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By J. E. Gill

"The Wasa Lake property is a block of about 1,640 acres in Beauchastel township, Quebec. Gold was discovered there first in 1936 in what is now known as the 'Southern area ( Fig. 1). Stripping and trenching were followed by 9,218 feet of diamond drilling in twenty holes. A number of good intersections were obtained, but their interpretation in terms of ore was difficult. A shaft was then sunk to 220 feet vertical depth. Thirteen hundred feet of lateral work and 543 feet of diamond drilling were completed at the 200-foot level before work was discontinued in 1938. Results up to this point had been just fair.Early in 1944, surface diamond drilling was resumed. The first hole cut material of ore grade in a strong shear zone 1,000 feet north of the old workings, and holes to the east and west showed this to be an extensive gold-bearing zone.In August 1945, a main working shaft was started. This has been sunk to 550 feet slope depth. A cross-cut at 390 feet vertical depth was driven to the ore zone and, by drifting and drilling at regular intervals, a length of 725 feet had been tested to December 31st, 1946 ( Figs. 1 and 2). The results check those from surface drilling quite closely.In the main zone, scoping widths from 8 to 40 feet and averaging about 25 feet are indicated. The grade after dilution should be about 0.15 oz. gold a ton and operation will probably start at 1,000 tons a day."

Jan 1, 1954

By R. B. Allen, E. E. N. Smith, B. C. MacDonald

"LocationTHE MAIN PROPERTY and townsite of Eldorado Mining & Refining, Limited, Beaverlodge Operation, is on the north shore. of Beaverlodge lake, •about seven miles north of the deserted town of Goldfields on lake Athabasca, and four hundred and ninety air-miles equidistant from Prince Albert, Saskatchewan, and Edmonton, Alberta. Both Eldorado Airlines •and Canadian Pacific Airlines maintain scheduled flights from Edmonton to the 4,000-foot airstrip on the north boundary of the Company property. Bi-weekly flights are made by Saskatchewan Government Airways from Prince Albert. Men and priority freight are flown by Company aircraft direct from Edmonton. From June to-September, when the waterways .are ice-free, freight is carried by barge from the •railhead at McMurray, Alberta, down the Athabasca river to Chipewyan, and across lake Athabasca to• Bushell, the terminal on Black bay, lake Athabasca, a total distance of 265 miles.The Eldorado Operation has been connected by a 12-mile, all-weather highway to Bushell. Where the highway crosses the park-land along the Fredette river, the townsite, Uranium, is under construction to provide services to the uranium-producing district. Spur roads have been built to the Ace, Fay, Eagle, •and Martin mines of the Eldorado operation, and to the properties of Pitch-Ore Uranium Mines, Ltd., and Rix Athabasca Uranium Mines, Ltd."

Jan 1, 1954

By Joseph J. Wallace

The 1952 Annual Report of the West Virginia Department of Mines lists seven beds that are being mined in Mingo County, of which the Cedar Grove is the most important. The Cedar Grove, Alma, and Coalburg contain the largest remaining known reserves. 2. Known measured and indicated reserves in all beds, based on a minimum thickness of 14 inches and 1,800 tons per acre-foot of coal in place, are estimated at 3,007 million short tons as of January 1, 1953. Of this total, 2,396 million short tons is in beds 28 inches and more thick. Areas in each bed were omitted from the estimate because available data relative to the bed characteristics are too meager for making an estimate that conforms with the definitions of measured and indicated coal adopted for this study. Should future drilling or development prove reserves in these areas, such reserves should be added to the total estimated reserves.

Jan 1, 1954

By R. R. Allen

[The utilization of Western coals and lignite is a continuing study of the Bureau of Mines, and storage of these fuels is of paramount importance. Since the seasonal demand for coal in the Western States triples between the summer and winter months, successful storage of these low-rank fuels would tend to level out this variation and should reduce fuel costs. The time is rapidly approaching when the West Coast States, particularly California, will annually import several million tons of coal from the Rocky Mountain States for additional power production. This will necessitate storage of large quantities of coal, and it is thefore important to develop experience in such storage. Dry surface storage, by virtue of its flexibility and low cost, is particularly adapted to the stockpiling of large quantities of coal. This report will, therefore be primarily concerned with such storage. Inasmuch as the problems of dry, open-pit storage are quite similar to those of surface stockpiling, experiences in pit storage will be discussed briefly. Although spontaneous combustion is a hazard in storing coal, experience has proved that heats can be avoided by proper stockpiling procedures. The purpose of this report is to show that the most reactive coals can be stored satisfactorily on the surface by avoiding segregation and by thorough compaction.]

