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By G. B. Thomas

Several methods of analyzing pressure build-up data in wells have been presented by various authors. This paper reviews the theory and method of D. R. Horner and presents example calculations performed on data obtained by testing several different types of wells. These calculations include, (1) graphical estimation of final static pressure, (2) determination of the productive capacity of the pay away from the well bore and. (3) the degree to which the formation adjacent to the well bore has been damaged by completion or other causes. The methods of testing and precautions which should be taken to assure the best data possible are discussed. Limitations and reliability of calculated results are also treated. INTRODUCTION Pressure testing of wells is generally limited to the determination of producing and static mean formation pressures. The so-called "static" pressure determination.. along with PVT, electric log and production data, enable the reservoir engineer to determine, within reasonable limits, the drive mechanism of the reservoir and in some cases. the amount of edge water encroachment. Producing pressure tests enable calculation of productivity indices and allow the engineer to plan the systematic production of a pool for optimum conservation of suhsurface energy. The radial flow formula advanced by Muskat' has been based on the assumption of incompressible radial fluid flow. It has been known that reservoir fluids do not behave in an ideally incompressible manner. For example, incompressible flow theory indicates a simple logarithmic relationship between the difference of the instantaneous and static well pressures when plotted against time. 'The latter stages of this type of plot of pressure build-up data generally show a marked devia-, tion from the earlier straight line trend. which deviation may be shown to be due to the compressible flow of fluids toward the well bore. Shut-in times of 24, 48. 72. or at most 96 hours are currently in wide use for determining so-called "static" reservoir pressure. Due to the continuation of compressible flow of fluids into the well bore long after this arbitrarily taken shut-in time, the determination of static pressures has almost invariably resulted in lower than equilibrium values. Materials balance calculations made early in the life of a reservoir often result in a calculated reserve which later observations prove to be too low. Failure to obtain reliable "static" reservoir pressures within the prescribed 24 or 48-hour build-up period has undoubtedly been a major factor in obtaining these low estimates. Comparison of theoretical and actual productivity performance has indicated that formation damage and not being able to attain true static pressures have been partially responsible for the observed discrepancies. Research into the theory of compressible flow behavior has resulted in methods of applying these theories to the testing of wells. Application of the developed theories to pressure build-up performance indicates that in many cases the following can be estimated. 1. The static reservoir pressure. 2. The in-place effective formation permeability away from the well bore. 3. The degree or extent by which the formation has been damaged adjacent to the well bore, either through completion methods or subsequent damage due to fluid entry. REVIEW OF THEORY Horner2 has shown that the pressure build-up within a "point source" well is approximated by the following formula: Equation (1) is the approximate "point source" solution to the radial compressible flow equation advanced by Muskat. The solution assumes the following: 1. A point source well is producing at constant rate from the center of an infinite reservoir with a constant pressure at its external boundary. 2. The fluid flowing is present in one phase only. 3. The compressibility and absolute viscosity of the fluid remain essentially constant over the range of temperature and pressure variation encountered. 4. The well i.; shut-in at the sand face arid there is 110 after production into the well bore. 5. The formation permeability is homogeneous in the direction of flow. From Equation (1) it can be seen that if an ideal well were shut-in while producing from a reservoir under the conditions assumed, the pressure build-up would be a logarithmic func-

