Search Documents

By E. J. Gleim

The construction of explosion-proof electric mine equipment for operation in gassy coal mines of the United States was begun as early as 1903. In 1.906, similar equipment with special protective devices was built for export. An investigation of explosion-proof motors was conducted by the United States Bureau of Mines during 1910 and 1911. In other words, this was among the first efforts of the Bureau directed toward the saving of life and property subsequent to its establishment by act of Congress in 1910. In these early motors, those with special devices for relief of explosion pressures predominated. At that time it was generally considered that, by providing for pressure relief, the enclosures could be made lighter in weight because of the lower explosion pressures developed within.

Jan 1, 1954

By W. P. Corking

"The purpose of this paper is to describe the relationship of the goldbearing ore to structure at the Pickle Crow mfoe. Information beyond this scope may be found in the reports and papers of Hurst( l), Bothwell(2), Thomson(3) , and Monette( 4).The property lies on Crow river, in Patricia district, 80 miles northeast of Allanwater, and about 125 miles northeast of Sioux Lookout on the Canadian National railway. From April of 1935, when the mill began operation, to the end of 1944, a total of 564,890 oz. gold and 49,275 oz. silver were recovered from 958,416 tons of ore.Mining on the property is at present carried on in two separate orebodies, the Howell and the No. 2 vein. Stoping has been carried on from every level on the Howell vein to a depth ot 2,450 feet below the surface and from several levels on the No. 2 vein through a mile long cross-cut on the 750-foot level to the Howell vein workings. A steadily increasing volume of ore is being obtained from the No. 2 vein.Each of the two orebodies consists of vein quartz with small amounts of other gangue minerals either in a single vein or a stockwork of veins. Both lie within sheared walls and carry gold in the free state. Fig. 1 shows the position of the veins with respect to the surrounding rocks. It will be noted that the delineation of the beds of iron formation shows that the lava flows and minor, interbedded sediments that make up most of the country rocks are areally folded and contorted.STRUCTURENearly all gold found to date in sufficient quantity and concentration to make ore in the Pickle Crow area is confined to these major quartz veins . Both strike a little north of east and dip north, and in both cases the attitude of the enclosing rocks is somewhat more north of east in strike and more steeply north in dip than that of the veins. As a result of this disposition both veins migrate across their enclosing rocks at a moderate angle in both dip and strike. (In this connection it should be noted that No. 2 vein dips at a considerably flatter angle than its host rock, but has a very pronounced rake to the northeast, effecting the same result as in the Howell vein.)"

Jan 1, 1954

By R. I. Jaffee, F. C. Holden, H. R. Ogden

ECENT papers dealing with the properties of unalloyed iodide titanium have been directed primarily at the determination of base-line properties for alloy investigations. Early work was limited to a few tests because of the limited availability of iodide titanium at the time. In the results of papers by Campbell et al.,1 Gonser and Litton,2 Jaffee and Campbell,3 inlay and Snyder,4 and Jaffee, Ogden, and Maykuth, data on mechanical properties are presented for unalloyed iodide titanium in the annealed and cold-worked conditions. Data are presented in this paper which show the effects of heat treatment on the structure and mechanical properties of commercially produced iodide titanium. Correlation is made between microstruc-tural variables and the mechanical properties. Experimental Procedures Melting Stock: The melting stock used was as-deposited iodide titanium, produced by New Jersey Zinc Co. The furnished analysis showed the following range of impurities: N, 0.004 to 0.008 pct; Mn, 0.005 to 0.013; Fe, 0.0035 to 0.025; Al, 0.013 to 0.015; Mo, 0.0015; Pb, 0.0045 to 0.0065; Cu, 0.0015 to 0.002; Sn, 0.001 to 0.01; Mg, 0.0015 to 0.002; and Ni, 0.003. Hydrogen content as determined by vacuum-fusion analysis was 0.0091 wt pct (0.44 atomic pct) after arc melting and fabrication. Nitrogen analysis on the arc-melted and fabricated titanium showed a content of less than 0.002 pct N. The average hardness of the furnished stock was Rf 70, or approximately 85 VHN. Melting Procedure: The as-deposited rods were rolled, sheared, and degreased in preparation for arc melting. The charge was arc-melted with a tungsten electrode in a water-cooled copper crucible under a positive pressure of high purity (99.96 pct) argon. The final ingot was approximately 2 in. in diameter and showed no increase in hardness over that of the initial stock. Fabrication: Heating for fabrication was done in air. It was begun by forging the ingot into a 3/4 in. diam rod, at an initial temperature of 1600°F. Scale was removed by sandblasting. The rod was then swaged to 1/4 in. diam at room temperature through a series of 20 dies, with approximately 10 pct reduction in area between each die. An anneal of 1 hr at 850 °C in air was given after the 1/2 in. die, such that the final cold reduction was 75 pct. Sections cut from this rod were used for test and microstructure specimens. Heat Treatment: Heat treatments were carried out in resistance tube furnaces with stainless-steel linings, under an atmosphere of gettered argon. As further protection against contamination, the specimens were packed in titanium turnings in a titanium sleeve. Control experiments have shown negligible hardness increases with this method, indicating that contamination from oxygen and nitrogen is slight. Three cooling rates were employed in this work; these have been designated as water quenching, argon cooling (to simulate air cooling under a controlled atmosphere), and furnace cooling. The cooling rate for an argon cool is 100°C per min for the first minute, with an average cooling rate of 35°C per min over a 15-min period. A furnace cool requires about 10 hr, with an average cooling rate of 3.6oC per min during the first hour, and an average cooling rate of 1.2°C per min over the 10-hr period. Microimpact Test: The specimen adopted was based on the cylindrical Izod Type Y specimen (ASTM, E23-41T). All dimensions were reduced to half scale, including the notch radius. Specifications are shown in Fig. 1. The specimen is held vertically in an adapter and broken as a cantilever beam. Impact tests were run on a constant-velocity (11.34 ft per sec) Tinius Olsen impact testing machine with a total available energy of 100 in.-lb. Tests were made to determine the correlation between this microimpact and the standard V-notch Charpy impact test. Curves showing impact energy as a function of temperature for both impact tests are plotted in Fig. 2. Transition temperatures, when they occur, are about the same for both impact tests. All three titanium-rich materials have the same conversion factor, 10. Tensile Testing: Tensile tests were conducted on Baldwin-Southwark testing machines using the 600, 2400, or 3000 1b range. Specifications for the test specimen were taken from the 1948 edition of the ASM Metals Handbook, and are shown in Fig. 1. Strain measurements were made using an SR-4 resistance gage (Type A-7) cemented to the reduced section in conjunction with a lever-type extenso-meter. Readings on the SR-4 strain indicator were

