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By S. J. Kidder, G. C. McCartney

The Helen Mine, of the Algoma Steel Corporation, in the Michipicoten district, Ontario, Canada, has produced more than 6,240,290 tons of iron ore. Prior to and during World War I, 2,823,369 gross tons of brown ore were shipped, and from 1939 to the end of the 1944 season, 3,416,921 gross tons of siderite ore were produced from open-pit operations. Brief descriptions are given of the nature of the ore, the mining methods employed and the manner of treatment of the ore. Shipment of the finished product, known as Algoma sinter, is made to blast furnaces in both Canada and the United States. The iron-ore bodies—i.e., the oxide ores of the Old Helen and the siderite of the New Helen—are a part of a tabular shaped area of iron formation that extends over a length of more than 100,000 ft. and has a maximum width of 1000 ft. The ores are considered to be a product of introduction of iron-bearing material accompanied by replacement of the rocks already present. It is shown that introduction took place subsequent to the development of the major folded structures. It is believed the iron-bearing material emanated from an igneous source. Introduction The Michipicoten mining district derives its name from the Indian word1 meaning "place of bold promontories" or "region of big places," and anyone who has visited this part of Canada will agree that the district is well named. In 1897, there was active prospecting for gold in the district and the Ontario Bureau of Mines2 opened a recording office at the Hudson's Bay post at Michipicoten River. In the summer of 1897, Ben Boyer, probably searching for gold, discovered an outcrop of brown iron ore at the east end of a small lake, less than 8 miles, in an air line, northeast of the nearest point on Michipicoten Harbour. In the following year, extensive surface exploration and diamond drilling disclosed a commercial body of nonbessemer brown ore, possibly 700 ft. long and 200 ft. wide, of about the following analysis:3 Fe (natural), 57.15 per cent; S, dried at 212°F.. 0.113; SiO2, dried at 212°F., 4.00; P, dried at 212°F., 0.125. The property (Figs. 1-3) was taken over by E. V. Clergue and ultimately was transferred to the Algoma Steel Corporation. The Algoma Central Railway shortly afterward built its line from Michipicoten Harbour to the mine, together with an ore dock at the Harbour. Thus, within three years of its discovery, the mine had adequate shipping facilities and a market for its ore. Production of Brown Ore and Pyrite The first shipment of brown ore was made in July 1930. By 1939, the mine was shipping at the rate of 170,000 long tons