Jan 1, 1953

By J. W. Smith

Our continuous search for underground productivity improvements has been brought about by the diminishing ore grades in existing underground mines. The need for more efficient mining methods is a result of the economic problems facing our industry today, and this has caused us to evaluate underground haulage methods which have traditionally been the "bottleneck" in the flow of material from the ore in the natural state to the surface processing facility of any underground mining operation. Small improvements in the face haulage systems have yielded much greater benefits as they relate to overall mine productivity so it's only natural that we are all concerned with the best method of moving ore from the face to the main line haulage. In a recent paper titled "Underground Haulage Trucks - Gaining Momentum Worldwide", Richard A. Thomas concludes that the use of trucks to haul ores in underground mines is on the increase spurred by the convergence of a number of technology advances and economic realities. Perhaps the most important stimulus for the growth of trackless haulage is the high degree of haulage flexibility in underground operations. On the economic side, the demand for higher productivity from underground mines has resulted in larger physical dimensions of haulage roads, that is, higher backs and wider drifts to provide more room for high capacity haulage units. In the process of determining the most effective type of equipment for haulage, the power source must be a major consideration. For the purpose of this paper, we will limit the comparison to rubber-tired trackless haulage vehicles and not try to make a comparison between rubber-tired haulage, continuous haulage systems and rail-mounted haulage. Cost is perhaps the only really measurable factor when making a comparison between electric and diesel haulage. You will find that some costs will be very well defined in absolute terms. In other areas of comparison, cost can be fairly well estimated, and yet in still others, the costs are totally arbitrary. Let's take a look at some of the cost considerations. (Figure 1) first of all, is the initial cost of the equipment. This capital cost quite often is a determining factor in the type of haulage vehicle to be selected, yet this initial cost is perhaps the most insignificant of all costs when evaluating an operation over the long term. Of much greater concern, is the cost of maintenance. This cost will often run three times the original capital investment during the life of a single piece of haulage equipment. This factor can include rebuild to extend the life of the original capital investment, but certainly includes the labor and materials necessary, plus the inventory to keep the equipment in good repair. Perhaps one cost which is now playing an even greater role in the rubber-tired haulage operation, is the cost of fuel. Conoco has recently come up with some rough estimates which indicate that diesel fuel will cost an average of three times the equivalent kilowatt output in direct electric power. Diesel fuel is almost twice the cost of stored electric power. (This of course relates to the efficiencies of charging and recovery of power from lead acid storage cells.) These particular figures of course will vary from one area to another but I think that there is enough significance here to certainly warrant the further study of fuel costs for each particular area or mine. Another cost is breakdown expense. This must be treated differently from maintenance costs because a potentially larger expense is involved, more than just parts and labor. Now we have to deal with the cost of lost production time, which can have a much greater overall effect. Mine plan economics are another cost consideration where we can't make a comparison without looking at specifics. Here you must look at the movement of power centers vs. the flexibility and freedom of movement of vehicles. The determination must be made as to what types of equipment will fit into any predetermined mine plan and if a change in the planned roadway dimensions for the mine plan itself would be more economical so that more efficient type of equipment could be utilized. Finally, two of the most important aspects to be considered with potential ramifications far beyond what we have mentioned previously, is the cost of health and safety, which is really the cost of meeting current and future government regulations, reasonable or otherwise. And of course, when making any consideration here it is impossible to come up with anything more than an educated guess on the cost of meeting the new regulations. Now let's take a look at some of the advantages of diesel vehicles as well as advantages offered by electric vehicles, both battery and cable powered versions (Figure 2). Much of the data used in this comparison is based on experience with three vehicles manufactured by Jeffrey Mining Machinery Division, Dresser Industries. Jeffrey manufactures all three types, each with approximately a 15-ton capacity, even though few of these Jeffrey vehicles are used in uranium mining operations. Much of our experience comes from the 4114 diesel powered RAMCAR which is a 4-wheel drive, articulated steering,vehicle powered by a Caterpillar 3306NA engine and using a powershift transmission. This will be compared with the performance of the Jeffrey 404H battery powered RAMCAR with articulated steering which utilizes a separate 35 HP DC drive motor on each of two wheels with solid-state speed controls, and the final comparison will be made on the Jeffrey 4015 cable-reel shuttle car which is powered by two 60 HP constant

Jan 1, 1982

By C. M. Nickerson

It is apparent that repressuring of the oil measures is becoming increasingly important to the oil industry, and is a matter that warrants the best efforts of the petroleum engineer charged with applying this method of operation to a producing property. Considerable attention has been directed to repressuring in various California oil fields, such as Dominguez and Seal Beach, which have served to sell the more progressive operator on this idea and also to keep him sold. The projects that have received most attention are in comparatively young fields where the fluid levels are high, the wells, perhaps, still flowing, where the wet gas production is of considerable volume, and where there is vital need of immediate conservation of gas, which otherwise would be blown to the air. The new State of California Gas Law, which prohibits the waste of gas, has served to interest the oil industry in means of conserving the gas that is now being blown to the air because of lack of markets. This vitally affects the operator in the older, well-depleted oil fields of the state, because in many cases he has a limited market for gas, on account of the flush production from the more active oil fields. While the waste of gas is relatively small in these older fields, such as the Midway-Sunset area, it now attains considerable importance, since under the new State Gas Law it must be put to some bcneficial use. Repressuring in the older fields of California has not received the attention of the petroleum engineers that is warranted by the present situation. This paper was prepared with the view of presenting the salient features connected with repressuring in the well-depleted oil measures of the older fields, and to show what results have already been obtained in these fields by injecting the unmarketable gas into the sands, particularly in the Midway-Sunset field in the San Joaquin Valley of California. Other projects in the older fields of the state offer some supporting data in this connection, such as the Shields Canyon and the Brea Canyon fields. Gas Storage and RepressuRing On account of the lack of market for a large portion of the dry gas produced in certain of the older, well-depleted oil fields of California, the

