Search Documents

By John Dasher

Getting spodumene to float quickly and cleanly can be a problem. The author has presented an excellent account of a valid and useful approach to the scale-up of such problem floats. This indicated advantages for doubling retention time, i.e., buying another set of flotation cells. The author states that 21 pct of the spodumene floated in 2 min in the laboratory cell, that the grade of the rougher was about 75 pct spodumene, and that the final tailing contained 4 to 12 pct spodumene in different sizes with the present cells and 1 to 9 pct spodumene with twice as many cells.

Jan 6, 1959

By Maurice Grieves

GC. Waterman Mr. Grieves' paper on "Shaft Sinking Today--A Boring Business Tomorrow" in the Dec. 1981 issue of MINING ENGINEERING is an excellent description of recent improvements in speed and costs of shaft sinking. However, shaft boring techniques are not a recent development. In 1935-36, the Idaho Maryland Co. (gold), Grass Valley, CA, bored the vertical 1.5-m-diam Idaho No. 2 ventilation and supply shaft to a depth of 346 m. The equipment was designed and perfected by Branner Newsome and was, I believe, described in "Transactions." Pickands Mather later bored a 1.8-m-diam shaft to a depth of about 305 m using Newsome's design. The Idaho No. 2 shaft cut through hard gabbro, diabase dikes, soft serpentine, very incompetent ankeritized serpentine and a strong fault zone. The completed shaft did not require timber support; the fault zone was cemented. Equipment consisted of a rotating 1.5-m or 3-m core barrel with a slotted bottom which cut a 76-mm kerf. The cutting agent was chilled steel shot introduced into the slot as needed. The driving mechanism and core barrel were lowered into the shaft, the former secured to the wall and the inshaft "engineer" operated the motors which controlled core barrel rpm and advance. After an advance of 1.5-3 m the equipment was lifted out of the shaft, a half(?) stick of powder cut off the drilled core, and it was hoisted to the surface with a cable attached to an eye bolt at the tope of the core. A "shift" consisted of three men: hoistman, equipment operator (down the shaft), and a surface laborer. Advance was variable as equipment and techniques were perfected. Near the end of the job advance was, as I remember, 1.5-3 m per shift. Costs were, as I recall, about $115/m. Up to 4 m cores were lifted out in one piece and swung to the dump by a stiff leg derrick. The Newsome equipment was relatively inexpensive to build and operate and his method should be (more than?) competetive with the methods described by Grieves. W.E. Hawes The author, in mentioning the South African developments of the cactus grab, ignores the parallel development of the Cryderman mucker in this hemisphere. Perhaps a slightly longer history of blind hold drilling would have been in order. Blind shaft boring got a major boost at the Nevada Test Site, as part of the nuclear weapons development, when it was essential to be able to quickly assess depth for weapons testing. One of the early civilian attempts to utilize this technology occurred when Kerr McGee drilled the shaft at the Section 19 Mine at Ambrosia Lake, NM. This twin masted drill rig belonged to a subsidiary of Kerr McGee, not Shaft Drillers. This was a difficult hole, due to bentonitic shales that decomposed. Later advances in drilling fluids eliminated this problem. The Conoco project referred to is not entirely correct, and needs minor amplification: The depth of the development (not pilot) shaft drilled was 684 m, at a diameter of 3.05 m, with a 2.16 m ID hydratatic casing being installed to a depth of 669 m. Conoco Inc., not Challenger, executed the drilling, however Conoco did use Challenger Drilling Company's Rig 14-S. Rig 14-S is unique in that it is specifically designed for shaft drilling, rather than being an oil field rig used for this service. All work was planned and directed by Conoco personnel. Deviation of the aforementioned shaft from vertical was 31 cm. Two smaller shafts (pilot shafts) were drilled with the same rig. These smaller ones were drilled 1.8 m in diameter, to depth of 665 m and cased with .91-m-diam casing, to depth of 650 m to be used as pilot holes for 5.5 m diameter shafts. The total elapsed time for all drilling, casing, cementing, and rig moves was 50 weeks. The constraints mentioned of vertical accuracy, torque, wall collapse, rig set up time, lining time etc., pumping out, etc., may not be as serious as the author implies. First, vertical accuracy: The record achieved by Conoco indicates that with reasonable care, good accuracy can be obtained, primarily by having a fair amount of weight in tension, rather than on the drill bit. Second, the amount of torque that can be applied to the drill stem has been greatly increased by the introduction of larger diameter drill stems, such as the new Hughes rig uses. Third, risk of collapse of shaft walls is minimal with the proper mud program. Basically, the walls are supported by the weight of the mud and it is not removed until some type of casing is installed and grouted in. Here in the real challenge of blind drilling shafts to devise a more economical method of lining bored shafts than using steel casing. The remaining issues were minor at Conoco, provided adequate planning and scheduling takes place. An excellent paper, "Shaft Drilling--Crownpoint Project" by Hassell H. Hunter, presented at the Fifth Uranium Symposium in Albuquerque in Sept. 1981 contains details of this project. A combination of blind boring, with enlargement by either mechanical means (boring machine, or tunneling type shield) or slabbing down, with muck removal through a larger bored shaft would seem to be the future trend in shaft development. The economics, especially for wet conditions, favor this concept over conventional sinking, as has been demonstrated at Kerr McGee --Churchrock, UNC--Churchrock and Conoco at Crownpoint.

