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By W. R. Davis

This paper presents a sunimary of the results of a study of the Dornick Hills-Springer sand reservoir of the Velma Pool. Stephens County, Oklahoma. The data are of interest because this type of reservoir is unusual in the mid-continent area. The reservoir is characterized by very steep dips and a relatively thick producing section. Its up-dip productive limit is defined by truncation of the productive sands, and the down-dip limit is defined by deterioration of sand quality rather than by the presence of an oil-water contact. The crude oil is of 29" API gravity, is relatively viscous and was saturated with dissolved gas at the original reservoir pressure. Consideratiqn of the physical characteristics of the reservoir indicates that natural water encroachment will be negligible, and that the producing mechanism will be a dissolved gas drive augmented by gravity drainage. Performance to date tends to confirm this. INTRODUCTION It is the purpose of this paper to present some of the features of the Dor-nick Hills - Springer sand reservoir which are unusual in mid-continent operations. The study was undertaken in an attempt to define the producing mechanism or mechanisms that will control reservoir performance, as an aid in determining well completion and production practices. GEOLOGY The reservoir under consideration is a series of sand bodies of Pennsylva-nian age found in the Lower Dornick Hills and/or Springer formations. For the purpose of this paper, the sand series is referred to as the Dornick Hills-Springer sand. This sand series lies on the northeast flank of the Velma anticline, and dips very sharply to the northeast in TWP 1S-5W, the average dip being about 45º. However, as shown in Fig. 1. the trend of the pool swings eastward in TWP 1S-4W with a coincident de- crease in dip angle. In this portion of the pool the average dip is about 25". Faults in Sections 30-1S-4W and 32-1S-4W are shown in Fig. 1. These faults show only minor displacements for the horizon contoured in Fig. 1, but they do show a pronounced displacement of the immediately overlying beds. Similar faults are thought to exist in TWP 1S-5W; however, since the detailed geological studies necessary to properly define these faults were not available to the author, no attempt to show them in Fig. 1 has been made. The up-dip productive limit throughout the pool is determined by the truncation of the Lower Dornick Hills-Springer sand. The down-dip productive limit does not appear to be defined by edge water, but by a gradual reduction in permeability of the sands. No dry hole has produced a significant volume of water from this sand section. There was a small gas cap in TWP 1S-5W. Most of the wells in TWP 1S-SW encountered an oil sand 70 ft to 150 ft thick, a shale 20 ft to 80 ft thick, and another oil sand 50 ft to 130 ft thick. A few scattered wells in this area encountered other and bodies below these

Jan 1, 1951

By G. M. Humphrey

It is a great honour and privilege to be here in this distinguished company of the outstanding men in the mining industry and the political life of Canada. I fully recognize that I am so honoured as a symbol of the combination of American and Canadian leaders in industry who are partners in the sponsor-ship of the Quebec-Labrador iron-ore mining development which we are undertaking, and as such I want to express our great appreciation of the co-operation and assistance we have met on every hand in Canada, and the genuine friendliness and never-failing courtesy which we are always accorded. We are quite aware that it springs primarily from the very high regard in which our partner and sponsor, Jules Timmins, and our other Canadian associates are held. Their high standing among you is well known and we quite naturally fall heir to the benefits of the great esteem in which you hold them. Jules Timmins and his long-time associate, John Knox, were the original inspiration for our participation in this venture. It was their enthusiasm and vision which caught our interest and led to our joining with them in the years of exploration and search which have brought us to the threshold of this great development where we stand today. So much has been written of the details of our exploration and plans for future development that it would bore you for me to review them further here, but there is a phase of this development, and an extremely important phase of it, which I think has passed almost unnoticed and which it may be well worth while for me to bring to your attention this evening.

Jan 1, 1951

By W. W. Bake

INTRODUCTION IN IN JUNE, 1943, the main pillars of the Beattie Mine glory hole failed, allowing over a million yards of wet clay and debris to flow into the mine, completely filling all workings to within seventy feet of the 3rd level. The resulting Job of clearing these workings and rehabilitating the mine represents a unique and hazardous operation, seldom, if ever, previously encountered in Canadian mining. The purpose of this paper is to describe one phase of this operation -that of cleaning and rehabilitating the mine from the 6th to the 9th level -to outline briefly the events which led up to this work and to describe the techniques and costs involved. It will be noted that such costs, methods, and performance figures as are dealt with in this paper refer only to work in the internal shaft and adjoining workings from the 6th to the 9th level, previous work having been done under other supervision and management. The Beattie mine is in Duparquet township, northwestern Quebec, twenty miles northwest of the town of Noranda and twenty-three miles southeast of the town of La Sarre. The mine and the town of Duparquet, which it?supports, are accessible by highway from Noranda throughout the year. Production at the Beattie mine began in May, 1933, and up to June, 1943, twenty-one million dollars in gold and silver had been recovered from five and a half million tons of ore.