Jan 1, 1930

By John E. Sweet

There have already been so many subjects of a purely mechanical nature presented to the Institute of Mining Engineers, that it is unnecessary for me to apologize for adding another to the list. When a new invention, or what is supposed to be a new invention, claiming simply an improvement on the well-known devices, is to be described, it is usual to point out the defects of the old and give a history of the object or the art up to date. But, as no doubt a large proportion of the members of this Institute know as much, or more, about the history and use of the steam-engine indicator than I do myself, I will not further preface the subject, than to remind you that what called for the Richards's improvement on the Watt's indicator was high speed; that the leading improvement was the reduction of momentum in the reciprocating parts; and that this was accomplished by making the parts lighter, and reducing the travel of the heaviest, of them, and consequently the speed and momentum. The use of the indicator has been extended, for the speed of steamengines and other prime movers has been increased, until one engineer at least, John Cooper, of Philadelphia, says, "I want an indicator that will work up to a thousand revolutions." .Mr. Thompson, of the Buckeye Steam-Engine Company, of Salem, Ohio, has improved on the Richards's indicator by another reduction in the momentum, which is done by putting the parallel motion nearer the fulcrum of the pencil arm, reducing both weight and travel. The characteristic feature of the indicator I am going to illustrate is a still further reduction in the same direction, which is accomplished by dispensing with the parallel motion altogether. As is often the. case, in putting a simple idea into shape, some difficulties

Jan 1, 1879

NEW YORK MEETING PAGE Winchell, H. V.-Geology of Pachuca and El Oro, Mexico. Discussed by Frederick MacCoy, Robert Hawxhurst, Jr 2 WILKES-BARRE MEETING Kynor, Herbert D.-Mechanical Mining of Anthracite. Discussed by C. Evans, Jr., S. J. Jennings, W. J. Richards, Geo. S. Rice, and the author 3 Stein, C. H.-Ashley Planes for Handling Freight Traffic. Discussed by Graham Bright, Douglas Bunting, S. J. Jennings, R. V. Norris, C. H. Strange, Edwin Ludlow, W. J. Richards, and the author 3 Markle, Donald.-Anthracoal: A New Domestic and Metallurgical Fuel. Discussed by William Griffith, Edwin Ludlow, E. M. Chance, Charles Dorrance, E. W. Parker, G. H. Ashley, C. A. Burdick, Geo. S. Rice, and the author :.. 7 Griffen, John.-Slush Problem in Anthracite Preparation. Discussed by Henry Flood, Jr 11 Eavenson, H. N.-Lynch Plant of United States Coal and Coke Co. Discussed by E. V. d'Invilliers, C. Evans, Jr., Graham Bright, Geo. S. Rice, Geo. Otis Smith, A. M. Rodgers, C. T. Starr, C. A. Burdick, and the author 15 Bunting, Douglas.-Mine Fires Extinguished by Sealing. Discussed by Geo. S. Rice, J. B. Warriner, H. H. Otto, Edwin Ludlow, E. M. Chance, F. B. Nold, H. M. Chance, W. L. Jacobus, J. A. S. Ritson, R. D. Hall, Geo. S. Rice, and the author 18 Bright, Graham.-Determination of Electrical Equipment for a Mine Hoist. Discussed by F. L. Stone, Geo. S. Rice, L. F. Mitten, and the author 34

Jan 12, 1921

By Henry E. Armitage

The object of this paper is to give a few useful hints on the concentration of low-grade ores. The machines that I shall speak of are, Cornish rolls, revolving screens, Hartz jigs, spitz-lutte, and the Linkenbach buddle (or Roberts's patent buddle, as it is called in Colorado). I think the reason why some of them have been discarded in favor of more recent inventions is that they have not been given a fair chance to do their work in a satisfactory manner. Cornish rolls are usually looked upon as a machine designed simply to crush the ore as a preliminary to sending it to the sizingscreens, beyond which they have nothing to do with dressing the ore. On the contrary, the rolls give the key-note to the whole process. It is well known that one class of rock will crush differently from another, and from the result obtained by passing the ore through the rolls the size of the first revolving screen and the surface required on the jig-screens must be determined. If it is found that the mineral crushes clear from the quartz, a large mesh can be used on the first sizing-screen, and the tailings will not require recrushing ; but if it is found that some of the tailings from the first jig are ragged (that is, are carrying off particles of mineral sticking on the quartz), it will be necessary to recrush them. Again, if a large percentage of the tailings carry off mineral, it will be more economical to commence with a finer mesh screen than to recrush the tailings. Having determined the best mesh to use on the first sizing-screen, the different sizes on the other screens can be determined from it. I have found for practical purposes that, commencing with a 6-mesh, the next should be 8, then 12, and finally 18. The holes in the jig-screens should be a trifle larger than in the revolving screens. For instance, if the revolving screen is made of No. 14 wire, the corresponding jig-screeus should be of No. 16 wire. The jig-screen surface required for each size made depends on the may