Jan 1, 1949

By John S. Brown

IN 1925 the writer undertook a detailed statistical study of all producing areas in the Joplin district as a basis for evaluating programs and measuring objectives. For this purpose, the published figures in the yearly volumes of Mineral Resources were used, supplemented for earlier years by publications of the Missouri Geological Survey and other local and less official sources. When all else failed, the available data were projected backward to hazard a reasonable guess as to the unrecorded early output of important areas. Fortunately, the proportion of such prehistory production is not a large factor in any of the totals. These results were used during the next few years to measure the relative importance of various producing areas and to predict the peak period of development of the all-important Picher field. For the purpose of this review, the charts have been completed to the end of 1950. During World War 11, the U. S. Bureau of Mines became interested in a similar study and issued comprehensive statistical tabulations of data up to 1945 ( Info. Circular 7383), which have been checked against the figures used herein. This tabulation, however, does not include all the earlier data used by the writer nor does it offer any estimates of the wholly unrecorded era in the beginnings of the earlier camps. The area covered in this study is shown in Fig. 1 on which are indicated the relative location and approximate outlines of the principal producing camps. This also shows the approximate yield to date of each major camp in terms of combined lead and zinc concentrates. The output of zinc concentrates is roughly seven times that of lead. Hence, the economy of the district has depended primarily on the price of zinc, with lead as an important byproduct. Over much of the productive period, lead concentrates averaged about twice the value of zinc concentrates per ton, and in certain mines or areas the proportion of lead to zinc was substantially above average. The Joplin district is largely flat prairie but is partly moderately dissected, partially wooded land with a relief generally less than 100 ft. The rocks are almost flat-lying, nearly parallel to the surface, and the chief ore formation is the Mississippian Boone limestone, including its cherty phases. This formation either outcrops in the producing areas or is covered by a thin veneer of Pennsylvanian shales. Virtually all the ore occurs within 400 ft of the surface, and a large part at less than 300 ft in depth. Most of the land was divided into small farms or town lots before mineral development; tracts seldom exceeded 160 acres, and averaged considerably less. Mineral rights followed the surface ownership, segregation was rare, and a system of leasing for mineral development became well established early in the region's history, many landowners deriving small to sizable fortunes from royalties. Because of the shal-lowness of the ore and other factors, prospecting and mining was cheaper than in almost any comparable mining district in the United States. This situation, coupled with the widely divided land ownership, offered a fertile field for promoters and speculators and led to the rise of many small mining concerns. Only in its later history, under stern economic compulsion, has control tended to centralize in a few companies. Under these conditions, any important new discovery or successful development had much the effect of a gold rush or an oil boom. Every property in the area was leased quickly, promptly drilled, and, if ore was found, it was soon on the market. Many companies and individuals participated, and the average producing lease-hold probably was about 40 acres in extent. Any important field thus was attacked by anywhere from 10 to 100 or more producers. Production zoomed, eventually steadied or wavered, and ultimately subsided, leaving a desolation of tailings mountains, cave-ins, empty housing, and wreckage. The object of this paper is to depict the pattern of this process, so far as metal production is concerned, and to note the way in which it reacted to economic and political pressures. Production Charts In Fig. 2 is charted the production record, in tons of lead and zinc concentrates combined, of eight of the principal camps, which together account for approximately 99 pct of the total district production, over the years from 1870 to 1950. This period covers all but the very minor beginning of mining history. Two important camps are divided by state lines; hence, it has been necessary to combine production records for the two portions, based on estimates that may be slightly in error. Certain camps are sub-dividable into important units for which separate figures are available in whole or in part and have been charted as fractions of the major unit. The corresponding price of zinc is shown above all the charts. Three camps, Aurora, Neck City, and Galena, show a remarkably symmetrical graphic pattern, which is interpreted as the norm. The curves rise steeply to a peak, level off for an irregular interval, and then drop sharply to zero on a slope corresponding roughly to that covered by the initial rise. The three portions of these charts seem appropriately characterized by the designations of youth, maturity, and decline. On the whole, with some irregularities, the production in each of the three periods seems to be almost equal. A fourth camp, Granby, fails to conform to the normal pattern. It exhibits a very long period of reasonably uniform, stabilized production corresponding to maturity, followed by a rather precipitate decline. Its youth is hidden in the era of prehistory. This habit of steady, long-continued production at an even keel is attributable to the fact that this camp, more than any other, was controlled largely by a single principal owner at any given period over most of its history and this permitted the imposition

