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By J. M. Seeman, R. A. Dodd

If some solution-quenched supersaturated aluminum solid solutions, e.g.., Al(Cu), are plastically deformed at room temperature and then aged at 200oC, the hardness may first decrease before increasing in the usual manner ' A similar transient softening is observed, even in the absence of plastic deformation, when a thin foil or small-diameter wire Al(Cu) specimen is rapidly quenched from the solution-treatment temperature and then aged.' Also, some solution-quenched and neutron-irradiated alloys undergo an initial hardness decrease during subsequent aging.3 The usual explanation of the transient softening in the above examples involves the effect of vacancies introduced by plastic deformation, rapid quenching, and neutron irradiation, respectively However, the mechanism of softening is not unequivo-cably established. In the first-noted case, Gane and parkins' observed that the softening was approximately proportional to the difference in the atomic diameters of solute and solvent, and it was argued that, if a large relative difference in atomic diameter promoted vacancy-assisted room-temperature solute diffusion to dislocations, dislocation climb occurring during subsequent aging at 200°C could cause transient softening. Some substantiation of this theory was offered by the additional observation that discontinuous yielding of solution-quenched polycrystalline Al(Cu) tensile specimens was suppressed by interrupting the test for a short aging treatment at 200°C. Of course, reversion (resolution of GPI) at 200°C would also cause softening, and Sil-cock4 expressed the opinion that the softening in the case of Al(Cu) must be partly associated with reversion. Also, Hirsch5 commented on the possibility that, if jogs cause hardening, an anneal at 200°C should straighten the jogged dislocations and result in softening. However, parkins' has re-expressed his belief that these mechanisms are not entirely consistent with the observed serrated stress-strain curves. Experimental. New data has now been obtained for Al(Cu), Al(Ge), Al(Mg), and Al(Zn) primary solid-solution alloys. These alloys were prepared from metals exceeding 99.99 pct purity, and specimens 4 by 4 in. diam were solution-treated, quenched, deformed 10 pct in compression at room temperature, and aged at 200'C in a cottonseed oil bath for various times. Generally, these operations could not be carried out in rapid succession, so the alloys were stored in a freezer between treatments to minimize aging. The variation in hardness with aging time, Figs. 1 to 4, was determined by microhardness testing with a 3.0-kg load. At least five readings were averaged to obtain each data point, and all measurements were taken about the center circumference of each cylinder (region of maximum deformation). Each specimen was used to obtain only one data point, i.e., specimens were not replaced in the aging bath after measurements were made. In

Jan 1, 1968

R.. B. T. KILIANI, New York, N. Y.-I do not care to discuss Mr. Taggart's paper in the light of theory, as that has been very well done by Prof. Bell,1 but I should like to criticize some of his conclusions, in the light of actual operating practice at plants all over the country. 1. Mr. Taggart's first conclusion is that in crushing average ore, the character of the gangue has little effect on the efficiency of the mill. This, I believe, is not in accordance with the usual practice, since in crushing a hard ore the capacity will be much reduced below what it would be with a softer ore, while the power consumed will be practically independent of the character of the ore, being proportional only, to the load of ore and balls in the mill. 2. Mr. Taggart's second criticism is that the Hardinge mill is not suitable for grinding soft, tough material. In answer to this I might mention that the mill is being used for grinding tough, dnctile material, such as metallic aluminum, and also for grinding licorice root. 3. He says that the ball-mill works more efficiently on material of intermediate size, say 1/2 to 3/4 in., than on either coarser or finer feed.. This is true as to coarser feed. For the most efficient work, I believe that a ball-mill should be fed with material not coarser than 1 1/2 in.; it' will handle material up to 3 and even 4 in., but the reduction from 3 or 4 in. to 1 1/2 in. can be done much more cheaply and efficiently by rolls or disk crushers than in the ball-mill. 4. As to Mr. Taggart's fourth conclusion, that a greater ratio of reduction can be expected with feed of an intermediate size than with a coarse feed, I have not enough information to express an opinion. 5. His next conclusion is that steel balls are much more efficient crushing media than pebbles. Steel balls are undoubtedly more efficient for crushing coarse feed. On fine material they are also more efficient as to tons per horsepower crushed to 10-mesh, but on fine material I think it will be found that flint pebbles are cheaper than cast-iron balls, per ton of ore, although there will be a saving in power per ton by using cast iron instead of flint. The increased cost of crushing with cast iron will be due to the higher cost of iron at the present time. 6. Mr. Taggart's sixth conclusion coincides with present practice, that steel balls will grind as fine or finer than pebbles when working on the same feed.

