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By B. Ladanyi, B. Hoyaux

A first requirement for assessing the safety of a rock slope, either artificial or natural, particularly if the rock behaves in an elastic-brittle manner, is a propw knowledge of the distribution of stresses within the slope and their development with time and advancement of workings. There are a uumber of different methods that can be used, in principle, for obtaining this information, such as (1) stress measurement in situ, (2) photoelasticity,1 (3) analytical methods, and (4) numerical methods. In cases where the stress distribution cannot be obtained by classical analytical methods, as in the one considered herein, either photoelasticity or numerical methods are norrnally used. However, actually, due to the rapid development of high-speed computers, increasing preference is given to the numerical methods. Among the latter, the most successful ones, to date, for determining the stress distribution in slopes have been (1) the finite difference method,2-4 (2) the lumped parameter method,5 and (3) the finite element rnethod.6 However, for determining gravity stresses in a natural rock slope, only the finite difference method has been used until now.7,8 A solution of the same problem by the finite element method will be shown here in which a particular reference is given to the development of tensile fracture zones within the slope. FINITE ELEMENT METHOD The stress analysis using the finite element method originates from an extension of the matrix analysis of structures.9 It consists in decomposing a continuous medium, assumed herein to be homogeneous and isotropic,

Jan 1, 1972

By S. C. Miller, R. D. Carter, H. G. Riley

A merhod based on theoretical considerations is presented for estimating srabilized performance of gas wells from data obtained from short-term flow tests. Calculated results are presented for II gas wells evaluated with four-to five-hour flow tests. For seven wells on which comparison war available, satisfactory agreement between measured and calculated performance was obfained. Although the final results are favorable, diflerences in flow capacity k,h obtained from flow tests conducted at different rates were observed. This inconsistency did not seriously affect the calculated results in the test wells. INTRODUCTION In order to negotiate the sale of gas from newly drilled gas wells and to plan otherwise for the orderly exploitation of gas reserves, it is necessary to predict the long-term withdrawal rates physically possible for each well. Since low permeability gas wells do not stabilize except after a lengthy period of flow, the operator is unable, without wasting large amounts of gas, to measure directly the stabilized performance prior to contracting for sale of the gas and connecting the well to a pipeline. This problem is not new, but is one which has recently received considerable attention because of the rising importance of natural gas in the economy and new emphasis on the development of gas reserves. Thus, a method is needed for estimating stabilized performance from information obtained from short-flow tests in which stabilized flow is not attained. Such short tests would be suitable for wells not connected to a pipeline since the tests require only small amounts of time, labor and production of gas. Other requirements of such a method are short calculation time and use of a model of the gas well and its reservoir drainage area which is as consistent as possible with present theoretical, experimental and field knowledge. The method presented herein was developed in an effort to meet these requirements. It has been named the "two-Row method". In addition to the section on the evelopment and description of the method, a description of field application of the method and a discussion of well test results are in- cluded. Finally, there is a section on an additional consideration prompted by some of the test results. The Appendix contains certain details of the derivation of the two-flow method and a sample calculation for Test Well 1. MATHEMATICAL DEVELOPMENT AND DESCRIPTION OF THE METHOD The method is based upon the following assumptions: (1) The system can be described as a finite cylindrical gas reservoir produced by a single cylindrical well located at the center of the reservoir. Flow is strictly radial. (2) Flow obeys Darcy's law except very near the well where a quadratic flow law is assumed. (3) The system is also characterized by a laminar flow skin factor. The use of a quadratic or turbulent flow term is more fully justified in Refs. 1 and 2. When such a system is opened at the wellbore to constant rate flow from an initial condition of uniform pressure distribution, the radius or drainage quickly passes through the non-Darcy flow and "skin effect" region imrnediately surrounding the well. After the drainage radius or "transient" has passed beyond this region and before it reaches the external boundary, drawdown can be expressed by'." After the drainage radius reaches the external boundary, stabilized flow exists and drawdown is given by'," Eqs. 1 and 2 form the basis for the method. The validity of these equations will now be examined in the light of some well test data previously available. It has been found by actual measurements on gas wells that the relationship between drawdown and flow rate, that is the performance curve, is closely approximated by' Eq. 3 is a straight line on log-log graph paper. Eq. 2 and, for any lixed value of time, Eq. 1 plot as a curve which is slightly concave upward on log-log graph

