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By A. L. Fentress, P. R. Bucky

The purpose of this paper is to present a scientific method for determining the proper span and shape of roof for safe and economical mining; at present, these two vital factors wherever pillars are left to support the roof are determined by local practice or general experience. To develop the method described here, experimental laboratory work was done. The arch adopted was a segment of a circle, a flat roof being considered an arch with an infinite radius. Hitherto, arches have been used underground without the background of stress analysis that generally precedes arch construction on the surface. In many cases results have not been as expected, because of the lack of knowledge of the factors affecting its strength and behavior. A theoretical and experimental study of arches should furnish information that will make it possible to design safe and economical surface structures as well as underground openings. For example, on the subject of mine drifts and tunnels, the present idea is that arching their backs increases their strength; this paper will indicate that this is not always true, and if it can be shown that under certain conditions arching becomes unnecessary an economic saving will result. In room-and-pillar work a question may arise as to whether arching the roof and gobbing the excavated material will result in timber saving, and a decrease in the dangers from falls of roof. It is to be distinctly understood that each mine must be treated as an individual problem, and that answers to these questions and many others can now be found after experimental laboratory data have been obtained. Mine-Arch Structure Fig. 1 shows the elements that affect the strength or behavior of the experimental arch. They are the span, crown thickness, radius of curvature, rise and restraint. The restraint in all cases was made sufficient to cause the arch to fail without displacing the restraining sides. Fig. 2a is a photograph of a model arch prepared for testing in the centri-

Jan 1, 1934

By H. I. Althulee

The mines operated by the Compañía de Real del Monte y Pachuca, Pachuca, Mexico, are in two districts, the Pachuca, and Real del Monte. The principal area of mineralization is within a rectangle roughly 10 miles from east to west and 6 miles from north to south. It contains a number of veins from 2 to 20 ft. wide, which are exploited by this mining company. The scattered position of the producing mines led to the construction of two mills between 1906 and 1908; the Loreto mill, in Pachuca, averaged 2600 tons daily, and the Guerrero mill, in Real del Monte, milled 1270 tons daily. On Feb. 1, 1929, the management decided to enlarge the Loreto mill to take the combined tonnage of both mills at that time. This economic measure required the centralization of ore delivery, which would also afford a material saving in the cost of ore transport. The centralization would give the company the following advantages over the former system of operation: 1. Beneficiation at a single mill, closer to source of materials and supplies than the abandoned Guerrero mill. The refinery is also in the Loreto patio. 2. Ore delivery to surface at one point (San Juan Pachuca shaft) instead of at four. 3. Elimination of aerial tramways. To show what would be required to effect this centralization, it is necessary to describe the means by which the ore was delivered to the two mills, prior to centralization. The heart of the Real del Monte district is roughly four miles from that of the Pachuca section. The producing mines in these two districts were: Pachuca district: Real del Monte district: Santa Ana mine La Rica group Camelia group Purisima group Santa Margarita mine

Jan 1, 1934

By C. S. Anderson

Georgia, since 1829, has produced nearly $18,000,000 from her gold mines, but in late years the output has dwindled to insignificance. In view of present universal efforts to increase gold production, it seems pertinent to record some recent observations in that field, briefly review its past and weigh probabilities of reviving gold mining there. Little in this paper can be claimed as new or original, and it does not present results of a detailed or profound investigation of the whole region. The Allatoona gold mine was reopened in January, 1932, and explored at greater depth during the period from January to June of that year. A number of other abandoned mines were visited, but since the workings, in common with most such mines in Georgia, were flooded or caved, underground conditions could not be examined. For a wider knowledge of the subject, a search was made through available pertinent literature and statistics, a bibliography of which is appended hereto. To all the sources obligations are gratefully acknowledged. Production The discovery of gold in Georgia, in 1829, led to an influx of 6000 to 7000 miners and a production of over $200,000 in the following year. A branch of the United States Mint, established at Dahlonega in 1838, accounted for $6,115,569 up to the time of the Civil War. The output fluctuated greatly from year to year, with an all-time peak of $582,782 in 1843. In only one other year, 1845, was the half million mark exceeded. After the war mining began again, but 1883 is the last year for which an output of $200,000 is reported. Placers, always the main source of production, were yielding less and less, and in spite of increased quartz-mining activities the downward trend continued. The last year in which there was a production of $100,000 was 1902. In 1914, $16,270 came from 26 placers and 11 deep mines, with the latter reporting 1750 tons of $3 average value. In 1925, one placer and four lode mines that mined 8100 tons assaying $1.19 per ton produced $9500. One placer and one quartz mine, operating in 1930, reported a total output of $4194. Table 1 pictures the waning output.