Jan 1, 1890

By H. R. Drullard

TO obtain the best results from hammer drills, close attention must be paid to two factors in drill maintenance, which are of equal importance; one is lubrication, the other is the shank. With the exception of stoping drills, most modern rock drills require both oil and grease. The oil ports are in the lubricator, on or behind the hammer cylinder; the grease port for the lubrication of the chuck sleeve and rotating mechanism is on the chuck end. Ordinary machine oil is not adapted to rock drills; a heavier more gelatinous oil such as castor machine oil or liquid grease should be used. The lubricators should be filled once for every 12 or 14 ft. (3.5 to 4 m.) of hole drilled. Grease guns should be used to fill the port in the chuck end once a shift; a medium (No. 3) grease is well adapted to this purpose. Hard grease must not be put into the lubricators, as it will not flow through that part. Contrary to the popular belief, oiling a machine once or twice a shift does not provide sufficient lubrication; the drills should be oiled once for every 12 or 14 ft. of hole drilled. The rotating handle of a stoping drill is an oil reservoir and is provided with a port for oiling. The rotating handle is packed with wick-ing, or similar material, which causes the oil to feed slowly from the handle to the other parts of the machine. Stoping drills require oil at least twice a shift; lighter oils than castor, such as Arctic Ammonia, may be used. Drills used in shaft sinking can be oiled satisfactorily by placing a good-sized, drop, sight lubricator on the station above the sinking operations and connecting it with the air-line supplying the sinking drills. If this lubricator is properly filled and adjusted, a uniform oiling of the sinking machines will be effected without the necessity of oiling the drills individually. The grease end should be filled at the station or surface before each drilling period. The shift boss should see that the machines are greased and also that the lubricators at the station function properly. The life of the air-drill hose is somewhat shortened by this method, as oil attacks the inner tube; but as some oil is always present in the compressed air, this is not a serious objection to the method.

Jan 1, 1922

By R. J. Arsenault

The yield drop that occurs in tantalum, Cu-AZ. and Ag-Al was investigated as a function of strain rate and at several temperatures. From the strain-rate dependence of the yield drop an activation volume was calculated and found to agree with that obtained from differential strain-mte tests. The temperature dependence of the yield drop can he explained in terms of the parameter temperature divided by activation volume. The yield drop is attributed to the rapid multiplication of dislocations during their motion. THE yield-point or yield-drop (upper yield stress minus the lower yield stress) phenomenon has been the subject of numerous investigations1-9 in both bcc and fcc metals. Also, there is no dearth of theoretical explanations10-20 to account for the yield-point phenomenon. The present explanations can be divided into two categories: the concept of dislocation locking and subsequent unlocking and the concept of multiplication of dislocations due to their motion. The Vottrell10,11mechanism of locking and unlocking, as applied to bcc metals, is the best known of the three proposed locking mechanisms. The two additional mechanisms, short-range order (SRO)7,16,17 and Suzuki-type locking,14,15 have been employed to account for the existence of yield points in fcc metals. The experimental evidence supporting the locking of dislocations, especially in bcc metals, is well-established. The dependence of the yield point on interstitial concentrations,' the kinetics of strain aging,2 and the observations by electron-transmission microscopy18 are some of these experimental results. The number of experimental observations supporting dislocation locking in fcc metals is not as extensive as in bcc metals. The unlocking of dislocations, which actually is supposed to account for the yield drop, is by no means well-established by experimental results for either bcc or fcc metals. The similar temperature dependence6 of the yielding and plastic flow is one of the experimental factors that cannot be explained in terms of the unlocking of dislocations. The mechanism based on the rapid multiplication of dislocations has been proposed by several investigators. 18-20 Johnston and Gilman19 have proposed a detailed model of yielding, which was later employed by Hahn21 to predict quantitatively the yield drop in iron. This mechanism was discussed in a previous paper22 and several limitations were pointed out. In the same paper it was shown that an activation volume could be obtained from the strain-rate dependence of the yield drop and that a correlation existed between the activation volumes obtained from the yield-drop tests and from differen-