By H. M. Fisher, R. E. Snow, S. C. Sun

RAPID colorimetric estimation of the amount of transparent minerals has been applied success- fully to instream mill products at the cryolite flotation plant of the Pennsylvania Salt Mfg. Co., Natrona, Pa. Apparatus used was a Bausch & Lomb monochromatic colorimeter. The optical assembly of the colorimeter, shown in Fig. 1, passes light from a 6-v, 32-cp, prefocus incandescent lamp through two condensing lenses, a heat-absorbing filter, an aperture-control diaphragm, an interference filter, and a sample container. The light finally strikes a barrier-layer photocell and the output current of the photocell is measured by a galvanometer. The engraved index line of the collective lens, located between the lamp and the galvanometer, is imaged upon the transmission scale of the colorimeter by the galvanometer mirror. Reflected light from the galvanometer mirror, the position of which is determined by the output current of the photocell, enables reading of percent transmission throughout the mineral sample being tested. The colorimeter, Fig. 1, was standardized as follows: 1-The colorimeter was connected to a Sola constant voltage transformer, A, of 6-v output operated from a 110-v ac supply. 2-Turning the toggle switch, C, illuminated the transmission scale, B, and the control knob, D, was adjusted with a blank in the optical system until the hairline on the scale read zero. 3-A No. 430 interference filter, E, was inserted into the filter channel, F. &--A test tube, G, containing 7 cc of distilled water was pressed into the sample space, H, and the fine adjustment knob,

Jan 9, 1954

Receipts, Jan. 1, 1914, to Dec. 31, 1914 Cash on hand, Jan, 1 1914 General Funds $1,843.54 Special Funds 1,039.00 $2,882.54 Initiation Fees $8,500.00 Annual Dues 44,060.84 Life Memberships 1,950.00 54,510.84 Sale of Transactions 4,798.87 Sale of Binding 10,029.06 Sale of Bulletins and Pamphlets 2,508.25 Sale of Special Editions 1,290.74 Sale of Advertising 8,783.29 Sale of Library Books 54.97 Sale of Pins and Fobs 347.65 Miscellaneous Sales 21.66 Space Released 5,943.00 Interest on Investments and Bank Deposits 230.15 34,016.64

Jan 1, 1917

By N. J. Themelis, A. W. Last

Autogenous grinding, broadly defined as the self-comminution of ore in a cylindrical tumbling mill without using auxiliary grinding media, was first applied in the early 1930s in the so-called Hadsel mills (Hardinge, 1955) as seen in [Fig. 1], mainly used in gold mines on the US West Coast. A number of these mills were built in sizes of 6.1 to 7.3 m (20 to 24 ft) in diameter and 1.5 to 1.8 m (5 to 6 ft) in length but were later abandoned because of operating and maintenance problems. Further development of autogenous grinding led to the Hardinge wet grinding mill, the Aerofall dry grinding mill, and others. A tabulation of all known autogenous and semiautogenous installations around the world was sent recently to the authors by G. J. Kushner of Koppers Metal Processing Division. In the size range of 7.3 to 9.8 m (24 to 32 ft) in diameter, it lists a total of thirty mills used in the processing of copper ores. An additional 120 units are listed in the same size range for the processing of iron ores and other minerals. The largest mills in existence are 11.1-m-diam X 4.6-m-long (36-ft-diam X 15-ft-long) mills in Minnesota, powered by 8950 kW (12,000 hp) motors and used in taconite beneficiation. It is indicative of the phenomenal growth of this technology that none of the above 150 mills was in operation prior to 1964. Grinding rates in autogenous mills are very sensitive to variations in ore breakage characteristics and even with fairly constant ore type can exhibit large fluctuations in throughput. Autogenous grinding also depends on availability of a fraction of the ore to be ground in the form of "competent" rock which can withstand to a certain degree the tumbling and impact action within the mill. If the ore is uniformly "soft", it will soon be crushed into small size particles and further grinding, which depends on the existence of large size rocks in the mill, becomes impossible. This apparently is the case with most copper ores, as discussed by Bachman, Last and Nabbs (1970) who described pilot testing of a large number of ore samples from Kennecott 's Chino Mines. Under circumstances where autogenous grinding is impractical, it is possible to realize most of its benefits by utilizing large grinding balls in place of the "competent" rock and practice what has become known in the mining industry as semiautogenous grinding. In semiautogenous grinding (SAG), a grate discharge tumbling mill is fed the product of a primary crusher or run-of-mine rock. Forged steel or alloy grinding balls, in proportions ranging from 3.10% of the total volume in the mill, are used to complement the self-grinding ability of the ore. Semiautogenous circuits can be either designed for single-stage closed-circuit grinding which produces a final product for recovery processes, or for two-stage operation in which the SAG mill product is fed to a second stage of grinding in a conventional ball mill.