Jan 1, 1952

By M. L. Brashears

IN the last 15 years industrial use of ground water has more than doubled, and in 1951 amounted to 5 billion gallons per day. A similar sharp increase in the utilization of ground water for irrigation and public-water supply occurred in the same period. In many areas rapid increase in withdrawal from wells has taken place almost entirely unhampered by regulatory control and with little or no integration of effort. As might be expected, the chief interest in many regions has been maximum production rather than sustained perennial yield. As a result, widespread depletion of underground reservoirs and deterioration of the quality of the water stored in them has taken place in many areas, even though total pumpage in the United States is far below ultimate potential. Of even more concern is the fact that excessive withdrawal has drawn salt water into the reservoirs beneath many heavily populated centers along the Atlantic, Gulf, and Pacific coasts, causing costly abandonment of pumping plants.' Many hydrologists expect that consumption of water will rise rapidly in the near future, and some predict that industrial requirements will more than double in the next decade.2,3 Thus it appears likely that the draft on many already heavily pumped underground reservoirs will be greatly increased and the search for additional sources of usable ground water intensified in years to come. In view of this, industry as a whole will be forced more and more to recognize the potentialities and limitations of ground-water reservoirs and to utilize them more effectively to prevent costly water shortages and disruption of production. Through painful experience, some industries are already well aware of the need for effective water utilization, and have managed individually or through joint effort to check trends threatening to deplete underground reservoirs completely or to impair the quality of the water. Various remedial measures have been used to bring about successful management of local or regional ground-water resources. Of these, replen-ishment of aquifers by recharge wells or basins has played an important role in overcoming some ground water problems. Artificial recharge of underground reservoirs by water spreading has been practiced successfully in the United States for many years. In the West it has become an important method of salvaging flood run-off for irrigation of crops and maintenance of public water-supply reserves, and it is used to some extent in parts of the East. Artificial recharge by means of wells, on the other hand, is a relatively new development. Until recently it was employed in only a few areas, principally along the East coast. For the last few years, in the ever increasing search for additional water supplies, industry has had greater recourse to this method. Utilization of recharge wells to control the temperature and quality of underground water supplies is also being considered seriously. Operation of recharge wells, like water spreading, is governed largely by local conditions. It requires water relatively low in turbidity, whereas in some areas water spreading has been used successfully with water of high turbidity and silt content. However, water spreading must be employed in large areas and can be carried on effectively only where aquifers crop out at the surface. Recharge wells can be used in limited space. Recharge wells are similar to production wells except that the water flows in the opposite direction. Thus any water-bearing bed that will yield water to wells may be recharged by wells. Often, however, the water available for recharge is of a different character and temperature from that existing in the ground-water reservoir and if transmitted directly underground from a recharge well to a production well might require expensive or difficult treatment before it could be used. Fortunately the physical characteristics of reservoir beds, which control the movement and behavior of ground water, are generally not homogeneous. Moreover, the movement of ground water is very slow because of the frictional resistance of the reservoir beds. By taking full advantage of hydrologic and geologic conditions, it is therefore possible in many instances to bring about favorable changes of temperature and dilution as the water moves from the recharge wells underground to the production wells. Furthermore, if the natural quality or temperature of ground water is unfavorable for industrial purposes, recharge wells may be used to introduce water of more favorable quality or temperature into the ground-water reservoir. When water is discharged into a recharge well, the head in the well is increased. Because of this, a cone of elevation is produced on the water table or the artesian pressure surface in the area surrounding the well. The cone of elevation is similar to the cone of depression produced around a pumping well except that the apex of the cone is above the water table or artesian pressure surface. Thus if a recharge well and a production well tapping the same water-bearing bed are close together, as would be the case at many industrial plants, some of the water discharged from the recharge well would be drawn into the production well within a short time. Under such conditions it is apparent that water of unfavorable temperature and chemical characteristics should not be used for recharging. The more important ground-water reservoirs in the United States often consist of alternating layers of impermeable beds and porous material that will yield water readily to wells. Physical characteristics of individual beds in a ground-water reservoir may not persist over great distances, the impermeable layers grading into beds that will yield large quantities of water. Thus the water-yielding material in underground reservoirs, whether large or small,

