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By James P. Savely, Victor L. Kastner

Inspiration Consolidated Copper Company is currently mining in four pit areas; Live Oak, Red Hill, Thornton and Joe Bush Extension, near Globe, Arizona. Small satellite orebodies lying outside the main pits have been mined out in a series of small open pits. Major slope instability has affected production in every pit. Some consequences of the instability have been; (1) Serious interruption of mining plans, which resulted in deferred ore production, lost ore, reduced mining rates and increased stripping requirements. (2) Forced replacement of crushing facilities. (3) Serious effects on other production facilities. The introduction of a slope stability monitoring and control function as an integral part of mine planning has been successful in reducing the impact of these failures and will eventually result in economic pit design based on probabilistic slope analysis. This case history presents a brief review of the problems associated with slope instability at Inspiration's mines followed by a more detailed report on one specific failure area in the Thornton pit where the use of horizontal drain holes for slope stabilization appears to be successful. Relationships between slope movement rate, rainfall, and mining activity are shown.

Jan 1, 1983

By J. Handin, S. J. Bauer, M. Friedman

Deformation mechanisms in room-dry and water-saturated specimens of Charcoal Granodiorite, shortened at 10-4s-1, at effective pressures (Pe) to 100 MPa and temperatures to partial melting (?1050°C) are documented with a view toward providing criteria to recognize and characterize the deformation for geological and engineering applications. Above 800°C strength decreases dramatically at effective pressures ? 50 MPa and water-weakening reduces strength an additional 30 to 40% at Pe = 100 MPa. Strains at failure are only 0.1 to 2.2 percent with macroscopic ductility (within this range) increasing as the effective pressures are increased and in wet versus dry tests. Shattering (multiple faulting) gives way to faulting along a single zone to failure without macroscopic faulting as ductility increases. Microscopically, cataclasis (extension microfracturing and thermal cracking with rigid-body motions) predominates at all conditions. Dislocation gliding contributes little to the strain. Precursive extension microfractures coalesce to produce the throughgoing faults with gouge zones exhibiting possible Riedel shears. Incipient melting, particularly in wet tests, produces a distinctive texture along feldspar grain boundaries that suggests a grain-boundary-softening effect contributes to the weakening. In addition, it is demonstrated that the presence of water does not lead to more microfractures, but to a reduction in the stresses required to initiate and propagate them.

Jan 1, 1982

Jan 1, 1965

By Thomas T., Read

ANY discussion of State-supported schools of mining and A metallurgy needs to be prefaced by a definition, since the first school to offer a mining curriculum, the Pennsylvania Polytechnic College, was established by a grant from the State. But, so far as can be determined, it received no further State support. It would be tiresome to go into precise details,. and it will be enough to stipulate that by State schools are meant those which derive their main financial support from regular appropriations made by a State legislature. The University of Michigan was the first of these to offer a mineral industry curriculum; its story, as well as those of Wisconsin and Illinois, has been related in the preceding chapter.* Arranging the others in chronological order involves making a decision as to what weight to attach to plans for giving instruction, as contrasted with its , effective initiation. This may appropriately be illustrated by the history of mineral industry education in California. California Among those who came to California in 1849 and the following years were many New Englanders who had a keen interest in education. Some of them, mostly Congregationalists and Presbyterians, sought to establish a school of higher education around San Francisco Bay and eventually succeeded in opening the "Scientific and Classical School at Contra Costa' in Oakland, at Broadway and 5th St., on June 20, 1853. This prospered so well that it was chartered by the State as the "College of California" on April, 13, 1855. Developing so rapidly that its quarters became cramped, in 1864 it not only bought a tract of