Jan 3, 1983

By Stephen Semeniuk

Discussion by Stephen Semeniuk Discussion As a valuator and SME member, I view Ross D. Lawrence's position as based on unproven and faulty assumptions made with regard to the pragmatic evidence presented in his paper. Moreover, his call for the inclusion of all mineral resources in discounted cash flow valuations is redundant, if not irrelevant, as there are procedures that already take all classes of mineralization into consideration for valuation purposes. In my experience as a valuator, I find that Canadian securities regulators are more than accommodating in allowing any valuations of mineral properties to be based on cash flow calculations. In other business areas, the same regulators usually reject discounted cash flow valuations of projects or businesses with no operating history. Instead, the regulators prefer valuations to be based on other transactions tested in the marketplace. Valuations based on actual market transactions are less subjective than discounted cash flow calculations and/or the capitalization of arbitrary, untested or undemonstrated earnings projections. In the mining and minerals sector, such projections may be made by persons not qualified to make such projections. In support of this statement, I direct readers to the article "Valuation of mineral properties which do not contain exploitable reserves" by Lionel C. Kilburn, which appeared in the August 1990 issue of the CIM Bulletin. To quote from the article: "Value of a mineral property depends basically upon metal prices, recoverable grades, mining and metallurgical practices, commodity supply and demand, and political-economic conditions, all of which vary with time. Five fields of expertise influence the value of a mineral property: •geological engineering, •commodity markets, •financial markets, •stock markets and •mineral property markets. The skills of geologists and engineers properly apply to the first field of expertise. Knowledge and consideration of the other four fields is important, but such advice is more properly sought from marketing experts, economists and stockbrokers." Securities regulators have created no hardship for valuators by excluding unsubstantiated resources in discounted cash flow valuations. In the field of engineering, best design practices apply. A margin of safety is built into the design criteria for any new equipment or structure that will exceed normally expected load levels and stresses, even though such considerations imply higher front end costs. In accounting, the comparable operating principle is conservatism. It explains why, for balancesheet presentation purposes, inventory is stated at cost or market, whichever is the lowest value. Similarly, when the engineering or geoscience-based valuator plays in the field of finance, the rules of accounting and applied economics apply. Regulators in their skepticism of income-based valuations are in tune with similar views already expressed in mining literature. To quote H. Van Tingley, in Mining Magazine, July 1996: "... the DCF method has inherent dangers when applied to undeveloped resources. Firstly, it presents an often-undeserved aura of accuracy. A DCF value is no more accurate than the least accurate of the assumptions incorporated into the calculation - and assumptions

Jan 1, 2003

By E. Arioglu

Discussion by E. Arioglu Y. Luo and S.S. Peng are to be congratulated for their outstanding paper providing new results from surface subsidence monitoring over longwall panels under different conditions. This writer found the authors' paper very interesting. In reference to this paper, the following contributions are offered: • In the paper, the maximum possible subsidence [(ASmax)] induced by chain pillar convergence has been related to a parameter (h/1- p) by linear regression. (Notations are expressed in this discussion's Fig. 1 (b).) It is a well-known fact that in a goaf the roof lowers onto broken material that gradually compacts. Then the physical behavior of goat may closely follow the compaction characteristics of broken materials. In light of this assumption, a mathematical means (Fig. 1(a)) of defining the expected compaction (i.e. surface subsidence) represents an exponential expression between [ASmax/m vs. h/1-p]. This is based on Luo and Peng's field observations from the Appalachian, Illinois and Alabama coal fields. (The coefficient of regression "r" under consideration was computed to be 0.883). • According to Wardell's (1969) findings for board and pillar mining, the maximum possible subsidence can be expressed in terms of maximum pillar loading (depth, ratio of extraction) and the height-to-width ratio of the pillars left. Using Luo and Peng's 1991 results, this writer discovered a regression equation between the above-mentioned factors. As displayed in Fig. 1(b), the functional expression approximates as closely as possible the observed subsidence data. It may be noted here that Wardell's approach (1969) can be successfully utilized for longwall panels having chain pillars. Reference Wardell, K., 1969, "Ground surface and control," Mining Congress Journal, January, pp. 1-7.