Jan 1, 1954

By E. T. Turkdogan, L. J. Martonik, R. J. Fruehan

Electrochemical measuretnents of the solid oxide electrolyte galvanic cells CY-Cr2O3 I ZrO2 (CaO) 1 O (in Fe alloy) CY-Cr2O3 I Tho2 (Y2O3)I O en Fe alloy) have been made at 1600°C (2912°F) in order to test the Performance of such cells at liquid steel temperatures. The oxygen pvobe (cell) consists of a disk of ZrO2 (CaO) or Tho2 (Y2O3) electrolyte fused at one end of a silica tube filled with a mixture of Cr-Cr2O3 which is the reference electrode. Upon immersion in liquid steel, the electromotive force readings achieve a steady value within a few seconds, and remain steady for 30 win or more. The perforwzance of the probes has been tested using Fe-O, Fe-Si-O, Fe-Cr-O, Fe-V-O, and Fe-Al-O alloys; the oxygen contents of liquid steel derived from the measured electromotive forces are in satisfactory agreement with those determined by arulysis. Use of the probe in the deoxi-datiorz of steel, in laboratory experiments, is discussed. The results indicate that there is insignificant electronic conductivity in ZrO2(CuO) at oxygen activities down to those corresponding to 10 ppm in steel. At lower oxygen activities, probes tipped with ThOn (Y2O3) disks perform satisfactorily at oxygen activities down to 1 ppm O or less. THE key to the control of deoxidation of steel is a sensing device to measure rapidly the concentration of oxygen in liquid steel in the furnace, ladle or tun-dish at any desired stage of deoxidation. The analysis of the cast steel by the neutron-activation or vacuum-fusion method gives total oxygen as oxide and silicate inclusions. This analysis is important for guidance to steel cleanliness; however, such a postmortem is of little value in the control of deoxidation of liquid steel. At the General Meeting of the American Iron and Steel Institute in New York, 1968, Turkdogan and Fruehan' presented a paper on the preliminary results of the work done in this laboratory on rapid determination of oxygen in steel by an oxygen probe. Details of the work done in this laboratory leading to the development of a galvanic cell for the determination of oxygen in liquid steel, and the results of the tests made are given in this paper. It was through Wagner's contributions, since the early Thirties, to the physical chemistry of semiconductors in general that it ultimately became possible to construct galvanic cells for application at high temperatures. In 1957, Kiukkola and wagner2 successfully demonstrated the use of several solid electrolytes in measuring the free energies of several chemical reactions, in particular, the use of lime-stabilized zir-conia in high-temperature oxidation reactions. Starting 7 years later, a number of papers appeared in the technical literature3-' demonstrating possible applicability of galvanic cells for the determination of oxygen in liquid steel. In the earliest work, Japanese investigators3j4 experimented with various types of reference electrodes, e.g., graphite-saturated liquid iron at 1 atm CO or Ni-NiO mixtures; the results obtained, though promising, were not of sufficient accuracy. Except for the work of Baker and West,6 all other investigators5,7,8 showed that ZrO2(CaO) electrolyte could be used for this purpose. The main part of the galvanic cell used by Fischer and ~ckermann' and by schwerdtfeger7 (the latter work was done in this laboratory), consisted of a ZrO2(CaO) tube, -1 cm ID, closed at one end, with a platinum contact wire fixed mechanically inside the closed end. The tube was flushed with a gas of known oxygen partial pressure, e.g., air, CO-CO2 or H2-CO2 mixtures; gas along with the platinum lead wire served as the reference electrode. The oxygen contents derived from measured electromotive forces agreed reasonably well with the oxygen contents determined by vacuum-fusion analysis. It is evident from recent investigations that the electromotive force technique using a solid oxide electrolyte is fundamentally well suited for the determination of oxygen in liquid steel. However, it is equally clear that the cell arrangement of the type as commonly used is in need of considerable improvement, as it exhibits several shortcomings for industrial and even laboratory use. 1) Because of its size, the zirconia tube, though stabilized, has a poor resistance to thermal shock. 2) Fine pores and microcracks, which are invariably present in zirconia tubes, are detrimental to the satisfactory operation of the cell, particularly when gas reference electrodes are used. 3) Air or carbon dioxide reference electrodes give rise to high electromotive force readings; as a result, the determination of oxygen in steel becomes less accurate. For higher accuracy, the oxygen partial pressure of the reference electrode should be in the range similar to that of oxygen in steel. 4) Even in laboratory experiments, difficulties are experienced when flushing the tube with gases and maintaining the proper gas flow rate. Fischer and Ackermann,' who used air as the reference electrode, reported that when the flow rate was too low, furnace gases would leak into the electrolyte tube, therefore lowering the oxygen potential and measured electromotive force. The required flow rate in order to avoid leakage depended on the tightness of the electrolyte tube which varied with different tubes, thus making it difficult to predict in advance the required flow rate. However, if the flow rate is too high the inside wall of the electrolyte tube would be cooler than the wall