Jan 1, 1941

By Orville R. Lyons

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

Jan 1, 1953

By Orville R. Lyons

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

Jan 1, 1953

By J. W. Pugh, J. D. Nisbet

F. B. Foley—The use of data published by Wever and Jellinghaus in 1931 to fix boundaries of the sigma phase in the Fe-Cr system, in the face of the author's own references to the suggestions of Bradley and Goldschmidt, Aborn and Bain, and Hougardy that the phase is much more extensive, and the very much more accurate work, which weighs heavily in favor of these suggestions, of Cook and Jones, published in 1943 and evidently disregarded by the authors, makes their derived ternary diagram, especially fig. 15 for 400°C, quite inaccurate on the Fe-Cr side and affects the extent of phase boundaries in the most controversial and important part of this ternary system. In commercial Fe-Ni-Cr alloys the occurrence of sigma has been observed time and time again at lower chromium contents than that of the 24 pct Cr, 16 pct Ni, 60 pct Fe alloy which is the lowest permitted by fig. 15 of practically carbonless metal. Newel1 reports sigma in 27 pct Cr-Fe even with some carbon present to be one of the greatest detriments to its extensive application, whereas Pugh and Nisbet set a low limit of 36 pct Cr for sigma in the binary Fe-Cr system. J. J. Heger—The diagrams presented by the authors do not agree with observations made on commercial Fe-Cr and Fe-Cr-Ni alloys, nor do they agree with two recent investigations made on the Fe-Cr and the Fe-Cr-Fe systems. I refer first to the investigation made by Cook and Jones1' on the sigma region of the Fe-Cr system. On the basis of their results which were published in 1943, Cook and Jones established new boundary limits for the sigma and the alpha plus sigma regions. These new limits have been accepted and are incorporated in the Fe-Cr diagram that appears in the 1948 edition of the Metals Handbook. This diagram is shown here as fig. 34. As will be noted, the boundary limits of the alpha plus sigma region in this diagram extend to much lower chromium contents than do those in the diagram presented by the authors in fig. 2. These new limits indicate that sigma phase should form in an Fe-Cr alloy containing 26 pct Cr, and experience with commercial alloys confirms this finding. Indeed, recent studies on a commercial Fe-Cr alloy containing 17 pct Cr have shown sigma phase to be stable in this alloy at 550°C. I recognize that the authors' work did not include studies on the sigma region; however, I believe this is a serious omission because sigma phase may profoundly affect the physical properties of these alloys and should be evaluated in any investigation which attempts to relate physical properties to the equilibrium diagram. The second investigation to which I refer is that made on the Fe-Cr-Ni system by Rees, Burns, and Cook,'' who used pure alloys and employed heating 17 W. P. Rees, B. D. Burns, and A. 3. Cook: Journal Iron and Steel Institute. (July 1949) 162, Part 3. p. 325. times that extend to 200 days. These investigators published their results in July 1949 and from these results constructed isothermal sections at 800" and 650°C. The isothermal section at 650°C is shown here as fig. 35. This diagram indicates sigma phase should form in a pure 18 pct Cr-8 pct Ni alloy. Although sigma phase has not been observed after 10,000 hr at 1200°F in commercial 18-8 alloys containing carbon and nitrogen, it has been observed under the same conditions in 18-8 alloys modified with titanium and columbium, both of which serve to reduce the effect of carbon and nitrogen. This section and the one at 800°C suggest that the constant iron sections which are presented by the authors, should be considerably altered, if they are to represent an accurate picture of the system. A few of the suggested alterations are as follows: In fig. 9, the section at 50 pct Fe, the gamma plus sigma region should be widened, not narrowed at 800" and 650°C. In fig. 10, the section at 60 pct Fe, the gamma plus sigma region should be widened at 800° and 650°C. In fig. 11, the section at 70 pct Fe, an alpha plus sigma, an alpha plus gamma plus sigma, and a gamma plus sigma region should be added. In fig. 12, the section at 80 pct Fe, the alpha plus gamma region should be widened at 800°C. In fig. 13, the section at 90 pct Fe, the alpha plus gamma region should be widened at 650° and 800°C. Undoubtedly, the chief reason for the discrepancies between the authors' data and those of Rees, Burns, and Cook is that the authors did not employ long enough heating times to allow for transformation. In this connection, I wish to warn that transformations in these alloys, particularly transformations at temperatures below 800°C, are extremely sluggish and may require a year or more to approach completion. Therefore, unless extremely long heating times are employed or steps are taken to accelerate these transformations by such means as mechanical working, the results will not yield an accurate picture of the alloy system. Certainly, an accurate picture is needed for the development of better high-temperature materials. E. J. Dulis-The purpose of this work is not clear. If an improvement of the ternary equilibrium diagram of the Fe-Cr-Ni system was wanted, it seems logical that testing techniques superior to those previously used would be a prime requirement. To approach equilibrium in this system, long holding times are needed; a fact established long ago but apparently ignored by the present authors, who used the continuous heating and cooling tests in equilibrium studies. A publication on the same system by Bradley and Goldschmidt6 was criticized by Monypenny in a

