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United Engineering Trustees, Inc. GEORGE A. EASLEY H. G. MOULTON A. L. QUENEAU The Engineering Foundation GEORGE D. BARRON E. DEGOLYER A. L. QUENEAU Engineering Societies Library Board H. K. MASTERS A. B. PARSONS, Ex-officio F. F. SHARPLESS H. H. SWEETSER ARTHUR L. WALKER

Jan 1, 1934

By Alan R. Grant

The Cascade Range of Washington can be considered metallogenetically to be primarily a copper province. Structural and chemical data compiled from examination of numerous copper occurrences in the Cascades indicate a marked parallel of physical conditions for sulfide deposition. In probable order of significance, the environmental criteria are as follows: 1) adjacent or subjacent, postregional metamorphism high-level intrusive activity, most commonly in the time range from 15 million years to 30 million years; 2) transverse northeast trending structural belts which obliquely cut the northwest regional trend of the range; 3) potassium-silica wall rock alteration; most important of the alteration suite of silicates are orthoclase secondary biotite and quartz; 4) the intersection of northwest trending and northeast trending structures. The varying geologic environments in the Washington Cascades present a complex setting for the sulfide deposits. The North Cascades are comprised of predominantly pre-Tertiary metamorphic rocks invaded by Tertiary calcalkaline intrustive rocks. The Central Cascades, lying south of the North Cascades, exhibit a younger, higher level geologic environment comprised of predominantly early to middle Tertiary volcanic rocks and subordinate sedimentary rocks which are intruded by Tertiary calc-alkaline igneous rocks. Although the Tertiary Cascade intrusions are dominantly quartz dioritic, copper mineralization is commonly found in rocks of quartz monzonite to granite composition. These acid rocks appear to be the product of late stage magmatic K alteration of the original quartz diorites. Sulfide concentrations are considered to be in part dependent on the degree of K-silicate activity and the intensity of transverse structural deformation.

Jan 1, 1973

By Orville R. Lyons

DISCUSSION Judson S. Hubbard (The Humphreys Investment Co., Denver)—In this very interesting paper several brief references are made to the Humphreys spiral, a device used for cleaning fine coal. In Table I, Plants 76 and 77, data are given on spiral performance treating Raton and Trinidad coal. The fine coal, as fed to the spiral in these instances, is actually a table middling, hence the more easily treated material was previously removed and a large number of particles were present which were difficult to clean. Mr. W. M. Bertholf of the Colorado Fuel and Iron Co. presented a paper in February, 1946, Cleaning Table Middlings from a Coal Washery with the Humphreys Spiral Concentrator, from which I quote: "In considering the results of our tests it should be noted that the feed was our table middling, and that any real separation is a 'moral victory' as there is little material that could properly be called coal and practically no heavy rock, the consequences being that previous attempts to clean the middling have not at all been successful." Referring again to Table I, Plants 72, 73, 74, and 75, these data were obtained by Yancey and Geer and others and presented at the February 1950 Meeting, AIME, in a paper entitled Laboratory Performance Tests of the Humphreys Spiral as a Cleaner of Fine Coal. Results shown for those tests involve all particles from 8 mesh through the colloids, which admittedly is not an ideal situation for spiral feed if much refuse is contained in the —80 mesh or —100 mesh size range. As an illustration of the effect of treating too broad a size range, let us consider Plant No. 75, Kentucky No. 9 seam. Spiral feed was 8 mesh x 0. Now had this been 8x100 mesh the percentage of misplaced material would have been 8.0 pct instead of the reported 15.26 pct. Similar comparisons can be made on the other data presented with respect to the spiral. Other types of equipment show a similar trend in that whenever too fine a size is treated in a given unit process the percentage of misplaced material increases. Since the spiral is working near the finer end of the size range, it will sometimes be advantageous to treat the entire range of —8 mesh material rather than to deslime and make a fair showing on, say, the +80 or + I00 mesh. Desliming is subsequently done in any case in the dewatering or thickening operation. Results obtained by spiraling any given coal depend on factors too numerous and complex to discuss here, but there are strong indications that proper preparation of feed to the spiral can improve results obtained on some of the raw coals tested. This is clearly pointed out at the end of the aforementioned paper by Yancey and Geer. "The spiral is an extremely simple device which involves no moving parts and is constructed almost entirely of unmachined castings. Since it is such an uncomplicated mechanism, operation is simple and virtually foolproof. These characteristics, which go far toward insuring low cost operation, are attractive attributes in any coal cleaning unit." Certain equipment used in conjunction with the spiral has resulted in a decrease in the percentage of misplaced material, notably in actual practice the launder screen which is used to remove objectionable high ash fines from a spiral-washed coal product. Private correspondence with the U. S. Bureau of Mines has intimated that an additional yield of coal is possible by flotation of the spiral middling. Possible future improvements and developments may result from other methods now under consideration. Finally, some compromise must be made between the best metallurgical performance and the best practical or economical results. Mr. Lyons emphasizes in his summary this objective of overall economy in selection of equipment. G. B. Walker (American Cyanamid Co., Stamford, Conn.)—I had the pleasure of reviewing the draft of this paper and my curiosity was aroused by the data given for Tromp plants, in that all of the examples shown appeared to be 2-product separations, whereas all the Tromp plants with which I am familiar have been 3-product units. The data given for plants No. 101 and 102 appear to be taken from Tromp's brochure on his process and represent the results obtained at the Dominale plant in Holland which was operated for many years by Mr. Tromp. The plant, which was designed to treat 58 tons per hr, was sampled while treating 35 tons per hr of 3Y4x-in. coal. Example 101 appears to conform to what would result if the middling product were calculated into the refuse product, while Example 102 represents the calculation of the middling into the coal product. It is believed that Examples 103 and 104 represent the operation of the Willem-Sophie Mine in Holland recalculated on the same basis. In checking the English examples given by numbers 14, 15, 16, and 17, the same procedure seems to have been followed. These results have, apparently, been taken from an article in Colliery Engineering in August 1941, describing the initial operation of the Williamthorpe Colliery of the Hardwick Colliery Co. Two vessels are employed in this plant, one to treat soft coal and one to treat hard coal. Example No. 14 presents the results that could be obtained from the soft coal bath if the middling were calculated into the refuse, and Example 15 the results when the middling is calculated into the coal. Examples 16 and 17 represent the same expedient in the case of the hard coal bath. Of interest to this discussion is the fact that during the past year the Simon-Carves Engineering Co. in England has installed in the Williamthorpe plant their new "Sim-Car" medium cleaning system which is based on magnetic extraction and control and which is licensed under the Heavy-Media Separation Processes patents by the American Zinc, Lead and Smelting Co. This system has been described in the December 1951, issue of Colliery Engineering. It is reported that since the Williamthorpe Colliery was changed from the Tromp system of medium cleaning to the Magneto-Motive method of medium control the opera-