Jan 1, 1970

By S. Horton, F. L. Stanley, P. J. Bethell

Introduction The McClure River Processing Plant is located at McClure in Dickenson County, VA. The plant was originally constructed in 1979 as a 600 st per hour facility. The plant is fed by coals from numerous mines in the area as well as a major longwall mine onsite. Coal is processed predominantly for the metallurgical coal market. The original plant was comprised of three separate processing circuits incorporating a heavy medium vessel for the coarse material (+10 mm). Intermediate sized material (10 x 0.6 mm) was beneficiated in heavy medium cyclones, with the remaining fines (0.6 mm x 0) passing to froth flotation for upgrading. The products from all three circuits were combined prior to thermal drying and then stored in silos or ground storage prior to shipment to the metallurgical coal markets. A middlings heavy medium cyclone circuit was also present to rewash the vessel and primary heavy medium cyclone rejects. The original circuit is as shown as Fig. 1 (Editor's note: All figures are reproduced at the end of this paper.) In reasonably close proximity to the McClure Plant is the much older, less efficient Moss 1 Plant, (Chance cone, table, water only cyclone, froth flotation, Fig. 2), which prior to the McClure Plant expansion was processing considerable tonnages of coal. At that time, the McClure River Plant was processing coal at the maximum possible feed rate, and the Moss 1 Plant was finding increasing difficulty in maintaining the stringent quality specifications on its major metallurgical coal accounts. Given the fact that many of the feeder coals to Moss 1 could be moved with minimal difficulty and cost to McClure, an exercise was conducted to evaluate the potential of expanding the McClure Plant. This would result in transferring coals from Moss 1, thereby increasing yield, reducing operating costs, and enabling customer specifications to be met more easily. After evaluating many possible ways of expanding the McClure Plant, we eventually settled on an approach based on the idea of Gallagher at German Creek in Australia (Gallagher, 1986). Our objective was to increase plant capacity by removing the coarser size fractions (0.60 x 0.15 mm) from froth flotation and at the same time eliminating the finer fractions (1.25 x 0.6 mm) from the heavy medium cyclone circuit, thereby allowing these circuits and therefore the plant as a whole to increase its capacity. These size fractions (0.60 x 0.15 mm from froth flotation and 1.25 x 0.6 mm from the heavy medium cyclone circuit) would be treated in a new spiral circuit. After desired evaluation of the various circuit capacities, including desliming screen, clean coal filter and thermal dryer capacities, etc., it was concluded that we should be able to maintain high levels of processing efficiency in the existing circuitry while at the same time expanding plant feed rate from a nominal 600 st per hour to 1000 st per hour, thereby achieving the objectives The circuit that was finally settled on is shown in Fig. 2. Several unique features were incorporated in this circuit in an attempt to maximize processing efficiency and reduce maintenance costs. First, the idea of a single-stage spiral circuit was abandoned in favor of a two-stage circuit incorporating rewashing the primary spiral middlings in a secondary spiral (Bethell, 88). The spirals selected were Mineral Deposits Model LD4 units. It was felt, after much previous testing, that spiral middlings rather than being a true middling gravity fraction were in fact a combination of true middlings with a preponderance of misplaced light coal and high-density rock. Consequently the insertion of this material directly into either clean coal or refuse streams (per normal practice) would only lead to loss of efficiency. The operating philosophy behind the design of the circuit was to produce the purest coal and rejects possible by adjusting the spiral splitters. Misplacement of refuse into the clean coal and vice versa would be minimized by taking a fairly large middlings cut. By reprocessing this material (10%-15% of primary spiral feed) in a secondary bank of spirals the misplaced material in the primary middlings would be reprocessed and its probability of reporting to the correct stream greatly enhanced. To minimize fines loading on the spirals, fine aperture sieve bends were to be included. A further novel feature of the circuitry is the desliming/ dewatering system. A philosophy of accepting inefficiency but forcing it to work in our favor was developed. The dewatering screens (Linatex Combis) were chosen not for their ultimate screening efficiency but for their ability to remove slimes (tolerating a fair quantity of oversize in the effluent) and their enhanced maintenance characteristics. Any oversize passing into the effluent, together with the fine coal centrifuge effluent, would pass to classifying cyclones, the objective of which was to reject to overflow (and hence froth flotation) as many fines as possible (-150 µm) without running the risk of losing coarse (1.25 x 0.15 mm) material to oversize. Hence, by a process of producing a "slime free" screen product coupled with a "coarse free" cyclone overflow, sizing efficiency could be maximized. Obviously, recirculating load had to be factored into the Combi Screen design capacity. The spiral circuit as shown in Fig. 2 was installed at McClure River in January 1988. The spiral circuit efficiency was very high immediately, and, after the normal debugging, plant capacity has been increased to the point at which we are currently processing of the order of 1100 st per hour. Test program The performance of the spiral circuit has been evaluated by two major samplings of the entire new circuit, coupled with routine bi-monthly tests on all other major components of the plant circuitry.