Jan 1, 1964

By H. R. Peiffer

Hardening of soft metals can be accomplished by dispersing finely divided hard particles in them. The dispersing of finely divided alumina in silver in the presence of oxygen yields a high strength material which is unusual in that its mechanical properties above the melting point of the continuous Ag-O alloy matrix are similar to other solids. The tensile strength is studied for two of these alloys, one of which contains 15 pct by weight alumina and the other 20 pct by weight alumina. The average fracture strengths above 960°C of these alloys were found to be 0.4 x106 dynes per sq cm and 3.8x106 dynes per sq cm respectively. The strengths appenred to be independent of temperature above 960° C. The creep behavior of the 20 pct alumina material was studied above 960°C. The initial creep rate, 6 , of this material can be represented by where s is the applied stress and E the activation energy for the process. This energy is of the order of 1.7 to 2.1 ev. HARDENING of soft metals by the addition of finely divided, hard particles has resulted in the production of materials with excellent creep resistance and with inhibited recrystallization.' It has been demonstrated that such composites have useful strength up to temperatures very near the melting point of the soft metals, but one might expect that the strengthening by the dispersed particles ceases to be of importance above the melting point of the softer phase. This, however, is not so.2 In this paper the strengthening of a liquid Ag-O solution by the presence of dispersed alumina is discussed and experimental data concerning the mechanical properties of such a material at temperatures above 960°C presented. EXPERIMENTAL PROCEDURE Specimens were prepared using powder metallurgical techniques. Finely divided silver oxide was mixed in the appropriate proportions with "Linde B" alumina. For these experiments the specimens contained 15 or 20 wt pct alumina in pure silver.* These alloys were used in these experiments since they were the ones that could be handled readily enough to yield experimental data. Silver oxide was used in preference to silver because it is readily available in a very finely divided form and because the presence of excess oxygen in the silver is important to the properties2 of the material. The powders were mixed wet, using alcohol as the medium, in an ordinary food blender until a uniform mixture was obtained. They were then dried and heated to reduce the silver oxide to silver. Only a fraction of the finely divided silver remains unoxi-dized upon removal from the oven since the finest particles of silver oxidize immediately upon contact with air. The remaining fine silver served as a binder during pressing. The mixture was then pressed at 20 tons per sq in. into shoulder-grip (reduced section 1/4 in. by l/4 in. by 1 1/4 in.) and rectangular bar specimens (l/4 in. by 1/4 in. by 2 in.). The green specimens were heated very rapidly in a globar furnace to 1000°C, held at temperature untic they had reached theoretical density and then furnace cooled. The strength of the two compositions at temperatures above 960 °C was determined by pulling shoulder grip specimens inside a furnace mounted on a tensile machine. Specimens were gripped by means of wires wrapped around the shoulders of the specimen. Temperatures were measured by means of a chro-mel-alumel thermo-couple placed in the vicinity of the specimen. Control of temperature (within ±°15C) was accomplished by means of a Weston Recorder-Controller. Before loading, specimens were held at temperature for approximately 15 min in order to insure uniform specimen temperature. The low loads were measured by a special load cell designed by Baldwin Lima Hamilton. All testing was performed on a FGT testing machine of the same com-pany. Creep studies were performed in bending utilizing rectangular specimens of the 20 pct alumina alloy. Specimens supported at both ends by knife edges of inconel were placed in a globar furnace, heated quickly to temperature, and the deflection of the beam measured optically as a function of time. The weight and size of the specimen were predetermined and the maximum stress on the beam calculated from

Jan 1, 1962

By H. T. Clark, H. K. Work

The instrument for the rapid determination of carbon in steel, described in this paper, was developed originally to facilitate the operation of the pilot-sized open-hearth furnace in the Jones & Laughlin Research and Eevelopment Laboratory. In this small furnace, with its shallow bath, the drop in carbon is accelerated, and a rapid, reliable test for carbon content is indispensable. This problem becomes rather acute when it is desired to finish heats at a carbon content higher than can be estimated with any degree of accuracy by the fracture test. Several independent methods for determining carbon were investigated, therefore, and it was found that a relatively simple magnetic device could be made to give reproducible results. A laboratory instrument was designed and set up for the use of the furnace operators. It was tested exhaustively in the laboratory and was used with considerable success for following the carbon content in the furnace bath. As soon as sufficient data had been compiled to indicate that the instrument might have a wider application, a more rugged model was constructed and placed in the mill to be tested on commercial heats. Finally, about a year ago, instruments were furnished to the plant open-hearth shops, and these have been in use since that time. As a matter of convenience, the instrument has been named the "Carbanalyzer," and will be referred to by this designation in the following pages. $ Eaklier Investigations While by far the greatest number of magnetic investigations in which the carbon content of ferromagnetic materials was considered have had to do with the evaluation of the magnetic properties of steels of known

Jan 1, 1940

By H. K. Work, H. T. Clark

The instrument for the rapid determination of carbon in steel, described in this paper, was developed originally to facilitate the operation of the pilot-sized open-hearth furnace in the Jones & Laughlin Research and Eevelopment Laboratory. In this small furnace, with its shallow bath, the drop in carbon is accelerated, and a rapid, reliable test for carbon content is indispensable. This problem becomes rather acute when it is desired to finish heats at a carbon content higher than can be estimated with any degree of accuracy by the fracture test. Several independent methods for determining carbon were investigated, therefore, and it was found that a relatively simple magnetic device could be made to give reproducible results. A laboratory instrument was designed and set up for the use of the furnace operators. It was tested exhaustively in the laboratory and was used with considerable success for following the carbon content in the furnace bath. As soon as sufficient data had been compiled to indicate that the instrument might have a wider application, a more rugged model was constructed and placed in the mill to be tested on commercial heats. Finally, about a year ago, instruments were furnished to the plant open-hearth shops, and these have been in use since that time. As a matter of convenience, the instrument has been named the "Carbanalyzer," and will be referred to by this designation in the following pages. $ Eaklier Investigations While by far the greatest number of magnetic investigations in which the carbon content of ferromagnetic materials was considered have had to do with the evaluation of the magnetic properties of steels of known