Jan 1, 1953

By Orville R. Lyons

DISCUSSION Judson S. Hubbard (The Humphreys Investment Co., Denver)—In this very interesting paper several brief references are made to the Humphreys spiral, a device used for cleaning fine coal. In Table I, Plants 76 and 77, data are given on spiral performance treating Raton and Trinidad coal. The fine coal, as fed to the spiral in these instances, is actually a table middling, hence the more easily treated material was previously removed and a large number of particles were present which were difficult to clean. Mr. W. M. Bertholf of the Colorado Fuel and Iron Co. presented a paper in February, 1946, Cleaning Table Middlings from a Coal Washery with the Humphreys Spiral Concentrator, from which I quote: "In considering the results of our tests it should be noted that the feed was our table middling, and that any real separation is a 'moral victory' as there is little material that could properly be called coal and practically no heavy rock, the consequences being that previous attempts to clean the middling have not at all been successful." Referring again to Table I, Plants 72, 73, 74, and 75, these data were obtained by Yancey and Geer and others and presented at the February 1950 Meeting, AIME, in a paper entitled Laboratory Performance Tests of the Humphreys Spiral as a Cleaner of Fine Coal. Results shown for those tests involve all particles from 8 mesh through the colloids, which admittedly is not an ideal situation for spiral feed if much refuse is contained in the —80 mesh or —100 mesh size range. As an illustration of the effect of treating too broad a size range, let us consider Plant No. 75, Kentucky No. 9 seam. Spiral feed was 8 mesh x 0. Now had this been 8x100 mesh the percentage of misplaced material would have been 8.0 pct instead of the reported 15.26 pct. Similar comparisons can be made on the other data presented with respect to the spiral. Other types of equipment show a similar trend in that whenever too fine a size is treated in a given unit process the percentage of misplaced material increases. Since the spiral is working near the finer end of the size range, it will sometimes be advantageous to treat the entire range of —8 mesh material rather than to deslime and make a fair showing on, say, the +80 or + I00 mesh. Desliming is subsequently done in any case in the dewatering or thickening operation. Results obtained by spiraling any given coal depend on factors too numerous and complex to discuss here, but there are strong indications that proper preparation of feed to the spiral can improve results obtained on some of the raw coals tested. This is clearly pointed out at the end of the aforementioned paper by Yancey and Geer. "The spiral is an extremely simple device which involves no moving parts and is constructed almost entirely of unmachined castings. Since it is such an uncomplicated mechanism, operation is simple and virtually foolproof. These characteristics, which go far toward insuring low cost operation, are attractive attributes in any coal cleaning unit." Certain equipment used in conjunction with the spiral has resulted in a decrease in the percentage of misplaced material, notably in actual practice the launder screen which is used to remove objectionable high ash fines from a spiral-washed coal product. Private correspondence with the U. S. Bureau of Mines has intimated that an additional yield of coal is possible by flotation of the spiral middling. Possible future improvements and developments may result from other methods now under consideration. Finally, some compromise must be made between the best metallurgical performance and the best practical or economical results. Mr. Lyons emphasizes in his summary this objective of overall economy in selection of equipment. G. B. Walker (American Cyanamid Co., Stamford, Conn.)—I had the pleasure of reviewing the draft of this paper and my curiosity was aroused by the data given for Tromp plants, in that all of the examples shown appeared to be 2-product separations, whereas all the Tromp plants with which I am familiar have been 3-product units. The data given for plants No. 101 and 102 appear to be taken from Tromp's brochure on his process and represent the results obtained at the Dominale plant in Holland which was operated for many years by Mr. Tromp. The plant, which was designed to treat 58 tons per hr, was sampled while treating 35 tons per hr of 3Y4x-in. coal. Example 101 appears to conform to what would result if the middling product were calculated into the refuse product, while Example 102 represents the calculation of the middling into the coal product. It is believed that Examples 103 and 104 represent the operation of the Willem-Sophie Mine in Holland recalculated on the same basis. In checking the English examples given by numbers 14, 15, 16, and 17, the same procedure seems to have been followed. These results have, apparently, been taken from an article in Colliery Engineering in August 1941, describing the initial operation of the Williamthorpe Colliery of the Hardwick Colliery Co. Two vessels are employed in this plant, one to treat soft coal and one to treat hard coal. Example No. 14 presents the results that could be obtained from the soft coal bath if the middling were calculated into the refuse, and Example 15 the results when the middling is calculated into the coal. Examples 16 and 17 represent the same expedient in the case of the hard coal bath. Of interest to this discussion is the fact that during the past year the Simon-Carves Engineering Co. in England has installed in the Williamthorpe plant their new "Sim-Car" medium cleaning system which is based on magnetic extraction and control and which is licensed under the Heavy-Media Separation Processes patents by the American Zinc, Lead and Smelting Co. This system has been described in the December 1951, issue of Colliery Engineering. It is reported that since the Williamthorpe Colliery was changed from the Tromp system of medium cleaning to the Magneto-Motive method of medium control the opera-