Jan 1, 1934

By C. H. Loux

This paper deals primarily with the advantages realized from the adoption of methods more suitable than those previously used. The improvements noted include: (1) modification of practice in shrinkage stoping, (2) use of sublevel caving over grizzlies, (3) improved haulage and loading, (4) use of contract wages wherever possible. The Washington mine is on Jonestown Hill, one mile south of Oxford, in northwestern New Jersey. It is producing 16,000 tons of concentrates a month, which are being shipped to the Alan Wood Steel Company's furnaces at Swedeland, Pennsylvania. History The first settlement in this region is reported to have been in 1735, and the magnetite ores were recognized immediately. The first furnace, built in 1741, began operation in 1743. During the Revolutionary War the output was chiefly used for making cannon balls, which, owing to the sympathies of the community, without doubt found their way into the hands of the British. This old furnace is still standing. Several important veins were soon developed, including the Martin, Strayley, McKinley, Adelaide, Welsh and Lanning. Mining was carried to a depth corresponding to the capacity of the pumps available. When these became inadequate, the workings were abandoned, and another vein was opened up. The Washington vein was probably discovered about 1850, and the first shaft started in 1854. In 1879, William H. Scranton made the first complete magnetic survey of the district and serious exploration of the deposit was begun. Rather indifferent success attended the earlier efforts of the many companies that attempted to mine and smelt the ore. In 1929 the property was leased by the Warren Foundry and Pipe Co. to the present operators, and the modification of practice then inaugurated is described in this paper. Geology These veins, like most eastern deposits of magnetite, are formed from, or are contemporaneous with, pegmatite intrusions into a gneissic country rock. The ore therefore follows the banded structure of the gneiss,

Jan 1, 1934

By T. C. Desollar

The group of iron-ore mines known as the Red Ore Mines, owned and operated by the Woodward Iron Co., is on Red Mountain approximately ten miles southwest of Birmingham, Ala. At the present time NO. 1 and No. 3 mines are active, No. 2 mines having been abandoned in 1928. Ore from this group of mines was first mined in 1882, and originally came from open cuts along the outcrop, but, when the "soft ore" (weathered hematite) was mined out and the harder ore was encountered, it became necessary to develop the orebody by means of slopes driven in the ore formation. The three mines to date have produced more than 22,000,000 tons, and the two operating mines now have a capacity of over 1,000,000 tons per year. Geology The U. S. Geological Survey has described the features of the district. It places the hematite of Red Mountain, commonly called "red ore," in the Clinton formation of the Silurian system. This formation varies from 200 to 300 ft. in thickness, and is composed largely of shale (slate) and sandstone. There are four iron orebeds in the formation, known as the Hickory Nut, Ida, Big Seam and the Irondale seams. The strike of the orebeds is northeast and southwest, and the dip is southeast, the outcrops being on the west side of the mountain. The two seams that have been worked at the Red Ore mines are the Big Seam and the Ida. The Big Seam, which is and has been the principal source of the ore mined, is approximately 20 ft. thick and consists of hematite and ferruginous bands of sandstone. It is separated into "upper" and "lower bench" ore by a slate parting that is about 30 in. thick. All present mining is confined to the "upper bench" ore, which is uniform in texture and drills and breaks fairly easily when the holes are placed so as to take full advantage of the face of the ore. Typical analyses are shown in Table 1. The presence of two water-bearing formations, the Fort Payne chert and the Hartselle sandstone, lying respectively 160 and 300 ft. above the Big Seam, make it essential that in mining the ore any movement of