Jan 1, 1954

By Geoffrey Gilbert

Rossland is in the West Kootenay district; 5 miles north of the International boundary and 4 miles west of Columbia River at Trail. The first claim in the district was located in 1887; but there was little activity until 1890; when the showings that later became important were staked and development began. Production increased rapidly from 1; 800 tons in 1894 to a peak of 360; 000 tons in 1903; and averaged 286; 000 tons for the next 13 years. It declined sharply in 1917; and continued to decline until the closing of the main mines in 1928. Between 1933 and 1941 leasers mined a further 137; 000 tons. Total output has been about 6; 200; 000 tons; with average recovery of 0.47 oz. gold; 0.6 oz. silver; and 1 per cent copper. The total value was $82; 500; 000; of which aboµt three-quarters was in gold; the camp still leads the province in total gold produced. Ninety-eight per cent of the tonnage came from a small compact area worked by four inter-connected mines : the Le Roi (39 per cent); Centre Star (25 per cent); War Eagle (24 per cent); and Josie (10 per cent); all now owned by the Consolidated Mining and Smdting Company of Canada; Limited.

Jan 1, 1954

By John C. Luck Jr.

"DEVELOPMENT of oil emulsion mud dates back to the early days of rotary drilling when oil was added to the drilling mud solely for the purpose of freeing stuck drill pipe. After the drill pipe was freed, and drilling resumed, this oil was considered a contaminant. Efforts were made to remove the oil but some of it remained and was emulsified with the drilling mud, thus forming a mechanical emulsion.During the late 1930's, operators found that adding crude oil to drilling muds seemed to alleviate 'sticky-hole' problems and so oil was allowed to remain in the mud. This practice was adopted when it was found that faster drilling rates and longer bit life were obtained. These observations led to the development of emulsion drilling muds. During the past five years their application has become more widespread because of the lower fluid losses obtained, greater lubrication qualities, faster penetration rates, and generally better hole conditions.An emulsion may be defined as a two-phase system consisting of a mixture of either finely divided oil droplets in water or of finely divided water droplets in oil. The surrounding liquid is known as the continuous phase. The emulsifiers reduce the interfacial tension, making the droplets tend to repel each other so that they remain dispersed and stabilized. These emulsions are classified as oil-in-water or as water-in-oil, depending upon whether water or oil, respectively, is the continuous phase. Briefly, close inspection shows that they consist essentially of globules of• the one liquid suspended in a continuous mass of the other liquid. Oil emulsion muds are usually of the oil-in-water type. Two basic types of emulsifiers •commonly used in Canadian muds are soaps and lignites."

Jan 1, 1954

By H. W. Wainwright

1. A schematic diagram and a description of the apparatus are presented. 2. Five runs were made using gas having a hydrogen-carbon monoxide ratio of about 2:1 and sulfur concentrations ranging from nil to 1.53 grains per 100 std. c.f. 3. At constant space velocity of synthesis gas the product of the sulfur concentration and days of operation, with conversion of H2 + CO above 97 percent, was constant. With a gas containing 0.04 grain of sulfur ac carbonyl sulfide per 100 std, cu. ft ,, the conversion of H2 + CO remained over 97 percent for 40 days. During that time the heating value of the product gas (on a C02-free basis) remained relatively constant at 940 B .t .u. per c.f., based on dry gas measured at 600 F. and 30 inches Hg.

