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By H. L. Gilbert

Zircon sand may be reacted with sodium hydroxide at 600° C. to form sodium zirconate and sodium silicate. Sodium zirconate is water insoluble and so may be washed free of sodium silicate. The sodium zirconate may then be converted to zirconyl chloride by the action of hydrochloric acid and the zirconyl chloride crystals separated by cooling the solution. This approach produces zirconyl chloride of satisfactory purity with less cost than is involved in buying anhydrous zirconium chloride and converting it. A proposed design for a plant is given, INTRODUCTION This work was conducted by the Rare Metals Branch as a part of research on zirconium at the Bureau of Mines Northwest Electrodevelopment Laboratories, Albany, Oreg., under the guidance of S. M. Shelton, Regional Director. The work was supervised by W. W. Stephens, head of the Rare Metals Branch (now with Carborundum Metals Corp.).

Jan 1, 1954

By R. M. Creamer

Attempts have been made to electroplate titanium and zirconium from aqueous, nonaqueous, and fused-salt electrolytes. Metals were produced only in fused-salt electrolytes, the major product being sponge metal. All other electrolytes gave only partial reduction of the metal ions. Fused alkali borate electrolytes containing titanium salts gave titanium boride powders and coatings of poor adherence. Some of the problems encountered in the deposition of these metals and some of the aspects of titanium and zirconium chemistry involved are discussed. INTRODUCTION Titanium and zirconium are both relatively light, noncorrosive metals with both military and civilian uses. This has led to intensive investigations into the metallurgy of these metals. Both metals are high in the Electrochemical Displacement Series and are difficult to prepare in a pure state. At present they are prepared by reducing their chlorides, using magnesium. Electrolytic reductions have resulted in the preparation of fine powders or sponge of variable quality. It is believed that a good process of electroplating these metals on a baser metal would result in a very desirable product. The objective of the research summarized in this report was to investigate methods of electrodepositing these metals, with the idea of producing uniform, coherent plates. Most of the work was concentrated on titanium, the more abundant of the two metals. A system developed for titanium should be applicable to zirconium owing to their chemical similarity. Information gained from these experiments may also be applicable to the electro-winning of the metal.

Jan 1, 1954

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. B. Taylor

"The Delnite mine is in the northwest part of Deloro township, Cochrane district. The property consists of 307 acres on the so-called 'South Break' in the Porcupine camp. The Company was formed in 1934 to take over the holdings of La Roche Mines, Limited. Gold has been produced continuously since June 1, 1937. The value of the gold recovered to date (December 31, 1946) has been $7,456,000, and the average grade of ore $6.81 a con.The ore at the Delnite occurs in relatively narrow, persistent, quartz-ankerite veins, which generally contain tourmaline and have a ladder-like structure. The gold occurs in the free state, and in association with pyrite and arsenopyrite, the two chief sulphides. The host rock is fine-grained, highly carbonatized andesite of Keewatin age."

Jan 1, 1954

By J. W. Evans

ONE of the most important and frequent calculations that the extractive metallurgist is called upon to make is that of the standard free energy change of a reaction (?F°). For many reactions of metallurgical interest, the lengthy arithmetical calculations have been eliminated by the use of free energy-temperature diagrams which have been constructed by Ellingham,1 Richardson and Jeffes,2 Kellogg,3 and Osborn4 (e.g., for oxides, sulphides, halides) from critical surveys of existing data. The speed and simplicity of making calculations with these graphs make them of great practical value and, this should be stressed, in the majority of cases no greater accuracy is obtained by the individual calculations from specific heat data, entropies, and heats of reaction. With the increasing industrial application of vacuum methods to the winning of metals, notably at present magnesium and the alkaline earth metals, it is necessary to calculate low metal vapor pressures above reaction mixtures. The existing free energy diagrams usually give values where the solid and liquid metals are at unit activity. To obtain the free energy change associated with the reaction Metal (solid or liquid) = Metal (gas) it is customary to calculate this from either vapor pressure equations or from the appropriate free energy functions. This paper presents graphically the free energies of vaporization of a number of metals of economic importance which can be used in conjunction with other free energy diagrams. Certain metals have been excluded on the following grounds: titanium, tungsten, molybdenum, and platinum have very high heats of vaporization and boiling points and the uncertainty of these values arising from difficulties of high temperature measurements makes free energy values of doubtful accuracy; the data for cobalt, vanadium, and zirconium are uncertain; finally, arsenic, antimony, and bismuth could not be conveniently included. because of the presence of mixtures of polyatomic and mona-tomic vapors in the temperature range considered. Free Energy Diagram In Fig. 1, the standard free energies of vaporization per mol (?F°T) have been plotted for the temperature range 0" to 2000°C for the reaction M (solid or liquid) = M (gas) The data have been obtained from the free energy functions — (F-H298)/T for the solid, liquid, and gaseous elements and the heats of vaporization at 298°K, i.e., ?H298 (vapor). The values of these functions have been taken without change from the recent critical survey by Brewer;' the bulk of this data is from Kelley7 with adjustments in the light of more recent work. Values of the function — (F-H298) /T at T = 298°, 500°, 1000°, 1500°, and 2000°K were employed. Where possible, a further independent value was obtained by using the boiling point, i.e., where ?F°T = 0. All calculations are based on the assumption that the metal vapors are monatomic and that the boiling points refer to those temperatures at which the partial pressure of the monatomic vapor above the metal is 760 mm. For the alkali metals, e.g., sodium, potassium, and lithium, diatomic gas molecules are present as minor species; details of these are discussed by Brewer." On the basis of the above values and allowing for the change in slope at the melting point, straight line graphs were constructed and extrapolated to 2000°C. The deviation from a straight line did not exceed ±500 cal, which in all cases appears less than the probable error in the ?F°T values.