Jan 1, 1934

By Robert K. Warner

When a new mine is opened, and often when an operating mine must adapt itself to physical or economic changes, a mining system must be selected in complete detail. In the past the plan chosen was usually the product of the local staff alone, however limited the training and experience of that staff might be, which often resulted in the introduction of methods that were familiar to the staff, but not so well suited to the conditions as others unknown to the local men. A new management, noting the difficulties with the existing layout, might make modifications or complete changes, so that many mines in their history have seen a whole series of mining systems, each somewhat laboriously grafted on to its predecessor. This should not be interpreted as a condemnation of the older staffs for sticking to the tried systems, such as open square-sets in the early days at the Homestake, because they knew it would work after a fashion, nor does the statement apply to the technically trained, and usually broadly experienced staffs of most present-day properties. The older miners did not have the variety of tools and methods, nor the present freedom of transfer of information as to what the other man is doing, and so did the best they could under the circumstances. Finally, the higher grade of the ores made the choice less vital. The ever-increasing need to mine low-grade orebodies, with a fixed or even lower price for the metal produced, has caused an increasing need for low-cost mining. In most new mines this has simplified the problem of selection by eliminating from consideration all of the high-cost systems, but it has also complicated the choice by the necessity for refinement in detail and a close adjustment of the plan selected to all the conditions at the property. In these low-cost problems any textbook division of mining systems as applied to "beds, veins or masses" avails little as a guide. It should be realized at the outset that to be successful under modern competitive conditions a system of mining must be worked out with elaborate detail, with careful time scheduling and control of the various elemental operations, and that after it is introduced it must be run with a large amount of technical direction.

Jan 1, 1934

By D. D. Homer, R. H. Cromwell

The Minas de Matahambre No. 2 shaft, vertical, four-compartment lined with steel sets and hardwood lagging, has been sunk 2057 ft. On account of rock structure, over-all cross-section dimensions are 23 by 6 ft. Certain sections have been concreted. Two compartments are for handling ore and men, one for supplies and men, and one for manway, cables and counterweight. Two 250-ton capacity skip pockets with air-operated gates and steel cartridge pockets have been installed underground. A steel "A" headframe, a 250-ton steel receiving bin, and skip-to-man-cage changing equipment have been provided. Skip capacity is 102½ cu. ft. The hoist is a Nordberg double-drum 10 ft. by 6 ft. 6 in. geared, electrically driven, with parallel grooving for 1-?in. cables. Maximum hoisting speed is 1600 ft. per minute. The motor-generator set, control panel, slip regulator and hoist accumulator are housed in the basement of the hoist house. The electrical equipment is General Electric. The direct-current hoist motor is 800 hp. Control is of the Ward-Leonard-Ilgner type. The motor-generator flywheel set, normal speed 720 r.p.m., consists of an 800-hp. induction motor, a 22-ton flywheel (dia. 117 in.) and a 650-kw. generator.

Jan 1, 1934

By P. R. Hanes

A description of the methods used in driving two tunnels, 2 miles long each, one a 9-ft. and the other a 19-ft. diameter section, for the Owyhee Irrigation Project, Oregon. One heading of the smaller tunnel was driven 1315 ft. in 73 shifts in one month and 10,966 ft. of this section completed in ll½ months using two shifts for eight months and three shifts the rest of the time. This is a good illustration of the proper use of a mechanical shovel and the importance of careful selection of supplernentary equipment to secure best results. Large cars, speedier and heavier locomotives, 36-in. gage track of 45-lb. rail, and a good switching method increased actual shoveling time and cut down muck removal period about 45 minutes from average. The larger tunnel, after trying out standard electric revolving shovels of both the rail mounting and caterpillar tread types, was excavated with a mechanical mucker. This is the first time that a tunnel of so large a section has been handled this way. This record work was accomplished without drill carriages.