Jan 1, 1953

By Malcolm J. McPherson, Michael Hood

Sunnyside Mines, owned and operated by the Kaiser Steel Corporation, are situated near the city of Price, Utah. The complex comprises three adjacent mines, named simply Nos. 1, 2 and 3, all connected underground. Two seams, the upper and lower Sunnyside have been worked. These dip at about 10 percent to the north-east. The surface cover is variable due to the mountainous nature of the topography. The Sunnyside upper seam varies from 5 1/2 ft (1.7m) to 9 ft (2.7m) In thickness whilst the lower seam remains at about 6ft (1.8m). The separation between the two seams has ranged from 7 to 45 ft over the mined area (2 to 14m). Longwall mining has been practiced at Sunnyside for over 20 years due to difficulties of roof control encountered when using the roan and pillar system. Number 3 mine is bounded on the north and south sides by mines Number 1 and 2 respectively. Whilst current production is concentrated into Number 1 mine, much of the future of the complex lies in the further development of deeper reserves in Number 3 mine. Workings in this latter mine were curtailed in 1978 due to difficulties in ventilation. Present developments are ventilated partially from the neighboring Number 2 mine where no workings are in progress. The layout of Number 3 mine is illustrated on the schematic Figure 1. Trunk airways extend down dip from the surface at No. 2 Canyon and the Water Canyon for a distance of some 9,600 ft. (2930m). The area between the two sets of trunk airways has been worked extensively in both seams as have the corresponding reserves on either side in the connected adjacent mines. At the present time exhausting fans are sited at the top of a shallow shaft in No. 2 Canyon and an 8 ft (2.4m) diameter shaft sunk to a depth of 1013 ft (310m) closer to the current developments (Figure 1). The current airflow system, even with an additional 116,000 cfm (55m3/s) entering from No. 2 Mine, is adequate only for the development work now in progress but will be unable to support new longwall faces further downdip. The basic ventilation problem of this mine may be stated quite simply. In a situation where all intake and return airways pass through extensive old workings, a ventilation system design was required that would be effective, efficient and economic for the foreseeable future of the mine. ORGANIZATION OF THE PLANNING PROCEDURE The procedure followed during the study is illustrated on Figure 2. Initial ventilation surveys established the current state of the airflow system and provided the necessary data for setting up a Basic Network File in a computer store. The data in this file was a mathematical model of the ventilation system of the mine. The basic network was analysed by a ventilation network analysis program in order to correlate the measured and computed airflows and to establish the basic network as a true representation of the mine as it stood at the time of the surveys. The network model could then be extended to simulate the future development of the mine and alternative ventilation designs investigated. The remaining sections of the paper outline the work involved in each of these main phases of the planning procedure. VENTILATION SURVEYS Conduct of Surveys Two types of measurements were conducted simultaneously throughout the air-carrying routes of the mine: (i) Airflow measurements were made by anemometer traverse or smoke tube at 221 selected stations. Anemometer traverses were repeated at each station until at least three gave results to within 5 per cent. (ii) Pressure drop measurements were made across ventilation doors, regulators and, wherever possible, across stoppings. Additionally, frictional pressure drops were measured along airways where such pressure drops were significant (above 0.01 inches of water gauge or 2.5 Pa over a 100m distance). The trailing hose method was used to determine these frictional pressure drops. This involved laying out 100m of abrasive resistant plastic tubing (3 mm internal diameter) with a 4 ft. pitot-static tube facing into the airflow at either end and a low range pressure gauge connected into the line. The trailing hose method was preferred to the alternative barometer technique for this study because of (a) the relative ease of access between measuring points and (b) the greater accuracy within individual airways. The anemometers used were Davis Biram Type A/2-3" (30 to 5,000 ft/min) and Airflow Developments AM-5000 digital (50 to 5,000 ft/min). The pressure gauges employed were Dwyer magnehelic instruments. These were preferred to liquid in glass manometers because of their portability and dependability under adverse mining conditions. A checklist of the equipment used in the survey is given in Appendix 1. The instruments were calibrated before and after the surveys in the mine ventilation laboratory at the University of California, Berkeley. The survey occupied two teams, each of three men, for ten working days. The work consisted

Jan 1, 1982

By John S. Crout

TWENTY-FIVE years ago the training of an engineer was of interest solely to the educator and to the student entering the field. At that time the engineer's position in society was relatively simple. His training emphasized design and operation and he expected to enter industry to handle the engineering problems of raw materials production, mechanical and product design, and the operation of power plants, factories, transportation and communication systems, and similar phases of industrial activity. He was not concerned with or particularly interested in the problems of management, administration, sales, publicity, industrial relations, or of legislation. To a large degree he was content to leave these matters to self-made businessmen, or men trained in law, finance, or commerce. That period in the evolution of the engineer is now history. It is generally conceded that the modern world is dependent upon engineering technology for its day-to-day survival. The continued operation of facilities on which mankind relies is dependent upon the functioning of mines, utilities, chemical plants, and other factories, all of which require the knowledge and the skills of trained engineers. Thus from a purely mechanical point of view, the engineer has become indispensable to