Jan 1, 1993

By A. M. Radigan, W. G. Prast

M.S. Azun The authors' paper did not explain the facts and reasons why state companies are established and what their role is in international industrial mineral trading. Also, I do not agree with the conclusion implying that the main goal for state companies is to capitilize natural resources. Almost all state companies are founded by certain laws and their activities are arranged by state regulations. Their only aim is to operate mineral deposits for their people and to protect the rights of the people. Although some ore deposits, such as nonstrategic or small, may be exploited by privately owned companies, the state has the right to control the operation and production. Activities of state companies may or may not be widened because of the character of the deposits and the importance of the mineral market involved. For some countries, although governments may occasionally change hands, the policy of the mineral trading system is not affected by this change. There is no doubt that any government desires to make more earnings to increase GNP, when it is involved in operating and trading mines. It wants to produce and sell the finished products, instead of unprocessed ones, if possible. When it is known that some mineral deposits are relatively rare, such as Turkish boron deposits, the government can nationalize the mines to control production and affect the market. If there is an abundance of a certain mineral in the market, the government can cut production and set a floor price, such as Canadian potash. Also, when production of a specific mineral is unique, such as Turkish chromite, the price of such mineral may be set by correspondant government. One may add that if some governments with their state companies produce a certain mineral in a larger amount, they may decide to establish a cartel to protect their production and price. These are the facts followed by government-owned state companies and partially government-owned companies. The reason being charge of many deposits by the state companies is always the same: "protect" the mineral rights of the people.

Jan 2, 1984

By J. C. Wynn

J.R. Hillebrand From the abstract, "A realistic ore deposition model to serve as a conceptual framework." Further, on page 246, "The first area of study was the Josephine peridotite, now identified as an ophiolite terrane, located in northern California and south-western Oregon (Fig. 3). Here, it was felt that we could test the models shown in Figs. 1 and 2 and, if successful, expand the model and test it in other chromite districts in the US and around the world." My comments refer only to chromite deposits in Europe and the Middle East. First, I question that may chromite deposits occur in the idealized ophiolite section shown in Fig. 1. Second, although some podiform deposits (referring to morphology) have been important (Guleman, Turkey, for example), the stratiform deposits are of greater importance in tonnage and frequently grade. Examples include Nada, Yugoslavia; Bulqiza, Albania; Xerolivado, Greece; Saranovsk, Russia; and Abdasht, Iran. Reply by J.C. Wynn Mr. Hillebrand raises two issues that deserve a serious response. One issue is basically subjective and the other is ultimately political. First, the question of where (and how many) podiform chromite deposits are found in the idealized ophiolite section of Dickey (1975). The Dickey model is still perhaps somewhat controversial (Panayoutou, 1980), but our experience in the Josephine ophiolite of northern California (Wynn, 1980; and Albers, 1983) shows that the model works well there. Virtually all the chromite is found in dunite lenses in the tectonite peridotite (harzburgite). Since there are few complete ophiolites exposed in the world, the model will probably never satisfy everyone. The degree of acceptance is a function of any individual's subjective experience. Second, it is well known that stratiform chromite contributes more than 95% of the chromite reserves presently known in the world, and it could be reasonably argued that there is no point in exploring for or exploiting podiform deposits further. The title of the paper, however, was in part 'strategic minerals geophysical research'; chromite is considered a strategic and critical mineral of the US and many other industrialized nations. Stratiform deposits in South Africa and Zimbabwe account for most of the chromite reserves in the world. Unfortunately, experience in modern times with substantial price fluctuations and even resource cartels dictates the wisdom of identifying and maintaining a national chromite reserve, irrespective of the cost. An improved exploration methodology is an essential element in this strategic objective. References Albers, J.P., 1983, "Geological and Geophysical Studies of Chromite Deposits in the Josephine peridotite, California and Oregon," US Geological Survey Bulletin 1546 A-D, 118 pp. Dickey, J.S. Jr., 1975, "A Hypothesis of Origin for Podiform Chromite Deposits," Geochimica et Cosmochimica Acta, Vol. 39, pp. 1061.1074. Panayiotou, 1980, Proceedings, International Ophiolite Symposium, Cyprus, 1979, 781 pp. Wynn, J.C., 1980, "Exploration Geophysical Methods and Strategies, and Their use to Locate and Assess Small Mineral Deposits," Chapter 12 of Meyer and Carmen, eds., The Future of Small Scale Mining, McGraw-Hill, New York, pp. 75-83.