Jan 1, 1992

By B. H. Caudle, L. G. Davies, I. H. Silberberg

This paper presents a method of predicting the recovery and performance of a five-spot steam injection project, in which a realistic approach to pattern sweepout efficiencies is made. Published methods for radial systems were modified for the five-spot pattern by approximating the stream lines with straight lines radiating from the injection well and then converging to the producing well. In each radial segment, the position of the steam front and the temperature profile ahead of the steam front were determined by heat balance equations, which included an estimation of heat losses to surrounding formations. The location of the saturations behind the cold water front was determined from a Buckley-Leverett solution to the material balance equation. Results from this program show steamflood recovery in a five-spot pattern to be considerably less than that predicted for true linear or radial flow systems. For a specific reservoir containing 900-cp oil, a steamflood in a purely radial flow system was predicted to recover more than 75 percent of the original oil in place when 2 PV of water had been injected as steam. A five-spot steamflood with otherwise identical properties was predicted to recover 10 percent of the original oil in place when 0.15 PV of water had been injected as steam and to recover almost no oil thereafter. A cold water five-spot flood in this system was predicted to recover approximately W percent of the oil in place with I PV of water injected. For a five-spot pattern in an example reservoir with 10-cp oil, steam injection similarly showed lower ultimate recovery than water injection but no improvement in recovery rule. Introduction The thermal recovery method considered in this study is steam injection in a five-spot pattern. Pattern steam injection has been given little attention in the past, possibly because of some rather obvious disadvantages. Unless steam temperature is maintained throughout the entire swept area, the process will revert to simple waterflood with all heat being lost prior to reaching the oil bank. Also, the oil viscosity reduction takes place near the steam front and not around the producing wellbore, so that low producing rates must still be endured. This is the reason for the success of the steam stimulation, or cyclic injection, in which the fluids are produced from the same well used for injection. On the other hand, steamflooding can offer some advantages. Oil displacement is by four methods: (1) mechanical displacement by the condensed water, (2) viscosity reduction of the crude oil, (3) swelling of the crude oil, and (4) distillation of the crude oil in the steam zone. Although dependent upon the crude, laboratory experiments have shown displacements up to 80 percent by steamflooding. In addition, steam is a good heat transport medium, since it is cheap and has a high heat content. Previous Investigations Previous publications have reported methods for predicting recovery in a steamflood for linear and radial systems. Since no pattern sweep efficiency is taken into consideration, the recovery even from a radial system must be greater than from more realistic geometries such as the very common five-spot patterns. Probably the broadest coverage is given by Willman who reported experimental results and offered a method of predicting recovery for a radial flow system. They concluded that both hot water and steam injection recover more oil than an ordinary waterflood, and that steam injection could yield recoveries "as much as 100 percent greater than by water flood." They also concluded that both the heat requirements for a reservoir and the residual oil remaining in the reservoir after steam injection were independent of the amount of oil originally in place, that short exploitation times were desirable, and that a high percentage of net sand in the reservoir with a high initial oil saturation was desirable. The method assumes that the flood occurs in three concentric cylindrical zones: (1) an inner steam zone, (2) a central hot water zone, and (3) an outer cold waterflood zone. Displacement in the steam zone is based on laboratory-determined residual oil saturations while the hot water and cold water zones use the conventional Buckley-Leverett equations. Although the results shown by Willman et al. are for a radial system flowing out from the well to an assumed external circular boundary, their equations did