Jan 1, 1981

By F. V. Tideswell

THE suggestion, which appears to find increasing favor, that the elementary composition of coals should be used as the basis of their classification, makes it important that our methods of expressing this elementary composition should be refined. It is doubtful whether many published ultimate analyses of coals are accurate within 1 per cent. carbon. Many are probably as much as 3 or 2 per cent. in error in carbon. It is, perhaps, not always realized that the extreme variation in carbon content over the range of bituminous coals is but 16 per cent. An error of 1 per cent. in carbon on analysis corresponds therefore with one-sixteenth of the whole range, an amount which is unjustifiable even for the application to commercial needs of a system of classification based on ultimate analysis. For the purposes of scientific research such limited accuracy is of little value. The object of this paper is to examine one important source of error in ultimate analysis, that due to the occurrence in coal of inorganic material, and to emphasize the need for the quantitative measurements (of carbon more particularly) on coals to be rendered independent of their contents of such material.

Jan 1, 1928

The following additional notes about prices of coal in the early days were collected from the sources shown. Several data about canal tolls and labor conditions are also included. 1796, November: "The inhabitants lay it (coal) in at about 3;d. (83¢) sterling per bushel." (Ref. 73: p. 147.) 1804, March: "In March, 1804, I sent a person accustomed to mining to open the vein of coal, and to load the ark, which was accomplished, and early in the month of April the ark, with the first cargo of bituminous coal passed safely the Conewago falls, and landed her coal at Columbia, Lancaster County. The coal was sold at 31¢ per bushel at Columbia." ($7.75 per ton). This was the first coal from Clearfield County near the town of that name. (Letter from William Boyd: Pa. Senate Jour., 1833-1834, vol. 2, p. 561.) 1807: The following notice appeared in the Bedford Gazette: "Huntingdon, June 4. "Stone Coal.-Such of the farmers as wish to make experiment with stone coal as a substitute for plaster, in manuring their Indian corn, may be supplied with coal gratis upon application to Peter Hughes, at Mr. Riddle's mines, on the Raystown Branch. The proprietor of the mines offers not only to refund the carriage, but to pay the expenses of applying the coal, if upon a fair experiment it is found to be inferior to plaster, which now sells at two dollars per bushel." (Ref. 735: p. 708.) 1809: Coal was delivered in Pittsburgh at 5¢ per bushel or $1.25 per net ton. (Ref. 444: p. 50.) Coal was delivered in Connellsville at 4¢ per bushel or $1.00 per net ton. (Ibid.: p. 185.) 1810: Pittsburgh.-"Coal may be obtained of a good quality, ready delivered at 5¢ a bushel." ($1.25 per net ton.) (Ref. 315: vol. 2.) 1811: "It (coal) costs 6¢ a bushel at your door." ($1.50 per net ton.) (In Pittsburgh.) (Ref. 277: 1811, p. 68.) 1814: At Columbia, now part of Donors, Washington County, Pa., an original sale of lots in 1814 gave the purchaser of each lot the right to mine stone coal gratis from a bank on the property. (Ref. 410: vol. 1, p. 547.) "A laborer is able to dig upwards of 100 bushels per day." (4 tons.) (Ref. 277: 1814, p. 109.) A paper about the early newspapers in West Virginia says: "Collier Wanting. From 10 to 20 steady and industrious men, who understand digging coal, may obtain high wages in Kenhawa for that business, if immediate application is

Jan 1, 1942

In Bulletin J 1, recently published by the Traylor Engineering and Manufacturing Co., it is stated that the largest jaw crusher ever built has a jaw opening of 66 in. by 86 in. and weighs 680,000 1b. It is said to be in successful operation and has a capacity of 490 tons per hour crushed to 12-in. size; 415 tons per hour crushed to 10-in. size; 330 tons per hour crushed to 8-in. size. The same company, in its Bulletin G 3, gives a diagram showing the percentage of a crushed product which will pass through screens of different sizes of perforation from a machine adjusted to a certain size screen. This diagram is reprinted below for the information of our members and the following notes apply to it. In order to prevent an excessive amount of fines, crushing machines are usually set so that 15 per cent. of the product will be larger than the maximum size desired, and this oversize is then recrushed by smaller

Jan 10, 1916

By Robert R. French

Beneficiation of low-grade nonmetalliferous mineral deposits by electronic color sorting is currently undergoing a period of rapid development. For those deposits in which the specific gravity, strength, soundness, and electric and magnetic properties of the ore and gangue are similar, the possibility of separation by color, reflectivity, or fluorescence offers hope for the utilization of much poorer quality ore than previously thought possible. For those ores in which consistency of color is required or for which a dry process is preferred, the electronic color sorter is particularly adaptable. Tests performed late in 1966 and reported herein are the result of the continued interest in the dry beneficiation of gypsum and the development of higher capacity color sorting machines. The ore used was a discarded fraction that consisted of – 1 ¼ -in, to + 3/8 -in. ore containing 63.1 to 71.4% (average 67.7%) gypsum. This material was contaminated by various amounts of both carbonate rock and shale, but no pretreatment was performed other than screening to closer size limits before electronic color sorting.