Jan 1, 1934

By A. Lee Barrett

SO commonplace that they are seldom noticed, mining-machinr, bits have a definite and important bearing on the cost of coal production. At the average mine many thousands of bits are used during the year. The actual cost represented by the purchase of the bits used, as well as that of sharpening and handling them, represents a considerable sum of money. From the point of view of indirect costs they are important, as the lack of good bits at a mining machine can seriously impair production. Good bits tend to ensure an adequate supply of coal. Classes OF Bits Mining-machine bits may be divided into three general classes, based on shape and method of use: (I) chisel, (2) pick-point and (3) one-use. The chisel and pick-point bits are shown in Fig. I. The so-called "chisel" bits were made of steel of various sizes and were sharpened to a wide chisel cutting edge. Few of this type of bit are used, except in places where coring is very bad. Sometimes chisel bits are used with standard bits to help cut out the core. The "pick-point" bit is widely used today. Normally it is made from ½ by I-in. steel, is about 4½ in. long when new, has a rake angle of about 25°, a front clearance of about 23O, a side clearance of approximsteji 8".' There are many variations in the shape of this bit, depending on the ideas of the individual that may be dressing the dies of the bit machine. This bit has been generally used for many years and at present, in one form or another, is by far the most popular bit used in coal mines. This bit usually is made from 0.85 carbon steel, although in recent years some alloy-steel bits have made their appearance. At present, carbon-steel bits are frequently tipped in. hard facing materials. A third class of bits might be called the L'one-use" bit; that is, a small bit that is used once and then thrown away. These bits have two or three points, are usually made from 0.85 carbon steel and sometimes from higher alloys. Recently some. of them have been tipped with hard facing material. Most of them were used at first with special

Jan 1, 1942

By A. Lee Barrett

SO commonplace that they are seldom noticed, mining-machinr, bits have a definite and important bearing on the cost of coal production. At the average mine many thousands of bits are used during the year. The actual cost represented by the purchase of the bits used, as well as that of sharpening and handling them, represents a considerable sum of money. From the point of view of indirect costs they are important, as the lack of good bits at a mining machine can seriously impair production. Good bits tend to ensure an adequate supply of coal. Classes OF Bits Mining-machine bits may be divided into three general classes, based on shape and method of use: (I) chisel, (2) pick-point and (3) one-use. The chisel and pick-point bits are shown in Fig. I. The so-called "chisel" bits were made of steel of various sizes and were sharpened to a wide chisel cutting edge. Few of this type of bit are used, except in places where coring is very bad. Sometimes chisel bits are used with standard bits to help cut out the core. The "pick-point" bit is widely used today. Normally it is made from ½ by I-in. steel, is about 4½ in. long when new, has a rake angle of about 25°, a front clearance of about 23O, a side clearance of approximsteji 8".' There are many variations in the shape of this bit, depending on the ideas of the individual that may be dressing the dies of the bit machine. This bit has been generally used for many years and at present, in one form or another, is by far the most popular bit used in coal mines. This bit usually is made from 0.85 carbon steel, although in recent years some alloy-steel bits have made their appearance. At present, carbon-steel bits are frequently tipped in. hard facing materials. A third class of bits might be called the L'one-use" bit; that is, a small bit that is used once and then thrown away. These bits have two or three points, are usually made from 0.85 carbon steel and sometimes from higher alloys. Recently some. of them have been tipped with hard facing material. Most of them were used at first with special