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 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 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 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 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

By H. L. Gilbert, H. A. Johansen, R. G. Nelson

High purity electrolytic chromium plate has been hydrogen-reduced and arc-melted under inert atmosphere to give sound ingots. These ingots may be hot forged to break the as-cast structure and then worked at 500°C to finished form. An intermediate and final high temperature anneal gives best ductility and grain refinement. HE increasing interest in refractory metals for I use in high temperature gas turbines, jets, and rockets has led to great efforts in exploiting the fifth and sixth groups of transition elements. Much of the interest has centered upon production of metals that can be worked and fabricated by normal means to produce various engine parts or components having the high temperature characteristics of oxidation resistance and high creep strength required for such uses. If these pure elements do not exhibit all of the desired characteristics, it may be possible to obtain a high purity alloy that will. Since, in all cases, development of these transition elements has depended to a large extent upon removal of oxygen and nitrogen as a means for producing ductile metal, the hypothesis has been accepted that oxygen is the normal embrittling material in a high grade electrolytic chromium metal. Various methods have been employed in an effort to remove this oxygen by direct or indirect attack.' Many workers have endeavored to produce pure chromium through reduction of the oxide by carbon or active metals, while others have recommended fusion electrolysis of chromium salts or reduction of these salts by active metals, such as sodium or calcium. Hydrogen reduction of the oxide or chloride and thermal dissociation of the iodide have been reported to give high purity metal. Experimental Work The U. S. Bureau of Mines, Northwest Electro-development Laboratory, has systematically re-examined the more promising methods and developed two which produce metal having ductility at temperatures lower than heretofore described. It has been found' that oxide-free chromium chlorides in a NaCl-KCl carrier salt may be mag- nesium-reduced and the byproduct salts distilled off in vacuo to give a high purity chromium sponge. Fusion analysis conducted under the supervision of W. W. Horton of Knolls Atomic Power Laboratory showed the oxygen and nitrogen both to be below 10 parts per million on a selected sample. A small amount of carbon is carried over in the final metal from the chlorination and reduction steps which are carried out in graphite crucibles. This trace of carbon is not detrimental to hot ductility. A more practical method for producing ductile metal is through hydrogen reduction of high purity electrolytic chromium. Normal electrolytic chromium will, if "developed" for a few minutes at 1100°C in high vacuum, be found by the method of Adcock:; to contain about 1 pct chromium sesquioxide (Cr2O3). This small amount of oxygen may be effectively reduced by hydrogen.2, 4 The method employed was to crush electrolytic chromium plate in a gyratory crusher to pass 60-mesh; finer crushing appears to offer no advantage. A nitric acid leach removed the iron picked up during crushing. The leached powder was rinsed until neutral with distilled water and vacuum-dried.

Jan 1, 1954

By L. J. Dijkstra, R. J. Sladek

IN earlier work the effect of manganese on the general behavior of nitrogen in iron was the subject of a careful examination by Fast.' Part of the investigation was made, in collaboration with one of the present authors (L.J.D.), by using the internal friction method. This technique, which has been described elsewhere,' showed some significant changes in the properties of Fe-N alloys by the addition of a small amount of manganese namely: 1—Under standard conditions the maximum of the nitrogen peak in pure iron appears around 20 °C and the half width is about 27 °C. By addition of 0.5 atomic pct Mn a general broadening of the internal friction peak occurs. The half width is increased by about 9°C. Moreover, a shift of the maximum of about 5°C toward a higher temperature takes place, Fig. 2. 2—The precipitation of nitrogen from the supersaturated solution, measured, e.g., at 200 °C, was inhibited by the addition of the amount of manganese given in effect 1. The broadening of the nitrogen peak was confirmed in more recent work3 on samples which were prepared from Westinghouse Puron iron. The effect of manganese on the rate of precipitation, however, was much less striking than in the earlier work' based on Philips high purity iron. Theory To explain the first change in properties just referred to, effect 1, the following two assumptions were made:' First, every manganese atom in sub-stitutional solid solution creates around it six interstices located between an iron and a manganese atom, as is evident from Fig. 1. These interstices, conveniently called Fe-Mn interstices, have for nitrogen a free energy level which is AG* lower than the normal Fe-Fe interstices between two iron atoms. For such low concentrations of manganese as 0.5 atomic pct, the possibility that two manganese atoms will be nearest neighbors can be disregarded. Second, it is assumed that in addition to the normal relaxation time TFe at least one new relaxation time, Mn, Fe, enters in the phenomena which is associated with the elementary diffusion jumps between a Fe-Mn interstice and an adjacent Fe-Fe interstice. Fig. 1 gives schematically the free energy for a nitrogen atom as a function of its position along the line ABC. A tentative analysis of the nitrogen peak in a 0.5 pct Fe-Mn alloy based on these assumptions has been made in Fig. 2. The dotted curve with a maximum at 22°C is caused by jumps between Fe-Fe interstices, the dotted curve with maximum at 32°C by jumps between an Fe-Mn and an Fe-Fe interstice. The two effects added together give the dash-dot internal friction curve for the alloy. In the following these two components will be referred to as the normal peak at 22°C and the abnormal one. (The small secondary peak at