Jan 1, 1951

By S. H. Williston

IT not often that bore hole surveys can be checked by actual. civil engineering methods. A recent Arizona survey was checked by normal surveying methods and the comparison of the results should be of value to both oil and mining men. During the summer of 1948 the Phelps-Dodge Corporation, at its Copper Queen property near Bis-bee, Ariz., drilled a 1245 ft, 8 in. diarn, churn drillhole in a mineralized area and cased part of it, intending to use it to transfer mill tailings for stope fill. The hole, as frequently occurs, was not straight, and, in endeavoring to locate the bottom in the underground workings, they found no evidence of the hole at the underground coordinates directly below the surface location. The noise of the drilling tools was reasonably clear, but the direction of sound was uncertain. Preliminary tests with available equipment were not successful in locating the bottom of the hole. Because of the mineralized character of the area and the fact there was some casing in the hole, any magnetic method of well surveying would give results of doubtful value. Sperry-Sun gyroscopic well-surveying instruments were finally used to locate the bottom of the hole. These instruments consist of a gyroscope to determine azimuth and either a pendulum or bubble inclino-meter. A multiple shot camera photographs both instruments on a single film and superimposes the photograph of a watch. Coordination of depth with time at the surface makes it possible to select the corresponding picture for any depth. After making several runs of the empty instrument housing from the top of the hole to the bottom to make sure there were no obstructions in the hole, three surveys on wire line were completed during the afternoon. The three surveys, in which readings were taken at different points in the hole on each survey, were computed and gave the following locations of the bottom of the hole in relation to the surface collar: survey No. 1—24.92 N, 30.30 W; survey No. 2—24.24 N, 31.11 W; survey No. 3—26.54 N, 27.72 W. Then the data from the three surveys were combined into a single set of calculations which gave a location for the bottom of the hole: combined surveys—24.27 N, 30.16 W. (Fig. 1.) Immediately upon the determination of the coordinates at the bottom of the hole, a drift on the 1300 ft level was started toward the indicated loca- tion some 38 ft northwest of the coordinates of the surface location, The bottom of the hole was located within the drift round in which it was expected, and the transit survey run to the actual location of the hole indicated N 27.18, W 29.71. This shows a discrepancy between the well survey and the transit survey of 0.45 ft in the westerly direction and 2.91 ft in the northerly direction. All surveys, both gyroscopic and transit, fell well within the width of an ordinary drift. While this is satisfactory for almost any and all mining requirements, a theoretical examination was made as to reasons for the discrepancy. A study of the course of the hole indicates that considerable right turn or spiral existed, and in all probability the surveying instrument was pulled out of alignment while traversing the turn by approximately 0.05 ft at the top and another 0.05 ft in the opposite direction at the bottom of the instrument. If such an allowance were to be made, the survey calculations would almost exactly correspond with those determined by transit. This sort of discrepancy would be minimized by the use of stabilizing guides. It is unfortunate that physical laws probably effectively prevent the use of gyroscopic instruments in EX and AX diamond drill holes. The directive power of a gyroscope falls off inversely at some rate between the third power and the sixth power of the diameter. Present instruments can be run in casing 53/4 in. ID or over and might be adapted to somewhat smaller diameters, but the difficulty of reducing these diameters to 11/4 in. or 2 in. is almost insurmountable at the present time. Acknowledgment The author wishes to express his appreciation to the Phelps-Dodge Corporation for permission to publish this article, and to the Operating and Engineering Departments for their cooperation on the survey; also to Donald Hering, of the Sperry-Sun Well Surveying Co., who actually made the survey and calculated the results.