Jan 12, 1983

By D. F. Lascelles

D. F. Lascelles (School of Earth Sciences, MacQuarie University, North Ryde, N. S. W., Australia)- The paper contains many misconceptions commonly held by both geologists and soil scientists with regard to soil processes. The analogy between supergene copper enrichment and "nickel laterite" deposits is quite apt and the use of nomenclature from the former in describing supergene nickel deposits is justified. However, care must be exercised with regard to the notable differences in

Jan 1, 1974

By A. F. Agnew

Discussion by A.E Agnew Strachan's paper summarizes a wealth of data on water-storage dam "incidents" published by USCOLD and ASCE in 1988, together with a review of tailings-dam incidents published by USCOLD in 1994. The data is summarized by text and 11 histograms. The histograms present a useful compilation of these data, although the patterns within the histogram blocks did not survive the Mining Engineering printing process - so their usefulness is impaired. Unfortunately, the text discussion is too cursory and does not highlight the significant relationships among the various causes of the incidents. Four paragraphs of text and Fig. 1 were devoted to water-storage dams, but the discussion is too sketchy, and Fig. 1 is unclear. Important statistics that could be uncovered are: •the total number of dam incidents tabulated were 500+, •earth embankments constituted 73% of those in operation and •incidents related to earth embankments were 75%, of which dam failures accounted for 24%, major repairs accounted for 42% and accidents accounted for 34%. Strachan concluded that "the majority of the dam failures are associated with overtopping (p. 49), and that the majority of incidents are associated with spillway and other facility structures" - although this is difficult to discern from Fig. 1. Confusion regarding Strachan's use of the terms "dam failures" and "incidents" is compounded by his use of "failures" and "accidents" in one paragraph of his Introduction section followed by his use of "incidents" in the next paragraph. Getting to the meat of Strachan's important paper, the tailings-dam incident survey (USCOLD, 1994), the connection between Figs 2 through 11 seems obscure. First, in the four paragraphs of introductory material, one learns that "incidents consisted of failures, accidents, seepage and other incidents of performance that did not meet design criteria" (p. 50). One does learn the helpful caveat that "structures not related to mill tailings were not included in the review" - so that the Buffalo Creek (West Virginia in 1972) and Aberfan (Wales in 1966) coal-mine failures, together with the ongoing metal-mine remediation work at Summitville (Colorado) do not muddy the waters of his paper. The all-too-brief discussion of the data shown in Figs. 3 through 11 (185 dams in USCOLD plus 26 dams in UNEP) reveals the following: •Figure 3: 57% of the incidents (failures and accidents) occurred with dams less than 20-m (66-ft) in height. •Figure 4: 63% of the incidents were due to failures alone (subset of Fig. 3). •Figure 5: Active dams had 82% of the accidents and 93% of the failures. •Figure 6: Types of dams (upstream, water-retention, downstream, centerline and "unknown") separated according to incidents. •Figure 7: Breaks out the incidents in Fig. 6 by failures, accidents and ground water. •Figure 8: Breaks out incidents by type (overtopping, slope stability, earthquake, foundation, seepage, structural, erosion, mine subsidence and "unknown"). •Figure 9: Separates Fig. 8 data into active and inactive dams. •Figure 10: Separates Fig. 9 data into failures and accidents. •Figure 11: Same as Fig. 8, but broken out by tailings-dam type (Fig. 6). Figures 6 and 7 show that most of the failures are associated with upstream or "unknown" types. Figure 9 shows that the leading causes of incidents for active dams