Jan 1, 1954

By H. C. Hamontre

The liquid-level gage described in this report is an instrument designed to locate liquid levels in wells rapidly and accurately and, with auxiliary equipment described, to follow moving liquid levels automatically. The depth at which the instrument makes contact with the liquid in a well is indicated at the surface by a depth meter on the cable to which the instrument is attached. When the gage is used to follow a surging moving level, the liquid depth is indicated continuously at the surface. This report contains a description and illustrations of the gage and outlines the operating procedures. The instrument is 1-3/4 inches in diameter and approximately 6 feet in length and may be operated through 2-inch tubing. The gage consists essentially of a float that moves upward with respect to the rest of the assembly, causing a change in the flow of electric current from the surface when the liquid level is reached. The instrument has been tested under a variety of conditions and has been found to be sensitive to changes in liquid level of 1/2 inch. Its accuracy is limited principally by the accuracy of the wire-line depth indicator used. It is not affected by pressure or temperature, and it can be run in open hole or under pressure through a lubricator. It employs only one electric circuit and can be run on an Amergraph single-conductor cable (type 1-H-0) ,

Jan 1, 1954

By K. F. Anderson

The increasing number of oil-productive reservoirs in the Ellenburger formation has emphasized the need for information pertaining to the hydrocarbon recoveries that may be obtained from such reservoirs. Most of the space available for accumulating hydrocarbons in the Ellenburger formation is derived from fractures, which makes it difficult to estimate reserves or probable recoveries. The Ellenburger reservoir in the Big Lake field was selected for study because it is one of the few near-depleted reservoirs in the area. The experience gained in operating and developing this reservoir has been and should continue to be a helpful guide to operators both in predicting recovery and establishing methods of development and operation In similar reservoirs. Development and production data were gathered from different sources. The writers have compiled and interpreted these data to estimate the ultimate recovery of hydrocarbons. In addition to the estimates of ultimate recovery, the report contains sample calculations and estimates of theoretical recovery based on different methods of operation. The calculations presented in this report indicate that the Ellenburger reservoir is 95 percent depleted and that the ultimate recovery of hydrocarbon liquid from the oil reservoir and gas cap combined will he 4-0 barrels per gross acre-foot of reservoir rock. An estimated 32 barrels of condensate per gross acre-foot of reservoir rock in the gas cap has been recovered, and the estimated ultimate recovery of crude oil will be 47 barrels per gross acre-.foot of rock in the crude-oil reservoir.

Jan 1, 1954

By Irving Hartmann

The record of safety in the use of permissible explosives in United States coal mines has been excellent, due in great measure to the persistence with which research has been carried on in the Bureau of Mines, to the development of permissible explosives, and to the cooperation of mine operators and workers in using permissibles in a recommended manner and in the effective though gradual replacement of black powder and other nonpermissibles. The demand for increased rate of coal production, changes in mining conditions and minim methods, and introduction of new mining machinery have resulted in modified or new blasting practices. To keep abreast of these developments the Bureau of Mines conducts investigations under actual mining conditions at the Experimental coal mine near Pittsburgh, Pa. Several years ago an extensive study was complete to determine conditions under which the allowable charge limit for permissible explosives per shot hole might be raised from 1.5 to 3 pounds without undue risk of gas or coal-dust ignitions. As a result of the study, this increase in the charge limit was authorized under certain specified conditions in 1948,5/ greatly facilitating production schedules.

Jan 1, 1954

By M. H. MacLeod

"IntroductionTHE ORE PASS raise system at the H.B. mine as contracted by the Patrick Harrison Company consisted of two raises 6 ft. by 10 ft. in section •at 60 degrees to the horizontal. No. 88-88 raise, the longer of the two, was a straight raise 690 feet long, with access only from the bottom. In addition, this raise had two branch knuckle-backs only one of which was driven by. the Patrick Harrison Company. The other, known as 88-89 raise, was 540 feet long with knuckle-backs at 280 feet and 410 feet. Its access, also, was from the bottom (see Figure 1)."

