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By M. K. Seguin

The study described in this paper is first centered on the determination of the physical properties (gravimetric, magnetic, electrical, and electromagnetic) o f metamorphosed iron ores of the Carol Lake-Labrador City District. Different geophysical surveys available in the Carol Lake-Wabush area were interpreted quantitatively and the dimensions, attitude, and tonnage of the ore bodies investigated were determined. A geophysical-statistical study was undertaken which permitted the determination of the grade of iron, magnetite, and, subsequently, specularite in the ore bodies. A discussion of the theory and logical progression of the study and the application of the results to actual geophysical data is presented. The conclusions, which are of real significance from the economic standpoint, may be predicted from the data obtained from geophysical surveys before any diamond drilling is undertaken. An integrated interpretation of this type is useful in planning future exploration work and selecting the most favorable geological structures. It also serves as a guide to minimize and determine tonnage drilling patterns. The general conclusion that a restricted range of residual gravity and magnetic anomalies are a key to exploration is valid. Indeed large depths to the top of deposits are absent since all the ore bodies are overturned synclinal structures with tops located close to the surface. Certain irregular geometries do occur but they are not common and the alignment of the iron ore bodies is sub parallel so that magnetic anomalies are evenly reduced in the earth's magnetic field vector.

Jan 1, 1975

The oil production in Bahrein for the year 1937 was 7,762,264 bbl., a monthly average of 646,855 barrels. Sufficient potential production had been developed in the Second Pay (Main Pay) at Bahrein by the end of 1936 to supply the anticipated crude demand, and the drilling program was changed from straight development to partial exploration during 1937. Two nonproductive wells were drilled to test the Third Pay at a depth of about 2900 ft., or some 900 ft. below the top of Second Pay. As a result of this exploratory work, the normal drilling program was slowed down and only eight new wells were completed in the Second Pay during 1937, making a total of 45 producers from the Second Pay. The deep test was drilling at 5081 ft. at the end of the year. A 550-ft. section (termed Fourth Pay) of porous limestone interbedded with anhydrite was topped at about 4150 ft. Numerous drill-stem tests indicated that the limestone was very permeable and that it carried large volumes of natural gas under high pressure together with a small amount of high-gravity oil or condensate. Drill-stem tests also indicated that the bottom 50 ft. of the 550-ft. section carried an encouraging amount of heavier oil, gravity about 37" A.P.I., with characteristics similar to the oil produced from the Second Pay. A well was started somewhat down-structure from the deep test to better determine the oil-producing possibilities of the Fourth Pay. Some refinements in drilling technique were adopted during the year. All wells started to the Second Pay were drilled with rotary from the surface. The top 1500 ft. of porous limestone section, formerly drilled with cable tools, is now drilled with rotary without circulation, using water to wash the cuttings into cavities. The shale section from this point to the top of the oil horizons is drilled with natural mud circulation. The productive horizons are drilled with oil circulation under pressure control. A second satisfactory method of completing wells was developed when the two Third Pay test wells were plugged back and gun-perforated through cemented casing opposite the Second Pay, with good results.

Jan 1, 1938

By M. H. Forester, John D. Cooner

ALTHOUGH the unmined anthracite will last for approximately A 150 years, most of the thicker and cleaner coal beds have been almost entirely first-mined and pretty well robbed, leaving much of the present and future production to come from thin and dirtier beds. Herewith is a section of one of the dirty beds being mined in the Wyoming region This section shows that 45.8 per cent of the bed is made up of impurities and that the 45 in. of coal in the bed is in five benches, the thickest of which is 19 inches. In some localities of the anthracite region there are 14 beds of minable coal. Orderly and proper mining of the coal makes it necessary to rob or second-mine-almost all beds have already been first-mined-in sequence from the top downward. Therefore the robbing of the few remaining portions of the thicker and cleaner beds is in many instances being held up by the mining of the overlying thin and dirtier coal beds. Impurities Impurities consist of fire clay, slate, rock, and bone. The latter, which is black, contains 40 to 60 per cent carbon and often resembles good coal. In some beds there are also bands of "blue" coal, which, as the name implies, is of a bluish slaty color. Although much of this "blue" coal has a carban content almost as high as the good anthracite, it is not

Jan 1, 1943

By A. M. Gaudin

A great number of hypotheses has been advanced to explain the complex phenomena that are encountered in flotation. In the days of bulk-oil flotation, when a large quantity of oil was employed, it was observed that the oil-mineral aggregate floated above the ore-pulp because of its lesser density. With the advent of froth-flotation this phase of the theory had to be modified to take into account the presence of air bubbles. The reason why the mineral-bearing froth of today rises to the top of flotation cells is, of course, because the air-mineral aggregate, with whatever reagent may be associated with it, is lighter than the pulp. In an attempt to explain froth-flotation the fact has been emphasized that certain minerals are wetted and therefore remain in water; while others are not wetted and therefore become associated in preference with air. Such an "explanation" does not explain facts, it merely states them; for instance, it does not explain why minerals are wetted under some circumstances and not wetted under others. Going somewhat further, the wetting or nonwetting of minerals by water has been ascribed' to differences in the angles of contact between oil droplets and mineral surface at the surface of the minerals. This explanation, however, was abandoned when soluble flotation reagents came to replace the so-called "insoluble" reagents. Another theory, sponsored about 10 years ago by Ralston,² involves the effects of electric charges at the surface of the mineral particles and air bubbles. This theory, unfortunately, was not developed as fully as it seems desirable. The belief that adsorption phenomena might play a preponderant role in flotation3 has been rather widespread for the last 10 years. Unfor-

Jan 1, 1928

By Byron M. Bird

JIGGING is the stratification of a mass of solid particles in upward pulsations of water or in alternating upward and downward pulsa¬tions. The stratification usually is effected in a rectangular open¬top container having a perforated bottom through which the pulsations of water are applied to the materials to be separated; in separating coal, to a mass of coal and refuse. Jigging usually is a continuous process involving three steps: (1) the reception of the raw coal, (91) the stratification of the coal and refuse so that the refuse settles to the screen plate, and (3) the overflow of the cleaned coal and the withdrawal of the refuse collected on the screen. The jig is the most universal washer ever devised. It is handling in practice a wide range of closely sized coals and ores, mixed sizes of coal from 8-in. to 0 through %-in. to 0, and mixed sizes of ores down to %3 O-in. to 0 and even smaller. It is treating materials ranging in specific gravity from that of gold at 19 to that of coal at 1.30. It is handling feeds in which the proportions of high-gravity materials range from 90 per cent or higher, as for some iron ores, to 10 percent or less for some coals. It is handling feeds containing all shapes of particles-round, rectangular, and flat. It is handling simple separation problems and difficult separation problems. Thus, it is an almost universal separating device. One might add that it has been one of the most misunderstood types of washers, from both a theoretical and a practical point of view. Although con¬siderable progress has been made in coal jigging in the past 10 years, the jig probably has more unexplored possibilities than any other type of washer. In this chapter the object has been to treat the jigging process in a practical manner and to give information that will help the reader to improve his particular operation. Inasmuch as performance data from operating plants would contribute little of value, none has been given. Instead, the emphasis is upon jigging principles as a foundation for attacking the problems that arise. These principles are in reality tenta¬tive conclusions from research in the field of jigging during the past 10 years and most of this chapter should be regarded as a progress report on an investigation still under way.