Jan 1, 1954

By W. A. Calhoun

This investigation of low-grade bauxites as potential sources of aluminum by caustic desilication and alumina extraction was undertaken by the Bureau of Mines to develop extractive methods for winning alumina from domestic deposits of low-grade bauxites. This report is the second of a series covering the work by the Bureau of Mines' Rolla, Mo., and Bauxite, Ark., laboratories. The first report published in the Journal of Metals, December 1951,3/ and dealt with kaolin and halloysite. The present investigation is concerned with low-grade bauxites only. The purpose of these investigations is to increase the domestic supply of aluminum ores. With eventual depletion of high-grade deposits, aluminous materials, such as low-grade bauxite, clays, shales, and high-iron laterites, will become our principal domestic sources of aluminum. The development of extractive processes for winning alumina from these low-grade materials is one of the many national security measures undertaken by the Bureau of Mines. A method has-been developed in the laboratory for successfully treating high-silica bauxites from domestic deposits. In this process the bauxite was calcined at 980° C. to form amorphous silica and alumina. The calcine was then leached by percolation with 15-percent sodium hydroxide solution at 940 to 96° C., removing 67 to 77 percent of the silica with a loss of 4.1 to 5.0 percent of the alumina. The sodium hydroxide in the spent liquor used in desilication was 95 percent recoverable, and comparable desilication results were obtained with 100-percent reclaimed liquor, The indicated cost of reagents, for sodium hydroxide and lime used in the desilication and regenera¬tion of spent liquor, was about $4.00 per ton of calcine treated.

Jan 1, 1954

By C. E. Cairnes

The Slocan mining camp in Kootenay district; British Columbia; has long been noted for the great number of its silver-bearing deposits; and for the success that has; in many instances; attended their development. The orebearing area lies east of Slocan lake and mainly within a radius of 6 to 8 miles of the mining town of Sandon. Within this limited area 135 or more properties have shipped ore in quantities of from 100 tons or less to more than 500; 000 tons; and with a total value in silver; lead; and zinc ranging from a few hundred dollars to well over $10; 000; 000. Of the three metals; silver has easily ranked first in value; with lead second and zinc third. Zinc production; however; made slow progress in the early years; and with . the exception of the Lucky Jim mine; and a few others where high-grade ore could be readily hand-sorted; was a detriment to the silver-lead ores until improved methods of separation enabled it to take an important place in the mineral production of the Slocan.

Jan 1, 1954

By J. W. Byrkit

Design features and operating methods at the new Ajo smelter are described in detail. Successful operation of a novel method of handling and charging wet concentrates to a deep bath type reverberatory furnace contribute to the daily production of 200 tons of anodes with good results from the standpoint of both metallurgy and economy. THE New Cornelia Branch of the Phelps Dodge Corp. is located at Ajo, Ariz. Large scale mining operations were started at Ajo in 1917, when a 5000 ton leaching plant was put in service to treat the copper carbonate ore that overlaid the sulphide ore-body. In 1924 a 5000 ton flotation plant was built to treat the sulphide ore and the leaching operation was abandoned a few years later. Changes in practice and additions to plant facilities have resulted in a gradual increase in the milling rate which now approaches 30,000 tons daily. Prior to the erection of the Ajo smelter, flotation concentrate was shipped by rail to Douglas, Ariz, where it was treated in the Phelps Dodge smelter. Construction of a one reverberatory furnace smelter to produce an average of about 200 tons per day of copper anodes at Ajo was started early in 1949. Initial heating of the reverberatory furnace was begun on June 21, 1950. Charging the furnace was started on july 8, and the first anodes were cast on July 14. Neither the anode furnace nor the reverberatory furnace has cooled since the initial firing. Fig. 1 is a photograph of the new smelter. Built primarily to eliminate the '300 mile haul of concentrate to the Douglas plant, the new smelter was designed to treat the Ajo concentrate, with a minimum of capital investment. A novel, and relatively simple, method of concentrate handling and furnace charging made unnecessary the installation and operation of expensive storage facilities and reclaiming equipment, and obviated the necessity of holding in storage large quantities of copper-bearing materials. (Patents are pending in the United States and abroad on the Ajo process.) In the selection and arrangement of equipment, consideration was given to the full utilization of all smelting facilities, reducing the amount of idle equipment to a minimum. The layout of the smelter is shown in Fig. 2. The important problem of internal transportation was minimized by arranging the reverberatory furnace, converters, and anode furnace in a compact group and providing storage bins inside the smelter building, within reach of overhead cranes, for fluxing materials and other supplies needed in the daily operation. This can be seen in the sectional view of the smelter given in Fig. 3. Incorporated in the design were many innovations in smelting equipment intended to facilitate operations and effect economies in maintenance expense and manpower requirements. Table I gives operating data of the entire plant. Metallurgy The metallurgical practice at Ajo is based essentially upon the use of a single, deep bath reverberatory furnace for smelting wet concentrate, without concentrate storage facilities. The large reservoir of slag and matte maintained in the furnace serves to equalize, to some extent, the day to day variations in the nature and grade of concentrate and permits