Jan 1, 1942

By McDowell Graves, Joseph Jensen, W. D. Goold, M. L. Gwin

During the present year the Santa Fe Springs field has proved to be the most important oil field under development in the United States. Its production will exceed that of any other field. Without it, production within the United States would have been almost in even balance with consumption. The field has forced to the fore and compelled, more than any other single factor, serious consideration of the need of conserving the gas resources of the State of California. It has also contributed much to the development of new drilling equipment and important features in petroleum engineering, and has required the development of a new cement for use in deep wells because of high temperatures. It is the only oil field in the country where wells 8000 ft. deep have become so common that they are no longer given special attention. The field covers a little less than 1400 acres of land. In it 196 producing deep-zone wells have already been drilled since August, 1928, and 224 wells are still drilling. More wells may be drilled. This indicates that probably more than 450 wells will be drilled in the present campaign. In hardly a year's time it has caused the outlay of nearly $50,000,000 for drilling, and within the year 1929 it will have returned practically an equal amount of revenue. In contrast to conditions during 1923, when this field was once before an important and disturbing factor, the California oil industry has not been forced to refinance to meet the situation presented by such a large production of oil. There has been practically no unit selling and no stock selling, nor any unnecessary development far beyond the limits of the field. Production is now about 275,000 bbl. per day and gives every promisc of remaining at this figure or increasing to nearly 300,000 bbl. Six months from now the field should still be capable of producing more than 200,000 bbl. per day. Obviously, the need for conserving gas that would be blown to the air, incident to the production of so large a volume of oil, represents a problem in conservation deserving of the most serious attention of the oil industry and of the State of California.

Jan 1, 1930

By Harold C. Price

Prior to Sept. 1, 1928, there had never been constructed what might be termed a long pipe line with electric-welded field joints. A year later more than 2500 miles of electric-welded lines had been completed which varied in diameter from 6 to 20 in., and in length from 50 to 700 miles. This paper discusses methods used on five lines with a total length of about 750 miles. The welding equipment used for the firing line or rolling welds, was 300-amp. generators driven with six-cylinder gasoline engines, and for the bell-hole or tie-in welds, 200-amp. generators driven by four-cylinder gasoline engines. A welding gang consisted of a foreman, 6 to 12 firing-line welders and 2 to 4 bell-hole welders, a mechanic and a tractor driver. Coupons about 1/4 in. thick were frequently cut from completed welds, which showed a cross-section of the weld and an indication of its size and penetration. For tests of tensile strength complete welds were cut out. The tensile strength of 52 welds cut from 420 miles of 12-in. line averaged 50,870 Ib. per square inch. The following results give a fair average of the number of welds made per man per day: on 157 miles of 8-in. line both firing-line and hell-hole welds averaged 15.3—the firing-line average was 21; on 39 miles of 10-in. line the firing-line average was 19.9; on 50 miles of 6-in. line, with unfavorable weather, the average was 17.7; on 424 miles of 12-in. line most of which was welded during the spring, 10.3, and with good weather the average rose to 13 welds per day. The average time for making a complete weld on 12-in. pipe 0.313 in. thick was found to be 30 min. for the rolling weld and 45 min. for tie-in welds. A comparison of costs between electric and gas welding shows that the cost of labor for electric welding is 25 per cent. more than gas, generating cost 61 per cent. less than gas, welding wire 40 per cent. less than gas, and supervision and overhead the same. Estimating 50 per cent, annual depreciation on electric welding equipment, and no depreciation