Jan 1, 1947

By Thomas L. Kesler

Essentially all of the barite produced in Georgia has come from the Cartersville district in the northwest part of the state. The earliest recorded shipment of ore, 60 tons, was made in 1894.1 With the exception of the four-year period 1931 to 1934, the separate yearly output has been recorded and published since 1915, whereas the total yearly output of the United States has been published since 1880. The separately recorded production of Georgia (Cartersville) through 1947 is 2,232,544 short tons valued at $14,900,746,2 but unpublished data on the period 1931 to 1934, plus the estimated unrecorded output prior to 1915, make the actual total about 2,400,000 short tons. This is about 21 pct of the total production of the United States, which, through 1947, is 11,328,440 short tons valued at $71,683,386.2

Jan 1, 1949

By John Chipman, Raymond L. Orr

Recently reported high -temperatlnre thermal data have been incorporated in a reuieu' of the therinody-r~ar,zic propesties of iron. The recent data permit mot corlsisterlt and reliuble choices to be made for many of the pvoperties, resnlling in bettev eslablished tabrtlations of the thermodynamic functions. Values of Cp, h°t- - H°298. s°t - S2sa, and (G°T -H298/T are tabulated for -the solid and liquid phases of iron in both stable and melastable regions. Consistency of the tabulations has been maintained to the precision necessr~vg to yield the Gibbs energy change belween the bcc (a, 8) and fcc (y) phases of iron at temperatures pertinenl to alloy studies. THE importance of the metal iron has stimulated many investigations of its thermodynamic properties at elevated temperatures. Reviews and summaries of the data have appeared from time to time. Among these the 1951 paper of Darken and smith' deserves highest recognition for the quality of its criticism and for the precision and thermodynamic consistency of its tabulations. These and the more recent table of Kaufman, Clougherty, and weiss2 reproduce the Gibbs energy of the (a, 6)--Y transformation with sufficient sensitivity for alloy studies. Other recent compilations of equal validity3-' make no attempt to do this. Since this Gibbs energy change is not greater than 22 cal per g-atom over the temperature range 1100° to 1800°K, it is obviously desirable to tabulate values of (G°t - H298)/T to four places of decimals even though the absolute accuracy would not warrant such precision. The necessity for reconsideration of the problem is emphasized by the recent publication of several important researches on the enthalpy and heat capacity of iron. The older work is well-covered in the reviews cited and will be mentioned here only when actually employed in the tabulations. GAMMA IRON Transition temperatures are taken from the selec -tions of Elliott and Gleiser,5 based on the work of Boulanger7 and earlier workers. For the a-y (A3) transition at 1184°K we retain the value 215 cal per g-atom adopted by Darken and smith.' It has received confirmation in the recent work of Dench and Kubaschewski,' and of Braun8 using adiabatic calori-metry, and of Wallace, Sidles, and Danielson10 using a pulse heating technique. The enthalpy data of Olette and ferrier" and of Anderson and Hultgren12 at higher and lower temperatures may be extended to 1184°K to give the following values which are adjusted to the required value of ? 1184 Ha,1184 Ha 298 = 8030 cal per g-atom [11 hr,1184-ha298 = 8245 cal per g-atom [2]

Jan 1, 1968

By J. E. Chappelear

We are concerned with the problem of calculating the forces produced by passing waves on a submerged object (in particular a drilling barge) located on the ocean floor. We employ the linearized theory of waves (the Airy1 theory), since solutions are generally more easily obtained in such a theory. It will be indicated how an approximate solution to the problem can be obtained for waves of finite height by use of a linear combination of Airy waves. We shall assume the submerged object has an infinite length, thus neglecting diffraction and making the problem two-dimensional. We find that there is not a considerable difference between our solution and the usual engineering solution. The force, as a function of time, is plotted in Figs. 4 and 5, together with the usual solution for easy comparison. So far, no experiments have been found to compare with the results. As will be shown, the calculations can be extended to waves of finite height (Stokes2-Struik3 waves) in an approximate fashion. This can be done by calculating the force produced by each Fourier component of the wave, separately, and omitting the coupling between the various harmonics produced by Bernoulli's theorem. The problem and its boundary conditions will be briefly stated. We seek a function, +, the velocity potential, which is harmonic, ?2?=0...........(1) and which satisfies the boundary conditions ??/?n =0 .......(2) (the normal derivative of + is zero) on the ocean floor and the sides of the barge, and s??/?y = ?..........(3) on the ocean surface, where s = g/w2........(4) The physical reasons for these choices and an interpretation of ? are given by Lamb.' As previously stated, we assume also that the problem is two-dimensional. We set up our coordinates such that x = 0 is the center of the barge and y = 0 is the ocean floor. The coordinates are shown in Fig. 1, together with the location of the barge on the bottom. Our solution may be decomposed into a function which is symmetric about x = 0,?e, and one which is antisymmetric, +,, ? = ? + ?o...........(5) Thus directly above the barge we have as boundary conditions that for x = 0, ?o = 0, ...........(6) and ??e/?x = 0........(7) Now, far from the barge, we have that (for x positive and assuming the waves are incident from the left) ?e ? Ae cosh ky cos k (x + a) , . . . (8) and ?e ? Ao cosh ky cos k (x + ß) . . . . (9) where + means ". .. approaches ... for x large." Let us put Ae, Ao, and k equal to one, and remember that, in accordance with the derivation of Eq. 3, a sinusoidal time factor is to be supplied. Then we may rewrite Eq. 8 for ?e, and we find that