Jan 1, 1918

By W. M. Rust, W. D. Mounce

The almost universal acceptance of electrical logging by the petroleum industry calls for a critical examination of the physical bases of the common methods. This is particularly needed for the natural potential log. The conventional explanations are based on the phenomena of electrofiltration and electro-osmosis. However, experimental data show that they fail to give a basis that is quantitatively sound. Moreover, they lead to interpretations of details of electrical logs that are erroneous. It was found experimenlally that when a plug of shale is placed in a chain of dissimilar electrolytes, a current flow results. Experiments on oil-well cores and fluids similar to those existing in wells show that potentials are set up which are in reasonable agreement with those encountered in electrical logging. Doubtless other sources of potential are important but this explanation not only accounts for much of the potential but also clarifies some fairly familiar examples of misinterpretations. Introduction The extraordinary value of electrical well logging as a tool for petroleum engineers and geologists is now so universally recognized that it is not necessary to point out the ways in which it can be useful. It is sufficient to mention only the saving that has resulted from the elimination of much coring that would have been necessary under former practices and the considerable additions to reserves resulting from the discovery of sands missed in coring. Paradoxically, the very remarkable success that has resulted from the application of electrical logging to correlation and completion problems has cost some of its most consistent users dearly.' The fact that some operators have placed unwarranted faith in erroneous interpretatioils of electrical logs does not reduce the value of the logs. It should, however, make one pause long enough to ask why experienced users should be misled. The answer to this question is not hard to find. In nearly every case, the fault has been oversimplification. This is illustrated very aptly by the familiar interpretation and theory of the natural potential log. Recording and Interpreting Logs The natural potential log is extremely simple to record precisely enough for all present-day uses. One needs only an electrode, a cable to lower the electrode through the drilling mud, a recording meter and a surface ground connection. The electrode can be elaborate but need not be. A simple brass rod 2 in. in diameter and 2 or 3 in. long is quite satisfactory. The cable need not be particularly well insulated, nor need it have a very low resist-

Jan 1, 1946

By W. M. Rust, W. D. Mounce

The almost universal acceptance of electrical logging by the petroleum industry calls for a critical examination of the physical bases of the common methods. This is particularly needed for the natural potential log. The conventional explanations are based on the phenomena of electrofiltration and electro-osmosis. However, experimental data show that they fail to give a basis that is quantitatively sound. Moreover, they lead to interpretations of details of electrical logs that are erroneous. It was found experimenlally that when a plug of shale is placed in a chain of dissimilar electrolytes, a current flow results. Experiments on oil-well cores and fluids similar to those existing in wells show that potentials are set up which are in reasonable agreement with those encountered in electrical logging. Doubtless other sources of potential are important but this explanation not only accounts for much of the potential but also clarifies some fairly familiar examples of misinterpretations. Introduction The extraordinary value of electrical well logging as a tool for petroleum engineers and geologists is now so universally recognized that it is not necessary to point out the ways in which it can be useful. It is sufficient to mention only the saving that has resulted from the elimination of much coring that would have been necessary under former practices and the considerable additions to reserves resulting from the discovery of sands missed in coring. Paradoxically, the very remarkable success that has resulted from the application of electrical logging to correlation and completion problems has cost some of its most consistent users dearly.' The fact that some operators have placed unwarranted faith in erroneous interpretatioils of electrical logs does not reduce the value of the logs. It should, however, make one pause long enough to ask why experienced users should be misled. The answer to this question is not hard to find. In nearly every case, the fault has been oversimplification. This is illustrated very aptly by the familiar interpretation and theory of the natural potential log. Recording and Interpreting Logs The natural potential log is extremely simple to record precisely enough for all present-day uses. One needs only an electrode, a cable to lower the electrode through the drilling mud, a recording meter and a surface ground connection. The electrode can be elaborate but need not be. A simple brass rod 2 in. in diameter and 2 or 3 in. long is quite satisfactory. The cable need not be particularly well insulated, nor need it have a very low resist-

Jan 1, 1946

By H. Carroll Weed

By caving a block into the workings of its open pit and using the block as an ore transfer, the lnspiration Consolidated Copper Co. has solved a transportation and sizing problem, making possible a great expansion of open-pit methods as applied to Inspiration ore-bodies. FROM 1915 to 1948 the entire production of Inspiration Consolidated Copper Co. was supplied from underground mining. The sole method used was block caving. Underground haulage and hoisting facilities were designed and geared to large-scale production. Beginning in 1948 open-pit mining was substituted for block caving in a portion of Inspiration's Live Oak orebody. The idea proved so attractive that before the Live Oak pit had come into production, another pit on the Colorado orebody (now the Thornton pit) had been laid out and stripping started. It should be noted that these orebodies were not new and the entire program was one of changed methods of mining. Since part of these orebodies had already been mined by block caving methods, it must be recognized that haulage levels had been established under the area or adjacent to it. The orebodies lay on the south side of Inspiration Ridge, while the shafts, crushers, and treatment plants are all on the north side of this ridge. The existing main crushing plant was designed to take a maximum of 12-in. material, and all ore was sized to this dimension by passing through grizzlies in the stopes. In the original planning for open-pit work much thought was given to the transfer of ore to existing underground levels for haulage and hoisting from the regular shafts. If this could be done the necessity of crossing the ridge could be eliminated. Final decision called for cutting a road through the ridge, installation of a primary crusher at an elevation of 3968 ft on the north side of the ridge, and railroad haulage with existing facilities to the main coarse crusher. This decision was brought about by the difficulties of properly sizing and transferring to underground haulage large tonnages of coarse breaking oxide which lay on the upper benches of the proposed pit. Trucking over the ridge on a 7 pct grade would give a cheaper and more flexible operation in handling this material. The benches in the pits are laid out at 50-ft inter-vals and designated according to elevation above sea level. The lowest bench in the original pit was laid out at 3500-ft elvation. All haulage roads are on a 7 pct grade and a vertical lift of 468 ft was considered about the maximum economically possible from a cost standpoint. After 2 years' operation the advantages of pit mining, both from the standpoint of costs and flexibility of operation, became more and more apparent. Studies were then started to develop the extension of open-pit mining to more of the ore reserve than had been planned originally. The idea of transfer raises was again explored. Obviously, as elevation of the pit was lowered, shorter transfer raises would be needed to reach the main haulage levels. Transportation from the lower elevations to the rim of the pit by belt conveyor or skip hoisting was also considered. However, it was recognized that sizing would be required both for conveying and regulation of feed size to the main crusher plant. This would necessitate a fixed or portable crusher located somewhere in the pit. It was known that the sulphide ore at the lower elevations was softer and easier to break than the overlying oxide of the upper benches. This sulphide ore is very suitable for grizzly sizing. Calculations indicated that for ore mined on 3650 elevation, costs for trucking to the primary crusher, crushing, and delivery by railroad to the coarse crusher would equal costs for dropping the same material to the 600 level, hauling underground, and hoisting directly to the coarse crusher bins, provided a suitable method of sizing and transferring could be developed. Above 3650 elevation costs would favor surface haulage; below they would favor underground haulage. The idea of caving an underground block from the 600 haulage level directly into the bottom of the pit