Jan 1, 2003

By J. D. Miller, R. J. Smith

Earl C. Herkenhoff Publication of the article "The Flotation Behavior of Digested Asphalt Ridge Tar Sands" by R.J. Smith and J.D. Miller has confirmed and again spotlighted a most significant and important development in the field of tar sand and heavy oil mining. This is the ability to recover 95 % of the bitumen in dry tar sands by a hot water process. The initial investigation by Miller and Sepulveda in early 1978, "Separation of Bitumen from Dry Tar Sands," was the first blockbuter. It is disappointing that there has been so little scientific and industry ackowledgement that the authors laid to rest the long entrenched belief that the sand grains had to be "water wetted" for a hot water process to be applicable. Dr. Miller and his researchers at the University of Utah used slightly different mineral dressing terminology in describing their process but it is essentially the old, established technique of high solids attrition scrubbing at elevated temperatures with a conditioner or modifier, followed by dilution and froth flotation. Similar techniques were described in detail for other nonmetallic minerals in the mid-1940s by Stuart Falconer of American Cyanamid Co. The authors used a 15-minute "scrub" period (or in their terminology "digester step") at 73% solids and 95°C (203°F) in a 0.3 m solution of sodium carbonate -about 4 kg/t (8 lb per st) of tar sand. Flotation was at 10% solids. Obviously, the scrub is an expensive step with respect to energy input but the high recovery of bitumen appears to justify it. Two basic questions then arise: • What are the effects of shorter scrubbing time, lower temperature, and less sodium carbonate? • Would a two-stage approach be advantageous? That is, could you get the major proportion of the bitumen with a mild scrub and apply the high power and temperature only to the refractory portion? These variables should be evaluated. Apparently, all results were only for rougher flotation. What would the grade and recovery of bitumen be if multiple cleaner flotation steps were employed? We must concede that a 60-62% grade of bitumen leaves considerable room for upgrading. Finally, it should be noted by all executives who approve funds for heavy oil recovery schemes, most of which are variations of in-situ methods, that mining can yield a 95% recovery of bitumen. That is correct, 95%. Imagine what that does to the total barrels recovered from a heavy oil resource compared with the scrawny 30-35% maximum recovery we are getting now. The conclusion-let's mine heavy oil and tar sands. Reply by R.J. Smith and J.D. Miller We, too, share Mr. Herkenhoff s enthusiasm for the mining of tar sands-heavy oil and continue the development of related processing technology. Specifically, it will be noted that by appropriate control of bitumen viscosity in the feed (probably the most important variable for successful separation) the recovery can be accomplished at low digestion temperature and low addition of alkali. A 91 t/d (100 stpd) pilot plant based on our technology is now being operated by Enercor in Salt Lake City (Hatfield, Oblad, and Miller, 1982). In these same pilot facilities, rougher concentrates containing more than 80% bitumen on a dry basis are being produced from Asphalt Ridge Feed. However, the processing of certain other tar sands results in incomplete digestion of bitumen/sand aggregates as discussed elsewhere (Miller and Misra, 1982). Such undigested aggregates are shown in Fig. 1. In this regard, it is intended that a recycle stream will be installed in the pilot plant to provide for redigestion of middling products. Further details regarding surface chemistry aspects of this tar sand technology have recently been published (Misra, Aguilar, and Miller, 1981). References Hatfield, K.E., Oblad, A.G., and Miller, J.D., 1982, "Pilot Plant Recovery of Bitumen from Oil Wet Tar Sands," Processing, Second International Conference on Heavy Crude and Tar Sands, Caracas, Venezuela, Feb. 7-17. Hatfield, K.E., Oblad, A.G., and Miller, J.D., 1982, "Pilot Plant Processing of Oil-Wet Tar Sands," Symposium, Rocky Mountain Fuel Society, Salt Lake City, Utah, Feb. 18-19. Miller, J.D. and Misra, M. 1982, "Hot Water Process Development for Utah Tar Sands," Fuel Processing Technology, Vol. 6, No. 1, April, pp. 27-49. Misra, M., Aguilar, R., and Miller, J.D., 1981, "Surface Chemistry Features in the Hot Water Processing of Utah Tar Sand, Proceedings, "Symposium of Separation Science and Technology for Energy Applications," Gatlinburg, Tennessee, May 5-8.