Jan 1, 1954

By W. J. Taylor, H. J. Mutz, A. F. Brock

THE International Kickel Co. of Canada Ltd. operates five underground mines and an open pit. Four of the mines, the Frood-Stobie, Creighton, Murray, and Garson, are on the south range of the Sudbury Basin, map above, within a radius of 10 miles of the city of Sudbury, Ontario. The fifth mine, the Levack, is on the north range of the Sudbury Basin, 30 miles to the northwest of the city. The open pit is operating on surface ores of the Frood-Stobie orebody. Two pits were opened at the Frood-Stobie mine; the Frood section was started in 1938 and the Stobie section in 1942. To meet the heavy demand for nickel during World War 11, open-pit production was accelerated until it constituted over 40 pct of Inco's total tonnage. As a consequence open-pit ore was depleted considerably alhead of schedule; the Stobie pit was completed in the summer of 1951, and it is expected that mining at the Frood pit will be completed in 1953. The major program launched 10 years ago to develop additional underground production as a replacement of open-pit tonnage thus had to be greatly speeded up. This conversion to all-underground mining will be completed within the next two years, and the underground mines will then have an annual hoisting capacity of 13,000,000 tons. One of the new mines in the underground expansion program, the Murray, was brought into production in 1950. Ore production from the new Stobie section of the Frood-Stobie mine was commenced on a limited scale in 1951 and will be increased as the program progresses. Delivery of ore to the new Creighton mill was started in the summer of 1951 from the caving project launched at Creighton mine. When mining was started in the Sudbury district 65 years ago the ore deposits were opened on a small scale and mining methods conformed to a general pattern of small open cast pits, followed by open stopes as mining was extended below the pits. The open stoping methods used were heading-and-bench and shrinkage. In certain cases a type of cut-and-fill mining was used with drywall tramways and dry-wall chutes in the fill. As the mines were deepened,

Jan 1, 1954

By J. F. Henderson, N. H. . C. Fraser

"Camlaren Mines, Limited, owns a group of claims near Gordon lake, in the Northwest Territories about 45 miles northeast of Yellowknife. The Hump vein on the property was explored continuously from the summer of 1937 to December, 1938, when all work was stopped. The two-compartment vertical shaft sunk near the vein is 380 feet deep and drifts and cross-cuts on the 200- and 350-foot levels total 2,241 feet . This exploration indicated 13,177 tons of ore with a gold content of between ½ and 1 ounce a ton1GEOLOGY AND STRUCTUREThe property is underlain by remarkably fresh greywacke and slate of the Early Precambrian Yellowknife group. The thickness of the b ~ds averages 1 to 2 feet, although that of the coarser greywacke may be 10 feet or more. Most of the greywacke beds show a gradation in grain from sandy at the base to slaty at the top. They are composed of chlorite, sericite; carbonate, and detrital quartz and feldspar grains.The structure of the greywacke and slate underlying a part of the property of Camlaren Mines, Limited, is shown in Fig. 1. The beds are isoclinally folded so that they dip at steep to vertical angles and at many places are overturned. The axes of the folds are 50 to several hundred feet apart, and generally plunge at angles of 45° or steeper. The beds along the axes of the folds generally meet to form a sharp V, and at the point of the V, or axial part of the fold, the beds are commonly ruptured . Minor faulting movements along the broken crests of the folds in many cases result in well-defined shear zones. Discontinuous quartz veins and irregular lens-like masses of quartz lying more or less parallel with the bedding are plentiful throughout the sediments, but tend to be concentrated along the ruptured and sheared, anticlinal and synclinal fold axes ( Fig. 1)."