Jan 1, 1954

By W. A. Jr. Slippy, E. P. Dahlberg, R. B. Reed-Hill

A large room-temperature mechanical-hysteresis effect under cyclic tensile loading was observed in zivconium specimens prestrained at 77°K so as to form large numbers of (1121) twins. The observed hysteresis loss was a maximum when the prestrain was just under 1 pct It also depended strongly on the maximum applied cyclic stress, but only moderately on loading rate and temperature (-72°to 100°C). The phenotnena have also been studied in terms of the elastic aftereffect where the data has been found to cotlforln closely to the functional relationship between strain and time, c = -RT/a In tank ft. t to)/Zr, for strains cts large as about 10-4 This eyication can be derived directly from the strain-rate equation There may be important commercial implications to the present discovery since it appears that within rather wide limits it may be possible to obtain any desired magnitude of damping capacity in zirconium melal. The damping propevties may also be given directional characteristics. It is shown that the observed anelaslic phenomena can he explained on the assumiplion that they are the resuilt of stress-induced twin-boundary movements in which the average twin increases its thickness by only a few percent. In polycrystalline zirconium, a grain is least favorably oriented for slip when its basal plane, containing the (1130) slip directions, lies nearly perpendicular to the principle normal stress. Mechanical twinning is favored by this orientation and under simple tensile deformation at room temperature the primary twinning mode is {1012).' The (1012) twinning shear is small (0.17) so that, during deformation, twin growth does not greatly enhance ductility. At 77°K, on the other hand, the predominant twinning mode shifts to {1121), and many thin (1121) twins can form at small strains.' Twins thus nucleated can grow readily during subsequent room-temperature deformation and permit easy deformation in unfavorably oriented grains. In addition, the (1131) twinning shear is large: 0.63. These facts have lead to a process' for greatly improving the room-temperature ductility of zirconium when it contains a sizable fraction of grains with basal planes nearly normal to the stress axis. Optimum results are obtained by slightly less than 1 pct prestrain. The effect is large so that a 50 pct elongation increase is readily attained. A large room-temperature mechanical-hysteresis effect has also been observed in zirconium specimens prestrained at 77°K. The present paper is concerned with some aspects of this effect which the experimental observations strongly indicate is primarily the result of strain-induced {1121) twin-boundary movements. This clearly suggests that mechanical twins formed by plastic deformation can cause anelastic effects similar to those previously observed2 for annealing twins and twins associated with phase transformations. EXPERIMENTAL PROCEDURES All specimens were prepared from arc-melted sponge zirconium plate stock as previously described,3 which has a preferred orientation with basal planes aligned parallel to and uniformly distributed around the rolling direction. Both lonti-tudinal and transverse tensile specimens, with axes parallel and transverse, respectively, to the plate rolling direction, were cut from this plate. In a longitudinal specimen most grains are favorably oriented for slip, while in a transverse specimen a large fraction are unfavorably oriented. Prestraining at 77°K was performed in two ways. In one, annealed zirconium plate stock was cooled in liquid nitrogen and then rolled in the transverse plate direction to strains varying from 0.5 to 12 pct. Cylindrical 1/8-in.-diameter by 1-1/4-in.-gage-length tensile specimens were machined from this material. In the other, previously machined tensile specimens were prestrained in tension at 77°K between 0.2 and 3.8 pct. Type FA 25-12 SR4 strain gages were cemented to all specimen gage sections. All tests were performed on an Instron testing machine of 10,000-lb capacity. EXPERIMENTAL RESULTS Fig. 1 shows a typical two-cycle room-temperature stress-strain diagram for a specimen prestrained 0.65 pct by rolling at 77°K. The maximum stress in both cycles was 14,500 psi. Note that in the first cycle the loading and unloading curves are unsymmetrical so that a residual strain, er, remains after unloading. In the second cycle the stress-strain curves form a nearly symmetrical hystere-