Jan 1, 1940

By C. T. Bischoff

This paper describes the results of testing with colorimetric methods, using "dithizone", soil samples taken over various known copper and zinc deposits covered by glacial till. Variation in results is shown for different depths and types of overburden as well as grade and type of deposit. Limiting factors for the practical application of this work are discussed. The method has proved successful for copper and zinc deposits under clay and fine sand with depths up to at least 30 ft but fails in coarse sand and gravel unless depths are very shallow. Swamps require penetration to the underlying subsoil for sample material. DURING the past few years geochemical testing for traces of heavy metals in water, soil, rock, and vegetation has aroused increasing interest. Various techniques for field and laboratory determinations have been described in the literature, as have some of the results of application. However, little information has been published regarding application to glacial terrain. Over 90 pct of the Canadian Shield is overburdened, resulting in high costs for mineral exploration once the outcropping ore deposits have been found. This has been largely responsible for the widespread application of mining geophysics in Canada, permitting concentration of effort on area or zones considered most favorable. However, geophysical interpretation has its limitations, since at least 90 pct of the structures indicated thereby contain no valuable minerals. A more direct approach is therefore highly desirable. If underlying base metal mineralization can be detected by sampling soil or vegetation at moderate cost this should prove of great help in locating new discoveries or extending known ones. A large proportion of Canadian ore deposits contain appreciable copper or zinc, even where these are not the principal economic metals. Consequently the field of application is a broad one. This paper deals with investigations along these lines with particular emphasis on soil sampling. The problems involved from the start were: 1—Which types of glacial soil, if any, would best lend themselves to testing, and conversely which types might be considered relatively hopeless? 2—Through what depths of overburden could significant mineralization be detected? 3—What sampling techniques would prove economically applicable? This included minimum depth of sample and maximum spacing that could be allowed.

Jan 1, 1955

Jan 3, 1953

By T. H. Alden

Using alternating tension-compression straining, the hardening of magnesium oxide single crystals was studied up to large stresses and strains. At 0.25 pct plastic strain amplitude, the hardening curve is approximately linear with slope 25,000 psi from the shear yield stress, 7 to 8000 psi, to 35,000psi. Above this stress, the slope decreases. The strain hardening behavior of MgO is considered qualitatively similar to that of metal single crystals. The relatively high stress attainable by strain hardening is associated apparently with the high yield stress on the cross-slip system, (001) <110>. Cleavage fracture during testing is uncommon. It is argued that the centers of high internal stress at glide band intersections, at which cracks tend to nucleate, are dispersed by cyclic strain. Special features of the glide band structure produced by cyclic strain and revealed by dislocation etch pits, support this view. Strain hardened MgO has mechanical properties greatly superior to the as-received material: yield stress, greater than 100,000 psi; elongation to fracture about 1 pct. A material is said to strain harden if the yield stress increases with an increment of plastic strain. This definition is usually applied for straining done in one direction, but is also applicable when the strain direction is periodically reversed, Fig. 1. For certain metal single crystals, data are available which permit a comparison of the hardening behavior for cyclic straining and for tension straining.&apos;-4 With certain qualifications, these data show that the same processes of hardening are operative in each type of test.5 Despite this fact, the importance of the technique is not immediately evident, although tension-compression studies of the common metals appear to suggest some deficiencies in theories of strain hardening developed exclusively on the basis of tensile tests. However, a recent observation suggests that the cyclic straining method may be very useful for studying semibrittle crystals in which large plastic strains are not accessible in unidirectional testing. The observation is that zinc crystals, when strained in tension-compression at -52°C, do not fail by cleavage at low stress (-500 psi)6 as they do in tension, but harden to a limiting stress of more than 5000 psi over a total plastic strain of about 600 pct.2 An important characteristic of the behavior of zinc crystals is the high stress, relative to the yield stress, attainable by strain hardening. By comparison, the hardening of aluminum single crystals tested by an identical technique saturates at 1100 psi. This difference is best explained by the cross-slip hypothesis of dynamic recovery.7,8 In zinc, cross slip is difficult because of the high yield stress for glide on planes other than the basal plane in the < 1120 > zone. The present work was undertaken in order to test whether these methods and ideas are applicable to other materials. Magnesium oxide single crystals, in common with most crystals of the rock-salt structure, deform plastically but fail by cleavage after a small strain when tested in tension. It was hoped that larger strains would be attained using tension-compression. There is, in addition, evidence 8a which shows that slip on the probable cross system, (001) < 110>, is difficult in magnesium oxide; it may therefore be possible to attain high stresses by strain hardening. 1) EXPERIMENTAL PROCEDURE Experimental methods used in this study were based in part on techniques reported in papers of Stokes, Johnston, and Li.&apos; MgO blocks, purchased from Norton Co., were used without further annealing. Specimens were cleaved to dimensions approximately 0.125 in. sq and 1 in. in length. The gage section, formed by chemical polishing, was sprinkled with 280 mesh silicon carbide particles in order to introduce fresh dislocations. The crystals were then cemented into cylindrical aluminum adapters and clamped in an Instron testing machine. One of two alternating straining programs was used. In the first, total cross-head travel was established and increased in steps after various numbers of cycles. In the second, a capacitance gage was used to directly measure the elongation of the specimen and the crosshead was controlled so as to keep the plastic strain amplitude constant. The straining was always symmetrical with respect to the initial, zero strain condition. While both procedures produce strain hardening, only the latter permits a measure of the total plastic strain so that hardening curves may be drawn. Constant plastic strain amplitude tests were done