Jan 1, 1954

By C. E. Schwab

A lead-zinc orebody, in fairly strong quartzite and with a dip of 35" to 60°, is block-caved by use of scrams in a stair-step pattern up the ore footwall. Scram linings to handle coarse muck and permit the use of folding scrapers are developed by the use of end-grain wooden blocks to reduce maintenance and keep operating cost to a minimum. THE Bunker Hill mine, since its discovery in 1885, has steadily produced a high grade of lead-silver-zinc ore. By the end of 1952 over 21,000,000 tons of this high-grade ore had been produced by square-set mining, and reserves in the mine continue to be very satisfactory both as to quantity and grade. For many years prior to 1941, mine production and mill capacity had been 1200 tons of feed per day. Closely adjacent to the mill, and stored behind dikes, coarse jig tailings had been impounded during the time preceding the advent of fine grinding and selective flotation. When manpower became short in 1941 and sink-and-float preconcentration was proved successful, mill capacity was increased to 1800 tons per day to treat these jig tailings economically. By 1946, because the supply of jig tailings was limited, underground exploration was started to discover and prove ore reserves of low-grade material which could be mined by an appropriate bulk mining method. During the years of square-set mining many possible areas of low-grade mineralization had been observed. One chosen for the first exploration work was sufficiently remote from active mining areas so that subsidence, if an ore-body were proved, would cause no problem. Also, old adits and workings were still open and in good enough condition so that exploration in the mineralized zone could be started with a minimum of preparatory work. In 1948 an orebody was proved of sufficient tonnage, of a grade about 2 pct Zn, 0.5 oz Ag, and 1.0 pct Pb. It was decided to use block-caving, the only appropriate mining method by which this grade of ore could be economically recovered. Exploration for additional reserves in other areas of the mine is continuing, but ultimate results are not known at this time. With more sink-and-float capacity, larger ball mills, and more flotation machines, mill capacity was increased to 3000 tons per day, permitting the mining of ore in the square-set area at a maximum rate not usually achieved, because of the scarcity of labor. Increased mill capacity also permits block caving and the mining of jig tailings at variable rates to keep mill feed up to 3000 tons per day. Fortunately the three types of feed are amenable to the same mill circuit and reagents for recovery of Pb and Zn. For example, during the first 10 months of 1952 square sets produced 827 tons per day, block-caving 1421 tons per day, and jig tailings 643 tons per day, an average daily production of 2891 tons for all three products. Exploration had proved the existence of an ore-body 1000 ft long and 165 ft wide in horizontal section, see Fig. 1. Company engineers were concerned only with the vertical extension, about 300 ft, from an old level to the surface. Much of this almost outcropped, Fig. 2. The ore lies in the hanging wall of a major fault of the Bunker Hill mine, standing at 65" in one end of the zone and separated from the fault by a wedge of waste, see Fig. 3. This wedge pinches out near the center of the zone, at which point the ore dips 45", lying nearly on the fault, Fig. 4. The remaining portion lies on the fault and conforms to the fault dip of 35", Fig. 5. Open-pit mining for the top of the ore was considered, but since the ore zone dipped into and under the mountains, adverse waste-to-ore ratios precluded use of this method. The ore occurs in massive quartzite of sufficient strength to support untimbered drifts, crosscuts, and raises. Zones of weakness in the quartzite are bedding, jointing, and small faults or slips. The mineralization, which occurs as small stringers of sphalerite and galena as well as pyrite, creates another line of weakness. The mineral veins or veinlets in themselves are high-grade. Their size and regularity and the amount of barren quartzite by which they are separated determined the limits of mineable ore, which are all assay limits except for the one determined by the major fault. Block 1 Without any background of caving in this type of quartzite, engineers selected the first block on the very steep end of the zone. Compelling reasons prompted this decision. The steep portion of the ore in Block 1 was of the lowest grade, so that if difficulties were encountered no very valuable ore would be lost, while the experience gained might be applied in mining the remaining blocks. A block 200x200 ft was laid out, with four scrams spaced 50 ft apart for drawing and placed at a right angle to the strike. Finger raises were placed in a 25-ft interval grid pattern, with flat undercutting done by crosscuts at the undercut level 25 ft above

Jan 1, 1954

By A. F. Banfield

"IntroductionELEMENTAL sulphur occurs in three types of deposits: in salt domes in the Gulf Coast region of the United States and Mexico, from which over 90 per cent of world native-sulphur production is obtained; in sedimentary deposits such as those in Sicily, which are thought by some authorities to contain the largest world reserves of elemental sulphur; and as surface deposits, associated now or in the past with some phase of vulcanism, the type which bas the greatest world-wide distribution.Deposits of the volcanic type are the subject of this paper.From antiquity until the successful development of the sulphur-bearing salt domes at the beginning of this century, important sulphur production for local needs was obtained from volcanic deposits in many parts of the world. Due to technical and economic factors, however, producers could not compete in price with sulphur from salt domes. •The result has been that, except in Japan and Chile, little sulphur was produced from volcanic deposits from about 1910 until the start of the Korean war in 1950. Shortly after the outbreak of the Korean war, an acute world sulphur shortage developed and prices skyrocketed in many countries up to several hundred dollars a ton. Interest in volcanic deposits was renewed and exploitation of many was undertaken. Few of these operations have been successful, mostly due to inability of management to solve technical problems.While important volcanic deposits of elemental sulphur have not been reported ( 1) in Canada, and it is doubtful that any will be found, data presented in this paper may be of interest to students of ore genesis.This paper is based on field examination of numerous deposits of the volcanic type in North and South America."