Jan 1, 1951

By W. C. Ellis, E. S. Greiner, M. E. Fine

C. H. Samans and W. R. Ham (Chicago, Ill., and Dix-field, Maine, respectively)-—For several years we have been studying transitions of this basic type in metals, alloys, glasses, etc. Usually, however, they are not so clearly marked as those which the authors have found, and hence are much more difficult to determine accurately. Since our studies indicate that most of them occur virtually unchanged (as far as temperature is concerned) regardless of the form in which the element appears, we believe that they are a characteristic of the atom. Specifically, we believe that there is an additional rotational degree of freedom possessed by the nucleus which has not been considered heretofore. This nuclear rotation is made up of several components, each related to the several quantum shells of electrons. In chromium there are four of these shells and hence four separate series of characteristic transition temperatures. The lowest temperature at which any transition occurs, based on the present state of our computations, is 125°K, 4° higher than the authors' value of 121°K. A convergence of this series, we believe, shows up at the higher temperature of 2085 °K, surprisingly close to the transformation temperature of 2103°K recently announced by N. J. Grant of M.I.T. for a body-centered to face-centered transformation in chromium. Likewise, our computations indicate a temperature of 311°K for the second transition temperature, reported by the authors as 310 °K. A convergence of this series, we believe, shows up at a higher temperature as the melting point at 2163°K. Although our work on these series must, in a sense, still be regarded as empirical, since we do not understand fully as yet just what the series mean, it is based on a reasonably firm picture. The individual constants, from which the various series are computed for each element, comes directly from the X-ray K absorption limit. Furthermore, the same basic method has accounted for transformation and melting temperatures in about 50 of the chemical elements, which is all we have tried thus far. In many cases the only known transformation is the melting point, but in others the occurrence of transformations or other transitions, equally as well marked as those of the authors, has been pointed out by others. These observations have assisted us greatly. Consequently we were very pleased to see the authors' excellent work in finding these two transitions in chromium. With these confirming data, our picture of this element is clarified considerably, so we expect that at least some of our work can be published in the near future. R. C. Ruder (E. I. du Pont de Nemours & Cu., Wilmington, Del.)—The authors' interpretation of these transitions in terms of 3d to 4s electronic structure transitions is most interesting. It would be interesting to have additional experimental evidence of such transitions from the temperature dependence of the Hall coefficient in the neighborhood of the property changes discussed in this paper. Simple theory15 suggests the Hall coefficient as a measure of the free electron (or s electron) concentration per unit volume. It has been shown that for paramagnetic'" and ferromagnetic1? metals the simple theory is in fact too simple. However, the existence of a discontinuity in the Hall coefficient would provide information which should aid both in our understanding of these transitions and the significance of the Hall coefficient in these metals. It was rather surprising that no significant paramagnetic effects were observed. In this connection the recent work of McGuire and Kriessman18 is cited. They measured the magnetic susceptibility of chromium from 20" to 1460 °C. They also observed no large change in the susceptibility although there might be a change in slope in the vicinity of the 40 °C transition. The existence of these 3d to 4s electronic transitions has been discussed in connection with the paramagnetic susceptibility behavior of nickel and nickel alloys.'"-" Assuming a correspondence principle between classical and quantum mechanical paramagnetic theory and using classical theory to calculate the effective Bohr magneton number from the Curie constant for substances obeying the Curie-Weiss law," it is found that the effective magneton number is a function of temperature. The process of calculation involves the inverse of the differential of the 1/x4 vs. temperature curve so that good and numerous data are necessary to obtain significant results. The data of Fallot23 show a discontinuous increase of about 12 pct in the effective magneton number between 850" and 900 oC, followed by a continuous increase up to the melting point. The data of Sucksmith and Pearce24 show a possible 8 pct increase. The older data of Terry25 and Weiss and Foex26 show a continuous increase. It is possible that small amounts of impurity atoms change these electronic transitions significantly. Fallot's23 data on a nickel alloy with 4.5 atomic pct Fe indicate that the discontinuity occurs around 1300 °C. Systematic investigation of the transition metals for transitions of this nature should provide information which would be very valuable for our understanding of these metals. The absence of antiferromagnetic structures in chromium has been shown by Shull27 using neutron diffraction techniques. M. E. Fine, E. S. Greiner, and W. C. Ellis (authors' reply)—The remarks by Drs. Samans and Ham are certainly very interesting, in particular those pertaining to the close agreement between the theoretically calculated values for transition temperatures in chromium and the experimental values reported by a number of investigators. This is a remarkable achievement and we shall look forward to a more detailed presentation of the method followed in their calculations. We do not believe that the transition in pure chromium near 40 °C remains temperature invariant with alloying, as was reported by Samans and Ham for a number of the substances that they studied. We have not done any work with alloys but base our belief on the results of earlier studies in which less pure chromium was included and considerably lower transition temperatures were observed. The transition tempera-

Jan 1, 1952

By Bunting S. Crocker

PEBBLE grinding at Lake Shore is not a temporary 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 1 1/4-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 1/4-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 1940.1 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 1/4-in. and 3/4-in. balls, the use of laboratory mills to evaluate plant changes, and several reports on classifiers and classification. In the following July the addendum report' 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 1/2-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 SiO2, 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, 1952

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 Mark David Rankin

This paper represents a general assessment of the primary factors driving the market for mobile offshore drilling rigs and the utility of those factors as choice variables for the offshore drilling program planner.

Jan 1, 1982

Andy H. Vassiliou – Introduction Ladies and gentlemen,.good evening and welcome to the Paul F. Kerr Memorial Symposium dinner. I believe we are all here to honor the memory of a great teacher and advisor, and a true friend -- Professor Kerr, or Dr. Kerr, or to most of us who were his students, "Pappy" Kerr. Most of us who attended the Paul F. Kerr Memorial Symposium earlier today at the New York Hilton had the opportunity to read, or listen to, or discuss scientific papers presented in honor of Dr. Kerr. In addition, all of the speakers had the opportunity to reminisce about their association with Dr. Kerr, who was credited as having contributed substantially to their graduate education and the development of their careers. Dr. Kerr's two daughters -- Ruth and Nancy -- who attended the Symposium at the Hilton are also here with us tonight. I think we should keep tonight's festivities on an informal basis since we are all old friends or former students of Dr. Kerr. Some of us may want to say a few words about our experiences at the time we were students at Columbia, and hopefully include some anecdotes relating to Professor Kerr. I would like to invite everyone to do so, but first we have arranged for two of us to be the lead speakers, whom I will introduce in a few moments.

Jan 1, 1985

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