Jan 1, 1930

By William L. Fink, H. R. Freche

This is the thirteenth paper of a series from the laboratories of the Aluminurrl Company of America presenting the results of investigations of equilibrium relations in aluminum-base alloys prepared from electro-lytically refined aluminum. The ninth paper, on The Constitution of High-purity Aluminum-titanium Alloys,' presented a type of system that had not previously been encountered in aluminum alloys. Subsequent work has shown that there are several similar binary aluminum diagrams including that of aluminum and chromium. The present investigation was limited to the cxtreme alnminum end of the aluminum-chromium system. The first systematic investigation of the aluminum-chromium equilibrium diagram was carried out by Hindrichs.? The resulting equilibrium diagram with some imaginary additions suggested by Guillet3 is shown in Fig. 1. Recently Goto and D6gane4 investigated the constitution of aluminum alloys containing up to 32 per cent chromium, and presented the diagram that is reproduced in Fig. 2. Wide discrepancies in the two diagrams are evident. The following intermetallic compounds have been reported in the literature: A1Crl5,6 AlloCrlzSi,5 AlCr4,6 AlCr3,7 AlaCr.8 Preparation of Alloys and Chemical Analyses With the exception of sample Y4206, which will be described under Determination of Hyperperitectic Liquidus, the alloys were prepared from electrolytically refined alurninum% Of which the analysis is given in

Jan 1, 1933

By William L. Fink, E. H. Dix, L. A. Willey

The results of a preliminary investigation of the aluminum-iron-manganese system showed that small amounts of iron (0.10 per cent) substantially reduce the solid solubility of manganese in aluminum. This is clearly shown in the photomicrographs of three alloys, Figs. 1 to 3. Fig. 1, of an alloy containing 1.04 per cent manganese after a heat treatment at 626° C. for 45 hr., shows only solid solution. This alloy contained less than 0.008 per cent iron. Fig. 2 shows the abundance of a constituent (which undoubtedly contains aluminum, manganese and iron) in an alloy containing 1.00 per cent manganese and 0.11 percent iron after a solution heat treatment at 625" C. for 63 hr. For comparison, Fig. 3 is included to show the relatively small amount of constituent (FeA13) in a similarly heat-treated aluminum alloy containing 0.10 per cent iron but no manganese. These data necessitated a reconsideration of the binary aluminum-manganese diagram that was previously presented,' inasmuch as the alloys used in that work contained from 0.02 to 0.05 per cent iron. This paper is the fourteenth of a series from the Research Laboratories of the Aluminum Company of America reporting the results of investigations of equilibrium relations in aluminum-base alloys of high purity. The purpose was to determine the phase relations in aluminum with aluminum-manganese alloys containing the smallest amount of iron that could be realized with the purest material available at the present time (less than 0.01 per cent iron). Since the previous publication from these laboratories, five investigations of the aluminum-manganese system have been reported. Krings and Ostmannz assigned the formula MnA14 to the compound at the aluminum end of the syst,em, and found a eutectic between this compound and aluminum at 649" C. and 3 per cent manganese.

Jan 1, 1933

By H. C. Otis

During the past year or two considerable time and money have been spent in developing equipment for tubing large-volume high-pressure oil and gas wells without loss of production. That the efforts have met with success is proved by the fact that any properly cased well, regardless of its production or regardless of its rock pressure or working pressure, can now be safely tubed with no appreciable loss of production, provided experienced crews handle the new equipment. It is conceded that a far better gas-oil ratio can be secured by flowing a well through tubing of a proper diameter—the diameter of the tubing depending, of course, upon the volume of fluid and gas and the depth of the well—than by the more usual method of installing a choke at the mouth of the well. Where possible it is becoming the practice to bring wells in through tubing, though in the majority of instances this is not feasible. In many instances it has developed, after a well has been brought in through tubing, that the tubing installed is either too large or too small for the most efficient operation or that the well had not been drilled deep enough. The equipment described in this paper, in the hands of experienced men, permits the tubing of these wells after they have been brought in, with the size of tubing then determined by actual observation to be best suited for each particular well, and with no danger of the well getting out of control, of catching fire, of drilling itself deeper, or of any loss of production. The deeper the flowing production with the consequent future higher lifting costs, the more essential it is that the size of tubing used be such as to give the best possible gas-oil ratio. The ability to bring in these deep wells through the casing and then, immediately after their completion, regardless of volume or pressure, to tube them with the size of tubing that permits the most efficient use of the gas present will doubtless save many dollars. Uses for the New Equipment The equipment permits a further step in the operation of wells where the conservation of gas is of major importance. Perhaps 3-in. tubing