Jan 1, 1958

By W. D. Forgeng, P. F. Wieser

MnSi alloys containing from 2 to 24 at. pct Si have been investigated by metallographic and X-ray methods. Contrary to published data, the temperature of the ß-manganese to a-manganese transformation is lowered by the addition of silicon. The existence of two additional intermediate phases E and ( is demonstrated. E, ranging in composition from 12 to 15.5 at. pct Si, forms by a peritectoid reaction between ß manganese and the second new phase, (. The latter phase is homogeneous at silicon contents from 16.2 to 18 at. pct and forms peritectically from the 0 phase and liquid. The eutectic reaction occurs between the ( phase and MnsSi rather than the ß phase and Mn3Si. As a result of this investigation, a modification of the manganese-rich section of the MnSi equilihrium diagram is proposed. The binary diagram of Mn-Si, according to Hansen,' Fig. 1, exhibits in the low-silicon region (below 24 at. pct Si) a peritectic transformation at 900°C of 0 manganese + Mn3Si to a manganese as determined by R. Vogel and H. Bedarff2 as well as a thermal effect at 980°C between 16.2 and 22.6 at. pct Si reported by the same authors. The nature of the latter transformation was unknown. K. Amark, B. Boren, and A. westgren3 reported indications that possibly four more phases existed in the range of 10 to 27 at. pct Si. The present investigation deals with the phase relations in the range of 2.1 to 17.7 at. pct Si. A few spot checks were conducted up to 23.84 at. pct Si. EXPERIMENTAL PROCEDURE Small experimental heats were prepared by induction melting in a vacuum furnace using argon as a protective atmosphere. The crucible was of commercial-quality magnesia. The furnace charge consisted of silicon metal and dehydrogenated electrolytic manganese. The typical analysis for these materials is shown in Table I. The molten alloy was tapped into 2-in.-square, 7-in.-long ingots. After visual inspection, a small quantity of alloy in the middle portion of the ingot was separated for use in this investigation.MnSi alloys containing from 2 to 24 at. pct Si have been investigated by metallographic and X-ray methods. Contrary to published data, the temperature of the ß-manganese to a-manganese transformation is lowered by the addition of silicon. The existence of two additional intermediate phases E and ( is demonstrated. E, ranging in composition from 12 to 15.5 at. pct Si, forms by a peritectoid reaction between ß manganese and the second new phase, (. The latter phase is homogeneous at silicon contents from 16.2 to 18 at. pct and forms peritectically from the 0 phase and liquid. The eutectic reaction occurs between the ( phase and MnsSi rather than the ß phase and Mn3Si. As a result of this investigation, a modification of the manganese-rich section of the MnSi equilihrium diagram is proposed. The binary diagram of Mn-Si, according to Hansen,' Fig. 1, exhibits in the low-silicon region (below 24 at. pct Si) a peritectic transformation at 900°C of 0 manganese + Mn3Si to a manganese as determined by R. Vogel and H. Bedarff2 as well as a thermal effect at 980°C between 16.2 and 22.6 at. pct Si reported by the same authors. The nature of the latter transformation was unknown. K. Amark, B. Boren, and A. westgren3 reported indications that possibly four more phases existed in the range of 10 to 27 at. pct Si. The present investigation deals with the phase relations in the range of 2.1 to 17.7 at. pct Si. A few spot checks were conducted up to 23.84 at. pct Si. EXPERIMENTAL PROCEDURE Small experimental heats were prepared by induction melting in a vacuum furnace using argon as a protective atmosphere. The crucible was of commercial-quality magnesia. The furnace charge consisted of silicon metal and dehydrogenated electrolytic manganese. The typical analysis for these materials is shown in Table I. The molten alloy was tapped into 2-in.-square, 7-in.-long ingots. After visual inspection, a small quantity of alloy in the middle portion of the ingot was separated for use in this investigation.

Jan 1, 1964

THE following statement is contained in an article entitled "Mortgaging the Future." by Walter C. Teagle, president of the Standard Oil Co. of New Jer-sey, appearing in the current issue of the Lamp, the official organ of the Standard Oil Co.: "During the last eight months the domestic daily production. of crude oil has risen 300,000 bbl.. which means that no less a volume than 2,500,000 bbl. now are coming to market each day. . . . Every twenty-four hours 250,000 bbl. more than the consumer required are being added to our surplus oil storage. The temporary advantage of cheap petroleum products and the availability of fuel .oil, for a widespread use as a substitute for coal are not being weighed against the time when scarcity may follow on the heels of plenty and the public of tomor-row may have to pay for today's excesses. The over-production of crude and its consequences constitute a problem of national import which far outweighs the temporary advantages either to a part of the industry or to sections of the public.

Jan 8, 1927

By R. G. Stickney

INTRODUCTION The Basalt Waste Isolation Project, conducted by Rockwell Hanford Operations under contract to the U.S. Department of Energy, is assessing the feasibility of the disposal of commercial radioactive spent fuel/high-level waste in deep geologic formations in basalts beneath and within the Hanford Site in southeastern Washington State. The Basalt Waste Isolation Project mission is to identify and characterize a potential geologic repository site in basalt and to develop associated facilities and technology required for the potential permanent isolation of radioactive wastes in basalt formations. In support of the Basalt Waste Isolation Project, a laboratory testing program was conducted to evaluate the performance of the Council for Scientific and Industrial Research's "Doorstopper" biaxial strain cell. To support site characterization studies and the design of repository openings, in situ stress measurements will be conducted within the reference repository horizon (~970 m). The Doorstopper is one of the instruments being considered for use in these measurements. It was selected because it requires only a small volume of work to achieve stress relief, which is important in a closely jointed basalt. Also, test durations are short, permitting a large number of measurements to be made within a given time to increase confidence in the results. Within the reference repository horizon, temperatures are expected in the range of 40º to 60°C with a humidity of up to 100%. For this testing program, test conditions were selected to closely resemble those expected within the repository horizon. The Doorstopper test, mainly developed by Leeman (1964), is a strain rosette relief method of measuring rock stress in situ. The Doorstopper consists of a four element, 45º strain gage rosette

