. The choice of rock-drilling equipment is a most controversial subject and one that is further complicated by unreliable and frequently misleading performance claims. Small operators without the means for making accurate evaluations of equipment frequently suffer from these over-enthusiastic claims. It is apparent from experience in rock drilling throughout the world that rock drillability is not alike in any two places, and that selection of proper equipment can only be made after conducting thorough trials of various types of equipment. Some recent drilling tests in tactite and hornstone at the Darwin, California mine of the Anaconda Co. present some interesting clues on rock drillability. Microscopic examination of thin sections of these rocks reveals that mineral composition and rock texture are equally important in governing drillability. The Darwin hornstone is at times so abrasive that the carbide bit cutting edges become flattened to 3/32 in. in 2 to 4 ft of drilling, and some carbide bits were dulled to this point after 9 to 10 in. of drilling. This wear was determined to be the proper point for resharpening to eliminate carbide insert breakage or breakage of the steel rod when drilling with 1½ to 1?-in. bits, with a drifter of 2 3/4-in. diam and 90 to 100 psi air pressure, see Supplement A. Before considering the merits of various bit designs it may be well to review the mechanics of drilling rock with percussion drills. A sharp bit cuts by penetration and chipping. The amount of penetration governs the amount of chipping and depends upon the contact area of the cutting edge, the foot-

Jan 1, 1954

By K. Adam

Grinding in steel mills adversely affects the flotation of pyrrhotite. The electrochemical nature of inter action between pyrrhotite and active metals was investigated by contacting pyrrhotite with platinum at different applied potentials. Flotation tests, scanning electron microscopy, Auger electron spectroscopy, and X-ray photoelectron spectroscopy were wed to examine the surface properties of pyrrhotite. The formation of hydroxide or oxide and sulfate species of iron through the oxygen reduction reaction at the cathodically polarized surface of pyrrhotite was shown to be the mechanism responsible for the reduced floatability of pyrrhotite.

Jan 1, 1985

By Blamey Stevens

(San Francisco Meeting, October, 1911.) IN this paper, which is a logical extension of my paper, The Laws of Intrusion, 1 the various pressures of emanation and their mechanical causes and effects on the large scale of nature are determined, .and a classification of emanations is deduced therefrom. The Subterranean Sea. The subterranean sea is here defined as the intercommunicating water contained in the trains of pores and fissures of the rocks down to great depths. Whether the actual amount of water constituting this subterranean sea be large or small, does not enter into the present argument, since the static pressure of this water at any particular depth is not a function of the quantity : it involves only the depth and the density of the water. Densities. The total pressure at depths down to 10 or 15 miles is made up of the rock-stresses investigated in my previous papers, and the pressure of the subterranean sea. The latter component must evidently be equal to the weight of a solid column of the overlying water of unit sectional area. As water expands with heat, it should really be necessary to sum the density at every level. A knowledge of the densities of water under a variety of conditions is therefore needed, but does not seem ever to have been ascertained for such combined pressures and temperatures as we have to consider. The specific gravity of water at 28° C. for various pressures up to 400 atmospheres, has been determined by Barus,2 who

Apr 1, 1912

By William B. Plank

ENGINEERING students to the number of 73,269 had been enrolled in United States and Canadian schools on Nov. 5, 1945, but, as shown in the following tables, even this sizable number will not greatly remedy the 150,000 deficit reported by Vannevar Bush, director of the U. S. Office of Scientific Research and Development, to exist in the engineering and scientific man-power supply. These figures were received in advance from the Society for the Promotion of Engineering Education, and are given in detail in the January 1946 issue of the Journal of Engineering Education. Although the enrollment on Nov. 5 was an increase of 76% over the low of 41,632 in 1944-45, it was still 36% below the peak enrollment in 1941-42 of 115, 228. These figures include both undergraduate students and graduate students in the schools of the United States and Canada and are therefore comparable. As was to be expected, almost half of this enrollment is in the first or freshman year and 25% of the total are veterans. Also, 1,850, or about 2.5%, are women. A measure of the seriousness of the engineering graduate supply in these two countries is indicated by the trend of the number of first degrees that have been conferred during this period of scarcity:

Jan 1, 1946

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Jan 1, 1885