Jan 4, 1969

By Otto H. Henry, Edward L. Badwick

Up to the present time, according to Hansen1 and Haughton,2 the constitution of the bismuth-indium system has not yet been published. The generally accepted value for the melting point of indium, as listed in most scientific journals and textbooks, is 155°C. Recently, however, H. M. Davis3 has determined a new melting point for indium of 257.3°C ± 0.1°C. Since bismuth has been used in preponderance as an alloying element in the majority of low-melting alloys, and indium has been used with considerable success in lowering the melting points of low-melting alloys,4,5 an interest has been developed in the application of these two metals, making a complete knowledge of the constitution of its alloys desirable. In this investigation, the melting point of indium reported by H. M. Davis has been found to be substantially correct. A value of 157.2°C ± 0.5°C was determined. The two eutectic compositions have been placed at 34 and 50 pct bismuth with their respective eutectic temperatures as 72.4'C ± 0.5°C and 90°C ± 0.5°C. The bismuth-indium phase diagram has been determined by thermal and microscopic analysis and is shown in Fig I. THERMAL ANALYSIS Granulated bismuth and indium shot, having a guaranteed purity of 99.99 and 99.95 pct, respectively, were melted in a carbon crucible by means of a laboratory heater. To minimize oxidation of the alloys, a heavy mineral oil (Nujol) having a high flash point was used. A platinum-platinum rhodium thermocouple 0f fine wire encased in a thin-walled Pyrex glass tube, about 1/8 in. in diameter, was connected to a type K potentiometer for temperature measurements. The thermo-couple was calibrated against the melting points of tin, bismuth, and the silver-copper eutectic and the boiling point of water. Since the temperature readings could be reproduced to 0.2°C, and the accuracy of the temperature standards in this range is slightly better than 0.5°C, it seems reasonable to assume that the temperature readings are reliable to approximately 0.5°C. The cooling rate could be controlled by varying the amount of insulation surrounding the crucible. Three or four layers of asbestos paper, 1/16r, in. thick, sufficed to give a cooling rate of from 1° to 2°C per minute. Undercooling was slightly noticeable. This condition was minimized by vigorous stirring, since the advantage of using slower rates was too small to justify the greater expenditure

Jan 1, 1947

By G. S. Hume

IN THE autumn of 1922, British Petroleums Ltd. found oil of 14° Baume in a sand 17 ft. thick in its No. 2 well at Wainwright, 120 miles southeast of Edmonton. This greatly encouraged drilling in the Wainwright-Irma area and this past summer several wells were started, three of which are now completed' and the others still being drilled. Of the three completed, No. 4 British Petroleums contains oil; No. 3 British Petroleums gave a show of oil but so far has not given much hope of becoming a commercial producer; and No. 1 Maple Leaf Co., as now finished, is a gas well with a daily flow of gas of about 2,500,000 cu. ft. at a closed pressure of 700 lb. An oil sand is reported . to have been encountered below the producing gas horizon, in the Maple Leaf well but has not been tested. Operations on this well have been suspended for the winter. The Geological Survey of Canada this past summer undertook a geological investigation of this area with the object of defining the structures with which the oil and gas are associated. The 'structure underlying this portion of the plains was supposed to-be a terrace with the western edge dipping into the Alberta syncline and with a monoclinal structure to the east. By accurate leveling on known geological horizons on this terrace, a number of gentle folds were located and it is in these folds that the oil land gas are found. At the British Petroleums Nos. 2-and- 4 oil wells, the western flank of the fold has a gentle dip of only a few feet to the mile to the southwest whereas the strata on the eastern flank dip to the northeast about 40 ft. in 1/2 mile, the strike of the fold being northwest and southeast. How far oil will be found down the gentle southwest dip is unknown but probably the productive part of the fold will not exceed 1 mile in width, and possibly less. No. 3 British Petroleums well is on the strike of the fold 1 mile southeast of the two oil wells. The sand that was productive in Nos. 2-and 4 wells was not present in this well or, if present, was very thin and unproductive. Two other sand horizons, however, were encountered. The upper of these