Jan 1, 1965

By Albert C. Harding

For several years barite (BaSO4) production has topped such better known minerals as feldspar and fluorspar in annual tonnage and is now well established among our foremost nonmetallic industries. Description Of widespread occurrence in the United States, barite is relatively soft (hardness, 2.5 to 3.5), inert, heavy (specific gravity, 4.5), and is also called "barytes" (sometimes pronounced "bear-i-tease"), "heavy spar" and "tiff." It occurs with granular and earthy structure but usually has coarse, orthorhombic crystallization. It is commonly white, and although occurring in various shades of red, yellow, and gray, barite always gives a white streak and can be scratched by the ordinary prospecting pick. Its softness and weight are usually sufficient to establish field identification. Another aid is the "rotten egg" odor of H2S which sometimes emanates from freshly fractured surfaces. Barite can be readily distinguished from two other minerals which have somewhat similar occurrences and characteristics. The relatively rare celestite, strontium sulphate, has a specific gravity slightly less than 4.00 and gives a distinctively crimson color in a flame test. The very rare wither-ite, barium carbonate, has a specific gravity of 4.3 but readily effervesces with dilute hydrochloric acid. Occurrence All known commercially important barite deposits in the United States occur as re- placements or cavity fillings in schists, shales, limestones, dolomites, and quart-zites, or as residual deposits from such formations. In Georgia, Tennessee, North Carolina, and Missouri, high-grade barite is recovered as fragments from residual clays covering quartzites and dolomites. In Arkansas, Nevada, and California, barite, usually of lower grade, is mined from replacement deposits. These seven states contributed the entire production reported by the U. S. Bureau of Mines for 1945 as given in Table I. Table i—Crude Berile Sold or used by Producers in 1945 State Short Tons Arkansas................... 260.660 Missouri.................... 225.467 Georgia..................... 110,393 Tennessee................... 32,812 Nevada..................... 28,919 Other States............... 37,811 Total..................... 696,062 ('other states,, included only california and North Carolina. Such a large part of this undistributed production came from California that that state probably outranked both Nevada and Tennessee. For the second consecutive year Arkansas was the leading producing state, particularly worthy of note because there was no commercial production of barite there prior to 1941 and because the entire production of Arkansas has come from one large deposit which is being operated by two companies. Consumption The 1945 consumption as reported by the Bureau of Mines was: 482,442 tons as ground barite 139,288 tons in making lithopone 99,173 tons in producing barium ------- chemicals Total 720,903 tons

Jan 1, 1949

By R. J. Stoehr, F. Howell

Since the ore mined in the Grants-Ambrosia Lake uranium area is taken from a water-saturated sandstone formation, part of the milling operation includes a drying process. The authors discuss the merits and disadvantages of two methods: natural drying and mechanical gas-fired drying; operating and cost data are included. In the Grants-Ambrosia Lake uranium area, approximately 7000 tons of ore are mined each day from a water-saturated sandstone formation. This ore is sampled and treated at four separate milling operations. Prior to the design and initial construction of the milling facilities, very little was known as to the physical properties of the ore which would be handled in the sampling and crushing facilities at the mills. All four mills initially were constructed without ore-drying facilities. As mining proceeded, it was determined that, even after drying the ore in place by use of drainage development underground, the moisture content of the ore as hoisted averaged approximately 18 pct. Actually, some of the mines produced ores which resembled cream-of-wheat in character and had a moisture content in excess of 24 pct. Each of the four mills had a differently designed sampling and crushing plant; therefore, the moisture content of the ores had varying adverse effects upon the operation of the plants. It was determined after some experience that most of the sampling and crushing facilities could not economically handle ore if the moisture content was greater than 10 pct—one operator considers 12 pct moisture as a maximum. In all of the plants, moisture content over 8 pct increases costs to some extent. In addition to the added sampling and crushing costs, some of the mills were located 15 to 20 miles from the mines and the high moisture content of the ores increased the haulage costs from the mine to the mill. It was generally conceded that it was necessary to employ some method to reduce the moisture content of the ore prior to sampling. Several methods were considered: 1) natural drying at the mine site, 2) mechanical gas-fired drying at the mill site, 3) mechanical gas-fired drying at the mine site, 4) radiant heating on concrete clabs, and 5) infrared shed drying. After preliminary investigation all possibilities but natural drying and mechanical gas-fired drying at the mill site were eliminated. NATURAL DRYING A cooperative feasibility study was conducted on the natural drying method. This study indicated the following: 1) During summer months May through September, ore piled 4-ft deep would dry from +20 to +10 pct HzO in 30 days with no mechanical turning. 2) In the first four days after piling a reduction of 5 to 7 pct total moisture content was accomplished by drainage and absorption of the water into the underlying ground on which the ore was piled, providing the ground surface absorption rates approached 10 gal per sq ft per day. After the first four days the decrease of moisture caused by drainage was practically nonexistent. 3) In the summer months drying by evaporation amounted to approximately 0.1 pct decrease in total moisture per day and mechanical turning of the ore could increase this as much as three times. 4) Climatic conditions of the area indicated that: a) Air temperatures averaged 64' May, June, July, August, September; 41" October, November, December, March, April; 31" January, February. b) Average rain fall was 10 in. with 40 pct of this occurring during July and August. c) Only on very rare occasions were there over five consecutive days with below freezing temperatures. d) The annual evaporation rate was approximately 75 in. On the basis of this study, two of the mill operations decided to attempt to meet the moisture specifications by natural drying supplemented by mechanical turning. The other two operations decided upon rotary gas-fired drying at the mill site. The natural drying method has been quite successful in practice. The ore running approximately 18 pct moisture is trucked from the head frame and piled in rows about 6 ft wide, 3 1/2 ft deep, 5 feet apart during the months of May through September. In approximately 30 days this ore will

Jan 1, 1961

By Rollyn P. Jacobson

The town of Pacific, in Jefferson County, Mo., is 27 miles west of St. Louis. Since the area lies entirely on the flood plain of a cut-off meander of the Meramac River, it was considered a likely environment for accumulation of commercial quantities of sand and gravel. Excellent transportation facilities are afforded by two major railways to St. Louis, and ample water supply for washing and separation is assured by the proximity of the river. As a large washing and separation plant was planned, the property was evaluated in detail to justify the high initial expenditure. An intensive testing program using both geophysical and drilling methods was designed and carried out. The prospect was surveyed topographically and a 200-ft grid staked on which electrical resistivity depth profiles were observed at 130 points. The Wenner 4-electrode configuration and earth resistivity apparatus* were used. In all but a few cases, the electrode spacing, A, was increased in increments of 1% ft to a spread of 30 ft and in increments of 3 ft thereafter.