Jan 1, 1951

By E. J. Roberts

The theory is developed that the tons ground through a given mesh per day in a wet ball mill is proportional to the percent plus that mesh in contact with the balls and the net power applied to the balls at this point. A grindability test is described. DURING the course of a study of the fundamentals of classification in 1937, the need for a more basic understanding of the action of a ball mill became acute. Unless one knows how classification affects grinding, one cannot hope to effectively improve on classification. The methods of evaluating grinding efficiency that depend on surface developed were studied but soon discarded for two reasons: 1. There was no apparent method which could be generally used to give a reliable figure for the actual new surface developed as a result of grinding. Subsequent papers have not changed this conclusion. 2. The practical evaluation of grinding in the main ore dressing applications was in terms of the percentage retained on a screen which passes 90 to 99 pct of the material and not in terms of surface area. The Probability Theory With the background of our experience in the field of closed-circuit grinding, together with the papers of Lennox,1 Gow,2 Gaudin,8 Fahrenwald,4 Coghill, and others, the approach of the theoretical physicist was then tried. The thought was somewhat as follows: When one grinds in a ball mill, a given expenditure of power leads either to a certain number of point to point blows per hp-hr or to a certain distance of line contact per hp-hr, depending on whether the action of the balls is considered to be cascading or rolling. It is also assumed that the balls actually come together on each blow or during the roll. Then a volume of slurry will be covered per minute which is some function of the size of the particle being considered (see fig. 1). All particles coarser than this size will be reduced through this size. This volume of slurry contains a certain weight of ore, depending on the percent solids and the density of the solids. If we fix the percent solids and the density of the solids and let w be this certain weight of ore in the volume covered, then, in mathematical terms, what we have just postulated is, w —— 8 hp (a) dt If W is the total weight of ore present in the mill, then we can write. W w/8 hp (b) W dt and if C is the cumulative percent plus the size chosen at the start of the time interval dt, w w c/dt W 8 hp x c (c) wc But wc/100 is the weight plus the size chosen which at 100 wc the close of time dt is finer than that size, and W is the decrease in the percent plus of the whole mass of ore or —dC. Then, —W dC/dt 8 hp x C. (d) In other words, the mesh tons ground through a given size per unit of time is proportional to the hp and the percent plus the mesh. A crude analogy would be to picture a 1-ft-wide steam roller going down the road at 1 ft per sec. If we place one egg on the road per square foot, one egg will be smashed per second. If we place a dozen eggs per square foot, a dozen eggs will be crushed per second. Similarly, if all the particles in w are plus the mesh, i.e., C=100, we should have a maximum rate of reduction. If only 10 pct of them are plus the mesh (C=10), we would have only one tenth the maximum rate; if only 1 pct are plus the mesh, the balls have a hard time finding anything to work on. This is where the term "probability theory" comes from. The chances of the balls crushing a particle through a given mesh depends directly on the concentration of particles coarser than this mesh in the general pulp in the mill. Giving W the units of tons and dividing equation (d) through by W, we obtain -dC hp ----- = k---— C [1] dt ton where k is a constant for any one size of particle, density of solid and moisture content of pulp. Eq 1 is the rate equation for a first order reaction and says that the rate of decrease of the percent plus a given mesh with time is directly proportional to the hp per ton applied to the body of ore and to the percent plus the mesh in the ore mass as a whole. Since it is a differential equation, it only

Jan 1, 1951

By H. B. Brunot

A complete list of permissible mine equipment, tested and approved by the Bureau of Mines previous to January 1, 1947, was published as Information Circular 7432.2/ A list of permissible mine equipment tested and approved by the Bureau of Mines during the calendar years 1947, 1948, and 1949 was published as Information Circular 7569. The present list supplements Information Circulars 7432 and 7569 and includes all of 'the equipment approved during the calendar year 1950. Information Circular 7465, which included a list of permissible mine equipment tested and approved by the Bureau of Mines during the calendar year 1947 therefore is superseded.

Jan 1, 1951

By Louis A. Turnbull

An investigation of coal deposits near Cerrillos Santa Fe County, N. Mex., was undertaken 14 the Bureau of Mines in 1943 as part of a general program to determine sources of metallurgical coal for western steel plants. Almost all of the coal used for metallurgical purposes west of the Continental Divide Is mined from the Lower Sunnyside bed in, Carbon County,, Utah. Coke used in western steel plants that is produced from Lower Sunnyside a coal is weaker, structurally, than coke used elsewhere in this country. For this reason, western coal operators and steel manufacturers are anxious to develop new areas of coking coal of better quality then the coal produced from the Sunnyside bed. The objective of investigations of western coals was to (a) develop additional reserves of coking coal to supplement proved reserves in the Sunnyside bed. and (b) develop; if possible, reserves of coal with coking quantities superior to those of Lower Sunnyside

Jan 1, 1951

By G. Carman Ridland

Front Range, Colorado: The majority of the rocks comprising the Front Range of Colorado are pre-Cambrian schists, gneisses, and intrusives which have been elevated to form part of the Southern Rocky Mountain physiographic province. The region is noted mainly for its production of gold and silver, but ores of tungsten, fluorspar, uranium, copper, and iron have been worked. Geological literature1-' describes pitchblende and other radioactive minerals occurring at Jamestown, Boulder County; Central City, Gilpin County; and Lawson, Clear Creek County (fig. 1). These three adjoining G. CARMAN RIDLAND, Member AIME, is a Consulting Geological Engineer, New York, N. Y. San Francisco Meeting, February 1949. TP 2738 IL. Discussion of this paper (2 copies) may be sent to Transactions AIME before Feb. 28, 1950. Manuscript received March 30, 1949. counties cover most of the highly mineralized section of the Front Range which extends from Jamestown to Breckenridge. As far as the writer is aware, uranium has been found only in the portion northeast of Lawson. Silver Deposits at Caribou Hill: The historic silver deposits of Caribou Hill, located approximately half way between Jamestown and Central City, and 35 miles in a straight line northwest of Denver, appeared particularly intriguing as an area to prospect for uranium for the following reasons: here exists an isolated group of high-grade silver veins in the center of a mining district containing occurrences of uranium, and, in two of the well-known uranium deposits of the world, those at Great Bear Lake, Canada, and Joachimsthal, Czechoslovakia, uranium is associated with silver in similar fissure veins. An exploration program, conducted in the spring of 1945, directed special attention to Caribou Hill and was rewarded with the discovery of heavy, canary-yellow stained specimens of'pitchblende. The veins are confined to shear zones cutting a north-south, elongated quartz-monzonite stock of Tertiary (?) age lying in the axis of a major anticline of the Idaho Springs biotite schist, the oldest pre-Cambrian formation in the district. The stock is one and three-quarter miles long by more than a mile wide, and Caribou Hill is approximately at its center. The ore zones vary in width from a few inches to 15 ft, and one structure has been traced for a length of over 4000 ft. The ore minerals are galena, sphalerite, ruby silver, native silver, chalcopyrite, pyrite, and pitchblende. They occur most common-