Jan 3, 1925

By S. R. Elliott

The Spies mine is located in the eastern half of the northwest quarter of section 2443-35, near the village of Iron River, Mich., and is operated by The Cleveland-Cliffs Iron Co. Speaking generally, the jasper hanging walls of most of the deposits in the Iron River district dip at a steep angle and are exceedingly strong, standing without caving, even after mining has been continued for years and after large excavations have been made. The ore is fairly hard and tough and contains no slippage faces. It occurs in large masses, which are often egg shape, with the smaller end down, and can be mined in open stopes without the use of timber. The common system employed in the district is the shrinkage stope, or some modification of it. The Spies system was developed to utilize the advantages of the open-stope method and to avoid the disadvantages of the shrinkage stope. The Spies shaft was sunk to such a depth that the bottom, or third level, was in rock at a short distance below the bottom of the orebody, the shape and extent of which had been previously determined by diamond drilling. This is shown in plan and by the ore contours at the first, second, and third levels in Fig. 1, and in section in Figs. 2 to 7. In order to provide the necessary mills and raises, three parallel crosscuts were driven on 25 ft. centers on the bottom level. Drifts on the second and first levels were driven to the orebody and followed the foot wall in a southerly direction to the end of the ore. In crosscut 3, Fig. 1, raises 30, 40, 50, and 60 were carried up on the foot wall and holed into the second level, and were ultimately carried up and holed into the first level. These foot-wall raises carry the pipes and ladders, provide a means of ventilation, and furnish safe communication to all working places in the stope. From crosscuts 1, 2, and 3, raises were completed to a height of 32 ft., or the elevation of the first sublevel. They were then connected by east and west crosscuts to the main foot-wall raises 30, 40, 50, and 60, shown in Fig. 4. The crosscuts in raises 1 to 5 were then driven west and the

Jan 1, 1923

By S. R. Elliott

The Spies mine is located in the eastern half of the northwest quarter of section 2443-35, near the village of Iron River, Mich., and is operated by The Cleveland-Cliffs Iron Co. Speaking generally, the jasper hanging walls of most of the deposits in the Iron River district dip at a steep angle and are exceedingly strong, standing without caving, even after mining has been continued for years and after large excavations have been made. The ore is fairly hard and tough and contains no slippage faces. It occurs in large masses, which are often egg shape, with the smaller end down, and can be mined in open stopes without the use of timber. The common system employed in the district is the shrinkage stope, or some modification of it. The Spies system was developed to utilize the advantages of the open-stope method and to avoid the disadvantages of the shrinkage stope. The Spies shaft was sunk to such a depth that the bottom, or third level, was in rock at a short distance below the bottom of the orebody, the shape and extent of which had been previously determined by diamond drilling. This is shown in plan and by the ore contours at the first, second, and third levels in Fig. 1, and in section in Figs. 2 to 7. In order to provide the necessary mills and raises, three parallel crosscuts were driven on 25 ft. centers on the bottom level. Drifts on the second and first levels were driven to the orebody and followed the foot wall in a southerly direction to the end of the ore. In crosscut 3, Fig. 1, raises 30, 40, 50, and 60 were carried up on the foot wall and holed into the second level, and were ultimately carried up and holed into the first level. These foot-wall raises carry the pipes and ladders, provide a means of ventilation, and furnish safe communication to all working places in the stope. From crosscuts 1, 2, and 3, raises were completed to a height of 32 ft., or the elevation of the first sublevel. They were then connected by east and west crosscuts to the main foot-wall raises 30, 40, 50, and 60, shown in Fig. 4. The crosscuts in raises 1 to 5 were then driven west and the

Jan 1, 1923

By Paul B. Crawford, Bobby L. Landrum

Electrical model studies have been made of the effect of unsymmetrical vertical fractures on production cnpacity. The study war restricted to the case of two vertical fractures originating at the well and extending out into the reservoir for various distances and orientations. The ratio of the short fracture length to the long fracture length was varied from 0.1 to 1.0 in successive studies. The angle separating the two fractures was varied from 0 to 180 degrees. The most effective fractured patterns were obtained when the fractures were oriented 180 degrees. It was found that when the total fracture length was equal to 0.2 of the drainage radius, the increase in production capacity may vary from approximately 1.55 to 1.95. depending upon the ratio of the short fracture length to the long fracture length and the angle separating the two fractures. If the total fracture length were 0.5 of the drainape- radius. poroduction cavacitv., increases varving 1.95 to 2.65 were observed, depending on the angle between the fractures and the ratio of the .short fracture length to the long fracture length. The fractured well production capacity decreased aPproxin?ately I to 10 per cent as the angle separating the fractures was varied from 180 to 90 degrees. It was dependent on the particular fracture. The commercial application of reservoir fracturing techniques has caused an interest to be expressed in the effect of vertical and horizontal fractures on production capacities. Some data have indicated that substantially vertical fractures may be created in some reservoirs. It is believed that the fractures are possibly unsymmetrical, that is, one fracture may extend due north 100 ft and a second fracture on the opposite side of the well may extend south or southeast only 50 ft. The total fracture length, ratios of the lengths and angles of separation may vary considerably. For this reason variations in these factors were considered. The study was restricted to the case of two vertical fractures originating at the well and extending out into the reservoir. It is not known that such will occur in a reservoir, but the results may be of value in interpreting or studying fracturing applications. PROCEDURE An electrical model was used in this study of the effect of unsymmetrical vertical fractures on production capacity, A wooden tank 30 in. in diameter and approximately 2 in deep was coated with paraffin wax. A copper band was situated at the periphery and a single wire electrode was positioned at the center of the tank. The ratio of the reservoir radius divided by the