Jan 1, 1984

By W. R. Hibbard Jr., R. W. Guard, N. G. Ainslie

Evidence is presented that copper-base solid solutions of different solutes having equal grain sizes, no preferred crystal-lographic orientation, equal electron-atom ratios, and, within experimental scatter, identical initial yield strengths, need not have identical stress-strain curves at strains larger than about 0.04. The stress-strain behavior is rationalized in terms of the proposed Suzuki chemical interaction between solute atoms and extended dislocations using what is thought to be a somewhat different means of representing stress-strain data. ALTHOUGH the effect: of alloying element upon the strength characteristics of sold solutions is a subject which has received considerable attention in the past, the exact relationships between the common deformation parameters and certain common variables are not really known in some cases. As a result some of the experiments reported in the literature in which these variables are inadequately controlled lose some of their persuasion regarding underlying principles. Nonetheless, facts are known which bear pointing up: When the true stress, a, and true plastic strain, E, of tensile deformatic~n are plotted on a double logarithmic coordinate system, one may observe a straight-line relationship at strains greater than 0.02. The form of the curve in the linear region is given by a = Ken! where a represents true stress, E, true strain, and K and tn, constants. If the relationship holds, K and m define the flow characteristics of the material being tested. m and K, however, may vary with other parameters. Hollomon found that in a-brass, m is influenced by grain size. French and HibbardZ found in alloys of copper that inverse relationships existed between m and 1) the solute concentration for a given solute, 2) the 0.01 yield strength, and 3) the constantK. Lacy and Gensamer3 observed (du/d~) (= U/Em) to increase with increasing values of K in systems of alloyed ferrites (although with considerable scatter of data which may be attributed to uncontrolled grain size). Brick, Martin, and Angier* deduced in copper-base alloys a straight-line relationship (with some scatter) between the change in the Dph number due to solid-solution strengthening and the change in the Dph number due to work hardening which suggested that copper-base alloys having equal yield strengths might have identical stress-strain curves in the plastic flow regions. French and HibbardZ concluded that the yield strength of copper-base solid solutions is the proper basis for comparing the effects of solute elements. Also, Allen, Schofield, and ate' showed that, within their experimental variation, copper-base alloys of zinc, gallium, germanium, and arsenic having the same electron-atom ratios have the same true-stress true-plastic strain curves. Dorn, Pietrokowsky, and ~ietz' also found that with aluminum-base alloys the stress-strain curves in the flow regions are approximately the same if "equivalent" concentrations of alloying elements are used. Solute valence and lattice parameter distortion were the parameters used to determine equivalency. The present report describes an investigation in which an attempt was made to obtain copper-base solid-solution alloys of four solute elements having within close tolerances equal grain sizes and yield strengths, and to see if the level of yield strength does indeed define the flow curve regardless of solute type. During analysis of the data certain unexpected features of the stress-strain curves became apparent which gave rise to some speculation and are discussed at length in the paragraphs that follow. EXPERIMENTAL PROCEDURES Alloy Preparation—Using the data of French and HibbardZ as a first approximation, four different binary copper-base alioys were designed so as to have the same yield strength. In addition, other alloys were prepared in which the solute aoncentrations varied slightly from those calculated above so as to span a range of yield strengths, see Table L The yield strengths of all alloys prepared except the copper-tin alloys were subsequently found to Lie fairly close to one another. The copper used in the alloys was produced by the American Smelting and Refining Co. and was of very high purity (99.999 pct). The alloy additions and their initial purities are as follows:

Jan 1, 1960

THE CHAIRMAN (H. W. MORSE, Los Angeles, Cal.).-I would like to open this meeting for a little while to the general subject of leaching. We ought not to hold back if we have any new schemes for the future. I know that those of us who are tangled up with this lei thing work will appreciate any suggestions that will make us work harder and, scheme, harder to pull through these past difficulties on leaching. F. S. SCHIMERKA, Clifton, Ariz.-Under the direction of J. W. Bennie, General Manager of the Shannon Copper Co., Clifton, Ariz., 'it has been my valued privilege to cooperate in the elaboration of a hydrometallurgical process aiming at the recovery of copper in mill tailing resulting from a gravity concentration of a low-grade semi-oxidized sulphide ore occurring in a porphyritic gangue. The experimental work which has been carried on with a 25-ton plant is completed, and the company is building at present a leaching plant on a larger scale, the first unit consisting of a 150-ton roasting furnace and accessories for leaching the calcine.

Jan 12, 1916

By R. G. Beverly

The presence of radioactivity in uranium mill wastes has resulted in somewhat unique waste disposal methods. In addition to the common problems of disposing of large quantities of solid wastes, neutralizing acids, minimizing dissolved heavy metals, and clarifying all liquid effluents, the uranium mill operator must sample and analyze liquid effluents for micro-micro-quantities of radio- nuclides such as Ra226, Th230, and Pb210. Special disposal methods or decontamination procedures are required to meet the stringent limitations established by the U.S. Atomic Energy Commission on the concentration of radionuclides in liquid effluents. Routine monitoring of the receiving stream is also normally required. In recent years there has been a growing concern about the radioactivity remaining in uranium mill tailings, particularly at the sites of shut-down or dismantled mills. The State of Colorado, in cooperation with the industry, has adopted regulations for stabilization of inactive uranium mill tailings.