Jan 1, 1954

By A. R. Byers

"Porcupine Peninsular mine is at the extreme southwestern end of North peninsula in Night Hawk lake, Cody township, Ontario. It lies approximately 10 miles east of the gold-producing district of Porcupine.Gold in quartz veins was discovered on Gold -island in Night Hawk lake in 1906, the same year in which the now famous Porcupine gold camp was discovered farther west. Most of the claims now held by the Company were originally staked by Charles Auer in 1907, In 1910 when the level of the lake was exceptionally low, gold was discovered in a quartz vein near the shore and close to the present mine shaft.The mine is developed by a main shaft to a depth of 440 feet, with levels at 80, 180, 300, and 425 feet vertical depth and by a winze to 630 feet. A mill of 200 tons daily capacity was built during 1923, Production commenced on January 4, 1924, and continued without interruption to May 31; 1926, when operations were suspended due to a decrease in the grade of ore. Total production amounted to 27,416 ounces of gold from 99,708 tons of ore, an average grade of 0.275 oz. a ton. Diamond drilling and re-sampling of the mine during 1934 and 1935 indicated approximately 1,480,000 tons, having an average grade of 0.123 oz. gold a ton, above the 625-foot level."

Jan 1, 1954

By L. J. Kane

Applications are described of various methods for determining the weight concentration and physical properties of solid and liquid impurities of gases in connection with problems encountered in synthesis-gas production. Of special interest among these methods may be listed: 1. A method developed for determining, the weight concentration of solid and liquid impurities in. hot, crude synthesis gas containing a high concentration of moisture. 2. A method developed for the same purpose but for highly purified synthesis gas or atmospheric air. 3. A method for obtaining a uniform dispersion of dry duet or, a microscope slide for particle-size determination or other purposes.

Jan 1, 1954

By A. H. Lang

"The Cordilleran region of western Canada includes most of British Columbia and the Yukon and pares of Alberta and the Northwest Territories. The western part of this region was for about 50 years the principal mineral producing area of the Dominion. It lost this distinction in 1907 when production from the mines of the Canadian Shield took the lead, but it continues to provide about 15 percent of the total Canadian production. In 1945 the total mineral production of British Columbia and the Yukon was valued at $65,302,900, of which about 80 percent was derived from metals.This paper is intended to serve as an introduction co the more derailed Cordilleran papers. Part I includes a brief synopsis of the complex geology and diverse types of mineralization in the region and Parr II collates the main structural principles exemplified in the derailed papers, and draws some conclusions regarding the application of this information to prospecting."

Jan 1, 1954

By J. W. Watkins

This report describes in detail methods of making corrosion tests and chemical analyses that have been used extensively and satisfactorily in a study of plants in which various surface, waters and brines are conditioned and used for water flooding, brine disposal, and reservoir-pressure maintenance in Kansas, Oklahoma, and North Texas. The tests and analyses described were designed and adapted for use in the field in determining the relative corrosive and plugging tendencies of waters from various sources conditioned by different methods. Although the described methods were developed and used in testing oil-field waters in the Midcontinent oil-producing area of the United States, these same methods should be equally applicable to testing oil-field or highly mineralized waters throughout the world. Many of the methods used have been discussed by the writer (1)2/ previously, and some of them have been used for similar work in other laboratories. Early in the development of the practice of the subsurface injection of brines for disposal and increasing the recovery of oil, it became evident that different methods of chemical treatment and conditioning equipment would have to be devised to sustain desired injection rates into water-input wells. These changes became necessary because the highly mineralized waters produced with petroleum from deep formations became chemically unstable when pumped to the surface, owing to loss of dissolved gases as pressure and temperature dropped to atmospheric and to the contact with atmospheric oxygen at the surface.

Jan 1, 1954

1. The following recommendations should apply to all mines, tunnels and underground quarries, but to opencast quarries only in so far as persons can be exposed to dust injurious to health. General In the present state of knowledge, all duets liberated in mining, tunneling and quarrying operations should be considered potentially injurious to health if breathed in sufficient quantity; dusts containing free silica should be considered specially injurious to health, the injurious particles being below 7 microns in diameter.

Jan 1, 1954