Jan 1, 1956

By G. Thodos, E. M. Phillips

A reduced density correlation for the gaseous and liquid regions of ethane has been developed from all the experimental PVT data available in the literature for this substance. Saturated vapor and liquid densities reported for methane, propane and n-butane were utilized to establish the rectilinear-diameter line for ethane. This line was then used with the saturated-vapor data for ethane to produce the saturated envelope for this substance. Most of the saturated liquid data reported for ethane were found to be inconsistent with the saturated-liquid curve resulting from this study. The correlation covers the regions 0.65 < TR <6.0 and 0 < PR < 15. For regions for which no experimental data have been reported, the isobars of the correlation were extended by the use of the theorem of corresponding states. This correlation is presented both on rectilinear and log-log coordinates. INTRODUCTION At the present time, there is considerable interest in the exact PVT behavior of pure hydrocarbons. Matschke and Thodos12 recently have developed a comprehensive compilation of experimental vapor and liquid density data for methane. Similar studies have been conducted for other pure substances.4, 7 In the present study the density data available in the literature for gaseous and liquid ethane have been critically reviewed and utilized to develop a reduced density correlation for this substance. SATURATED LIQUID AND VAPOR DENSITIES The rectilinear-diameter rule proposed by Cailletet and Mathias5 states that the sum of the saturated liquid and vapor densities of a pure substance is a linear function of temperature which passes through the critical point. This law, although empirical, has been well established by continued verification since its conception in 1886. This rule has been applied to the saturated density data of methane,6 propane 17 and n-butane8 to produce the straight lines presented in Fig. 1 which pass through the critical point TR = 1.00, Pr = 2.00. The rectilinear-diameter relationship for ethane was established from the lines for the other substances in Fig. 1 by plotting the quantity p~, + PR against zc for different reduced temperatures. A critical analysis of the saturated vapor and liquid densities reported for ethane by Kuenen,10 Maass and McIntosh,11 and Sage, Webster and Lacey18 revealed that the values of PR1 + pr resulting from these studies were inconsistent with the rectilinear-diameter line obtained from the reliable density data of the other substances, as shown in Fig. 1. In the analysis of these data, reduced quantities were calculated using the critical con-

By Park Wright

THE fact that granite has been encountered by the drill by those in search of oil and gas in Kansas is becoming more and more a matter of interest, not only to the oil producer but to everyone directly and- indirectly connected with the oil business. It is, therefore, time that someone should offer an explanation for its presence and form a definite opinion based on all the data available in regard to the effect it will have on the future prospects of the area affected as probable oil- and gas-producing territory. At present there seems to be no consensus of opinion. To the writer&apos;s knowledge, only one attempt has been made to explain; by Erasmus Haworth, State Geologist of Kansas, in a paper entitled Bulletin, Number Two on Crystalline Rocks in Kansas. The purpose of the present paper is to offer a theory for the presence of the granite and to draw conclusions which will aid in determining the prospective oil and gas value of the area in question. In taking up this question it will be well to consider first the tests in which granite has been reported, with a general history of each. The accompanying map (Fig. 1) shows the location of all wells described. The first well in which granite was reported was that drilled just south of Zeandale in section 27, township 10 south, range 9 east, by a local company which had taken up a block of leases in that vicinity. At a depth of 958 ft. (291.99 m.) granite was encountered. Drilling was continued to a depth of 1,093 ft. (333 m.) in the same formation, at which depth the hole was abandoned. Another well was started 1 mile east of the first well and the same rock was found at a depth of 945 ft. (288 m.). Drilling was continued to a depth of 1,200 ft. (365.76 m.) with the bit in the sane formation. Samples of the drill cuttings have been examined and they are undoubtedly crushed granite. The next well in which granite was found was that drilled by DeLaat and Shepard, independent wildcat operators, in section 34, township 19 south,. range 7 east, near the town of Elmdale. Granite was struck at 1,707 ft. (520 m.) and the hole was abandoned in the same formation at 2,525 ft. (769 m.). Mr. DeLaat watched the drilling of this well very closely and he has described to me in some detail the formation immedi-