Jan 11, 1982

By Stephen Semeniuk

Working with junior mining companies as a valuer, my experience with regulators is that simplicity works and the sales comparison method is the way to go. What I like about Mr. Ellis?s approach (as in his Table 7) is the integration of commodity prices with the timing of acquisitions and the implication of price changes in gold on the adjustment of individual project valuations. However, the sample gold-related appraisals in Ellis?s paper are pre-2008 transactions and readers should be made aware of the current market for such projects. Before moving on, to be noted is that the effective date of the Las Brisas appraisal (Ellis?s sample Appraisal 3) is Feb. 10, 2006; given that the subject value of Transaction 9, Choco 10 in Ellis?s Table 6 was listed as $353 million for a 95% interest in the con-cession, I am uncomfortable accepting that the acquisition was suddenly worth the adjusted value of $1,205 million (as given in Ellis?s Table 7). In some cases, illiquidity discounts can apply, and commodity-wise, in the one-month period between the acquisition in January 2006 and the adjustment to the Las Brisas conditions as of February 10, 2006, the indicated gold price change was only 41 cents/g ($12.60/oz); i.e., from $17.50/g to $17.91/g ($544.40/oz to $557/oz).

Jan 1, 2012

By L. R. Flint, W. L. Freyberger, W. J. Howarth

W. L. Freyberger (Director, Institute of Mineral Research, Michigan Technological University, Houghton, Mich.)-The authors are to be complimented on their theoretical analysis of a column flotation process, and the comment I wish to make is not related to that analysis, but to the use of the word "countercurrent" to describe that process. Strictly speaking, the process the authors have described is countercurrent only in the sense of particle-bubble contact. That is, as solid particles and air bubbles pass through the column, the leanest flotation feed (the tailing) comes in contact with the greatest concentration of mineral-free bubbles. However, air bubbles alone do not comprise the complete extracting phase in a flotation system. A second component of this extracting phase is the collector, which is really the extractant inducing transfer of mat-

Jan 1, 1972

By S. G. Malghan

W.H. Moyer In reference to the section “Excess pyrite, ” the author discusses difficulties in separating chalcopyrite from pyrite. A technique does exist, and has been used commercially, for the flotation, separation, and recovery, if desired, of both chalcopyrite and pyrite. The method was described in Mining Engineering, March 1966, “An Improved Flotation Method for Cornwall Sulfides. ” While it was not used on a porphyry copper ore, it was used on a chalcopyrite - pyrite-containing ore and might well have application. Both sulfides, chalcopyrite and pyrite, were floated simultaneously using xanthate as a collector and MIBC frother to produce a cleaned sulfide product. This concentrate was then thickened and conditioned at high solids with aeration and lime to a pH above 12.0 for a period of 1.5 to 2 hr. A supplemental suppressant, such as cyanide, could be used if necessary to sharpen the subsequent separation; however, it was not required at Cornwall. The conditioned sulfides were floated with two stages of cleaning to produce a highgrade chalcopyrite froth product and a pyrite sink product barren in copper. The separation at Cornwall gave a 28.6q. Cu grade chalcopyrite product and a pyrite analyzing 0.0670 Cu. A desliming step on the raw ore initially used at Cornwall for a period of time was later discontinued. Advantages to be found in using a strong sulfide collector, such as a xanthate, is the rapid flotation it induces and, generally, an ability to float at a coarser grind. A following regrind circuit could be used to give maximum liberation and produce a clean sulfide concentration. The “differential” separation is concerned primarily with grade and with assuring a barren pyrite for maximum recovery.

Jan 11, 1986

S.L. Barua and Y.C. Kim M. Armstrong's paper is very interesting and can be regarded as a pioneering work of applying geostatistics on washed coal. The density levels used in her study are uniformly spaced, which facilitated curve fitting through a proper experimental design. Most of the existing coal washing plants in the US, however, gather washability data at non-uniform intervals which, in turn, affects the goodness of curve fitting. We would like to emphasize the importance of having a proper experimental design, if a similar geostatistical application is to be conducted on washed coal. We, at the University of Arizona, tried the variogram on recovery (Kim, Barua, and Baafi, 1982, SME preprint 82-143) which showed a similar pattern as Fig. 4 of Armstrong's paper. In both cases, the first experimental variogram value is invariably and inexplicably higher than the second variogram value. Although Armstrong ignores the first point because of "small number of sample points," we did not do so because there were 17 sample pairs. As a consequence, we concluded that the obtained experimental variogram was unsatisfactory and that further research is warranted. In other words, this disagreement in variogram interpretation between M. Armstrong and the University of Arizona begs the question whether the experimental variogram of Fig. 4 truly reflects the spatial correlation in washed coal, assuming there is such a correlation. Finally, we wish that Armstrong could have done a similar study on clean sulfur, which is a vital interest to the coal mines in the US. Her implied remark on sulfur in the concluding section of the paper may turn out to be incorrect because our experimental variograms on clean sulfur at different density levels were even less satisfactory than those on recovery.