Jan 2, 1955

By Frank C. Roe

Two cogent and ubiquitous facts have supported a constant search by the ore processing industry for better wear resistant materials in equipment. First, designs and types of equipment change or improve continually to meet the economic and physical demands of different and/or leaner ores. Second, practically all equipment is subject to impact and abrasion conditions that probably are more severe and extensive than in any other industry. Development and testing by operators and equipment manufacturers alike, has produced an extensive variety of longer lasting metal alloys, rubber, plastics, ceramics and other materials to improve equipment performance under different specific conditions. The need for diversified materials becomes obvious upon the simple realization that no two beneficiation operations are identical. Among the variables in the handling of ore are: 1. Size and weight of ore particles. 2. Hardness and sharpness of ore particles. 3. Concentration of ore particles in liquid or gas suspension. 4. Velocity and direction of particle impact toward equipment surface. 5. Coefficient of friction between ore particles and equipment surface. 6. Temperature of operation. 7. Chemical environment of operation.

Jan 12, 1967

By J. P. Manning

UNQUESTIONABLY, the outstanding feature of the piping for the sulphide plant is the large amount which had to be done in almost every size from instrument tubing to 84 in. OD pipe. In this article the piping will be considered under the main headings of Yard Piping and Building Piping. Tables I and II give data concerning the individual systems. In the case of yard piping the figures pertaining to lengths and weights are approximately correct; those for building piping are given only for the concentrator and smelter buildings where the amounts were the most extensive. Yard Piping The installation specifications for yard piping called for all sizes larger than 12 in. to have welded joints made with outside welding rings. This type of joint was selected over butt-welded joints because less time would be needed for lining up, damaged pipe ends would be less of a problem, and deflections up to 6° per joint could be made when necessary. These advantages were thought to outweigh the disadvantage of the greater amount of welding required. Yard piping in sizes of 3 to 12 in. was specified to have butt-welded joints made with inside welding rings which were intended to facilitate lining up and keep welding material out of the lines. Screwed connections were specified for sizes under 3 in. No protective coatings or wrappings were required on pipe to be buried because of the dry climate at Chuquicamata.

Jan 1, 1952

By R. T. Pepper

DURING an investigation into the effect of heat-treatment on the creep properties of the magnesium alloy ZW1, (1 pct Zn, 0.6 pct Zr), the previously published methods of final polishing were found to be unsatisfactory. A new diamond-polishing technique has been developed to facilitate the study of fine precipitates in this and other magnesium alloys, and is described in this note. The sectioning and coarse grinding techniques are those of Fox and Hall,' and Hess and Cieorge.' Pre- liminary mechanical polishing is done with 6-µ, and 1-µ( grades of diamond compound on 'Microcloth' using a lubricant of 'Hyprez' oil. When the conventional technique is used with a 1/4- µ diamond wheel for the final polishing of magnesium alloys, scratching is likely to result from the abrasive action of the 'Microcloth' nap on the soft metal surface. This problem has been overcome by the use of a cream on the wheel in order to separate the specimen surface from the polishing cloth. To prepare this cream, a solution of 6-ml tri-ethanolamine in 75-ml water is stirred into 12.5 g stearic acid at a constant temperature of 80° to 90°C. The cream is allowed to cool, and is then 'whipped' to a smooth consistency with 100-ml 'Hyprez' oil. Before polishing, the diamond loaded 'Microcloth' is moistened with a few drops of 'Hyprez' oil, then a small amount of cream is worked into the nap. For the best results a wheel speed of 500 rpm should be used for polishing, and the specimen should be held by hand with a light pressure while being rotated in the opposite direction to the wheel. After polishing it is necessary to swab the specimen thoroughly in methylated spirits to remove residual cream and oil from the surface in order to avoid uneven etching and drying stains. Fig. 1 shows the surface finish obtained on a specimen by polishing on a 1/4- ° diamond wheel using the cream, and Figs. 2 and 3 the precipitate observed in ZW1 after polishing, and etching for two seconds in nitric acid 15 pct in ethylene glycol. This new diamond-polishing technique has made a valuable contribution to the programme of metal-lographic research on ZW1 and has become a routine laboratory practice for other low alloy content magnesium alloys.