Jan 1, 1954

By B. F. Decker, J. H. Keeler, W. R. Hibbard

Textures of hot-rolled, of cold-rolled, and of cold-rolled and annealed zirconium sheets were determined by use of an X-ray spectrogoniometer. All textures showed a tilt of the basal planes 240" from the rolling plane about the rolling direction. Deformation textures showed the [1010] direction to be approximately in the rolling direction. After recrystalliza-tion, the 111201 direction was approximately in the rolling direction. This texture was sharpened by annealing above the allotropic transformation. Textures of sheet produced by alternately cold rolling and then annealing were less strongly oriented than sheets that underwent severe deformation without intermediate anneals. THIS study of the deformation and annealing textures of high purity zirconium was carried out to supplement an investigation of zirconium and zirconium-base alloys. The spectrogoniometer method of pole figure determination was used and yielded more detailed and precise information than the photographic techniques used by previous investigators. Earlier investigations of the rolling textures of hexagonal metals titanium,1,2 beryllium," and zirconiumL having a c/a ratio slightly less than 1.6 disclosed a rotation of the basal planes from the rolling plane about the rolling direction so that the basal poles were inclined toward the transverse direction, and with one exception,& the [1010] direction was found parallel to the rolling direction. Similarly, hot rolling produced a spread of the basal poles in the transverse direction,2,3,5 With the [1010] direction parallel to the rolling direction. The textures of these metals after recrystallization differ somewhat from each other although still exhibiting the transverse rotation of the basal poles about the rolling direction. Zirconium" was found to have the [1120] direction in the rolling direction, beryllium" continued to have the [1010] direction in the rolling direction, and titanium' was shown to have the [1010] direction in the rolling direction (the cold-rolling texture) when annealed at 1000°F, changing to the [1120] direction in the rolling direction when annealed at 1500°F. It was suggested2 that the titanium retains the deformation texture after re-crystallization, and preferred grain growth at higher temperatures produces the second texture. Annealing above the allotropic transformation temperature did not appreciably change the high temperature recrystallization texture for titanium.' Coarsegrained zirconium showed more spread" in the transformation texture than in the recrystallization texture. The material used in this investigation was crystal-bar zirconium prepared by the iodide process. The major impurity was hafnium which was less than 0.3 pct. The composition obtained by spectro-graphic analysis is given in Table I. The hot-rolled specimen was prepared from a 50-g button of zirconium arc-melted in a water-cooled crucible under an argon atmosphere. This button was enclosed in a welded sheath of soft iron and the ensemble was rolled at 825°C. After each pass the sheath was put back into the furnace to allow it to return to the desired temperature. The zirconium ingot was reduced in this manner to a thickness of about 0.035 in., a reduction of approximately 90 pct. The sheet after removal from the sheath was alternately polished and etched to a thickness of 0.001 in. Cold-rolled specimens were prepared from crystal bar by a series of reductions of approximately 0.010 in. until the sheet reached a thickness of 0.010 in. after which it was cold-rolled to 0.001 in. thickness between sheets of alloy steel. The total cold reduction amounted to about 99.8 pct. Specimens were annealed in an evacuated Vycor or quartz tube. The transmission pole figures were obtained by the quantitative method of Decker, Asp, and Harker6 modified by Geisler7 using an X-ray spectrogoniometer with Cu Ka radiation. Unless otherwise specified all four quadrants were examined with 5" increments along the radius of the pole figure. Absorption corrections were made for changes in length of the X-ray path through the sample as inclination of the