Jan 1, 1985

Jan 1, 1985

By A. D. Hughes

C. W. MERRILL*-Mr. Hughes' paper not only is very well presented but is most timely in that it covers a subject of vital interest to the United States. Tin is one of the strategic metals which has not been found here in appreciable quantities. The total United States tin output since the first recorded production over 100 years ago would not supply current consumption for a fortnight. On the consumption side, tin is extremely important in national defense because of its use in tin plate which, in the form of tin cans, makes it possible to feed properly the armies that the United States recruits from its civilian population in emergencies. Moreover, the internal combustion engines so important in powering tanks, planes, and other war machines, require large quantities of tin alloys for bearings and for soldering. These considerations, along with the fact that Malaya, historically, has been the leading source of tin to the United States, make the kind of information that Mr. Hughes has presented so ably, a matter of vital interest.

Jan 1, 1949

This is a most interesting paper and some data in it are corroborated by work I did at Watertown Arsenal some time ago with iron powder. I would like to ask Dr. Wulff to say just a little more about the compressive and tensile stresses on the surface of compacts. JOHN WULFF (author's reply).-Our work has not yet been completed; although enough data have been obtained with cylindrical compacts of iron to permit some conclusions. Briefly, in a compact 3/4 in. high and 1/2 in. in diameter pressed at 48 tons per sq. in., the residual stresses on the surface measured by X-ray diffraction techniques are as follows: The radial stresses on top center and bottom of compacts are practically zero; the longitudinal residual stresses on the side walls are near the top quarter of the compact about seven tenths that of the yield point of mild steel in tension; at the center surface of the compact they are in compression and in magnitude about 0.3 that of mild steel in tension; near the bottom, they are practically zero. F. V. LENEL.?- I was particularly interested in the work that Messrs. Wulff and Seelig reported on the influence of lubrication on the density of parts. In connection with the work of the American Society for Testing Materials on determining the green density of a compact, we ran into considerable difficulty in getting reproducible results of green density determinations. It was much easier to get a reproducible green density when the lubricant was mixed with the powder than when the lubricant was applied to the die walls; particularly when the die was not perfect-i.e., if it was somewhat rough or out of round. In this case, it was almost impossible to get reproducible green densities when the die walls were lubricated, while it was not too difficult to get reproducible

Jan 1, 1946

MR. DICKINSON: There is an increase of dust in machine loading, especially with the McKinlay loading and cutting machine which was extremely dusty. MR. FORRESTER: Declared more dust must be expected because when loading by machinery, the coal is shot harder in order to make it easier for the machine to pick it up. He believes however, that there was less hazard because of the closer supervision possible by reason of the concentration of operation. MR. McAULIFFE: Machine loading more dusty than hand loading. D. HARRINGTON: The hazard is greatly increased by modern methods of loading. Not only do the methods increase the quantity of dust, but they also increase the rapidity of mining and the increased quantity of coal broken down causes large quantities of gas to be released. The number of motors on loading machines and conveyors add to the hazard. No more dangerous places can be found for such motors than at the working face where there is excessive quantities of pure coal dust and where there is more gas than along any other point. Regarding Mr. Herres remarks that the machine loading was safer on account of the concentration and possibility of close supervision, Mr. Harrington said he believed concentration likely to prove a source of danger rather than its preventive. At Dolomite, Alabama, a number of men had assembled, and as a result what was a relatively small explosion killed a large number of men. Concentration is not without its dangers to mine operations. MR. McAULIFFE: These arguments do not appeal to me. Few places are better sprinkled than the working places in a machine loading mine. There is. more effective supervision, and there can be more ventilation. I believe machine loading will increase safety by that very concentration that Mr. Harrington deplores.