Jan 1, 1961

By J. M. Lommel

DURING an investigation into the effect of heat-treatment on the creep properties of the magnesium alloy ZW1, (1 pct Zn, 0.6 pct Zr), the previously published methods of final polishing were found to be unsatisfactory. A new diamond-polishing technique has been developed to facilitate the study of fine precipitates in this and other magnesium alloys, and is described in this note. The sectioning and coarse grinding techniques are those of Fox and Hall,' and Hess and Cieorge.' Pre- liminary mechanical polishing is done with 6-µ, and 1-µ( grades of diamond compound on 'Microcloth' using a lubricant of 'Hyprez' oil. When the conventional technique is used with a 1/4- µ diamond wheel for the final polishing of magnesium alloys, scratching is likely to result from the abrasive action of the 'Microcloth' nap on the soft metal surface. This problem has been overcome by the use of a cream on the wheel in order to separate the specimen surface from the polishing cloth. To prepare this cream, a solution of 6-ml tri-ethanolamine in 75-ml water is stirred into 12.5 g stearic acid at a constant temperature of 80o to 90°C. The cream is allowed to cool, and is then 'whipped' to a smooth consistency with 100-ml 'Hyprez' oil. Before polishing, the diamond loaded 'Microcloth' is moistened with a few drops of 'Hyprez' oil, then a small amount of cream is worked into the nap. For the best results a wheel speed of 500 rpm should be used for polishing, and the specimen should be held by hand with a light pressure while being rotated in the opposite direction to the wheel. After polishing it is necessary to swab the specimen thoroughly in methylated spirits to remove residual cream and oil from the surface in order to avoid uneven etching and drying stains. Fig. 1 shows the surface finish obtained on a specimen by polishing on a 1/4- µ diamond wheel using the cream, and Figs. 2 and 3 the precipitate observed in ZW1 after polishing, and etching for two seconds in nitric acid 15 pct in ethylene glycol. This new diamond-polishing technique has made a valuable contribution to the programme of metal-lographic research on ZW1 and has become a routine laboratory practice for other low alloy content magnesium alloys.

Jan 1, 1961

By John F. Abel

Personal gain seems to be the most powerful incentive to economic progress. One of my first exposures to this phenomenon was the comparison between company stopes and leaser's stopes in a vein gold mine. The company stopes were 4 to 6 ft wide, whereas the leaser's stopes were so narrow that in some places one could not turn around. The company leased only that ground it felt was marginal to mine with day's- pay miners, but by skillful, careful and energetic mining, the leasers managed in most cases to make money for themselves and the company. The Hurting Truth This capitalist system, which puts a premium on successful risk taking, is under attack. Not that it hasn't been under attack all along. What hurts and angers us at this time is that the attack comes from the children of the system-its youth. Such blasphemy should, of course, be eradicated, root and branch. Or should it? Are we hurt in part because some of the things they say are true?

Jan 1, 1970

By M. P. Tixier, A. Poupon, M. E. Loy

Simple qualitative methods are explained for identifying those shaly sands in a well that are most likely to contain oil. A need for more precise measurement of the variables that enter shaly sand analysis is indicated. Field examples are given to illustrate the methods. A theoretical discussion of the quantitative interpretation of shaly sands is given as a basis for discussion and as a guide for the future. While not generally capable of practical application at the present time because of the lack of sufficient accuracy of the electric log data, these methods may become more feasible in the future as the result of the improved logging methods now being introduced. INTRODUCTION Experience has shown that for porous formations containing only a negligible amount of clayey material, reliable information on the fluid saturation and porosity of the reservoir rocks can usually be derived from the electrical logs. The interpretation is based on empirical formulae relating the true resistivity of a porous formation to its lithologic character, to the resistivity of the interstitial water, and to the proportions of water and hydrocarbons in the pores.' If the resistivity of the interstitial water is not known, its approximate value can be derived from the SP curve.2,3 The porosity can usually be determined to a good approximation from a MicroLog, or a MicroLaterolog.4,5 When the reservoir rocks contain an appreciable percentage of clayey material, an additional factor is introduced into the analysis. In a clean formation, the matrix is an electrical insulator, so that the ability of the formation to conduct current is due only to the conductivity of the electrolytes in the pores; in a shaly formation, the shale constitutes a part of the rock matrix able to conduct current, and influences the resistivity of the formation. On the other hand, the spontaneous emf's and the circulation of the SP currents are also responsive to the presence of the shaly material. All other factors being the same, the deflection of the SP curve opposite a shaly permeable formation is smaller than the deflection which would be observed if the formation were clean. The effects of the presence of shale on the resistivity and on the SP curves in a shaly formation are functions of the amount, of the specific resistivity, and of the geometrical distribution of the shaly material within the formation. The presence of shale therefore constitutes an important complication in electrical log interpretation. It is easy to realize, for example, that the presence of a large amount of clayey material in a sand may have such a predominant effect that the resistivity of the sand will not change very much, whether the pores are entirely filled with connate water, or contain a high saturation of hydrocarbons. Important research projects have been undertaken to investigate the influence of shales and to find a way to allow for this influence in the interpretation of a log. The problem of the SP in shaly sands was analyzed mathematically by H. G. Doll"' for the case of sands with shale laminations. Formulae expressing the resistivity of shaly formations were proposed by Patnode and Wyllie8 and de Witte.9 Other quite interesting theoretical and experimental results have been published by various authors.l0,ll,12,l3,14 The purpose of the present paper is to discuss an approach to electrical log interpretation in shaly sands. This approach is based partly on some of the theoretical and experimental data available in the literature and partly on observations made on actual field logs. Comparatively simple methods for qualitative and semi-quantitative analysis are described. These methods depend principally upon information provided by the electrical logs.