Jan 1, 1954

By J. P. Nowlan

"This paper is an attempt to outline the structural problems at the Cochenour Willans mine, in so far as their details have been tentatively worked out. It must be emphasized that there are considerable data that can be interpreted in more than one way, and that further work may modify or completely change some of the hypotheses presented herein. Certain features, however, are sufficiently constant to warrant reliance on them when searching for extensions of orebodies.As the individual ore shoots usually have relatively small dimensions, but are often very high grade, structural details are of more than usual significance in laying out a mining programme. An illustration of this is a small shoot recently located and mined as a result of studying a minor slip cutting off a part of high grade gold ore in No. 319 stope. The shoot, which is bounded on three sides by slips, and on the fourth by the original stoping, was roughly triangular in plan, 20 feet on the base, and 40 feet in length. The average width mined was 4.5 feet, but the gold obtained totalled over 1,800 oz.The property lies in Dome township, east of McKenzie island, and consists of an original group of thirteen claims acquired from the Cochenour Willans syndicate, plus one additional group of five claims obtained by purchase from the Kelson Red Lake Gold Mines, Limited, and one other claim that rounds out the southeast corner of the block.The claims of the original group were staked by W. M. Cochenour, Dan Willans, and Horace Young, during 1926 and 1927. The Cochenour Willans Syndicate was formed in 1928, and until incorporation of the present Company work was carried on at varying intervals by several different interests, including Ventures Limited, Hollinger Consolidated, who sank a shaft to open up the 150-foot level, and by the Syndicate itself. The first ore from the property was shipped to the Gold Eagle mill by the present Company in 1939, and amounted to 4,959 tons, from which 2,360 oz. gold was recovered. The first brick was poured in December, 1939, and production has been continuous since that date. To the end of June, 1946, gold having a value of $5,982,000 has been produced from 345,033 tons of ore treated."

Jan 1, 1954

By W. A. Haley

This is the second of a series of publications that describes the progress of mechanical longwall coal mining in the United States in which a German coal planer was employed.3/ Details of the equipment, a description of its operation, and the results obtained in mining the experimental panel were published in the preceding report. Experimental mining with the planer in the first panel showed an improvement in performance over conventional mining methods at the coal company's Stotesbury 11 mine. The work, which was begun on an experimental basis, soon proved successful, and the coal company purchased the equipment for continued use. Two panels, adjacent to the first, were developed and mined with the planer. Each of these panels was shorter than the preceding one because water seepage from a surface creek limited development. On completion of mining in the third panel, the planer equipment was moved by the cooperating company to a nearby mine operated in the same coal bed where a large area of coal suitable for planer mining is available.

Jan 1, 1954

By Franc R. Joubin

"Gold has been found in many parts of the Bridge River district of British Columbia; but it is from a relatively small area in the valley of Cadwallader creek that most production has come. The ore-bearing area measures 1/2 mile in width by 3 miles in length; and for convenience it is called the Cadwallader gold belt in this paper. In this belt are found the Bralorne and Pioneer mines and in addition several prospects under development.The gold of the Cadwallader gold belt occurs in quartz veins averaging about 3 feet in width and containing about ½ oz. of gold a ton. Except for a relatively small amount of silver, no other metals are produced from these veins. The ore-bearing veins occur in a variety of rocks. They occupy reverse faults and tension fractures belonging to a distinctive fracture system. Some of these structures are ore-bearing over lengths exceeding 5,000 feet and depths exceeding 3,700 feet. The production of the Bralorne mine (1932-1945) is 996,383 oz. gold and 190,709 oz. silver; that of the Pioneer mine (1930-1945) is 777,588 oz. gold and 135,698 oz. silver. These two mines have together distributed about $25,000,000 in dividends.The Bridge River district first attracted attention in 1863* when prospectors were attracted by its small but rich placer gold deposits. Most of the important lode deposits of the Cadwallader gold belt were discovered in 1897 (11). Many years of desultory development followed and it was not until 1928 and 1932 that the Pioneer and Bralorne mines, respectivdy, entered into profitable production."