Jan 1, 1930

By William H. Bassett, C. J. Snyder

Cable sheaths made from common lead have been used to protect underground cables since the beginning of the power-cable industry. Only within the last few years, however, has a systematic study of the production methods and of the causes of failures in cable sheathing been undertaken. Recent increases in the use and demands for service made upon power cables have emphasized the importance of high-quality cable sheaths. It was customary for the users to attribute the cause for their sheath failures to the lack of purity of the lead, but this contention has not been borne out by field and laboratory inspection of the sheaths. During the past two years the writers have had an opportunity to examine various types of sheaths that were removed from power cables after failure. These samples were obtained from various parts of the country and represent, it is supposed, the production of the principal cable manufacturers. Aside from corrosion, the common field difficulties can be classified either as mechanical or structural, as outlined below, and when structural irregularities are found they can all be attributed to manufacturing causes and in no way reflect the purity of the lead. Clasxification of Cable Sheath Failures Other Than Corrosive Mechanical Irregularities A. Splits caused by excessive internal pressure B. Injuries such as nail holes, cuts, tsars, etc. C. Intercrystalline cracks caused by vibration Structural Irregularities A. Splits C. Voids in wall of sheath a. Caused by imperfect welds a. Blisters 13. Holes through sheath b. Laminations a. Caused by yellow oxide inclu- D. Questionable structures sions in sheath wall a. Prominent flow lines b. Blowouts by entrapped gases b. Large areas filled with suboxide during manufacture c. Scattered yellow oxide or dirt C. Caused by inclusions of dirt and

Jan 1, 1933

By C. Stansfield Hitchen

The paper describes the application of the logarithmic wedge sector method of quantitative spectrography to the estimation of commonly occurring minor impurities in crude and refined tin. Both arc and spark methods of emission were employed and with the former a new device was used to overcome difficulties arising from the low melting point of the metal. The impurities investigated included antimony, arsenic, bismuth, copper, iron, lead, and zinc, their concentrations lying, in general, between 0.01 and 0.50 per cent. The results obtained show that the method may be usefully employed for the estimation of bismuth, copper, iron, lead, and zinc, in amounts ranging from 0.50 per cent to well below 0.01 per cent. Arsenic and antimony, however, could not be estimated in amounts less than 0.03 and 0.07 per cent respectively. Out of a total of 116 estimations by arc and spark emission, in 96 the errors were less than 5 per cent of the total quantity of the element present, whereas in the remaining twenty, the errors lay between 5 and 25 per cent. Spectro-analytical data for the several impurities studied are given in the form of tables, and curves correlating line-intensity with concentration have also been included. DISCUSSION (R. G. Waltenberg presiding) D. M. Smith, London, England (written discussion).—I am interested in Dr Hitchen's account of his investigations on the application of spectrographic methods to the assay of tin and congratulate him on his valuable contribution to the literature on the subject of quantitative spectrum analysis. Some of the results obtained in the course of investigations carried out for the British Non-Ferrous Metals Research Assn., although obtained under somewhat different experimental conditions, may be of interest in connection with this paper. Arc and spark spectra of standard samples of tin containing various known amounts of antimony, bismuth, cadmium, copper, lead, silver and arsenic have been investi-