Jan 1, 1943

By GEORGE L. CHRISTIAN

T HE Tonopah Extension assay office is a two- story, concrete structure on a solid foundation of andesite, situated about 100 yd. from the company's mill, so that it will not be affected by the stamps or the tube-mills; also it will not be affected by under- ground blasting. On the second floor are the mine superintendent's office and drafting and blueprint rooms. On the ground floor are the assayer's office,' mixing room, furnace room, crusher room, chemical laboratory, parting room, store room, and change and shower room for the mine superintendent, assayers, and engineers. The crusher room is outside of the main building, and has four concrete walls, thus eliminating the jar of the crusher; it is so situated that the least possible amount of dust spreads through the main building. An 8 by 10-in. jaw crusher, which crushes to ¼ in., was made by a local foundry. Attached to the top is a small galvanized-iron hopper covered with a ¼ in. screen, on to which the sample is dumped. The fines-pass through the screen into the hopper, while the oversize is dumped into the crusher. The crushed sample and fines from the screen drop down to a set .of splits suspended from an iron rod below the crusher jaws. The small sample is caught below in a small pulp pan. The rejects drop from the splits through a 2-in. pipe on to a bucket conveyor, which dumps into a galvanized-iron hopper in an upper corner of the room. The rejects drop through a 2-in. pipe to a reject bin outside the build- ing. This bin is constructed high enough for a truck to be loaded from it by means of a chute door on the side.

Jan 1, 1921

By S. J. Rothman, L. T. Lloyd

S. J. Rothman and L. T. Lloyd (Argonne National Laboratory)— Hagel's paperz0 discusses the question of the mechanism for self-diffusion in bcc transition metals. Recent work21"23 has led us to agree with him in reabandoning the ring mechanism in favor of a defect mechanism, most likely vacancies. On the other hand, we think that Hagel's attempt to group together all self-diffusion in bcc transition metals is wrong. We prefer the scheme of Peart, Graham, and omlin," who separate them into two groups. One group contains the anomalously behaved metals P titanium,', y uranium, and perhaps zirconim.' The other group contains a iron," this alloy was held at 1800°C for 2 hr before quenching, its structure was all a. B. C. Giessen, I. Rump, and N. J. Grant (Authors' reply)—We very much appreciate the discussion by Deardorff and Kato regarding the a to /3 transformation temperature of hafnium. In the W-Hf program, as well as in a current study of the system Cr-Hf, standard precautions were exercised against contamination, including vacua of less than 10"5 mm Hg and a weight control check of all specimens to insure an impurity rise of less than 100 ppm. Following receipt of the above discussion, a revision was made in the annealing process. Arc-cast specimens were etched with a 10 pct HF-90 pct HNO solution until 1 mm of surface, including a possible oxide film, was removed. Annealing was done in a hafnium container which provided a long diffusion path for gaseous contaminants. Finally, the heat treating time was shortened. Structural changes were noted to occur quite slowly. A 99.5 at. pct Hf alloy showed 70 pct /3 after 30 min at 1700°C and was 100 pct /3 after 10 min at 1715°C. After 24 hr at 1740C, the alloy was found to be 10 pct/3. This change is due to contamination; it was estimated that it corresponds to an increase of 60°C in the transformation temperature over a 24 hr period. The corresponding change in 30 rnin would be slight. Using the same hafnium as in the above publication, and three new alloys at 98, 99, and 99.5 at. pct Hf, the transformation temperature was indicated to be 1715" 15°C. Since this hafnium contained 4.4 at. pct Zr, a correction of 40 5C, based on an assumed approximately linear transformation temperature curve, must be made for pure hafnium. This yields a value of 1755 20°C in good agreement with Deardorff and Kato's value. Because the slope of the boundary is small, this permits a good extrapolation, and we regard the W-Hf alloys as well suited to check the Hf transformation temperature. niobirn,'' and molybdenum,22 whose activation energies correlate with the melting temperature and whose Do values follow Zener's predictions.28 We have no quarrel with Hagel's placing chromium into the second group; we disagree with his interpreting our data to make uranium fit into that group. Hagel considers that the measurements of self-diffusion in uranium below about 950°C represent combined volume and grain boundary diffusion. This contention is based on the curvature of the penetration plot for our 891°C run,4 which also has a nearly straight line plot of log c vs x. Hagel has extrapolated these characteristics to our 803.5C run and fitted his Arrhenius plot for self-diffusion in y uranium to our four highest temperature points only. We think this is wrong for the following reasons.