Jan 1, 1954

Small scale' table concentration tests have many critics. Many metallurgists consider that such tests are of problematical value because of the difficulties involved in conducting and interpreting them.Many kinds of small-scale ore dressing tests are difficult to conduct, and there is, perhaps, good reason for thinking that table concentration tests are amongst the most difficuIt.Interpretation of results from small-scale tests is the responsibility of the metallurgists and engineers in charge, and it is often held that small-scale table concentration tests are particularly difficmlt to interpret.TESTING FOR TABLE CONCENTRATIONOn analysis it appears that the difficulties encountered in conducting small-scale table concentration tests are of three fairly distinct kinds. Firstly, there are difficulties due inherently to the smallscale nature of the operations; for example the smaller width of all mineral bands on the table and the less complete separation due to the shorter length of travel between the feed and discharge points.Secondly, there are the effects of batch operation owing to the fact that the mineral particles behave differently during the initial period when the sample is just beginning to spread over the table, the middle period when feed and discharge are even and continuous, and the final stage, when the last of the sample has been added and the table is beginning to empty itself.

Jan 1, 1954

By H. G. Iverson

As a part of the program of investigation of domestic sources of manganese ore, mining engineers of the Bureau of Mines examined and sampled three properties in the vicinity of Winnemucca, Nev. As the ores were too low grade for direct production of ferromanganese, they were tested for their amenability to concentration by ore-dressing methods. Two samples of ore were procured by examining engineers of the Bureau of Mines from several outcrops in the Pollard Canyon area, Pershing County, about 30 miles southeast of Winnemucca. The manganese occurs in relatively small lenticular bodies as secondary replacement deposits in chert and quartzite. The first sample represented average run -of -mine ore as exposed in the Pollard Canyon deposits. Sample 2 was composed of hand-sorted, high-grade ore.

Jan 1, 1954

By R. G. Dressler

The design and operation of the Koppers powdered-coal gasifier and its auxiliaries at Louisiana, Mo., are described. This unit was designed to convert 1 ton of coal per hour with oxygen and superheated steam into synthesis gas - primarily carbon monoxide and hydrogen. The first run was made on May 4, 1949. After this, 44 short runs (less than 8 hours each) and 1 long run (lasting 11 days) were made. Total gasification time was 420 hours. The gasifier was not cooled down during the year when these runs were made. After an initial period of test operations and equipment proving, during which 12 runs were made, the effect of numerous variables on gasifier performance was studied. The long run was made primarily to check the equipment in sustained operation. Detailed data from these runs are presented.