Jan 1, 1951

By T. S. Smith

INTRODUCTION IN PREPARING this paper for the District 4 Meeting, three factors had to be considered. First, the paper must fit the time alloted; second, it must give a general picture of the geology for those unfamiliar with the property; and third, it must contain some new material in order to be of interest to local delegates. The most comprehensive extant discussion of the geology of the property of McKenzie Red Lake Gold Mines, Limited, is contained in Horwood's report on the general Red Lake area ( 1). This report is the 'bible' of the area, and a general paper on the geology of this particular mine would almost certainly appear to be a condensation of it. It was thought, therefore, that the most suitable procedure would be to base the general outline on Horwood' s report, to extend that outline where possible from information gained since its publication, and finally to leave dangling for the contemplative geologist some of the ideas and problems which ?are currently under consideration. LOCATION AND HISTORY The property of McKenzie Red Lake Gold Mines, Limited, is in Dome township, in the Red Lake mining division of northwestern Ontario. The mine is on McKenzie Island, about four miles north, by water, of the Hasaga mine and old Howey mine.

Jan 1, 1951

By J. Cooperman

THE INVESTIGATION was carried out in three stages, depending on the objectives of the work and the kind of equipment used. The first stage involved testing of 11 representative Alabama coals, the source of which is given in table 1. The objective was to find whether these coals could be gasified raw by the Lurgi method, and the equipment used was a 4-inch batch generator (to be described later). This generator proved too small and difficult to control to provide adequate amounts of products under continuous operating conditions for testing. For the second stage, a 6-inch batch generator was used, which was provided with facilities for superheating the steam-oxygen mixture and for synthesizing additional methane from the residual carbon monoxide and hydrogen in the primary Lurgi gas. Selected Alabama coals, Pittsburgh-bed coal from West Virginia, "Disco," and anthracite were used. For the third stage of the work, a continuous generator of 1 square foot grate area, following closely the design of the industrial generators operated in Germany, was provided. This generator was designed by 0. Hubmann, engineer for the Lurgi company, and was expected to give results closely simulating industrial-scale apparatus. The purpose of this equipment was to test prepared materials shown to operate smoothly in the 6-inch generator under conditions approximating as nearly as possible those prevailing in industrial practice. Work of the first stage showed that Alabama coals could not be used in the Lurgi process without first destroying the coking power.

Jan 1, 1951

By Gordon Smith

ABSTRACT Gold-bearing quartz veins in the northeastern Ontario and north western Quebec mining camps are shown to have been mineralized in part by epithermal solutions. The vertical gradient of pressure during this stage of mineralization is calculated, using decrepitation temperatures of quartz and thermoelectric potentials of pyrite. The base of the Cobalt sediments, which lie uncomformably over part of the mining region, may be taken as datum level in calculating the depth of erosion into the basement below that level. The vertical extent of the epithermal gold mineralizaiton in this region is at least 9,000 feet below the datum level. INTRODUCTION IT WOULD have been difficult a few years ago to convince economic geologists that after making a series of measurements in a laboratory in Toronto, the depth of erosion which has taken place below the base of the Cobalt sediments at any place in the Ontario-Quebec mining region can be calculated to the neare.st 500 feet, or that the same series of measurements would also serve to allow a prediction of the minimum vertical depth of probable ore in a producing mine in the same region, with the same degree of certainty. Possibly it is just as difficult today, but the following description of the data and calculations may help to demonstrate the reliability of the method.