Jan 8, 1917

By A. Boltax

The effect of cold work on the electrical and magnetic properties of solution-treated and aged Cu-Fe alloys was studied. The electrical resistivity of solution-treated and of aged Cu-1.7 wt pct Fe samples increased by up to 25 and 140 pct, respectively, follozcing a 55 pct reduction of area. The principle contributions to the increase of resistivity of the solution-treated and aged samfiles are examined. Measurements of the .saturation induction, coercive force, and stacking fault Probability are also reported. 1 HIS paper is the second in a series which examine the behavior of the electrical and magnetic prop- erties of Cu-Fe alloys. The first paper1 was primarily concerned with the effects of thermal treatment whereas the present paper examines the effects of cold work on solution-treated and aged Cu-Fe alloys. The effects of cold work on the transformation of the coherent fcc iron precipitate to the ferromagnetic form2 and the coercive force and remanence3 have already been described, but the unusually large increases in resistivity (up to 220 pct) which occlir with cold work have not been reported previously. EXPERIMENTAL METHOD The methods used for alloy preparation, heat treatment, and electrical and magnetic measurements are described elsewhere.1 The specimens, initially 4-in. lengths of 0.0625-in. diam wire, were cold worked by swaging in hardened steel dies. Following each reduction, the samples were etched lightly in hydrochloric acid to remove surface con-

Jan 1, 1962

By J. B. Cohen, S. L. Sass

The nucleation process for y&apos;, Ni3Ti, is shou&apos;n to change from heterogeneous to uniform as the undercooling within the phase boundary increases. As unifornz nucleation beconres copious, (100) alignment is observed in the early stages of aging; howecer. whether or not the alignment is due to spinodal decomposition could not be determined. The presence oi a large-scale reversible segregation of titanium and not the intermediate precipitate y&apos; may be the cause of the initial strengthening in this alloy system and the discrete regions detected magnetically by Ben-Israel and Fine. Evidence is presented for the role of dislocations in the formation at long aging tines oJ the equilibrium precipitate . ThE nickel-rich portion of the Ni-Ti phase diagram consists of a terminal fcc solid solution of titanium in nickel called y.&apos; For temperatures below 1290°C if the solubility limit is exceeded and the temperature is high enough, the excess titanium will combine with nickel to form q, Ni3Ti, a four-layer hcp ordered structure with stacking sequence ABAC."~ Since a change in structure is involved, formation of q is slow and a metastable precipitate referred to as y&apos; forms. This metastable precipitate is an ordered cubic phase in the unconstrained state having a structure similar to Cu3Au (LIZ) with the composition Ni,Ti."&apos;5 From X-ray studies,8, 7 the following sequence is indicated for the precipitation reaction at low temperatures: 1) The appearance of satellites around low-index diffraction spots and splitting of high-index spots, indicating the formation of a slightly tetragonal phase in an imperfectly periodic arrangement. The tetragonal phase is y&apos; or the depleted matrix, depending on the volume fractions of matrix and precipitate (and hence, the alloy composition); tetragonality apparently arises from coherency strains. 2) The appearance of diffraction spots from the equilibrium second phase, q. Ben-Israel and Fine&apos;" used magnetic methods to measure the changes in matrix composition during aging, and to estimate the precipitate composition in a Ni-10.1 at. pct Ti alloy. They observed that the alloy, solution-treated at 1270°C and quenched, contained heterogeneities with discrete compositions which gradually vanished upon aging. It was suggested that y&apos; may initially form as a defect structure with several possible compositions deficient in titanium as compared to Ni3Ti; the composition Ni3Ti develops with prolonged aging. The initial composition was Ni6Ti which would require a very large deviation in stoichiometry for y&apos;. It was noted that at 700°C the precipitation reaction was 80 pct complete after 2 hr and that the volume fraction of second phase was 0.2 after 1 hr. Mihalisin and Decker&apos; suggested that the equilibrium precipitate, besides being nucleated at grain boundaries, could also form at stacking faults. ~errick" examined the formation of in Ni-Cr-Ti, and suggested that a collapsed vacancy cluster within y&apos; acted as a nucleus for intragranular q at high temperatures, 820" and 900°C. After long aging times at lower temperatures, 650" and 750°C, diffraction patterns taken from regions containing ribbons of stacking faults showed spots associated with q. Aging greatly enhances the mechanical properties of these alloys, even at high temperatures, and the metastable precipitate y&apos; is thought to be responsible.&apos; At 600°C the hardness of a Ni-10.1 at. pct Ti alloy increases rapidly during the first 20 hr of aging and continues to increase slowly thereafter.&apos;" At 700°C the hardness reaches a maximum at 1 to 2 hr and then decreases.l1 In this paper the kinetics of formation of y&apos; and the mechanism of transformation to in a Ni-Ti alloy are described. Evidence is presented that suggests that large-scale regions of titanium segregation are present initially and these may have an important influence on the initial strengthening of the alloy: these regions may be the heterogeneities detected by Ben-Israel and Fine. I) EXPERIMENTAL TECHNIQUES The Ni-10.3 at. pct Ti-0.6 at. pct A1 alloy was prepared by vacuum melting. Its chemical analysis is given in Table I. A 0.007-in.-thick strip was sandwiched between foils of the alloy and placed in an open Vycor tube with titanium chips (to get oxygen) and solution-treated for ; hr at 1150°C in a Globar Furnace under a flowing argon atmosphere. Following a quench into an ice-brine mixture at -5" to -2"C, a sample to be aged at temperatures up to and including 700°C was sealed in an evacuated Vycor capsule before being put into a fused salt bath. Aging above 700°C was carried out on bare samples in a salt bath. When removed from this bath and quenched into an ice-water mixture, the sample was normally coated with a thin layer of oxide easily removed with emery