Jan 1, 1955

By D. T. Van Zandt

IN the late 1920's rotary drills began to replace the churn drills in the petroleum industry, but until the middle 1940's the churn drill was the only widely accepted means of drilling large-diameter blastholes for quarry operations. The Calcite plant of the Michigan Limestone Div., U. S. Steel Corp., was one of the first to experiment with rotary drills for quarry blasthole drilling, and the first to employ compressed air on a fully rotary rig to cool the bit and raise the cuttings to the collar of the blasthole. The Calcite plant operates a limestone quarry near Rogers City, Mich., in the northern part of the lower Michigan peninsula. The formation quarried, a portion of the middle Devonian series, is the Dundee limestone, which is uniform, seldom massive, and characterized by definite bedding planes. The dip is southeast, 40 ft to the mile. Quarry faces vary from 20 to 116 ft in height. Vertical blastholes are used entirely, from three to five rows of holes being drilled parallel to the working face, spaced 18 ft apart with 18-ft burden and drilled 6 to 8 ft below shovel grade. Quarry operations coincide with the navigation season on the Great Lakes, as the bulk of the stone is transported by lake carrier. The normal operating season runs from April to December, the remaining time being devoted to stripping operations and plant and equipment maintenance. In the followirig discussion drilling rates mentioned refer to overall drilling time and include all operations such as moving from hole to hole, penetration and extraction of tools, and routine maintenance. Time consumed by such factors as power delays and major machine repair is not included in drilling time unless otherwise stated. Figures cover only operations at this one plant in the formation mentioned. Needless to say, a very different set of figures could be obtained in a different formation. However, the comparison of footage obtained with churn drills and rotary rigs in this particular formation has been used as an indication of what might be the expected performance of rotary rigs in other formations. Prior to 1950 the bulk of the blasthole drilling at the Calcite plant was done by electrically powered churn drills. Both crawler and wheel-mounted rigs were used. These machines, which mounted a 22-ft drill stem of 4½ in. diam and a spudding type of bit 2 to 4 ft long, drilled a hole of 5 ?-in. diam. Average drilling rate of these rigs in the Rogers City formation was 8 % ft per hr. In 1946 one of the first rotary blasthole drills offered to the quarry industry was put into use on an experimental basis. This machine, known as the Sullivan Model 56 blasthole drill, Fig. 1, was on 16-in. crawler pads and electrically powered at 440 v. The drill bit, a Hughes Tri-Cone roller bit of 5?-in. diam, Type OSC, was threaded into the end of the 4-in. square hollow drill rod or stem. These drill rods were 20 ft long with female threads on one end and male on the other to allow for addition of the desired number of rods for drilling holes of various depth. Rods were handled by a single drum hoist geared to the main drive motor and racked by a 30-ft derrick or mast when not in use. The cable from the hoist drum fed through a crown block on the top of the derrick back to the water swivel mounted in the top end of the drill stem in use. This cable remained attached during drilling operations and was used to hoist the tool string from the hole. Down pressure was applied to the tool string by means of a pair of 4-in. diam hydraulic cylinders acting on the drill chuck holding the drill rod. The first chuck consisted of flat jaws which gripped the flat sides of the stem. These jaws were controlled by set screws forcing them into contact with the drill stem. As these set screws had to be loosened and tightened by hand with each stroke of the hydraulic feed cylinders, there was great delay. For this reason the semi-automatic chuck was developed which automatically gripped the stem on the downward stroke but released for retraction of the hydraulic feed cylinders. Rotation was imparted to the tool string by a rotary table acting on the chuck and geared to the main drive motor through a separate gear train and clutch. A positive displacement water pump, mounted on the drill, fed water through a system of pipes and hose into the water swivel mounted on the top of the drill rod and through the rod and bit, washing the drill cuttings to the collar of the hole. Where water was scarce, provision was made to settle out the cuttings coming from the collar of the hole and re-use the water. Where water was abundant the stream coming from the hole was wasted. Drilling rate with this machine was about 20 ft per hr and bit life 1600 ft of hole. While this rate was more than twice that obtained with the churn drills employed, the problem of water supply and drill cuttings disposal rendered the machine impractical from an operating standpoint. Consequently it was used only in that part of the operation for which water was easily supplied, when the character of the formation made it least difficult to wash cuttings away from the collar of the hole. In October 1949 it was suggested that drill cuttings be removed by compressed air, long used for this purpose on pneumatic drills, and collected at the collar by suction. Thereafter, the water pump on the Sullivan 56 was replaced by a 500-cfm air compressor and a trial run made. Air pressure at