Jan 1, 1963

By Y. Kondo, A. M. Gaudin

A combined method for sizing in the sub-micron range was developed using centrifugation and gamma emission. Its essential features are: 1) Centrifugation of suspended mineral in water carried out at 4°C where water has its maximal density and zero temperature gradient of density 2) Gamma activition of the sample suspension induced by irradiating it in a nuclear reactor prior to centrifugation 3) Measurement of the activity of a beam of suspension as it is sedimenting in the centrifuge, using a scintillation counter provided with a discriminator. Albite, the sodium feldspar, was used as feed to the comminution step, and the induced gamma activity of Na24 was measured. Dissolved sodium ion caused a a non-settling activity in the measurement. Other "noise" is due to scattered gamma radiation from albite particles in the centrifuge outside of the line of sight. These two activities were obtained sep arately and subtracted from the total activity counted. The relative solid concentration in the suspension under the collimating slit is proportional to this net activity. The cumulative weight per cent of undersize particles was calculated and plotted against size on logarithmic scales. The relationship was found to be linear down to about 30 mµ. The specific surfaces calculated from particle size distribution agree with those measured by krypton gas adsorption. The measurement of particle size distribution of comminution products has been of interest in clarifying the mechanism of the process. The Gaudin-Schuhmann equation1,2 relates the cumulative percentage undersize, y, to relative particle size x/k. The particle size, x, and the size modulus, k, together with the distribution modulus, m, characterize the particle size distribution, m is the slope of a plot of log (y) against log (x/k). This equation has been verified for ground particles down to 1 micron in size.3'4 For particle sizes smaller than a few microns, the precision and facility of sedimentation and microscope techniques become questionable. Use of a centrifuge which shortens settling time from weeks to hours is attractive. Many centrifugal methods have been proposed; however, determination of the solid concentration of the suspension at a reference level as a function of time is difficult, and sampling the suspension is likely to introduce errors. Measurement of particle size distribution by X-ray absorption in gravitational5 and in centrifugal6 sedimentation have also been tried. In these methods, the absorption coefficient for the solid is independent of size, since the wave lengths of the X-rays and gamma-rays are orders of magnitude smaller than the particles. Radiometric techniques7 in size distribution measurement have been developed within the last several years. They offer advantages over usual gravimetric and optical methods. PRINCIPLE OF METHOD The combined method of centrifugation and gamma emission for sizing in submicron range described in the present paper has the following essential features. 1) Centrifugation is carried out at 4 ± 0.1°C. Convection currents in the water phase due to density changes caused by temperature variation may disturb completely the settling system. To avoid this difficulty, the unique property of water having zero temperature gradient of density at 4°C is used. 2) Prior to size measurement the sample suspension is irradiated in a nuclear reactor and gamma activity induced. The gamma activity of the particles sedi-mented radially is measured by a scintillation counter provided with a discriminator. By this method, interruption of centrifugation by withdrawing samples is avoided and the counting rate at any time represents the solid concentration in the suspension at the level of the collimating slit of the centrifuge at the particular time at which the measurement is made. 3) By proper selection of the feed material and of the comminution device, gamma activities of foreign

Jan 1, 1964

By Bunting S. Crocker

EBBLE grinding at Lake Shore is not a temporary BlE wartime substitute. The tube milling plant, with a 1000 ton per day capacity, grinds a hard siliceous ore to 90 pct — 325 mesh. The plant, prior to using pebbles, was consuming 4.3 lb of 11/4-in. grinding balls per ton of ore, which amounted to 785 tons of balls per year. At September 1951 prices, $132.60 per ton, this steel cost amounted to $104,400 per year, or $0.285 per ton milled. By the Lake Shore method of substituting screened rock for this steel, all of this cost is saved. This is one of the major economies in Lake Shore mill practice. Regardless of the ultimate price of grinding balls, a change back to steel balls is not considered. The present pebble plant is more flexible than a steel ball plant and equally efficient. For example, if it is desired to change the size of grinding media, the pebble charge in any mill can be changed completely in 4 to 5 days, as against 76 days to change completely a 1-in. steel ball charge. To change pebble size it is necessary only to change two sets of screens and clean out the rock feed storage bin. This will take 8 to 10 days as against 3 to 4 months to clean out the customary supply of grinding balls kept on hand. Also it has not always been possible to purchase all desired sizes of balls at any price. In an analysis of savings effected by the use of the pebble mills, the flexibility of the Lake Shore grinding plant should be discussed, as it has a direct bearing on these savings. The plant has always used several units to handle tonnage rather than sending all the tonnage through a single unit. This principle may result in using small diameter mills, but no objection to that is seen. At no time has any advantage been found in cost per ton ground in large diameter mills. Both the capacity and the power of any mill varies as the diameter raised to the 2.6th power. Consequently large or small mills are equally efficient, and a plant, should be designed to use as many units or combination of units as is consistent with reasonable operating practice. Mills under 5 ft diam are harder to reline, etc. To define the case at Lake Shore, 2600 tons per day formerly were milled in seven 7x6-ft ball mills and twelve 5x16-ft (and two 6x16-ft) tube mills.* This gave an excellent test plant and an extremely efficient one. In this plant the ratio of ball mills to tube mills was 1:2. When the much cheaper pebble mills were substituted for the tube mills, this ratio was changed to one ball mill to four pebble mills to take the greatest possible advantage of the cheaper operating mill, i.e., the pebble mill. This flexibility without loss in efficiency has been an important item in the cost savings. It is interesting to note that the use of pebbles for fine grinding was proved first in the laboratory in a 12-in. ball mill. In fact, since 1934 all testing on — 8 mesh material has been done in this 12-in. mill. Scope of the Tests A paper on fine grinding at Lake Shore Mines was published in July 1940.' This paper covered 7 years of intensive research on fine grinding as well as sizing methods and equipment, plant scale grinding tests on 5x16-ft tube mills with and without grate discharges both with 1-in. and -in. balls, the use of laboratory mills to evaluate plant changes, and several reports on classifiers and classification. In the following July the addendum report2 was added in which the idea of series-circuit grinding was introduced, and the results of running five stages of tube mills and bowl classifiers were shown. Since 1940, the ball milling end of the plant has been altered extensively as a result of tests on the use of 3-in. rods in 7x6-ft mills and the use of the Tyler repulping screen with from 7 to 14 mesh screens. These tests are lengthy and may be covered in a separate report later. The scope of this report is confined to ore ground by rod milling and ball milling until it passed through an 8 mesh Tyler Ty-rod screen. The —8 mesh screen undersize then was pumped to a primary bowl classifier in open circuit and the sands from the bowl sent to the primary pebble mills, see Fig. 1. In the pebble mill circuit the ore is ground to 90 pct —325 mesh (24 pct + 28 microns). In studying the flowsheet, attention should be paid to the efficiency of the classification equipment used. The Tyler repulping screen is an efficient machine on the 8 to 10 mesh separation, and the bowl classifier is equally efficient at the 325 mesh separation. Efficient classification is a necessity for series-circuit stage grinding. The ore is hard siliceous porphyry, 60 pct SiO,, 80 pct insoluble. Its grindability at different meshes has been shown near the top of the list in F. C. Bond's grindability tests.' Lake Shore is not shown, but an adjacent mine with identical ore, Wright-Hargreaves, is. Reasons for the Changeover Since 1936 grinding balls have been rising steadily in price with no sign of stopping. For a mill that used 4.5 to 5.2 lb of grinding balls for every ton of ore ground this rise represented an alarming increase in grinding costs. In many cases the quality of the grinding balls fell off as the scrap steel became more difficult to obtain. The ratio of tube mills to ball mills increased with the use of the Tyler repulping screen in the ball mill circuit. Originally only 3.0 lb of balls per ton were used in the tube mills, and this