Jan 1, 1951

By R. W. Fenn, H. A. Lepper, W. R. Hibbard

THIS paper reports the results of static tension and torsion tests made on single crystals of alpha brass for the purpose of determining its elastic coefficients. 70-30 alpha brass was chosen because it has been used extensively in the study of the plastic deformation and because it has the following desirable characteristics: (1) face-centered cubic structure, (2) marked elastic anistropy, (3) low elastic modulus. In 1928 Masima and Sachs' using static methods reported the following values for the elastic coefficients of 72-28 pct alpha brass: S11 = 19.4 X 10-13' cm2 per dyne S12 = —8.35 X 10"' cm2 per dyne S44 = 13.9 X 10-13= cm2 per dyne These values were obtained from the measurements shown in fig. 1. The experimental points for both tension and torsion lie consistently below the lines drawn to represent them. Fig. 2 shows the stress-strain curves for their specimens. Only a few points in the elastic range of these curves were available to obtain their values for Young's modulus. Therefore, it was considered that a re-evaluation of these data was desirable. In addition a variation of from 1.13 to 1.485 kg per mm2 was found in the values of the critical resolved shear stress for 70-30 alpha brass reported by Burghoff,² Treuting and Brick: Martin,4 Maddin, Mathewson and Hibbard5,6 and von Goler and Sachs.' Therefore, it was desired to evaluate the conclusion of Boas and Schmid8 that a constant resolved shear stress and not a constant elastic shear initiates plastic deformation. Theoretical Considerations*: The elastic deformation of cubic crystals may be defined by three elastic coefficients (S11, S12, S44,) or three elastic parameters (C,,, C,, C,,). These coefficients or parameters are sufficient to define the stress and strain relations for cubic crystals in the generalized form of Hooke's law referred to Cartesian coordinates. Elastic parameters and coefficients in the cubic system are related by the expressions: C11 —S11+S12/(S11-S12)(S11+2 S12) C12=-S12/(S11-S12)(S11+ 2 S12) (S11 — S12) (S11 + 2 S12) [2] C44 =1/S44 PI It is possible to determine all the elastic coefficients for a cubic metal by measuring two of the three elastic constants: Young's modulus (E), torsion modulus (G), or compressibility (S). The following relations exist between crystallographic orientation and E, G and S for cylindrical crystals of the cubic system: 1/E = S11 — 2 [(S11 - S12) — Su] (?1² ?2² + ?2²?83 + ?3² ?1²) [4]

Jan 1, 1951

By C. J. Osborn

In oral discussion at the Columbus midyear meeting, September 26, 1949, these pertinent questions were asked: Was a satisfactory separation of metal and matte obtained? The matte was quite fluid and in most cases the separation was satisfactory. Were the mattes chlorinated after aging? The mattes were stored in sealed glass mason jars, and chlorina-tion experiments made after standing for several weeks gave satisfactory results. Was ferric chloride volatilized? Most of the iron content of the matte was converted to ferric chloride. However, because of the low chlorination temperatures used (150" to 250°C) the major part of the ferric chloride remained with the residue. Undoubtedly, some ferrous chloride was also formed, because of the reducing conditions within the reaction vessel, but no effort was made to separate it from the ferric chloride. Could the reaction be stopped before making metallic iron? No attempt was ever made to stop the reaction before making metallic iron. Generally, 20 to 60 pct of the iron in the charge was obtained as metallic iron which could be readily separated from the matte. In certain cases, metallic iron was dispersed through the matte in the form of globules, which could be removed by magnetic separation. Could other sulphides, i.e., calcium sulphide or sodium sulphide be used in place of iron sulphide? No other sulphides were investigated. It was felt that titanates of calcium and sodium would be formed should these sulphides be used to replace the pyrite. What were the sulphur volatilization losses during matte smelting? In the course of this laboratory investigation, a rather wide variety of variables was introduced from highly oxidizing to reducing conditions and various ratios of pyrite, rutile, and coke. These and other factors all have a definite bearing on the sulphur losses. In general, the higher the sulphur in the charge the higher were the sulphur losses and the percentage of titanium in the matte decreased. With sulphur additions more nearly the theoretical amount, the sulphur losses averaged approximately 10 to 15 pct. With a higher percentage of sulphur on the charges, the losses were in the order of 15 to 30 pct.

Jan 1, 1951

By C. S. Gibson

INTRODUCTION I WELCOME the opportunity to speak on Programmes in Safety. The invitation to do so shows that those guiding affairs of the Institute in this District realize the practical importance of this subject. My remarks are intended primarily for mines having 100 to 400 employees, approximately. With minor modifications, however, they apply also to larger or smaller operations. Others, I hope, will derive benefit from this paper, for they, too, have a part to play in accident prevention; for example, contractors such as shaft-sinkers; manufacturers of equipment; and consultants and designers of mining, milling, and metallurgical methods. These troubled times stress, as never before, the importance of efficient operations in mining. The days are gone when neglect can be afforded. Safe working conditions and satisfied employees are essential for efficient production at reasonable cost. The mining industry of Ontario long has been keenly aware of the human aspects of this subject, and as a whole it has done much to decrease the frequency and severity of disabling accidents. The number of fatalities per 1,000 employees has been reduced from a high of 5 some two decades ago to an all-time low of 0.89 in 1948; the incidence was slightly higher in 1949, being 1.17 per 1,000. The reduction in non-fatal compensable accidents also has been outstanding, frequency having ?been reduced from over 200 per 1,000 employees some two decades ago to 54 per 1,000 at the producing mines in 1949. These statistics demonstrate progress. Nevertheless, the general picture still leaves much to be desired when viewed from two other angles.