Jan 1, 1954

By Garth M. Crosby, F. McIntosh Galbraith, Bronson Stringham

THE Hercules mine is located in the northeastern section of the Coeur d'Alene district, approximately 1 1/2 miles north of the town of Burke, Idaho. Surface indications of the ore deposit were first discovered in 1886, but regular mine production was not started until 1902 and was continuous until April 1925, when the known ore had been extracted. Incomplete records show that from 1912 until operations were suspended the mine produced 2 1/2 million tons of ore containing 9.4 pct lead and 7.7 oz of silver per ton, together with an estimated 2 pct zinc, 0.3 pct copper, and 20 pct iron. This operation was the first in. a series of mining enterprises culminating in October 1947 with the consolidation of Day Mines, Inc. In the same year it was decided to unwater the levels below the collar of the Hercules shaft in the hope of finding some indication of a recurrence of ore. The unwatering operation has been described in a. previous paper.' The initial exploration, following recapture of the workings, showed sufficient promise to warrant a detailed study of the mineralogy with modern techniques. The general geo1ogy of the Coeur d'Alene district, including a detailed description of the rock types encountered, has been comprehensively treated by Ransome and Calkins' in their classic paper, and only local background description, therefore, is felt to be appropriate here. The Hercules deposit transects a portion of the trough of a broad south-trending synclinorium which has been greatly complicated by faulting. More locally, it lies within a block of ground bounded on the east; by the O'Neil Gulch fault, a steep north-south overthrust of considerable magnitude, and on the west by a monzonite stock, the outcrop of which is 1/2 mile or more wide and 5 miles long. The country rock is composed of thin to medium-bedded argillites and argillaceous quartz-ites of the Prichard and Burke formations, the oldest members of the Pre-cambrian Belt Series of sediments in the area, believed to be of Algonkian age. The contact between them is a conformable gradation. The argillite is colored gray to tannish-gray and is fine-grained, compact, and generally massive in structure. Under the microscope the unaltered argillite is seen to be composed principally of anhedral quartz and a few feldspar grains which were at one time presumably partly rounded sand grains, but as a result of recrystallization and cementation by silica, the interstices are now almost obliterated and quartz grains show crenulate boundaries. The sizes of these crystals vary from 0.5 mm down to 0.1 mm in greatest dimension. In all specimens sericite comprises 10 to 20 pct of the rock and is present abundantly between most of the grains as flakes or shreds which vary considerably in size. Sometimes they form a fine felt-like mat or aggregate, and sometimes flakes are seen which appear to be good muscovite. In some specimens, separated rhombic-shaped carbonate grains are abundant, and in some instances these have been changed to sericite. Mining operations to date have explored the Hercules vein to a maximum vertical depth of 3600 ft below its outcrop, and along a maximum strike-length of 3600 ft on certain of the lower mine levels. The main orebody is irregular in outline, extending over a variable strike-length of 400 to 1500 ft; and it is intersected by a strong transverse fault that has been traced from the surface to the bottom level. This has been named the Hercules fault, and apart from the vein itself, it is the most prominent structural feature in the mine. There is good evidence that it existed prior to the introduction of ore solutions and may have influenced ore deposition, but it was also the locus of important post-ore displacement and shows a progressive right-handed horizontal component reaching 200 ft on the deeper levels. Its vertical component is not definitely known but may be considerably greater. The fault strikes 20° N to 50° E and dips westerly at angles of 70" to 45", flattening in dip where it crosses the original orebody from east to west between 1000 and 1600 ft below the surface. At about 3000 ft in depth the Hercules fault is joined by a vertical fault of similar strike, and the major post-ore dis-placement below their junction is taken up along this vertical branch of the structure, now called the Mercury fault. Recent work has been concentrated in this vicinity. Another structural feature of special geologic interest, though of little economic importance, is the occurrence of a porphyritic dike in this area. This lies a short distance above the Hercules fault, essentially parallel to it, and is 5 to 15 ft in thickness. It appears at first glance to cut the mineralization, suggesting push-apart relationship, but small stringers of the vein minerals have been observed to penetrate the dike for a matter of inches at several points. The dike is thought to be related to the monzonite intrusion. A vertical longitudinal projection of the mine is shown in Fig. 1, which illustrates most of the features discussed above. The Hercules vein was deposited along the course of a strong, persistent shear zone that now appears as a braided network of gouge seams running through more or less crushed and shattered country rock. It strikes 70° N to 80° W and dips southerly at an average of 75". Barren parts of the structure vary in width from less than 1 ft to more than 15 ft. The width of mineralized segments may be double that. Although the evidence is not conclusive, pre-mineral, normal movement along the zone may be 1000 or 1500 ft. The horizontal component is unknown. Post-ore movement appears to have been

Jan 1, 1954

By N. A. Gokcen, J. Chipman

A revised treatment of the authors' published data eliminates the complex relation previously proposed between concentration of silicon and activity coefficient of oxygen in liquid iron. Revised values of the thermodynamic properties of the liquid solution are presented. IN a recent experimental study of the reaction SiO2 (s) = Si+ 2O; Kf, = [% Si] [% O]² [1] the authors' found a substantially constant equilibrium product in liquid iron at 1600°C of 2.8x10-5 They also reported extensive data on the reactions: SiO2 (s) + 2H2 (8) = Si- + 2H2O (g); K'2= [% Si] (H2O/H2 [2] and H, (g) + 0 = H2O (g); K'3 =( H2 O [31 (H2) [%O] From the results on reaction 3 and earlier data of Dastur² on this same reaction in the absence of silicon, they determined the activity coefficient of oxygen, f0, on the basis of the definition K3 = (H2O)/ (H2)f0 [% 0] where K, is the equilibrium constant and f0, is taken as unity in the pure Fe-0 system. Similarly values of fsi were deduced from results on reaction 2. In a more recent study" of analogous reactions in the system Fe-A1-0, it was found impossible to reconcile the results on reaction 3 with Dastur's data; accordingly the latter were ignored and the equilibrium results were extrapolated to find a value of K, at zero concentration of aluminum. This procedure failed to locate the cause of the discrepancy but it did yield reasonable values of activity coefficients. It also avoided introduction of the complex empirical relation between the oxygen activity coefficient and the concentration of the added element. The same type of discrepancy exists for system Fe-Si-0.' In the earlier paper an attempt was made to fit both sets of data by a single curved line (Fig. 6 of ref. l), the form of which is contrary to the theoretical requirement of a finite slope at infinite dilution. In the light of experience on the Fe-A1-0 system the discrepancy must be recognized as one which can be resolved only by more refined measurements. Accordingly Figs. 6 and 10 are retracted. It is pointed out also that until the discrepancy is resolved Figs. 7, 8, and 11 are subject to some uncertainty. Qualitatively the following conclusions still appear valid: 1—The activity coefficient of oxygen is reduced by addition of silicon. 2—In dilute solutions the activity coefficient of silicon increases with its concentration. 3—With respect to equilibrium in reaction 1, the above effects are approximately compensating. The discussion of K'1 in the previous paper requires no revision. It was pointed out that the constancy of the product [% Si] [% 0]² ndicated a compensating effect of the activity coefficients of silicon and oxygen. Therefore, as a very good approximation, K1 = K'1 and the following average values are suggested both for K, and K', at the temperatures 1550°, 1600°, and 1650", respectively, 1.0x10-", 2.8~10-" and 5.5 ~lo-'. Revision of the thermodynamic treatment is necessitated by the recent appearance of new data, based on a combination of combustion and solution calorimetry,' which yields for the heat of formation of low-cristobalite from the elements, the value —209,330 ±250 cal per mol at 25°C. This is about 4000 cal larger than the value previously accepted. The new value for cristobalite is used, together with Kelley's tables of high-temperature heat contents" and entropies and with Korber and Oelsen's' heat of fusion of silicon to obtain the following equation for the standard free energy of cristobalite in the temperature range 1700" to 2000°K: Si (1) + 02 (g) = Si02 (crist.); ?F° = -217,700 + 47.OT [4] The free energy of solution of 0, in liquid iron is:8 O2 (g) = 20 (in Fe); AF° = -55,860 - 1.14T [5] and these two equations are combined to give: Si (1) + 20 = SiO, (crist.); AF° = -161,840 + 48.14T [6] ?F°1873 = -71,700 cal. From the experimental value of K, = 2.8x10-5, Si + 20 = SiO2 (crist.); ?F°1873 = -39,000 cal. [7] The combination of Eqs. 6 and 7 yields the free energy change when liquid silicon dissolves in iron to form the dilute solution of unit activity (1 pct). Si (1) = Si; ?F°1873 = -32,700 cal. [8] The heat effect in this process according to Korber and Oelsen' is an evolution of 28,500 cal per gram