Jan 1, 1953

By Bunting S. Crocker

EBBLE grinding at Lake Shore is not a temporary BlE wartime substitute. The tube milling plant, with a 1000 ton per day capacity, grinds a hard siliceous ore to 90 pct — 325 mesh. The plant, prior to using pebbles, was consuming 4.3 lb of 11/4-in. grinding balls per ton of ore, which amounted to 785 tons of balls per year. At September 1951 prices, $132.60 per ton, this steel cost amounted to $104,400 per year, or $0.285 per ton milled. By the Lake Shore method of substituting screened rock for this steel, all of this cost is saved. This is one of the major economies in Lake Shore mill practice. Regardless of the ultimate price of grinding balls, a change back to steel balls is not considered. The present pebble plant is more flexible than a steel ball plant and equally efficient. For example, if it is desired to change the size of grinding media, the pebble charge in any mill can be changed completely in 4 to 5 days, as against 76 days to change completely a 1-in. steel ball charge. To change pebble size it is necessary only to change two sets of screens and clean out the rock feed storage bin. This will take 8 to 10 days as against 3 to 4 months to clean out the customary supply of grinding balls kept on hand. Also it has not always been possible to purchase all desired sizes of balls at any price. In an analysis of savings effected by the use of the pebble mills, the flexibility of the Lake Shore grinding plant should be discussed, as it has a direct bearing on these savings. The plant has always used several units to handle tonnage rather than sending all the tonnage through a single unit. This principle may result in using small diameter mills, but no objection to that is seen. At no time has any advantage been found in cost per ton ground in large diameter mills. Both the capacity and the power of any mill varies as the diameter raised to the 2.6th power. Consequently large or small mills are equally efficient, and a plant, should be designed to use as many units or combination of units as is consistent with reasonable operating practice. Mills under 5 ft diam are harder to reline, etc. To define the case at Lake Shore, 2600 tons per day formerly were milled in seven 7x6-ft ball mills and twelve 5x16-ft (and two 6x16-ft) tube mills.* This gave an excellent test plant and an extremely efficient one. In this plant the ratio of ball mills to tube mills was 1:2. When the much cheaper pebble mills were substituted for the tube mills, this ratio was changed to one ball mill to four pebble mills to take the greatest possible advantage of the cheaper operating mill, i.e., the pebble mill. This flexibility without loss in efficiency has been an important item in the cost savings. It is interesting to note that the use of pebbles for fine grinding was proved first in the laboratory in a 12-in. ball mill. In fact, since 1934 all testing on — 8 mesh material has been done in this 12-in. mill. Scope of the Tests A paper on fine grinding at Lake Shore Mines was published in July 1940.' This paper covered 7 years of intensive research on fine grinding as well as sizing methods and equipment, plant scale grinding tests on 5x16-ft tube mills with and without grate discharges both with 1-in. and -in. balls, the use of laboratory mills to evaluate plant changes, and several reports on classifiers and classification. In the following July the addendum report2 was added in which the idea of series-circuit grinding was introduced, and the results of running five stages of tube mills and bowl classifiers were shown. Since 1940, the ball milling end of the plant has been altered extensively as a result of tests on the use of 3-in. rods in 7x6-ft mills and the use of the Tyler repulping screen with from 7 to 14 mesh screens. These tests are lengthy and may be covered in a separate report later. The scope of this report is confined to ore ground by rod milling and ball milling until it passed through an 8 mesh Tyler Ty-rod screen. The —8 mesh screen undersize then was pumped to a primary bowl classifier in open circuit and the sands from the bowl sent to the primary pebble mills, see Fig. 1. In the pebble mill circuit the ore is ground to 90 pct —325 mesh (24 pct + 28 microns). In studying the flowsheet, attention should be paid to the efficiency of the classification equipment used. The Tyler repulping screen is an efficient machine on the 8 to 10 mesh separation, and the bowl classifier is equally efficient at the 325 mesh separation. Efficient classification is a necessity for series-circuit stage grinding. The ore is hard siliceous porphyry, 60 pct SiO,, 80 pct insoluble. Its grindability at different meshes has been shown near the top of the list in F. C. Bond's grindability tests.' Lake Shore is not shown, but an adjacent mine with identical ore, Wright-Hargreaves, is. Reasons for the Changeover Since 1936 grinding balls have been rising steadily in price with no sign of stopping. For a mill that used 4.5 to 5.2 lb of grinding balls for every ton of ore ground this rise represented an alarming increase in grinding costs. In many cases the quality of the grinding balls fell off as the scrap steel became more difficult to obtain. The ratio of tube mills to ball mills increased with the use of the Tyler repulping screen in the ball mill circuit. Originally only 3.0 lb of balls per ton were used in the tube mills, and this