Jan 1, 1933

By Harry J. Morgan

One of the greatest improvements in drilling equipment has been the development of hard facing metals which are welded by electricity or by acetylene to form a facing which protects the steel of the bit from abrasive wear. The main desirable characteristics of hard facing metals are resistance to abrasive wear, resistance to shock and perfect bond with underlying metal. Resistance to abrasion offered by various metals shows many anomalies. Compounds containing carbon, silicon and boron are the most important hard substances. Carbon, as diamond, silicon carbide and especially the carbides of iron, chromium and tungsten are the outstanding hard products in common use. Tungsten carbide is probably the most commonly used. It is made in the electric furnace and it is important to get the right amount of carbon and the correct crystalline structure in order to obtain the greatest hardness and greatest resistance to abrasion. It may be obtained in three main divisions: wrought, cast and casehardened. Insert hard metals are common. The method of application varies, but it is generally welded to the bit, placed in channels or in drilled holes and covered with a hard surfacing metal. Robinson' has given a very complete article on the use of inserts. Gregg and Küttner2 has discussed the metallography of tungsten carbides and given a good bibliography of the subject,. The time and expense of large-scale testing, together with variations in results, led to the development of a laboratory testing machine for determining the abrasive-resisting qualities of different metals. This machine consists of a motor-driven 40-grain, alumina-base corundum wheel, 8 in. dia., rotating at 30 r.p.m. It is run wet and dressed with

Jan 1, 1930

By J. T. Eash, E. M. Wise

In a previous communication1 the improvement in slrength, harden-ability and color, occasioned by replacing gold in a typical gold-silver-copper alloy by various amounts of platinum, or palladium, was discussed. In the present paper the results of further studies upon the constitution and the nature of the hardening transformation in these alloys are presented. This research may be summarized as follows: 1. The liquidus and solidus have been determined for the alloys of the palladium and platinum series. Cooling curves were utilized for determining the liquidus of the palladium alloys but this method proved unsuitable for the platinum alloys so that a novel quenching method was utilized for the latter. Both the liquidus and solidus of the palladium alloys increase steadily with the palladium content. The liquidus of the platinum alloys increases rapidly with the platinum content, while the solidus temperature increases with the platinum content up to about 14 atomic per cent platinum, but suffers no further increase with additional increments of platinum up to 20 atomic per cent at least, owing to a peritectic reaction. 2. Palladium, and even more effectively platinum, reduces the rate of grain growth. Owing to this fact and to the increase in the solidus temperature, "overheating" does little or no damage to alloys containing reasonable quantities of these elements. This is confirmed both by tensile tests and metallographic examination. 3. The excess phase present in one of the high-platinum-content alloys has been isolated and chemically analyzed and was found to consist almost wholly of platinum, gold, and copper, evidently derived from the compound AuCu, by the partial replacement of gold by platinum. 4. The variation in the solubility of this platinum-content phase in the alpha solid solution with the temperature has been determined. This

Jan 1, 1933

By L. W. Kemph

The thermal expansion characteristics of aluminum alloys have been examined by several investigators, among whom may be mentioned Shakespear,' Schulz,2 Hidnert13 Hidnert and Sweeney4 and McCullough.' Shakespear compared the expansion of a series of aluminum alloys containing 14 per cent copper, with separate additions of about 1 per cent each of manganese, tin and zinc with that of pure copper. Schulz examined the expansion of aluminum and zinc and some aluminum-zinc alloys. Hidnert determined with a high degree of precision the exp:tnsion characteristics of pure aluminum, some aluminum-copper, aluminum-zinc, aluminum-silicon and some commercial alloys.' Hidnert and Sweeney examined the expansion characteristics of aluminum-beryllium alloys. McCullough, by an interferometer method, determined the expansivity of some aluminum-copper, aluminum-magnesium, aluminum-iron, aluminum-nickel, aluminum-silicon and aluminum-manganese alloys. Incidental to an investigation carried out during the past six years and leading to the development and exploit,ation of an aluminum piston alloy of relatively low thermal expansivity, the expansion characteristics of a wide variety of alloys were examined. Some of the data which might be of general interest are gathered in this paper. Experimental Procedure The apparatus used was essentially similar to those developed by the Bureau of Standards6 and Howard Scott,? with a few minor changes for expediency. The specimen was approximately 4 in. long and the differential length changes between specimen and the quartz tube were

Jan 1, 1933