Jan 6, 1968

By A. J. Strauss

The solubilily of lead in PbS as a funclion of lemperature has been investigated by means of Hall coej'icient (RH) measurentents OIL samples prepared by annealing and quenching Bridgman-grourz single crystals, which contain excess lead incorporated during /heir growth from the mell. For sufficiently long annealing times and sufficiently rapid quench rules, the samples obtained consist of n-lybe PbS snturated with In the semiconducting compounds PbS, PbSe, and PbTe, the point defects associated with the presence of lead in excess of the stoichio metric composition are shallow donors. The defects associated with the presence of excess nonmetal are shallow acceptors. lead at the annealing teuiperatiire, together luilh a dispersed second phase of lead microprecipitates. Due to the retrograde solubility of lead, the measured free carrier concentration (n = l/eR77) decreases with decreasing annealing tewperature (T) from- 2.4 X 10 em-3 for 1l85OK to 1.3 X 1018 cm-3 for 623°K. The dala are well-represented by the equation n =(1.08 x 102) exp[ - (0.37 ev.)/kT] cm-3 For most undoped samples of these compounds, the carrier concentration at room temperature and below is determined primarily by the stoichiometry, since the concentration of either excess lead or excess non-metal is large compared to the concentrations of intrinsic carriers and electrically active impurities. Therefore the dependence of composition in temperature and partial pressure can be investigated by means of Hall coefficient measurements, even though the deviations from stoichiometry do not exceed 0.03 at. Pct

Jan 1, 1968

By K. G. Black, B. P. Faulkner

Laboratory batch flotation rests are used for comparison of the floatability of ores and for comparison of the effect of various treatments on ore floatability. Although it is difficult to predict exact recoveries for continuous flotation on the basis of batch tests, considerable quantitative information can be obtained through their use. A simple method is presented which provides a convenient means of quantitatively evaluating results from laboratory batch tests. The information obtained is in the form of a discrete flotation rate distribution, which, if desired, can be used directly in a simulation of continuous flotation. Equation of Batch Flotation The validity for batch flotation of the first-order kinetic equation is accepted: [ ] where k is flotation rate of a class of articles and W (t) is weight of particles of the class in cell at time t. Thus, at time t, the fraction of the original amount of the class of particles still remaining in the cell is [ ] where W (0) is weight of class of particles in cell at time zero. If there are several classes of particles in the cell, each with its own flotation rate, k4, the fraction of the particles remaining is [ ] where W (t) is now the total weight of n classes, a, being the original fraction of W (0) belonging to the ith class. Accordingly, [ ] [Fig. 1] shows the result of a typical batch test. Such result can be characterized by a set of particle classes, each set described by a flotation rate, [ki,] and a corresponding fractional magnitude, [ai.] [ ] Multiple Regression Equation Eqs. 3 and 4 may be combined to give

Jan 1, 1973

By K. Huessener

Discussion of the paper of K. HUESSENER, presented at the New York. meeting, February, .1916, and printed in Bulletin No. -110, February, 1916, pp. 443 to 474. KARL NIBECKER, Youngstown, O.-It has been contended that the method of measuring the operation of a blast-furnace boiler plant by means of the amount of steam generated per pound of coke used in the blast furnace is not of any value and even that it is quite misleading. It has been found, in the plant with which I am connected, that, in general, a blast furnace will operate during a month so close to average. conditions that the average monthly measure of the plant efficiency is indicated by the number of pounds of steam made per pound of coke put into the furnace, neglecting the amount of iron made and the amount of gas used for heating the stoves. For some time, it has been our custom to keep a record of this figure and use it as a general indicator of the operating practice. In comparing this figure from two different plants we have found some difficulty, due to differing conditions. It is conceivable but not altogether probable that a slight difference in burden, kind of ore or coke, would account for this difference. The method of using the gas, whether washed or unwashed, and the plant arrangement, may affect this number. For several plants on which we have worked up this number, we have found that most of the ordinary running irregularities are eliminated during a run of a month or more and that the number is, therefore, satisfactory. In connection with a plot of the number of pounds of steam generated per pound of coke, we are also plotting the production of the furnaces and the pounds of coke used per ton of production. With these three curves plotted on the same sheet, we believe that the operating man has sufficient information to judge of the efficiency of his plant' for any month.

Jan 5, 1916

By Vard H. Johnson

IN 1943 the U. S. Bureau of Mines undertook drilling in an effort to develop new reserves of coking coal in an area near Paonia, Colo., as a part of an attempt to alleviate the shortage of known coking coal of good quality in the western United States. Geologic mapping of the area was undertaken by the U. S. Geological Survey with the purpose of first furnishing guidance in location of drillholes and later aiding in interpreting the results of the drilling. The drilling program was under the general supervision of A. L. Toenges of the U. S. Bureau of Mines. J. J. Dowd and R. G. Travis were in charge of the work in the field. Geologic mapping was started by D. A. Andrews of the Geological Survey in the summer of 1943 and was continued from the spring of 1944 to 1949 by the writer. The first few holes drilled failed to locate coking coal, but in the summer of 1944 coking coal was discovered by drilling 6 miles east of Somerset, Colo., the site of present mining. In the succeeding years, 1945 to 1948, 100 to 150 million tons of coal suitable for coking were blocked out by drilling. The ensuing discussion of the geologic controls on the distribution of coking coal in the area is based on the geologic mapping as well as the drilling done in the Paonia area, more complete descriptions of which have appeared or are in process of publication."' In order that the possible geologic controls affecting the present distribution of coking coal may be considered, it is necessary to discuss briefly the indicators of coking quality coals. Coking Coal Coal that cokes has the property of softening to form a pastelike mass at high temperatures under reducing conditions in the coke oven. This softening is accompanied by the release of the volatile constituents as bubbles of gas. After release of the contained gases and upon cooling, a hard gray coherent but spongelike mass remains that is referred to as coke. This substance varies greatly in physical properties and, to be suitable for industrial use, must be sufficiently dense and strong to withstand the crushing pressure of heavy furnace loads. Western coals have a generally high volatile content and therefore form a satisfactory coke only when they attain a rather high fluidity during the process of heating arid distillation in the coke oven. When this high degree of fluidity is developed, the volatile constituents escape and leave a finely porous coke. On the other hand, when the degree of fluidity is low the product is an excessively porous and therefore physically weak mass that is called char." Small quantities of oxygen present in coal are believed to decrease the fluidity of the material during the coking process and to favor the development of char rather than coke. In consequence, coal chemists have for some time considered the possibility of developing an index to coking qualities by inspection of chemical analyses of coals.' A formula has now been developed that does permit a rough preliminary estimate of the cokability of coal on the basis of the analysis on an ash and moisture-free basis. Coals may be eliminated as possible coking fuels if the oxygen content is greater than 11 pct. Similarly the ratio of hydrogen to oxygen must be greater than 0.5 and the ratio of fixed carbon to volatile constituents must be greater than 1.3. If the coal, on the basis of these limiting factors, appears to have possible coking qualities, the following formula permits determination of the coking index: a+b+c+d Coking index = -------- 5 a equals 22/oxygen content on ash and moisture-free basis, b equals two times the hydrogen content divided by oxygen content on moisture and ash-free basis, c equals fixed carbon/l.3 x volatile matter, and d equals the heating value on moist, ash-free basis/13,600. Coking indices higher than 1.0 suggest that the coal will coke, and indices above' 1.1 indicate good coking tendencies. Although generally usable, this formula 'is not completely satisfactory because the percentage of oxygen shown in ultimate analyses is derived only by difference; i.e., by subtracting the sum of the percentages of the constituents determined analytically from 100 pct. Although the coking index indicates the coking tendencies of coal, it is necessary to make physical tests of coke before its industrial value can be determined. The U. S. Bureau of Mines has developed a standard procedure for determining the approximate strength of coke that would be formed from a given coal. In this test one part of ground coal, mixed with 15 parts of carborundum, is baked to form a standard briquette. The weight, in kilograms, necessary to crush the briquette is termed the agglutinating index. This test determines the relative fluidity attained in the coking process by measuring the cementing strength of the coal in the briquette. A