Jan 1, 1954

By P. W. Gilles, B. D. Pollock

THE pioneering work of Steinitz1 and Steinitz, Binder, and Moskowitz2 has shown conclusively the existence at high temperature of two additional phases in the molybdenum-boron system and thus brings to a total of six the number of structures appearing in this system. To the structures Mo2B, MOB, and Mo2B5 they have added MO3B2, a new -MOB form, and have shown that MOB,, which has the same range of composition as Mo2B5, is only a high temperature structure of the latter. This solid solution, interestingly enough, includes neither of the compositions corresponding to the stoichiomet-ric compounds, MOB, or Mo2B5, but rather at all temperatures has intermediate values of composition. These workers have also, in the course of their work, measured melting points, transition temperatures, eutectic and peritectic points in the system and have shown that Mo3B2, because of its dispro-portionation at low temperature to Mo2B and MOB, is stable only in a limited high temperature range. During the course of the present work on the vaporization properties of the molybdenum-boron compounds, a few transition temperatures were observed. When the report of the other workers appeared, it was decided to repeat, in part, their study of the system. As a result, considerable evidence has been obtained that substantiates the specific kinds of melting processes they report as well as the general features of their diagram. However, a marked difference was found between the temperatures they report and the ones observed in this study, with the latter being higher. The purposes of this paper are to present the evidence obtained in this laboratory that verifies their diagram of the system, to give some important temperatures in the system, to compare them with those previously published, and to seek an explanation of the difference. Samples The metal starting material was 400 mesh molybdenum powder with a purity stated by the manufacturer to be 99.9 pct. The initial treatment, designed to remove volatile contamination, consisted of heating in a vacuum for 10 min to a temperature of from 800" to 1000°C during which a loss of 0.3 to 0.4 pct occurred. An assay following this treatment showed it to be 99.4 pct pure, with the principal impurity probably being oxygen. The boron starting material was obtained from the Cooper Metallurgical Laboratories and the Fair-mount Chemical Co. as 325 mesh powder with manufacturers' analyses of 99 pct or better. Initial treatment consisted of heating in molybdenum in a vacuum at about 1700°C for 10 min. During this time a loss of 3.5 pct occurred. An assay following this treatment showed the different samples to have purities ranging from 95.5 to 99.0 pct with iron and carbon as the principal impurities. Following the initial treatment, the elements were combined to form stocks of Mo2B and MOB by heating pressed mixtures in a vacuum to 1100" to 1200°C to accomplish reaction and to 1500" to 1900°C for a few minutes to evaporate the more volatile impurities. Analysis of the two compounds for boron by a modification of the method of Blumenthal3 and for molybdenum by the lead-molybdate method indicated them to have purities greater than 99 pct. The individual samples to be studied had compositions in the Mo2B-MOB range and consisted of mixtures of the stock compounds. Procedure As is usually the case in high temperature work the selection of containers for the samples posed some problems. For vapor pressure studies tantalum crucibles, allowing little contact with the pressed samples, were used and some of the observations made during these experiments are pertinent to the study of the phase diagram. Most of the experiments, however, were performed in graphite containers, as were those of the previous authors. Two kinds of spectroscopic grade graphite crucibles were used. One was a % in. cylinder, 3/4 in. high, containing seven 3/16 in. holes drilled 1/2 in. deep into which were packed samples of the different mixtures weighing 250 to 500 milligrams. The other, consisting of separate crucibles, was prepared by drilling 3/16 in. holes, 1/2 in. deep into 1/4 in. graphite rods % in. long. The 7/8 in. cylinder was heated directly by induction while the small crucibles were packed in a tantalum heating element for induction heating. All heating was done in a high vacuum system in which the pressure was generally less than 1x10-5mm and never rose above 2x10-5mm when the samples were hot. The general pattern of the heating in graphite was to heat rapidly to a temperature somewhat below the desired one, then to raise the temperature slowly. The samples were held for 2 to 5 min at the maximum temperature, which in all cases was far higher than that needed to produce reaction. The short time was employed to reduce possible contamination by the crucible material and to reduce composition changes that would occur because of vaporization. After examination following the heating, the samples were reheated to a higher temperature. Temperatures were measured with a Leeds and Northrup disappearing filament optical pyrometer, certified by the National Bureau of Standards, by sighting through a window at the top of the vacuum

Jan 1, 1954

By C. W. Sill

[In the original Norris and Fay method for determining selenium, selenious acid is reduced by an excess of standard thicsulfate in accordance with the following reaction: H2Se03 + 4Na2S203 + 4HC1 = Na2SeS406 + Na2S406 + 4NaC1 + X20. (1) The excess thiosutfate is then back-titrated with standard iodine solution. The disadvantages of ouch a method are obvious. On an unknown sample, the quantity of thiosutfate required would be difficult to determine; a definite excess is required, but large back titrations would be undesirable. Inasmuch as thiosulfate is decomposed by acid, considerable error may be introduced in this method, the amount depending on the acidity, temperature, concentration of excess thiosulfate, duration of contact between thiosulfate and acid, and other factors, In a critical study of the Norris and Fay method, Coleman and McCrosky3/ show that the method is accurate to less than .1 part per thousand if the excess of this sulfate Is not greater than 5 ml. of 0.1 N solution (in 1.0 ml. of 0.x4 N sulfuric acid) whether the titration is made at 0° or 20°C. A greater excess produced unsatisfactory values at either temperature. Coleman and McCrosky regulated the excess thiosulfate by adding 5-ml. portions of thiosulfate solution followed by the addition of 1 to 3 drops of 2.5 percent potassium iodide in the presence of starch. If a starch-iodine blue color formed locally around the drops of potassium Iodide, the reduction of selenious acid was not complete. On stirring the solution, the local blue color fades and another 5-ml. portion of thiosulfate is added. When addition of potassium iodide fails to produce a local blue color, reduction of selenious acid is complete, and the excess thiosulfate is then back-titrated with standard Iodine solution. The number of equivalents involved in the reduction of selenious acid by either thiosulfate or iodide is the same, and therefore no correction need be made for the potassium iodide. This procedure is much better than the original Norris-Fay titration, but acidity and time would stilt be rather critical factors subject to variation with individual technique. Also, the method still requires the lose of two standard solutions, a rather undesirable factor from the standpoint of routine application.]