Jan 1, 1953

By R. C. Gifkins, K. U. Snowden

The value of the index n in power ktivs for the stress sensitivity of minimum creep rale at lead is derived front results drawn from lite literature and from previously unpublished nork on commercial (99.99 pel) and high-purity (99.999 pel) lead polycryslalline and hi crystal specimens. In tests al 0.5Tm (22 to 50'C) for slvesses below 250 psi, n fulls to mily for poly-cr,yslullirle specimens, single crystals, and bicrystals, and toy grain boundary sliding in bicrystals. Between 250 and 500 psi there is a tyansilion range and, for the polycrystalline specimens al least, u high-stress regime for 500 to 1500 psi where n is 5 or 6. These results are such that none of the suggested stress taws can he used to extrapolate to very loir stresses. Meclianisms for deformation in Ike grains and toy grain boundary sliding in the low-strcss regime are outlined. In particular, the mechanism for grain boundary sliding involves sliding controlled by diffusion from tensile to compressire Ledges. predictions based on this model give promising numerical agree-nicul with experiment. II is suggested that the mechanism for grain boundary sliding may he important in other circumstances, e.g., in creep of ceramic materials or superplaslicity, because it can operate with lille or no dislocation aclirity. WhEN the stress dependence of secondary creep rate is represented by a power law the values taken by the stress exponent n are found to vary between 1 and 7, see, c.g., Garofalo.' In Eq. 1 11 u is the applied stress and A a constant. The case for accepting values of v = 1 rests principally upon tests at high temperatures (>0.9Tm) and there have only been isolated tests at lower temperatures (rather than plots of creep rate against stress) for which the creep rates are such that some form of Nabarro-Herring creep with n = 1 is applicable. Hanson5 and Greenwood and cole6 plotted creep rate against stress for tests made in the vicinity of 0.5Tw, and found a "creep yield", so that below a certain low stress the creep rate was linearly related to stress: these observations appear to have been subsequently ignored. Various models have been proposed which agree with a large body of experimental results in which n equals -5 at high stresses (e.g., weertman7 and McLean8). Various alternatives to Eq. [1] have been proposed principally as a means of unifying creep data, but sometimes, it seems, in order to obviate the necessity for assuming a series of regimes of deformation associated with the various values of n; further refer- ence to these will be made in the section on stress laws. This paper calls attention to extensive results which show n = 1 for the creep of lead at room temperature (0.5 T,). RESULTS Fig. 1 is a composite diagram based on figures in the literature6,9,10 with additional results obtained by the present authors and their colleagues11,12 in the course of a number of investigations not directly concerned with stress sensitivity. In Fig. 1 In creep rate is plotted against In stress. The curves show results for two kinds of material. Lead I is of commercial purity (99.99 pct Pb) and lead IT a specially refined material, hydrogen-purified to remove oxygen; it contains 99.999 pct Pb. Curve A is for lead I in polycrystalline form (grain diameter 0.02 to 0.05 cm) at room temperature (22' to 26°C). Curve B1 is for lead I, polycrystalline, tested at 50C. Curve B2 is for lead 11, polycrystalline (grain diameter 0.02 cm), tested at room temperature (22 O- Curves C to F relate to tests made of bicrystals of lead 11: all of the same nominal orientation ("Y" in Ref. 12). Curves C and V are for creep within the component crystals of the bi crystals at 22°C (AB and CD in Fig. 2). Curve E is for creep measured in a gage straddling the boundary, approximately 0.15 mm either side (BC in Fig. 2). Curve F shows rates of grain boundary sliding (as distinct from strain rates in the other curves), obtained by averaging the offsets of six marker lines on the bi-crystal boundaries. The principal feature of the results is that the stress sensitivity falls into two distinct regimes, according to whether the stress is below approximately 250 or above 500 psi (there is a transitional range between 250 and 500 psi). The new result is that below -250 psi two of the curves (A and BI) give a value of n 1, and the other curves also drop below n - 2 to -1.5 but do not extend to low enough stresses to show clear evidence of n 1. At very high stresses the curves tend to give values of n - 6, as previously determined in other investigations. Curve R2 for the high-purity lead in polycrystalline form at room tenlperature almost merges with B, and thus gives new significance to the values for stresses of 150 and 200 psi; previously'3 these points had seemed to fall away from the curve of slope -3, but they now can be seen as almost completing a transition to n = 1. It should be noted that there was originally considerable apparent scatter in the results for the high-purity lead. However, this was found to be associated with annealing conditions or variations in residual content of oxygen and curve B2 was constructed using the highest rate (i .e., the rate for the purest lead) determined for each stress during the course of various investigations over the years: altogether, some forty tests were used in constructing curve R2.

Jan 1, 1968

By J. C. Scully, G. Sanderson

The electrolytic formation of y hydride in thin foils of Ti-5Al-Z.5Sn alloy and four Ti-A1 binary alloys has been studied by electron microscopy and electron diffraction. Two orientation relationships have been found: THE adsorption of hydrogen by titanium alloys from aqueous solutions has been widely reported. This can occur, for example, in both alkaline' and acidic2 pickling baths, in chemical polishing374 and electro-polishing Solutions,3,4 during cathodic polarization5" and under certain corrosion conditions,7,8 Etching reagents may also produce visible precipitates, sometimes after an aging period.9 During an investigation into the stress corrosion susceptibility of a titanium alloys by thin foil electron microscopy many observations have been made of hydride precipitation which has been deliberately promoted by techniques that are described below. The characteristics of hydride precipitation are reported and the observed crystallographic orientation relationships are discussed with reference to distortion in the titanium lattice created by the presence of the hydride. EXPERIMENTAL TECHNIQUE One grade of commercial titanium and five alloys were examined. Their composition and hardness values are shown in Table I. All the materials were in the form of thin sheet, 0.05 to 0.065 cm thick, which had been annealed after cold rolling. Foils suitable for transmission electron microscopy were made in two ways: a) by chemical polishing; and b) by elec-tropolishing. The solution used for chemical polishing was a mixture of 44 pct HF and 73 pct HNO3 in the volumetric ratio 2:3. During this polishing specimens were manually moved up and down in the solution. The reaction was veFy rapid, specimens thinning to foil thickness (2000A) in 5 min. Since chemical polishing solutions tend not to cause hydride contamination if they are kept cool3 no attempt was made to prevent the solution from becoming hot during the thinning operation. Electropolishing was performed at -30° C with a mixture3 composed of 295 cu cm methyl alcohol, 175 cu cm n-butyl alcohol, and 30 cu cm perchloric acid with an applied voltage of 40 v after chemically thinning specimens to approximately 0.017 cm thickness in a cool solution of 44 pct HF, 40 vol H2O2, and water in the volumetric ratio 1:6:3. EXPERIMENTAL RESULTS 1) Ti-5A1-2.5Sn Alloy. The Ti-5A1-2.5Sn alloy was severely contaminated by hydrogen during the chemical polishing. Foils obtained in this way contained large amounts of the y hydride phase. A typical micrograph is shown in Fig. l. The laths were identified as the hydride phase by selected-area diffraction of a large number of foils examined in a tilting stage that could be turned through 40 deg. Using the lattice spac-ings of the alloy obtained by X-ray diffraction as a standard the a value of the hydride was determined fromo the electron diffraction patterns to be 4.38 to 4.40A. The accuracy of this figure is only about ±2 pct. Further evidence of the identification was obtained by vacuum annealing chemically polished foils of the alloy at 500°C. This resulted in the dissociation of the hydride with the production of dislocations and possibly vacancy loops as shown in Fig. 2. The hydride