Jan 1, 1925

INTRODUCTION Data available on prophyry copper occurrences in the Andean, Caribbean, Appalachian, and Cordilleran orogens may be categorized in a manner to suggest a hypothetical evolution for deposits that occur in the western hemisphere. A classification of de- posits is also suggested. A classification dependent on internal structure of a porphyry copper deposit would lead to two broad categories, breccia pipe and stockwork models. A classification of this type is examined herein but it is not universally accepted in the literature. Modeling based on igneous petrography as well as metasomatism attendant with the sulfide mineralization is more widely accepted by economic geologists whose evaluation of porphyry systems is influenced by the presence or absence of a phyllic zone. Lowell and Guilbert model deposits ordinarily include a phyllic zone, whereas those of the diorite model do not. Position of the phyllic zone in the alteration sequence is a significant ore guide. Should the phyllic zone be absent, dependence on alteration zoning to identify metallization is weakened. Presence or absence of the phyllic zone therefore is a significant parameter in the economic evaluation of a porphyry copper deposit (Hollister. et al., 1975). The Lowell and Guilbert (1970) model commonly includes a phyllic zone in quartz- bearing rocks. Parameters of both calc-alkalic and the quartz monzonite models mentioned elsewhere in this volume are sufficiently close to the Lowell and Guilbert (1970) model that it should be expanded to encompass them. The diorite model remains separate and distinct because deposits of this type are characterized by an absence of quartz in the pluton, commonly highly gold: copper ratios, low molybdenum content, and regular absence of sericite in the alteration zones. Therefore, for the benefit of those

Jan 1, 1978

BENEDICT SHUBART: I am rather surprised at one statement of Mr. McKeehan, that he found it satisfactory to bond only one track. It has always been my idea of good practice to bond both tracks and cross bond frequently, so that in case of any break in the bonding of one track, which frequently occurs, there will be no interruption in the circuit.

Jan 1, 1923

By David A. Zegeer

Since 1980, more than 40 miners have lost their lives by being trapped and suffocating in stock-piles, bins, silos, and similar facilities. Mine Safety and Health Administration (MSHA) is concerned about these accidents, most of which could be prevented if management and miners appreciated the hazard. In February 1986, the coal industry had a suffocation accident that drew national attention. A group of engineers were standing on a coal stockpile. It gave way under their feet. Five of the engineers were buried in the coal and suffocated. As with all fatal accidents, MSHA investigated the cause of this tragedy. I also asked my staff to look into the history of similar accidents. I found out it is rare to have multiple fatalities due to suffocation in material. But, over the years, there has been a long series of individual fatalities from this type of accident. Here are some examples. In Oct. 1985, a coal mine employee in Virginia entered a hopper to collect a coal sample. The feeders were started while he was standing in the hopper. He sank into the coal and suffocated. In Sept. 1985, an employee at a Texas sand company climbed onto a surge pile, to check a problem with sand flow. When other employees found him, the sand had swallowed him. Only his face and one hand were showing. Though the victim's face was clear of the sand, he suffocated before rescuers could free him. Many people do not realize you can suffocate in this kind of situation, though your head is free. Every time you exhale, the material has room to shift and press harder against your chest. Then you have less room to inhale. If this goes on long enough, it gets to the point where you cannot inhale at all. That is what happened to the victim in this accident. In May 1985, a plant operator at a Wisconsin sand and gravel operation purposely slid into the drawhole of a surge pile. He tried to break up large rocks that were clogging the drawpoint. Material piled on the side of the surge pile slid down and buried him. The operator was wearing a safety belt and line. But he was covered by 3 to 3.7 m (10 to 12 ft) of material and could not be pulled free. This incident illustrates another point. A safety line can help. But it is not enough. Workers should never put themselves in a position where loose material can fall on them from above. In Jan. 1985, a young Utah coal miner was waiting for a ride home from a co-worker whose job was at the tipple. A jam developed in the coal bin. The men could not clear the jam from the outside. So the young miner jumped into the bin. He started digging with a shovel at the toe of the coal pile. It caved in on him and he suffocated. The young miner had only three months of mining experience. He had not been trained for work around the coal bin. Since 1980, 16 coal miners and 26 nonmetal miners have lost their lives in loose-material suffocation accidents. Suffocation accidents occur in many areas around mines - bins, hoppers, stockpiles, surge piles, and silos, among others. One recent accident occurred at a surface coal mine in Illinois. The victim was walking over an area that had been drilled and blasted the week before. He fell into a void that had been crusted over with earth. He too suffocated. And many materials have been involved in suffocation accidentscoal, crushed stone, sand, salt, different kinds of ore, ammonium nitrate pellets used in blasting, and a powdery byproduct of iron ore milling called calcine. In addition to fatal accidents, there are near misses. We do not know how many. Most of them are not reported. A person who gets trapped in material, but is rescued, is usually unhurt. MSHA sometimes hears about these accidents by chance, or because they are reported in the local media. We do know that near misses are happening far more often

Jan 1, 1987