Jan 1, 1951

By E. A. Has, E. Mcl. Tittmann

Double roasting of sinter carrying a high percentage of lead concentrates, gave rise to the problem of removing the sheets of metallic lead formed in the wind boxes. The solution of the problem has been found in the water sealed wind boxes. The metallic lead is granulated and can be removed without seriously interfering with the operation of the machine and eliminates necessity of men going in wind boxes with air guns. THE practice of double sintering the charge for lead blast furnaces, which is now standard throughout the lead smelting industry, has introduced the rather difficult problem of cleaning the wind boxes on sintering machines used for the final pass. Two factors have more recently accentuated this problem: (1) the growing demand of workmen for better working conditions and less arduous labor, and (2) the growing dearth of crude lead ores and the improved grade of lead concentrates demand a charge carrying the maximum percentage of lead. One to 2 pct of such high lead charges (30 to 40) when processed on machines in reasonably good condition will pass through the grates and report in the wind boxes. One-half to three-quarters of this material will be metallic lead reduced by the coke breeze or other fuel mixed with the charge. This lead forms large cakes or chunks which are hard to remove and which substantially increase the time the machine is out of operation for cleaning the wind boxes. A high lead fall can reduce the operating time 30 pct. As a specific example, a 30 ft machine having 63 in. pallets, and equipped with three conventional wind boxes, requires 3 to 4 hr per day for removing the accumulation in the wind boxes when operated as a final pass machine. This same machine when operated on raw charge requires less than an hour per day for cleaning. To eliminate formation of large chunks of lead in the wind boxes of the final pass machines and time lost in cleaning the boxes, R. C. Rutherford of the Chihuahua smelter of the American Smelting and Refining Co., devised and patented* a wind box having a water bath to granulate the lead as it falls from the grate bars. Later, C. H. Harris of the San Luis plant made and patentedt an improvement on Rutherford's device in which the water bath seals the wind box in a manner permitting the latter to be readily cleaned without destroying the vacuum. The first machine to our knowledge to use the water bath box was at the Chihuahua plant along the lines of the Rutherford patent. In 1942, a machine was placed in operation at the El Paso plant which incorporated the features of both Rutherford and Harris. This machine is 40 ft long equipped with pallets 63 x 24 in., and having 4 wind boxes 10 ft 4 in. x 7 ft 2 in. x 51 ft 3 in. in size (fig. 1). The air from the fourth box is returned to the first three, and the gas from the first three discharged to a baghouse. The machine has conventional drive, discharge hopper,

Jan 1, 1951

By L. Guttman

THE equilibrium diagram of the indium-thallium system was of interest to us in connection with a study of the superconducting properties of metallic solid solutions in progress at this Institute. For this purpose, a series of alloys was prepared covering the range from pure indium to 75 atomic pet TI, and X-ray diffraction patterns were obtained. Roughly speaking, the results indicated a region of continuous decrease in the axial ratio of the face-centered tetragonal structure from the value ca. 1.08 in pure indium to unity at about 23 atomic pct TI. From this concentration to the solubility limit, the structure was face-centered cubic, a form not reported by Hansen.1 At this point we became aware of the work of Valentiner,2 who had found the same structures at room temperature but had not investigated further the relationship between them. To study the transformation f.c.c.??f.c.t., and to clarify the phase diagram, we undertook the work reported in this paper and that which follows." Preparation of Alloys Pure indium and thallium* (both about 99.9 pct, as estimated from spectroscopic analyses) were weighed out in quantities sufficient to prepare 10 to 25 g of alloy of the desired concentration. (Occasionally indium was mixed with a previously pre- pared solution.) The metals were melted over a flame, in an open graphite crucible, stirred with a graphite rod, and poured into a graphite trough. The rough slug was hammered and swaged into a rod of 2 to 3 mm diam, and annealed in a drying oven at about 125°C for 1 to 2 days. Samples were taken from the rod for analysis, as well as for diffraction and other measurements. The thallium was determined by electrometric titration, using a modification of the method of Beale and co-workers.' The precision of analysis was generally about 2 0.2 pct of the thallium content. The rods were sometimes found to vary in thallium content as much as 1.5 pct from end to end, but since analyses were also made close to the segments used in the various studies, the composition was usually known as accurately as the analyses could be made. Where one less significant figure is given, the error is estimated to be ±0.4 pct of the thallium content. X-Ray Diffraction Measurements Three different cameras were used: (1) A 7 cm Debye camera, with oscillating specimen (manufactured by the Picker X-ray Corp., Cleveland, Ohio); (2) a 9 cm diam Debye camera of the type designed by Bradley;" (3) a 10 cm diam symmetrical focusing back-reflection camera (manufactured by Geo. C. Wyland, Ramsay, N. J.). Samples were prepared from the homogenized rods, either by filing under liquid nitrogen, or more often, in the form of wires or foils made by working at room temperature or liquid nitrogen temperature. Sharp patterns were obtained from specimens annealed at room temperature, although higher annealing temperatures were also used (cf. below). Generally, copper radiation was used; a few patterns were obtained with