Jan 1, 1955

By Whitman E. Brown, Louis J. Erck

L. A. ROE*—This paper is one of great value to the iron ore industry. The Humphreys spiral is a relatively new tool and gives promise of being quite useful in solving certain problems of iron ore beneficiation. Spirals have been tested on a martite ore at the Benson mine of Jones and Laughlin Steel Corp. in northern New York State. Our New York ore is considerably different, physically and mineralogically, than the one mentioned in this paper, and contains only 25 pct iron. This ore can be processed on spirals to give a concentrate containing 62 pct iron with an 85 pct recovery. On our particular ore we found operating difficulties when the spiral feed was coarser than 14 mesh. These were chiefly due to excessive "build-up" of locked middlings in the spiral circuit. In extreme cases these middling particles would accumulate to such an extent that the plant had to be shut down and the spiral surge tanks cleaned out. It is interesting to note that there exists a close relationship between the results of tabling a given iron ore and concentration of this same ore on Humphreys spirals. The size range of the ore must, of course, be within those limits acceptable to spiral concentration. Comparative tests on several of our mar-tite ores showed tabling results to be the same as spiral results. The authors make no mention of the use of a tailing stream splitter (now available from the spiral manufacturer) which is a useful tool in studying iron losses in the tailings stream. 011 our particular martite ore we found a considerable accumulation of fine-sized iron ore particles in the outside portion of the tailings stream. This fraction may be amenable to further treatment. E. H. ROSE*—The authors have concluded an interesting piece of work and this paper is an excellent factual account of a rapid and somewhat unusual transi-;ion in practice in the plant described. Perhaps it was modesty on their part which caused them so casually to limit to one sentence the fact that "several other methods of concentration on the fine ores representative of the Hill Trumbull group . . . failed to produce consistently an acceptable grade and recovery of finished product." The fact is that they were confronted by a difficult mineral-dressing problem and they are to be congratulated on their courage and persistence in staying with it until a satisfactory solution was evolved. A visitor to the plant might have been mildly astonished, as I was, to see one type of concentrator handling that part of the load early in the 1947 season, its experimental replacement by another a little later, and then in 1948 to see that both types had simply disappeared and their place taken by the spirals which were operating as placidly as though they had been there all the time. In working out economic methods of beneficiating Alabama red ore, most of which development is still ahead of us, it is likely that we also will pass through a period of successive disappointments, for we too have a problem where it is next to impossible either to "guess 'end right in the first place," or, because full-scale operating cost is such an important factor, to lay the ultimate answer on the line in advance by means of laboratory experimentation. Iron ore being what it is instead of what it used to be, it is encouraging to have the example the authors have given us today that a tough nut that will not crack on the first blow or the second is apt to do so on the third or fourth.

Jan 1, 1950

By William Kelly

I AM-glad to have the opportunity at this time to say that I consider it a very great honor to be elected President of the American Institute of Mining and Metallurgical Engineers. It fulfills the proudest hopes of my ambition. There is, ,however, one drawback to my personal gratification-a feeling of unworthiness to have my name placed on the list with former Presidents of the Institute, with the brilliant Raymond who guided our society for so many years, with the versatile and lovable James Douglas, my first employer and my friend, a most generous benefactor of our Institute, with the skilful Holley, who gave to the manufacture of steel in this country, its first great impulse, to mention only three of the accomplished men who have preceded me. I assume that the honor of my being elected President of the Institute is not merely a personal one but that it should be considered a tribute to the wonderful progress in the art of mining, both underground and in open pits, and in the beneficiation of native copper and iron ores, which have been shown in the Lake Superior District and also a recognition of our members there and elsewhere who are living at the mines and directing the operations, upon the success of which the whole mineral industry depends.

Jan 1, 1924

By Roger L. Goin

Maintaining smooth haulage roads is a key to significant cost savings at Kennecott Copper Corp.'s Bing- ham Canyon copper mine, located near Salt Lake City, Utah. The truck operations section of the mine moves an estimated 300,000 tpd of waste plus some ore to maintain a balanced production cycle. A three-year truck replacement program has produced a dramatic new look in both size and capacity of the truck fleet. Currently there are 95 haulage trucks in the Bingham fleet with capacities up to 150 t. The trend towards the larger haulage units places greater importance on overall truck availability. The impact of a 150-t truck being out of service is far more critical than with the former 65 or 85-t units.

Jan 12, 1974

By N. J. Grant, Hans Brunner

ThE materials used in this investigation were furnished by the Aluminum Co. of America and consisted of high-purity alurhinum (99.995 pct) and two aluminum-magnesium alloys containing approximately 2 and 5 pct Mg. The compositions of these alloys are listed in Table I. The materials were received as hot rolled 3/8-in. diam rods and were made from the same heats as the materials which were previously used in this laboratory.'j3 The rods were cold swaged to 1/4-in. diam by 1 in. long, and then had two parallel flats machined over the gage length 0.085 in. apart. The test bars were annealed for 10 min at 700° F, and strained 1 pct in tension at room temperature. They were electropolished (80 pct acetic acid; 20 pct perchloric acid, 70 pct concentration) at a temperature of 60°F at approximately 20 v. A grid, transverse or square, spaced 0.2 mm apart, was impressed by rolling a fine threaded screw (128 threads per in.) across the flat surfaces. The depth of the grooves was less than 5 p. For final annealing and grain size control the specimens were enclosed in aluminum foil and sealed in Vycor tubes. Annealing conditions varied from 1 hr at 700° F to 5 days at 1050° F, depending on the alloy content, to give for most of the specimens a final grain size of about 1 mm. This treatment was followed by electropol-ishing as described above. For the test series presented in Figs. 4 and 5 round specimens were used instead of flat ones, and fine machining marks served as marker lines instead of the impressed grid lines. Electropolishing was omitted. The major advantage is that the spacing of the machining marks is only about 1 u as compared to the 200 u of the impressed grid lines, permitting higher precision in strain measurements. Unfortunately the strain annealing necessary for grain size control also resulted in some preferred orientation of the grains. The preferred orientation of three arbitrary specimens was measured by X-ray diffraction methods and the averaged results are shown in Fig. 1. The possible effects of the preferred orientation on the test results will be discussed later. The testing temperatures were 410°, 515", 725", and 940° F, with one exception in the case of the 5.1 pct Mg alloy. Since the solubility limit of this alloy is 520°F, a test temperature of 545°F was cho-se sen instead of 515°F. The selection of the test temperature was based on two considerations: 1) at 900°F and above, extensive grain boundary migration made it more difficult to measure grain boundary sliding; 2) at 400°F and below, there was so little grain boundary sliding that it was again difficult to make precise measurements. The stress was chosen to give a creep rate of approximately 2.7 pct per hr in connection with Figs. 2

Jan 1, 1961