Jan 1, 1968

By Antoni Kidybinski

Mechanical properties of the nether roof rocks play a substantial role in maintenance of mine openings. On longwall faces too low strength is a cause of rock- falls which bring about delays in face advance, and hazard to the miners, while too high strength may result in over hanging of roof rocks in gobs, sometimes over a large area, initiating dynamic periodical pressures. Hence, efforts have been undertaken to develop methods of caveability evaluation for proper selection of roof control system and of longwall-face supports. DESCRIPTIVE CLASSIFICATIONS One of the basic qualitative criteria for classification of roof types with different caveability is the criterion of geological structure. The 26-stages scale worked out for conditions of Donetz Coalfield. USSR, by E.T. Proyavkin et al.(l) may be an example of that type of classification. This scale is presented graphically in Fig. 1. The first types of roofs according to this scale are easily caving immediately after exposure of a small surface, while the last be- long to the strongest roofs which over-hang in waste and require artificial fracturing. Observations and measurements carried out also in Donetz mines lead W.T. Davidyanc (2) to following classification: Class I - weak and densely bedded rocks very easily caving after each advance of supports. The roof subsidence over a waste edge is: [ ] where: h is roof subsidence, m M is thickness of the seam being mined, m S is width of an opening, m Class II - medium caveability and small main roof movements [ ] Class III - hardly breakable rocks, thicker beds with higher strength: caving by large blocks. After exposure of larger roof surfaces the movements of the main roof strata may affect an opening. [ ] Class IV - very hardly breakable rocks - thick, monolithic sandstones: they cause considerable difficulties in working with caving. [ ] Class V - thin-bedded rocks featured by the capacity of a smooth subsidence without disintegration into loose blocks with keeping contact between the particular blocks in a bed. In this classification mining with full roof caving is recommended for class I-III and only exceptionally for class IV. In many coal fields, qualitative and quantitative observations of roof rocks behavior have been carried out: as a result relations and characteristics found in certain mines are given. As an example, Kizielov Coalfield data in the USSR given by B.S. Vinokur (3) may be mentioned. In working thin pitching seams man- less system of undercutting the seam by means of rope saw is used there. The author separated four classes of roofs and gave, for every class, the empirical relation between the extracted area and the time of roof stability. The size of the stable area of exposed roof for particular classes is as follows: 120 to 150 m2 (1291 to 1614 ft2 ), 170 to 200 m2 (1829 to 2152 ft2), 250 to 280 m2 (2690 to 3013 ft2 ) and 350 to 400 m2 (3766 to 4300 ft ). For Kuznetz coalfield in the USSR the following classification of roofs was developed at the WIMI Institute in Leningrad: Class I - completely unstable: impact strength in- dex of rock f = up to 1.5 - the strip of roof 1.4 to 2.0 m (4.59 to 6.56 it) wide, exposed after shearer's cut holds stable during the time not longer than 5 min. Class II - low stability: f = 1.5 to 2.5 - the strip of roof 1.4 to 2.0 m (4.59 to 6.56 ft) wide and up to 4.0 m (13 it) long, exposed after shearer's cut holds stable during 5 to 20 min. Class III - medium stability: f = 2.51 to 3.5 - the strip of roof 1.4 to 2.0 m (4.59 to 6.56 it) wide and 4.0 to 25 m (13 to 82 ft) long, exposed after shearer's cut holds stable from 20 min. to 2 hours. Class IV - stable: f = 3.51 to 4.5 - the strip of roof 1.4 to 2.0 m (4.59 to 6.56 ft) wide and 25 m (82 it) long exposed after shearer's cut is stable during the time longer than 2 hours. Class V - entirely stable: f = over 4.5

Jan 1, 1982

By E. C. Wright

FOR many years the cost of making pig iron and steel in the Birmingham district has been about the lowest in the United States. The close proximity of the important raw materials such as coal, iron ore, limestone, and the large available scrap supply made it possible to assemble all of the important ingredients necessary for steel- making with only nominal costs for transportation. Both iron ore and coal were being mined in this district at one time for about $1 per ton from underground operations. At that time mining wages were low everywhere and the labor rate in the South was considerably below that of other mining districts. This situation persisted until the emergence of labor unions as an economic and political force, emigation of southern labor to northern plants, and continued growth of industry in the South in the years following the depression of 1932. Since that time union organizing of miners and steel workers throughout the nation has increased labor costs in mining and steelmaking. This influence has been felt particularly in underground mining operations since the proportion of labor cost to total cost is still high. The previous advantage in basic labor rate in the South has been almost entirely eliminated by the equalization of labor rates in basic industries by national unions. Since 1945 these influences have been especially noticeable in increasing the cost of mining coal and iron ore in the South. The northern steel districts are still being largely supplied with iron ore from the Minnesota open-pit mines and the cost of this ore has not increased nearly as much as the cost of ore obtained from underground mines in Birmingham. A similar situation exists in coal mining; the thinner seams in Alabama are not so amenable to mechanization as the northern mines and, as a result, the output per man-day in southern mines is only about three tons of coal as compared to five or six tons per man-day in the large captive mines of West Vir-