Jan 1, 1951

By E. A. Gulbransen, K. F. Andrew

THIS paper. will present the results of our studies on the kinetics of the gas phase reactions of co-lumbium and tantalum with O2, N2 and H2. Studies on zirconium and titanium have been previously reported.9,11 Due to their high melting points and their many excellent physical and chemical properties, colurn-bium and tantalum have been finding their places as useful metals in science and industry. Chemically the metals are very inert to gas and liquid phase corrosion at room temperature. However, at moderate temperatures of 250°C and higher the metals react readily with oxygen and hydrogen. It is of interest to study the gas phase reactions of these metals for the following reasons: (1) to compare the rates of the several reactions with those of other metals and to correlate the rate data with theory and structure predictions, (2) to add basic information on the practical difficulties of the reduction, refining and working of columbium and tantalum, (3) to anticipate new uses for the metals and (4) to understand the role that the metals play as components in high temperature alloys. Literature Survey A number of papers exist in the literature on the gas phase reactions of columbium and tantalum with O2, N2 and H2; however, only fragmentary data have been reported on the kinetics of these reactions. Although the free energy of formation of the oxide Ta2O520at 25°C has been determined and some thermodynamic work has been reported on columbium and its oxides, the data are not sufficient to make equilibria calculations at elevated temperatures. Oxidation: 1. Kinetics: The rate of oxidation of columbium and tantalum has been studied by McAdam and Geil23 using the interference color method. Tantalum is reported to oxidize more slowly than columbium, zirconium and iron. The amounts of oxidation of columbium after 20 hr in air for several temperatures are reported by the Fansteel Metallurgical Corporation.36 Columbium starts to oxidize in air at about 200°C. The oxide is said to be adherent and prevents further oxidation until the temperature is raised. Columbium is reported not to become brittle on heating in air for short periods like tantalum because the oxide film prevents further reaction. The columbium oxides dissolve into the metal when heated in a vacuum at temperatures between red heat and 1200°C. Above 1200°C the oxides are reported to evaporate. Columbium, in the oxide free state, is an active getter in vacuum tubes. 2. Preparation and Structure: G. Grube, O. Kuba-schewski and K. Zwiauer15 have studied the decomposition of Cb2O5 in argon. They find that CbO2 is formed slowly at 1150° and rapidly at 1300°C. Cb2O5 can be reduced to CbO2 by moist H2. The reduction of this dioxide to the lower oxides is rapid. By varying conditions CbO2, Cb2O3, CbO and Cb2O can be formed. CbO2, CbO and Cb2O have independent lattices while Cb2O3 is a stoichiometric mixture of CbO and CbO2. In a later paper O. Kuba-schewski" has shown that CbO is cubic and that CbO2 has a rutile type of lattice. The most exhaustive work on the oxides of columbium is that of G. Brauer.8 The number and constitution of the solid phases in the binary system Cb-O were investigated by the preparative, analytical and X ray methods. He indicates that only the oxides Cb2O5, CbO2 and CbO exist with limited regions of homogeneity. The solid oxide Cb,O, exists

Jan 1, 1951

By D. W. Akins

The case history of a combination gas and water flood instituted early in the life of a field is described. It was recognized from the beginning that recovery would be low, under normal production methods, from the several separate, thin beds of porous limestone which had such poor native permeability that the wells would not produce without acidizing. It is thought that the combination flood gives excellent recovery efficiencies for this type of reservoir. The chief function of the gas flood now is to maintain bottom hole pressure in that part of the reservoir which will not be affected by the water injection program until later. Present estimates are that ultimate recovery will be approximately double the recovery expected under the original solution gas drive mechanism. Wide (80-acre) spacing has enhanced the economics of the over-all field program by the elimination of more than $5,000,000 worth of unnecessary wells. INTRODUCTION The Haynesville Pettit Field was discovered in November, 1941, with the completion of T. L. James Co.'s Callie Akin No. 1, located in the W 1/2 of the NE 1/4 Section 27-23N-8W, Claiborne Parish, La., at an approximate depth of 5,300 ft. The center of the field is about two miles north and a mile west of the town of Haynesville, La. A total of 189 producing wells have been drilled on 14,500 acres. An additional nine dry holes were drilled. The Louisiana-Arkansas state line divides the field so that 34 wells are in Arkansas and 155 are in Louisiana. Governmental sections are divided equally to contain eight production units making 80-acre spacing where the sections are regular. Along the state line in Louisiana there is an irregular tier of sections, each of which contains approximately 456 acres. When divided equally into eight production units the spacing is 57 acres per well for these three and a fraction productive sections. All Haynesville Pettit wells are identified by their section number followed by their position in the section as determined by numbering the production units counterclockwise beginning in the upper right corner of the section. All of the producing wells in the field except two are unit-ized and operated by the Haynesville Operators Committee. These two wells are located in the extreme northern part of the field in Arkansas. Efforts to bring them into the Unit are being continued. The Pettit operators organized the Haynesville Operators Committee and the field became a unit operation May 1: 1944. GEOLOGY The Pettit is Lower Cretaceous in age and is defined as the uppermost porosity in the Sligo formation. The unitized area known as the Haynesville Pettit Field is composed of two separate zones of porosity approximately 40 ft apart and without connection between the two, except in wells. These zones are known as the Upper Pettit and the Lower Pettit. The Upper Pettit is separated into two sections throughout most of its area by a relatively thin but quite impervious shale streak. These are called the "A" and "B" sections of the Upper Pettit. All of these lenticular zones are composed of coquina limestone and a small percentage of normal oolites. The large