Jan 1, 1954

By Albert E. Long

Preliminary experiments conducted in 1949 with a few diamond bits in which the diamonds in the bit crowns were set in hard-vector directions indicated that a substantial saving could be effected when a high percentage of the diamonds were oriented in this-manner. It was realized, however, that many additional tests would have to be performed before a positive trend could be established and the possibility of coincidence eliminated. Therefore, a long-range series of experiments was planned to investigate various phases of the problems related to the orienting of diamonds in diamond bits. This report (the third in a series) 2/3 presents the information obtained from tests conducted to investigate more fully the effects of orienting high-grade drill bort in diamond bit crowns. In these tests, the drill bort were oriented in both hard-vector and soft-vector directions in the crowns of AX-size bits. Test drilling was done in a dike of saussuritized dacite porphyry similar to a fine-grained granite in hardness e drillability.

Jan 1, 1954

By Carville E. Sparks, Rollin Farmin

TRIAL installations of rock bolts, of the slit-rod-and-wedge type, were under way at several units of Day Mines, Inc., when Korean hostilities interrupted the already slow deliveries of steel bars to the Coeur d'Alene district. Factory-made bolts had not yet been put on the local market, so the program was halted for lack of supplies. Interest was revived by a visitor's description of wooden roof bolts. These were said to have been used briefly with apparent success in a coal mine, until apprehension voiced by the U. S. Bureau of Mines caused the practice to be suspended. To make wooden bolts for trial in ground support, Day Mines acquired a second-hand doweling machine equipped with two cutting heads, one to turn out the desired round rods of 2-in. diam, the other to turn out 1-in. rods to be used as powder-tamping sticks. This machine was installed in the all-weather sawmill of the Hercules mine unit at Burke, Idaho, where fabrication of the wooden bolts commenced early in 1951. Most of the mining in the Coeur d'Alene district is along steeply dipping veins in shaly quartzite and argillite of Algonkian age. Ground support commonly is required in zones where the rocks have been sheared, brecciated, and hydrothermally altered. Pressure from the sidewalls is more troublesome than weight overhead, but both increase with the size of the mine opening. Caving may come from a progressive sloughing of irregular rock fragments or from an exfoliation and buckling of the layered wall rocks. The disintegration is thought to develop from an initial elastic expansion of the rock toward the newly-created mine opening, followed by the dilation of many tiny partings in the rock by absorption of water. As the partings widen, masses of rock develop weight and become free to fall. The function of rock bolts is to prevent or retard widening of partings in the rock supported. Wooden Bolts, Wedges and Headboards Bolt assembly used by Day Mines consists of a bolt 4 or 6 ft long, two wedges 16 in. long, and a headboard 30 in. long, Fig. 1. All four pieces are made of local red (Douglas) fir, either green or well-soaked in the mill pond before it enters the sawmill. Bolts are fabricated from cants, 2 1/4 in. sq, cut from relatively straight-grained timber with a minimum of knots and trimmed to 4- and 6-ft lengths. The bolt then is turned in the doweling machine from 21/4 in. sq to 2 in. diam round, except for a 4-in. length at one end which is left full square to provide the striking head and the shoulder that holds the headboard in position for wedging. The foot end of the bolt is slit with a thin saw for a length of about 16 in., thereby making a slot to receive the wedge against which the bolt is driven for anchorage at the bottom of the rock hole. A similar slit, 12 in. long, is made in the opposite (head) end of the bolt to receive the second wedge, which crowds the headboard against the ground at the collar of the rock hole and puts the bolt in tension. The second slot is aligned 90" from the plane of the first slot to avoid Longitudinal splitting and is notched out slightly to allow easier insertion of the collar wedge after the bolt has been driven to bottom. To prevent splitting the headboard by spreading action of the head wedge, this slot is oriented at 90" to the grain of the headboard when the pieces are assembled, Fig. 2. The wedges are similar to standard mine wedges, but more slender; they are cut 1 7/8 in. wide and 1 in. thick at the heel and taper out in 16 in. of length. The headboard, or bearing plate, is not necessary for some types of ground but generally is desirable because it helps the bolt to support an area of loose, friable rock and reduces the tendency for the rock at the collar of the hole to split away from the wedged head by distributing the pressure over a wider rock surface. The headboard may be a 24-to 30-in. length of 3-in. plank, 8 to 12 in. wide, but a similar length of rounded sawmill slab serves equally well at 20 pct of the cost. A hole of 2-in. diam is bored or punched through the center of the headboard, either at 90" or at various high angles to its surface. The bolt is inserted to its shoulder through this hole, then driven into the rock hole. Bolts, wedges, and headboards are given a full timber preservative treatment to inhibit rot. Bundles of each are immersed in a warm saturated solution of Osmose salts in water for 48 hr, removed, dripped dry, and stored in a relatively humid underground depot to cure. Most wooden rock bolts used by Day Mines are 4 ft long. Holes to receive them, about 42 in deep and 2 1/8 in. in diam, are drilled into the rock' to be supported, nearly normal to the periphery of the mine opening. The type of drill used is dictated by convenience: stoper, jackleg, or jumbo-mounted drifter. Correct depth of the hole is assured by use of a measuring stick that has been cut to the proper distance from drill chuck to the ground at the collar of the hole when a standard length drill rod is at the bottom. The bolt is seated to the shoulder through the hole in the headboard, the foot-end wedge is placed in its slot, and the assembly is inserted into the rock hole. Then the bolt is driven until it is seated solidly on the wedge against the bottom of the rock hole. Driving may be by hand with a sledge, or