Jan 1, 1963

By Linn Bradley

F. H. WILLCOX, Pittsburgh, Pa. (communication to the Secretary*). -We must keep in mind, in balancing the savings-to be anticipated by the most efficient combustion of gas, the best heat absorption by stoves or boilers, and by heat conservation of the thermal equivalent carried in. blast-furnace gas-against the cost in operation, maintenance, depreciation and interest charges of the equipment required to realize the savings, that in boilers and hot-blast stoves we are basing our economies upon two different, combustion phenomena. In a boiler, evaporating water at a heat efficiency of 70 per cent., approximately 47,550 B.t.u. is consumed per boiler horsepower. Taking the blast, furnace gas cited in the above paper, we are using about 38 ¾ lb. of cold-washed gas per boiler horsepower. Using the hot-cleaned gas, the same equivalent heat is furnished by 34 1b. of the gas , This leaves of the 38 ¾ lb. required in the case of cold-washed gas, 4 ¾ lb. or 6,650 B.t.u. available for raising more steam when dry-hot gas is used, which means that an equivalent amount of coal firing can be dispensed with. Per ton of iron, this figure 2/3 X 12,000/38.75 X 4.75 = 980 lb. gas or 1,372,000 B.t.u. or 100 lb. coal. At $1.50 per ton of coal, the net saving per ton of pig iron produced by the use of dry-hot over wet-cold cleaning will, therefore, be about $0.06 2/3. At a stove heating 8,500 lb. of air per toil iron from 150° F., one-quarter of the dry-hot cleaned gas, or 3,000 lb., would heat the air to 1,450° F. 3,000 X 1,400 X 0.70 = (8,500 X 0.255) = 1,300. The cold-wet gas would heat the same amount of air to 1,280° F. 3,000 X 1,225 X 0.70 = (8,500 X.0.255) = 1,130.

Jan 4, 1917

By H. F. Bishop, F. A. Brandt, W. S. Pellini

M. S. Fisher and D. R. F. West (Imperial College of Science and Technology, London, England)—It may be of value to compare certain features of the results recorded in this very interesting paper with those of an investigation11 involving determinations of temperature gradients and rates of solidification of steel blocks of 3 in. sq section, cast horizontally in sand molds, some of which contained chills. Castings in plain carbon steels containing up to 0.8 pct C were made and, throughout this range of compositions, both inverse-rate and direct cooling curves, obtained from thermocouples placed within the mold cavity, showed a major arrest, the temperature of which was taken as the "effective liquidus" for the particular steel in question, solidifying under the conditions of the investigation. At the completion of the arrest, the curves showed a comparatively rapid de- crease in temperature. Some of the steels investigated lay within the peritectic region, and a second arrest at 1485°C was detected. In Fig. 15, which shows the effective liquidus temperatures, the limit of the 8 + liquid field has been placed at approximately 0.5 pct C. In inverse-rate curves corresponding to the thermal center of each casting, there was always a discontinuity, viz: an abrupt change of slope, at a temperature about 10 to 20°C below the "liquidus arrest," after which there seemed to be no significant irregularity in the rate of cooling. Curves corresponding to other positions in the casting generally showed some irregularities below the main arrest, probably as a result of heat flow from the solidifying interior. Even the reproducible discontinuity in the curves for the central positions appeared to have no theoretical or practical significance, though it probably corresponded to a stage in the solidification process at which only a relatively small amount of interdendritic liquid remained. Its temperature, rela-

Jan 1, 1953

By Jacob H. Masliyah

The dynamic simulation of a gravity separation vessel (a high throughput classifier) has been developed. The feed slurry was assumed to contain light (oil droplets) and heavy (mineral solids) particulates of various size. The feed slurry was characterized by the particle size distribution of both the light and the heavy particles. The equations describing the slip velocity for each specie and those defining the time variation of each specie concentration were solved implicitly using Runge-Kutta algorithm. The dynamic simulation was capable of predicting the movement of the froth layer and the density variation of the under- flow. Field data are in agreement with the simulation model and it is shown how this model could be used to develop a control strategy.