Jan 1, 1951

By E. M. Elkin, J. H. Schloen

All known methods of treating and recovering the various components of copper refinery slimes are discussed. The slimes treatment processes presently used by five copper refineries are described and flowsheets are given. A bibliography is appended. THE recent increase in anode copper production at the Noranda smelter placed an additional load on the Montreal East plant of Canadian Copper Refiners Limited. The consequent increase in the amount of anode slimes led to a careful consideration of the various methods of slimes treatment. A critical review of these methods is presented here in the belief that it would be of interest to copper refiners. No such review has yet been published, although a number of individual flowsheets have been described. To make the information of greater value, a questionnaire was sent to all copper refineries. While the overall response to the questionnaire was not entirely satisfactory, the foreign refiners gave much detail in their answers and expressed willingness to cooperate further if needed. Although the data below do not cover all refineries, it is thought that examples of all flowsheets currently in use are included. A comparison of these flowsheets should prove of interest. The paper is divided into two parts. One part deals with a discussion of all known methods of treating and recovering the various components of the slimes. The second describes the processes now in use at various copper refineries for the treatment of raw slimes up to the production of Doré bullion. A brief description of the two main electrolytic methods of parting Doré bullion is appended. Since the first commercial application of electrolytic refining of copper by J. B. Elkington in 1865, the treatment of anode mud or slimes has undergone many changes. The object of treating slimes has always been the recovery and separation of precious metals. With the passage of time, increasing attention has been paid to improvement in recovery and quality of byproducts, mainly selenium and tellurium. The original methods of treatment were crude, although relatively direct. For example," the slimes were fused on a dolomitic clay bed in a reverbera-tory furnace; the slag was removed and the resulting metal cupelled to silver and gold. In another process, the dried slimes were oxidized and scorified on the lead bath of a cupelling furnace. Later, these pyrometallurgical methods were modified by sulphuric acid leaching, followed by fusion with niter in a reverb eratory furnace. A preliminary oxidizing roast improved the leaching step. Straight hydro-metallurgy has not found practical application. No matter what method is employed, the result is Dorb metal—a silver-gold alloy, with or without the metals of the platinum group. Characteristics of the Slimes Anode slimes are made up of those components of the anodes which are not soluble in the electrolyte. They contain varying quantities of copper, silver, gold, sulphur, selenium, tellurium, lead, arsenic, antimony, nickel, iron, silica, etc. The color of raw slimes is grayish black and the particle size is —200 mesh. However, Baraboshkin and Gaev (Ref. 1, p. 37), describe light gray colored slimes from the electrolysis of secondary copper which are composed predominantly of sulphates and of basic salts. In anode furnace practice primary and secondary metals are not usually segregated and such slimes do not commonly occur. Copper in the slimes originates to a large extent in the cuprous oxide of the anodes and is therefore dependent on the oxygen content of the latter: Cu2O + H2SO4 + CuSO4 + H2O + Cu [I] Opinions have been expressed that some of the copper is formed by decomposition of cuprous to cupric sulphate Cu2SO4 ? CuSO4 + Cu PI The product of either reaction is very finely divided. Depending on the slimes composition, much of the copper is combined with sulphur, selenium, tellurium, etc. The balance of the copper is in the free metallic state.

Jan 1, 1951

By I. Haskett

ABSTRACT The Redwater pool, in north-central Alberta, was discovered in October, 1948, and at the end of 1950 was about 85 per cent developed. Production is from a Devonian coral limestone reef and is found at an average depth of 3,150 feet. The pool has an estimated .productive area of 37,000 acres, and a maximum oil zone thickness of slightly over 200 feet. Wide variations exist in both type and values of porosity, with the average value determined for the reservoir being 6.16 per cent. The total stock tank oil in place is about 1.24 billion barrels. The crude has a saturation pressure of 485 psig, com-pared to the original value of 1,050 psig for the reservoir, and the principal recovery mechanism so far has been the natural water drive. Cumulative production to January 1st, 1951, has been about 15.5 million barrels. These withdrawals resulted in a reservoir pressure decline of slightly over 100 psi. Predictions of future behaviour indicate that the water drive will not be of sufficient magnitude to maintain the reservoir above saturation pressure past 1953, and that, with natural depletion, the pool will be produced largely by a solution gas drive. Pressure maintenance by water injection has been considered and the possibility of raising the pressure to the original value of 1,050 psig is also discussed. INTRODUCTION WITH the discovery of the Lewduc pool, early in 1947, the Devonian limestone became the major geological formation contributing to Western Canada's rapidly growing petroleum reserves. Redwater, discovered in October, 1948, is now the largest known Devonian pool in Canada, with a recoverable reserve estimated at approximately 500 million barrels. Until recent years, the knowledge of the nature and characteristics of this type of limestone reservoir has been comparatively limited.

Jan 1, 1951