Jan 1, 1954

By W. A. Cole

This report covers the preliminary phase of an investigation of some clays and shales and other materials in North and South Dakota to determine the feasibility of lightweight -aggregate manufacture in the two States. The background of the light - weight -aggregate industry, on a national basis, and results of preliminary laboratory testing of samples of North and South Dakota clays and shales from various localities are presented herein. The countrywide search for lightweight aggregates was necessitated toward the end of the 19th century by the change in building design from thick, heavy, load- bearing walls to a framework of structural steel .beams and columns with thin walls, followed shortly by the use of reinforced concrete and structural concrete as materials for the load-supporting frame. Ordinary concrete, with sand and gravel aggregate, weighs about 150 pounds to the cubic foot. A lighter concrete was needed to relieve the dead load and make it possible to construct higher buildings or add stories to existing buildings.

Jan 1, 1954

By R. K. McGeary, B. Lustman

Orientation relationships and rates of annealing of 97 pct cold rolled zirconium have been studied by X-ray techniques, metallography, and by hardness measurements. The process of annealing occurring by the formation of domains, and their subsequent growth without recrystallization, is demonstrated. NUMEROUS investigators1-' have revealed sub-grain formation after annealing cold worked metal single crystals and have generally described the results in terms of a dislocation mechanism of polygonization.5,6 in heavily cold rolled polycrystal-line aluminum the first stages of heating cause an apparently spontaneous subdivision of regions of the deformed matrix into a great number of slightly disoriented domains followed or accompanied by rapid growth of certain domains to form a recrystallization texture.7,8 In the present investigation the isothermal rate of annealing of cold worked polycrystalline zirconium has been studied by X-ray techniques, metallography, and by hardness measurement. It will be shown that, rather than by the usual nucleation and growth mechanism,"12 the process of annealing can best be explained by the formation of domains early in the annealing treatment and the subsequent growth of these domains without recrystallization. Experimental Procedure All the experimental work was performed on an arc-melted ingot of high purity zirconium crystal bar made by Westinghouse in a production scale deBoer unit. The weight percentage of impurities in the zirconium were as follows: C, 0.04; Fe, 0.05; Hf, 0.014; W, 0.01; Ti, Ni, and Al, 0.003; Si and N, 0.002; Cr and Cu, 0.001. Metallographically the zirconium showed only a single phase. The 4 in. diarn ingot was forged at 760°C to a 1 1/2x3 in. plate and was then rolled at 760°C to 0.150 in. thickness and annealed at this temperature for 20 min to produce an equiaxed grain structure of about 0.1 mm average grain diameter. The hot rolled plate was roller-leveled, surface ground to 0.135 in., cold rolled 83 pct reduction in thickness on a two-high mill, and then to a total reduction of 97 pct on a Sendzimir mill. The 0.004 in. strip material so produced received no subsequent mechanical treatment. Specimens % in. square were sheared from the rolled strip and isothermally annealed at 104 mm Hg in an all glass vacuum system equipped with a Vycor furnace tube. The temperature data presented are the average temperatures during the runs and did not vary more than about k3"C. X-ray Diffraction Methods: When zirconium is drastically cold rolled, a high degree of preferred orientation results in which prism poles, {1010}, are perpendicular to the rolling direction. After fully annealing this material an entirely new texture develops which corresponds to a 20" or 40" rotation of the deformation texture about hexagonal axes of

Jan 1, 1954

By O. C. Blade

This report on the properties of aviation gasolines and aviation jet fuels produced in the United States during 1953 was made under a co-operative agreement between the American Petroleum Institute and the Bureau of Mines of the United States Department of the Interior. Analytical data are presented for 101 samples of aviation gasoline which include, for the second time in this series of reports, data on some samples of grade 108/135 aviation fuel. Data also are given for 32 samples of aviation jet fuel. The analyses were made on samples typical of the manufacturers' production in their laboratories and the results sub-mitted to the Bureau of Mines for compilation. The products of 19 manufacturers of aviation gasolines and of 19 producers of aviation jet fuels are represented.