Jan 1, 1970

By B. J. Harlan

The trying times experienced by the steel industry during the past four years have emphasized the necessity of producing pig iron at the lowest possible cost. The trend in both design and operation of blast furnaces has been toward this end rather than toward securing maximum production from a unit or group of units. In the following discussion the trend in design of stacks and auxiliary equipment will be taken up before operation is considered. Although few furnaces have been constructed in the past few years, the trend is definitely toward larger furnaces with a hearth diameter of 24 or 25 ft. Such a furnace is very flexible, being able to produce efficiently any desired daily tonnage between 500 and-1000 or more. The detail of the design of furnace stacks, with the exception of the top, has changed little in recent years. To decrease the velocity of gas through the throat of the furnaces, thereby reducing flue-dust losses, the stock-line areas are constantly being made larger. The top dimensions of the larger furnaces are usually as follows: Stock-line diameter, 19 to 20 ft.; large bell diameter, 14 to 15 ft.; straight section of throat, 5 to 10 ft. high; inwall batter, 1 to 11/4 in. Pe ft. The ratio between the area of the stock line and the area of the large. bell is usually about 1.85 to 1. The stock-line diameter of some smaller furnaces has been increased to that of the hearth diameter of the stack at the expense of the stack batter. Remarkable decreases in flue-dust production are claimed for the so-called " Venturi top," which is achieved by flaring the stock line, in the form of an inverted cone, from a diameter of 17 ft. at 10 ft. below the large bell closed to a diameter of 19 ft. 6 in. at 3 ft. below the large bell closed. Time alone will tell which of the two designs is the better. Because of differences in raw materials, one top may be better for some plants and the other top for other plants. In the search for lower costs and higher efficiencies all operators are, adopting stoves that will maintain maximum hot-blast temperatures with a minimum gas consumption. To meet these requirements, stove capacity of approximately 7 sq. ft. of heating surface and 1.8 cu. ft. of brick volume is required per cubic foot of wind blown per minute.

Jan 1, 1934

By C. G. Dunn

IDENTIFICATION Of the kinds of orientation relationships that may exist among crystals is an important problem in the metallurgical field. As an aid to its solution standard orientations of several orders of twins would be very useful for rapid analyses to discover high-order as well as low-order twin relationships. Although data have been published that give orientations for three generations of twins, these data are not readily useful in the form presented; i.e., in a stereographic projection with one pole per orientation.1 However, accurate data based on calculations of the positions of all {100} planes for such groups would be very useful when presented in tabular form. Therefore, calculations have been made for a group of four orders of twins and the data are listed in Table 3. Illustrative applications are considered briefly, with the idea of pointing out what may be done with large groups of orientations. METHOD OF COMPUTING ORIENTATION OF A TWIN It is common practice to describe the orientation of a crystal in terms of the positions on a stereographic net of the poles of certain planes of the crystal. A single position on a net may be given it terms of two angle variables, two sets of which appear in the following equations: [Polar NetWalff Net x=rsinOocos4pyx=rcos0. y=rsinB,sin4,ny-rsin0.cos¢,, z=rcos0,,z=rsin0.sin4,,] The three numbers x, y and z determine a length r as well as a direction [(0, 0)] in an X, Y, Z coordinate system. In the present case, r will be set equal to (h2 + k2 +12)1/3, where h, k and 1 are the usual Miller indices for a crystallographic plane, a restriction that facilitates determination of the positions of poles of new planes from those of old or known ones,2 as will become evident later. Furthermore, since twinned lattices have certain planes in common, it is an easy matter to determine the orientation of a twin of a crystal if the orientation of the crystal itself is known. We start therefore with a crystal of known orientation (positions of three {100} poles given) and determine where the cube poles occur in first, second, third and fourth-order twins. Methods for identifying twinned lattices and for making the calculations to find the orientation of a twin have been published.3 4 Reference 4 will be illustrated in the calculation of the orientation of a twin, even though this method differs from the one used by the writer in obtaining the following data on four orders of twins. Consider a crystal S and its twin A1, A1 being one of the four {112} or {III}

Jan 1, 1945