Jan 1, 1950

By G. S. Cole, G. F. Bolling

A series of Al-Cu alloys has been cast from constant superheat to solidify either with a hot top or with a free liquid/air interface. All the other variables which affect relative fluid motions were kept constant except the imposition of slow rotations m oscillations of the mold. Depending upon the exact circumstances, evi-dence could be found for the nucleation of equiaxed grains from dendrite bits detached from the columnar zone, or from the free surface, or from solid bits induced at pouring; a special experiment showed the solute buildup during ingot solidification and by inference the reality of constitutional supercooling. All methods of grain nucleation are thus supported. When solidification is arranged in a static mold the columnar- to-equiaxed transition occurs by one of these nucleation paths, subject to the natural decay of temperature gradients which promotes nucleation or the survival of nuclei. But when solidification is forced under special fluid flows it may display odd transitions. In one of these the equiaxed zone is synihesized from. dendrite bits detached from the solid/liquid interface. These survive under the concurrent lowering of temperature gradients and eventually cause transition by impeding columnar growth. GRAIN structure control is desired because real experience shows that specific properties are associated with specific cast grain structures. But since control is thus a matter of need and does not require understanding, much of casting remains art. For example, the manipulation of relative fluid motions in simple ingots provides a control of grain structure development,1-4 but theoretical insight is not required in order to broaden its potential applicability. Fortunately, however, once a variable like fluid flow is recognized it can also provide paths toward new insight. New observations using this variable have provided an insight which is curious. All the theories or descriptions of the columnar-to-equiaxed transition (CET) seem to be supported as long as the appropriate experiment is carried out. In order to develop this idea, let us ignore segregation, the chill zone (until a later section of this paper), and the origin and development of the columnar zone.= The details of questioning how equiaxed material forms in a casting may be divided in two. The first part is simply a questioning of the origin for equi-axed grains: are there fresh nuclei from heterogeneous nucleation, or dendrite bits broken off or melted from the columnar zone or elsewhere,6 or other solid bits surviving from the outset of freezing? The second is more complex, a questioning of the CET: does columnar growth slow down and allow the dominance of equiaxed grains, however they are formed, or do equi-axed grains take over and force columnar growth to stop? This "chicken-or-the-egg" second question can be quite useful, especially because it allows us to delineate the ordinary or NATURAL EQUIAXED ZONE and the induced or SYNTHETIC EQUIAXED ZONE. A natural equiaxed zone occurs if solidification proceeds in the absence of disturbances or extraneous forces such as ultrasonics, stirrings, and so forth. Both the CET theory proposed by Winegard and Chal-mers, 7requiring constitutional supercooling, and the CET theory suggested by Chalmers,8 requiring nuclea-tion only on pouring, pertain to the natural equiaxed zone. The best collating evidence to date for both these theories is contained in the work by Biloni and Chal-mers;9,10 the former theory7 is supported by special versions of the unidirectional growth experiment performed by Plaskett and winegardl1 and the latter theory' is supported in detail for the solidification of small ingots. An equiaxed zone can be synthesized when disturbances caused by forced fluid flows2-4 or other artificial means9,10 break off or melt off dendrite tips from an advancing solid/liquid interface,* if (and only if) temperature gradients in the melt are (forcibly) reduced enough to ensure the survival of these bits of dendrites, 2'3 Willful disturbances even so trivial as those produced by rapping a mold might be included here; as will be shown for example, this follows Southin's suggestion that the free liquid/air interface

Jan 1, 1970

By George B. Clark

THE " Munroe effect" of shaped explosive charges was discovered by Charles E. Munroe more than 50 years ago (in 1888), but it was not until World War II that it was put to any practical use. Both Allied and Axis armies used the principle effectively in bazooka projectiles, artillery projectiles and in fixed shaped charges. The first two were used primarily to combat tanks and other armor and the latter to assault pillboxes and other fixed concrete emplacements. Many engineers and others who saw the military application of the jet effect of shaped charges believed application would be possible in many industrial fields where blasting is used, such as mining and quarrying. Some work has already been done on the problem of making shaped charges applicable to mining operations. Experimentation conducted at the National Tunnel and Mines' was directed primarily along the lines of use of shaped charges for secondary blasting in connection with long-hole stoping methods. Experimentation initiated earlier and conducted by the author2 at the University of Utah on a more scientific basis was directed at revealing the basic principles of the functioning of shaped charges, as well as the feasibility of using them in mining operations after the principles of their operation had been clearly defined. It was found that: (I) a given design of charge required an optimum standoff to make it most effective, either in penetrating solid rock or steel (see Fig 3); (2) for a given design of charge there exists an optimum ratio of charge diameter to charge height above which the increase in effectiveness of the charge decreases rapidly (see Figs 4 and 5); (3) partial confinement has the effect of making charge performance much more consistent; (4) the higher power and higher velocity explosives tested are much more effective than relatively low-strength explosives; and (5) shaped charges may be adaptable to mining operations in secondary breakage and drilling blasting holes. Fig I is a diagrammatic sketch of the action of a shaped charge for two types of cavities, while Fig 2 shows the approximate manner in which the jets are believed to be formed when an explosive wave encounters the cavity, approaching it from the top. The experimentation already completed has laid the foundation for further research in the study of shaped charges. Standoff effects, effect of the strength of explosives and other features have been fairly well established. Such elements in the design of shaped charges as (I) geometry of cavity liners, (2) relative horizontal dimensions of explosive column, (3) effect of

Jan 1, 1947

By R. T. Torgerson, W. A. Baginski

Alloying studies were conducted on modifications of Cb-10W-1OHf (C-129) alloy as part of a cooperative alloy development effort between The Boeing Co. and the Wah Chang Corp. The purpose of these studies was to develop a fabricable, high- strength columbium alloy with superior melding properties. The initial investigations of alloy modifications were conducted using laboratory-melted button ingots. An alloy with the composition exhibited excellent veld ductility, elevated-ternpevature strength, and fabricability. The yttrium-modified alloy was desigrzated C-129Y and was selected for further developrnetzt. Three prodziction heats of C-129Y weve processed into sheet. Welding, forming, and tensile proper-lies are presented for this material. Tungsten-are welds of. C-129Y exhibited a minimum bend radius of 1/2T in the as-welded conditioti compared to 8T for C-129 alloy. An investigation was conducted on Disil coating of C-129Y and test data are presented on oxidation resistance and effects of re-entry environments. The excellent combination of strength, fabricability, weldability, and low transition temperature of C-129Y alloy offers potential design advantages for use of this alloy in structures of aerospace vehicles. It is atiticipated that C-129Y will rapidly supplant C-129 and similar alloys clue to the versatility provided by its impvoved welditzg properties. COLUMBIUM alloys with improved properties are required to achieve better performance and efficiency of radiation-cooled structures for re-entry vehicles and recoverable launch systems. Cooperative work by The Boeing Co. and the Wah Chang Corp. resulted in the development several years ago of C-129 alloy, a high-strength Cb-W-Hf alloy.' This alloy has received favorable consideration for many advanced designs of high-temperature structures due to its excellent combination of properties. However, this alloy exhibits low weld ductility unless given a high-temperature stress-relief treatment. For this reason the development effort was directed towards modification of the C-129 alloy composition (Cb-1OW-1OHf) to eliminate the need for post-weld heat treatment. This paper describes the successful development of an alloy meeting this requirement. This new alloy was designated C-129Y and has a nominal composition of EXPERIMENTAL PROCEDURES Alloying Studies. Various Cb-W-Hf compositions were investigated using are-melted button ingots. These buttons were hot-forged using a protective glass coating and rolled into sheet specimens. Metal-lographic examination: tensile testing, and welding studies were conducted to evaluate properties of the experimental alloys. Sheet Fabrication. Three pilot production heats of C-129Y were fabricated into sheet to establish processing techniques and to determine sheet properties. The three heats consisted of 6-in.-diam ingots weighing from 140 to 250 lb. The ingots were prepared by electron-be am melting followed by vacuum-arc melting. The hafnium and yttrium additions were made during the are-melting operation by tack welding alloy strip additions to the Cb-W master-alloy electrode. The final ingots were machined into billets with samples of the machine turnings collected for chemical analysis. Ingot breakdown methods for the three heats are as follows: Ingot Ingot Heat Size Weiaht___Breakdown Method A 6 in. 250 lb Hot extrusion B 6 in. 140 lb Hot forging C 6 in. 250 lb Hot extrusion Sheet bars from the extrusion and forging operations were conditioned, recrystallized, and rolled into sheet. Initial rolling was conducted at elevated temperatures and final rolling was conducted at room temperature.