Jan 1, 1954

By E. B. Fitch, E. C. Johnson

The operating behavior of liquid-solid cyclones is outlined, together with the nature and range of the process results obtainable, to serve as a background for engineers wishing to consider application of this new process tool. BY now most engineers are familiar with the liquid-solid or Dutch State Mines cyclone. However, it should be helpful to know exactly what it is that the equipment does and what its limits are. Without going into cyclone theory, this paper will describe the operating characteristics of Dutch State Mines cyclones. These are manufactured under license in this country and sold under the trade-marked name of DorrClone. The physical construction of the liquid-solid cyclone has been covered in many papers,'-' the DorrClone in particular being described in some detail by Weems. Fig. 1 shows the unit in cross-section. The feed enters at C. The coarse, heavy particles are thrown centrifugally to the periphery and make their way down the wall to the apex where their rate of discharge as underflow is controlled by an adjustable rubber apex valve. As the apex diameter is decreased the solids build up behind the valve, producing a denser underflow. Meanwhile the fine particles are swept into the upward flowing vortex stream which exits as overflow through the vortex finder, F. Flexibility to produce the specific result desired in a particular process is achieved by providing means for varying the areas of the entrance, vortex discharge, and apex discharge. The entrance area may be varied by insertion of special shims. Vortex discharge area may be changed by use of different-sized vortex finders which are interchangeable. Similarly, the different sizes of apex valves are interchangeable and in addition each apex valve is variable down to about 60 pct of its maximum diameter. A most significant primary distinction to make is that although liquid-solid cyclones have been sometimes called thickeners, they actually are classifiers, and very potent ones. They are almost never thickeners in the special sense that many metallurgical engineers understand the term. There would be no profit in quibbling over the definition of a word, but when the application of cyclones is considered, it will help to understand the difference be- tween two mechanisms, one of which will be called classification, and the other thickening. In what is called thickening the fine solid particles present in the feed hold together by surface attraction during the sedimentation process. The loose network of particles thus held together constrains all particles to settle at approximately the same rate, the larger ones dragging the smaller ones down. As a result, pulp settles with a sharp line of demarcation between solids and a relatively clear supernatant liquid. Essentially all the solids, regardless of their fineness, pass into the thickened underflow, and a clear overflow is separated. In classification, on the other hand, the interparti-cle forces are relatively insignificant as compared to the settling force on the individual particles, and are insufficient to prevent independent movement of the particles. The coarsest, heaviest particles settle most rapidly through the pulp, passing more slowly settling fines. Particles coarser than the mesh-of-separation essentially all settle into the underflow, but if the feed contains any particles finer than the mesh of separation, at least part of them will appear in the overflow. A clear supernatant or overflow can be obtained only if there are no undersize particles present in the feed. Thus it will be seen that classification is impossible under ideal thickening conditions. The finer particles are pulled down at essentially the same rate as coarser particles, and there is no separation on the basis of particle size. The surface attraction holding the particles together in a thickener is usually feeble. Whenever the sedimentation force on any particle is strong with respect to the interparticle forces, that particle can pass through the tenuous structure and settle independently. There are at least four ways of making the sedimentation force strong, with respect to the interparticle forces, and obtaining classification. First, and most obvious, the particles may be large and heavy. Thus coarse sands settle out in a beaker or Dorr thickener ahead of the rest of the thickening solids. Second, the interparticle forces may be altered by physicochemical means; i.e., it is often necessary to add dispersing agents to destroy the interparticle forces and permit classification to take place. Third, the interparticle forces may be reduced by dilution of pulp. It is well known that to obtain the most efficient separation of

Jan 1, 1954

By G. F. Carr

"IntroductionFLUORSPAR has been found at several •localities in Newfoundland, but all the commercial deposits discovered to date are in the vicinity of St. Lawrence, a town on the southeast coast. Most of the veins are within six miles of St. Lawrence harbour, and between sea-level and an altitude of 400 feet. The St. Lawrence Corporation of Newfoundland, Limited, and Newfoundland Fluorspar, Limited, a subsidiary of the Aluminum Company of Canada. Limited. are the two producers.With the arrival of 1953, the St. Lawrence area entered upon its twenty-second •consecutive year as a producer of fluorspar. Production has increased more or less steadily during this period, as is shown by the fact that approximately one-half of the total shipments to date have been made during the past six years. Between four and five hundred men are employed by the industry, and more than $1,250,000 was paid out in wages and salaries during 1952. Shipments during that year reached an all-time high of 81,283 tons valued at $2,484,943, compared with 67,925 tons valued at $1,966,477 in 1951 (1).As a result of an agreement made in July, 1952, with the United States Defence Materials Procurement Agency, production of the St. Lawrence Corporation, Limited, and St. Lawrence Fluorspar, Incorporated, an affiliated company, will •be increased by .approximately 50,000 tons of acid-grade fluorspar per year. The contract extends over a period of four years, or until 150,000 tons of acid-grade concentrates have been produced."

Jan 1, 1954