Jan 1, 1953

By Vard H. Johnson

IN 1943 the U. S. Bureau of Mines undertook drilling in an effort to develop new reserves of coking coal in an area near Paonia, Colo., as a part of an attempt to alleviate the shortage of known coking coal of good quality in the western United States. Geologic mapping of the area was undertaken by the U. S. Geological Survey with the purpose of first furnishing guidance in location of drillholes and later aiding in interpreting the results of the drilling. The drilling program was under the general supervision of A. L. Toenges of the U. S. Bureau of Mines. J. J. Dowd and R. G. Travis were in charge of the work in the field. Geologic mapping was started by D. A. Andrews of the Geological Survey in the summer of 1943 and was continued from the spring of 1944 to 1949 by the writer. The first few holes drilled failed to locate coking coal, but in the summer of 1944 coking coal was discovered by drilling 6 miles east of Somerset, Colo., the site of present mining. In the succeeding years, 1945 to 1948, 100 to 150 million tons of coal suitable for coking were blocked out by drilling. The ensuing discussion of the geologic controls on the distribution of coking coal in the area is based on the geologic mapping as well as the drilling done in the Paonia area, more complete descriptions of which have appeared or are in process of publication."' In order that the possible geologic controls affecting the present distribution of coking coal may be considered, it is necessary to discuss briefly the indicators of coking quality coals. Coking Coal Coal that cokes has the property of softening to form a pastelike mass at high temperatures under reducing conditions in the coke oven. This softening is accompanied by the release of the volatile constituents as bubbles of gas. After release of the contained gases and upon cooling, a hard gray coherent but spongelike mass remains that is referred to as coke. This substance varies greatly in physical properties and, to be suitable for industrial use, must be sufficiently dense and strong to withstand the crushing pressure of heavy furnace loads. Western coals have a generally high volatile content and therefore form a satisfactory coke only when they attain a rather high fluidity during the process of heating arid distillation in the coke oven. When this high degree of fluidity is developed, the volatile constituents escape and leave a finely porous coke. On the other hand, when the degree of fluidity is low the product is an excessively porous and therefore physically weak mass that is called char." Small quantities of oxygen present in coal are believed to decrease the fluidity of the material during the coking process and to favor the development of char rather than coke. In consequence, coal chemists have for some time considered the possibility of developing an index to coking qualities by inspection of chemical analyses of coals.' A formula has now been developed that does permit a rough preliminary estimate of the cokability of coal on the basis of the analysis on an ash and moisture-free basis. Coals may be eliminated as possible coking fuels if the oxygen content is greater than 11 pct. Similarly the ratio of hydrogen to oxygen must be greater than 0.5 and the ratio of fixed carbon to volatile constituents must be greater than 1.3. If the coal, on the basis of these limiting factors, appears to have possible coking qualities, the following formula permits determination of the coking index: a+b+c+d Coking index = -------- 5 a equals 22/oxygen content on ash and moisture-free basis, b equals two times the hydrogen content divided by oxygen content on moisture and ash-free basis, c equals fixed carbon/l.3 x volatile matter, and d equals the heating value on moist, ash-free basis/13,600. Coking indices higher than 1.0 suggest that the coal will coke, and indices above' 1.1 indicate good coking tendencies. Although generally usable, this formula 'is not completely satisfactory because the percentage of oxygen shown in ultimate analyses is derived only by difference; i.e., by subtracting the sum of the percentages of the constituents determined analytically from 100 pct. Although the coking index indicates the coking tendencies of coal, it is necessary to make physical tests of coke before its industrial value can be determined. The U. S. Bureau of Mines has developed a standard procedure for determining the approximate strength of coke that would be formed from a given coal. In this test one part of ground coal, mixed with 15 parts of carborundum, is baked to form a standard briquette. The weight, in kilograms, necessary to crush the briquette is termed the agglutinating index. This test determines the relative fluidity attained in the coking process by measuring the cementing strength of the coal in the briquette. A

Jan 1, 1953