Jan 1, 1967

NAME "Utah" is derived from the name of the Indian tribe, variously spelled "Yuta," "Ute," "Youta." "Uta." "Eutaw," and finally "Utah." It means "in the tops of the mountains," or "on the heights." The state emblem is the "Beehive," typical of industry. State nickname is "Beehive State." AREA Utah is the tenth state in area. It contains 84.990 square miles, of which 2.806 miles are water. When admitted as a territory in 1850. Utah included all the area from the east line of California to the summit of the Rocky Mountains between the 37th and 42nd parallels, which embraced practically all of the present state of Nevada, all of western Colorado, part of Idaho and part of Wyoming, in addition to the present area of Utah.

Jan 1, 1925

By Frank Firmstone

When in charge of the Glen don Iron Works, the importance of good methods of filling was forcibly brought to my attention, and it occurred to me that the first step toward the discovery of the best plans was the accumulation of exact information as to the distribution actually effected by various chargers; also, that something could be done in this direction by careful measurements in the furnace, to be afterwards recorded for comparison on drawings. The figures shown herewith have been prepared from such measurements, and nearly all of them refer to Glendon, where the fuel used was anthracite and the ore-mixture about threequarters of New Jersey magnetic ores and one-quarter of brown-ores from the vicinity of the works. Every occasion furnished by the blowing-in of furnaces was utilized, but the data accumulated rather slowly in this way; and, opportunity oEering to make special trials at no great expense, the figures from 2 to 7 were thus obtained. The trials were made by dumping a hopperful (6 barrows) of material at once into the furnace, and, when all the fuel of a round (12 barrows) had been put in, the profile of the upper surface was obtained by stretching a steel tape across the top on a diameter and at a known height. By measuring down from this with a graduated rod, the profile was easily obtained with sufficient accuracy, especial pains being taken to get correctly the crests of the ridges and the bottoms of the hollows, as many intermediate points being included as seemed necessary. This was repeated on another diameter when there was occasion. The same course was followed for the limestone and ore.

Jan 1, 1899

By J. C. Trantham, J. W. Marx

From 1955 to 1958 the Phillips Petroleum Co. conducted a series of small scale counterflow combustion field tests in a tar sand about 60-ft deep and 6 to 12-ft thick near Bellamy, Mo. A total of seven different well patterns, including conventional five-spots and seven-spots plus a 15-well line drive pattern and a 10-well radial pattern, were employed. Electric heaters, gas burners and wellbore fuel packs were tested as counterflow ignition devices. Direct drive ignition followed by reversal to counterflow operation was not possible in this case because of formation plugging by the semi-solid tar during the direct drive phase. Sustained counterflow burning was achieved with wellbore fuel packs under conirolled conditions which permitted close correlation of fire front velocity, air velocity, air-oil ratios and yields for a wide variation in process parameters. The tar in place was about 10' APl with a reservoir viscosity greater than 500,000 cp. The composite oil prodlcct was 26' API with a viscosity in the range of 5 to 15 cp. Vertical sweep efficiency in the line drive pattern was essentially 100 per cent, with no evidence of gravity segregation. Original tar in place ranged between 800 and 1,000 bbllacre-ft, specific rock permeability averaged about 800 md before burning and effective air pemeability with residual oil and water in place (at the time of ignition) averaged 250 md. The volume of oil recovered during stabilized line drive counterflow burning represented about 67 per cent of the volume of tar originally in place in the burned out zones. A minimum was found when air-oil ratios were plotted against formation air velocities, and a limiting air velocity existed below which the counterflow fire front reversed and burned back over its own trajectory, using the residual coke left behind. INTRODUCTION Early in 1955 the Phillips Petroleum Co. began preparations to field test counterflow underground combustion as a method for producing oil from high viscosity hydrocarbon deposits, such as bituminous (tar) sands. At that time all published information on the use of underground burning io increase oil recovery was based on the direct drive process, in which the fire front moved through the formatim in the same direction as the injected air stream. Such direct burning is not applicable to many heavy oil sands or to most tar sands because the heat-thinned native hydrocarbon congeals in the cold rock ahead of the fire front forming a gas permeability block which smothers the combustion. Laboratory experiments first demonstrated that this permeability blocking would not occur if the fire front and the air stream moved in opposite directions. In this counterflow (reverse) burning process the heated heavy oil or tar is driven back through the combustion zone. Here thermal cracking produces relatively light hydrocarbon products which pass through the heated rock behind the fire front (largely in the vapor phase) and are incapable of causing a gas permeability block.' Exploratory wring revealed that consolidated bituminous sandstones with a combination of tar saturation, permeability, thickness and depth favorable for experimental field testing existed at several points in western Missouri. By the fall of 1955, a test site for determining technical feasibility of the counterflow process under field conditions had been selected near Bellamy, Vernon County, Mo. — about fifty miles north of Joplin. The pay zone was too thin and too shallow to constitute a commercial prospect, but these same factors permitted controlled experimental operation with a small compressor installation and with a large number of monitor wells. These were needed for adequate observation of the underground phenomena at reasonable cost. RESERVOIR CHARACTERISTICS The reservoir chosen for the experimental field tests was a 124 thick section of tar sand extending from 49 to 61 ft subsurface, with shale and siltstone laminations sealing the top and bottom. This section was part of a larger 30-ft thick tar sand deposit which extended both above and below the test zone. Core analyses showed that the 12-ft sand was effectively subdivided into two approximately equal sections, with the lower zone being consistently more permeable than the upper zone. The most significant laboratory core data are summarized in Table 1. The line drive test which is the principal subject of this report utilized only the lower 6-ft thick zone, with the upper zone regarded as part of the 55-ft overburden. WELL PATTERNS The over-all field test program involved a total of seven different well patterns, including conventional five-spots and seven-spots plus a 15-well line drive pattern and a 10-well radial pattern. In some of the tests the full 12-ft zone was employed; in other cases only the more perme-

Jan 1, 1967