Jan 1, 1954

By L. Delisle

This work represents the first step of an attempt to test the applicability of the electron microscope to the study of subgrain structures in copper. Observations on annealed and deformed single crystals and polycrystalline samples of copper are described. IN the course of study of the structure of fine tungsten wires and tungsten rods with the electron microscope, well defined subgrain structures were observed. The size, size distribution, and orientation uniformity of the etch figures varied widely in different samples. Figs. 1 and 2, electron micrographs of a tungsten wire and of a tungsten rod, respectively, are illustrations of the difference in size and size distribution of the etch figures in different samples of the same metal. The observed differences, as pointed out in a previous paper,' appeared to be related to the heat and mechanical treatments of the samples. They were also consistent with the results reported in the literature on the mosaic structure of metals.' For that reason a program of research was initiated in an effort to obtain more systematic evidence of the possible relation of heat and mechanical treatments to the subgrain structure of metals as observed in the electron microscope. The purpose of this paper is to present observations made on the effect of cold work on the subgrain structure of copper. Procedure Starting Materials: Copper was the metal studied because it can be obtained in a high degree of purity, much information is available in the literature on its properties and its response to cold work and heat treatment, it shows no allotropic change, and it is sufficiently hard to be handled without great difficulty. Two groups of specimens were used: 1—single crystals cast from spectroscopically pure copper and 2—polycrystalline samples of oxygen-free high conductivity copper. Single crystals were studied because it was hoped that the elimination of a number of variables, such as grain boundaries, orientation differences, degree of purity, would simplify the problem and perhaps permit a better understanding of the phenomena that would be observed. The polycrystalline samples were designed to give a general picture of the changes considered. The single crystals were made of copper which analyzed spectroscopically to better than 99.999 pct Cu. They were cast in vacuum, by the Bridgman method, in crucibles made of graphite with a maximum ash content of 0.06 pct. The mold design is shown in Fig. 3. It permitted casting crystals of the size and shape required for the experiments, so that the danger of introducing cold work in the original samples by cutting or other machining would be eliminated. The polycrystalline samples were pieces, 3/4 in. long, cut from a rod of oxygen-free high conductivity copper, % in. in diameter. A flat surface, 1/4 in. wide, was milled along the rods, polished, and etched. The samples were then annealed in vacuum at 850°C for 1 hr. Polishing and Etching: Work previously done on tungsten,' polished mechanically and etched chemically," had shown that: 1—the general appearance of the etch figures of a given sample was not altered by repeated polishings and etchings under similar conditions; 2—variations in the time of etching and the concentration of the etchant changed the definition of the etch figures, but did not alter their general size nor orientation distribution within the limits of observation. Further work confirmed the reproducibility of the subgrain structures observed in, 1—single crystals and polycrystalline samples of copper when polishing and etching were repeated under similar conditions, and 2—specimens of tungsten and polycrystalline copper when electrolytic polishing and etching were substituted for mechanical polishing and chemical etching, respectively. On the strength of these observations, it was felt that, if conditions of polishing and etching were kept constant, changes observed in the subgrain structure of a sample upon deformation and annealing would be attributable to such treatments. For that reason the conditions of polishing and etching were kept as constant as possible. The single crystals were polished electrolytically in a bath of orthophosphoric acid in water, in the ratio of 1000 g of acid of density 1.75 g per cc to 1000 cc of solution, under a potential drop of 1.6 to 1.8 V. Electrolytic polishing was selected to prevent the formation of distorted metal in polishing. The same samples were etched by immersion in a 10 pct aque-

Jan 1, 1954