Jan 1, 1984

By Frank H. Reed, P. W. Henline, Harold W. Jackman

A SUMMARY of the consumption of coal in 1945 shows that the coke industry accounted for 17 pct of the total coal used. No substitute for coke and the blast furnace in the reduction of iron ore has gained sufficient prominence so far to be given serious consideration. However, while the same slot-type ovens are being built and used now as were in use a few years ago, the problem of obtaining suitable coking coals, both high- and low-volatile, has forced many operators to revamp coal-blending facilities. Until the beginning of World War II, high-temperature coke plants were able to secure the required high-volatile and low-volatile coals from two or three seams and from three or four mines. Thus, many of our coke oven plants were able to operate with storage bins for two coals only. The variance in raw materials was not great, and uniformity of blend was a relatively simple problem. Since the beginning of World War II, many and important changes have taken place. Demand for all coals increased greatly. Production, though kept as near maximum capacity as possible, could not meet the demands for coal generally, and particularly for the various kinds of coal which have proved to be satisfactory for specialized uses; coals were allocated as nearly as possible in the order of greatest need; storage piling was reduced to the minimum necessary to keep plants in operation; uniformity of quality could not be maintained, and with this last, uniformity of composition of blends for coking could not be maintained. Whereas before the war producers of metallurgical coke could fill their needs from three or four mines, by the end of the war it was not unusual to find shipments to a single coke oven plant coming from as many as 30 different mines, in a single month. Since many of the mines producing both high- and low-volatile coal for high-grade metallurgical coke are rapidly nearing depletion, the problem of seeking new sources and of blending a number of coals in each plant for the production of high-temperature coke is facing practically every producer in the United States at the present time, and is one which will probably continue for many years to come. Midwestern by-product coke ovens in the Chicago and St. Louis areas use annually from 12 to 15 million tons of bituminous coals which are transported 500 to 700 miles from the Appalachian coal fields of Pennsylvania, West Virginia, and eastern Kentucky. Approximately two-thirds of this coal is high-volatile bituminous. The critical transportation problem confronting the nation in 1943, and the growing scarcity of the best Appalachian coking coals, prompted the Illinois Geological Survey to propose a research program in which would be studied the coking properties of blends of low-sulphur, high-volatile Illinois coal with the high- and low-volatile coals from the eastern fields. Such blends containing Illinois coal, if substituted for the all-eastern blends normally coked,

Jan 1, 1948

By J. B. Newsom

During July and August, 1931, the Bloomington Limestone Co., at Bloomington, Ind., ran a single wire saw on ledge No. 2 at Maple Hill quarry. The ledge was hard rock, much harder than the average Indiana limestone. It was opened at two ends for a depth of about 20 ft., leaving a 130-ft. section between. During the tests 21 cuts were made in this ledge, to an average depth of 10.42 ft., for a total effective cutting area of 28,445 sq. ft. The actual area sawed was about 20 per cent higher, as we did not level down in the center, preferring to cut deeper on the ends. The various cuts were carefully logged, and the results plotted. A study of these results is interesting, as it enables a forecast of probable future costs to be made, and also because it points out the limitations of wire saws, and the lines development should logically take. Some of the inferences to be drawn are discussed below. Best Length of Cut.—No trouble was due to length of cut, except that we could not easily level down a long cut, and it is probable that longer cuts, say up to 200 ft., can be made. Obviously, longer cuts will result in a higher sawing rate in square feet per hour if they do not cause stoppages. Working Higher Ledges.—The cuts averaged 10.42 ft. high by 130 ft. long. A study of several logs indicated that during only about 15 per cent of the cutting time was the saw cutting to its full efficiency; that is, during about 85 per cent of the time the saw was bedding in at the ends and not cutting well in the center, or the ends were below the floor line and only part of the wire was doing effective cutting. This does not mean that the wire was cutting to only 15 per cent of its capacity, but it means that during about 85 per cent of the time part of the cutting was being wasted. If higher ledges had been worked, the wire would have cut to its full efficiency a large part of the time, as it would have had a greater distance to saw after it had bedded in to the ledge before the ends went down below the floor level. The exact saving to be expected cannot be calculated, but it would be large if really high cuts were made.

Jan 1, 1932

By R. Fahn

Bentonites are best suited for the preparation of acid activated clays with efficient adsorption properties. They are weathering products of acid volcanic ashes and natural deposits are spread worldwide. The main constituent of bentonite is aluminum hydrosilicate montmorillonite with a high ability to swell. Physical and Chemical Properties of Montmorillonite The clay mineral montmorillonite is named after its place of discovery, near Montmorillon in southern France. The structure of this clay is shown in [Fig. 1]. The montmorillonite crystal consists of about 15-20 silicate plates. Each silicate plate is composed of three layers. The central one is formed by aluminum or by ferric hydrate octahedrons, the two outer ones consist of silicon dioxide tetrahedrons. In the presence of water, the montmorillonite crystal swells and water molecules become deposited between individual silicate layers. Two types of natural bentonites are known: the Western type, which is a natural sodium bentonite (high swelling grade), and the Southern type, an alkaline earth or calcium bentonite (low swelling grade). Southern bentonites have efficient adsorption properties of their own. They are called natural active earths or fullers earths. Acid Treatment of Bentonite Under the influence of a mineral acid, the exchangeable alkaline earth ions of bentonite are substituted by hydrogen ions. Moreover, depending on the acid concentration, temperature, pressure, and reaction time, a more or less pronounced leaching process takes place. Thereby Al-, Fe-, and Mg-ions are extracted from the octahedron layers. This results in a change in the crystalline structure of montmorillonite and in an increase of its specific surface area and porosity, as shown in [Fig. 2].

Jan 1, 1983