Jan 1, 1969

By D. A. Munro, G. P. Contractor

Fourteen 50-lb laboratory melts were investigated to determine the effect of uranium on the tenpering characteristics of loo-carbon (0.06 to 0.1 pct C) steels. It was found that uranium additions, particularly in the range 0.30 to 0.45 pct, enhanced the hardness and both ultimate and yield strength of the experivzental steels in the quenched and tempered condition. The structural and morphological chazges indicated that uranium retarded tempering of the tnartensite, thereby hindering the normal formation of polygonal ferrite formed in the late stages of tempering. The effect of this was to make possible the re-tension of the acicilar ferritic structure in the uranium-bearing' steels. The iraniuin-bearing steels also showed IVidnzanstatten-type growth of ferrite plates and had large prior austenite grains containing assenzblies of fine ferrite grains, mainly acicular in geometry. The fine-grained ferrite structure and the presence of more numerous and apparently smaller precipitates in the uranium-bearing steels are thought to he principally responsible for the itnproved tensile strength and hardness of the experinzental uranium-bearing steels. At ternperirzg temperatures above 455% (850'F) the ferrite in the higher-uraniun steels nzaintained acicularity and, hence, its strength and resistance to tempering. Uranium did not produce a secondary hardening peak. However, it retarded softening during the third stage of tempering because of its effect of inhibiting the grouth of cementite particles and of retaining the acicularity of ferrite plates. The resistance to coalescence accounted for the slow grocth of the ferrite grains in the uranium-modified steels and, hence, fov the persistence of the acicular ferrite structure. IT had been found previously1 that uranium additions up to about 0.45 pct had no significant effect on the tensile properties of low-carbon steel (0.06 to 0.10 pct C) in the as-rolled and normalized conditions, Fig. 1. On the other hand, it was observed that uranium in excess of about 0.30 pct had an embrittling effect as revealed by Charpy V-notch impact results. It was also noted that, as the uranium content increased, the morphology of pearlite changed from lamellar to feathery and the ferrite grains showed an etching effect resembling striated or dashed markings, suggestive of precipitation. The sharp drop in the impact properties shown in Fig. 2 warranted an assumption that the uranium content of about 0.30 to 0.45 pct may produce some secondary hardening reaction on tempering, analogous to that associated with a Cr-Mo-V steel, which shows very poor CVN toughness at the secondary hardness peak in the tempering curve.1' With this background and the reported findings of Hasegawa and noda that low-carbon uranium-treated steel showed signs of secondary hardening, the present investigation was undertaken to determine the effect of uranium additions on the mechanical properties of 0.10 pct C steels. No attempts were made to investigate in detail the mechanisms of hardening, although some suggestions based on the experiments are made. MATERIALS AND PROCEDURES A series of 50-lb induction-furnace melts was made using AISI 1008 rimming steel billets as the melting stock. The melting, forging, and rolling techniques proven satisfactory in previous projects'-3 were employed as a guide for this investigation. The steel was deoxidized with aluminum (2 lb per ton) prior to the addition of high-purity uranium. The analysis of each melt is given in Table I. Properties were evaluated as a function of heat treatment and are presented in terms of hardness and tensile strength vs tempering temperatures. The variation of hardness with the tempering temperature was studied on the quenched and tempered specimens, some of which measured 0.50 by 0.25 in. diam and the others 0.40-in. cubes. Before quenching, the specimens were vacuum-sealed in glass tubes and normalized at 900°C (1650°F) for 20 min. Following this treatment, the sealed specimens were hardened by austenitizing at 955°C (1750°F) for 20 min and water quenching, and then tempered for 1 hr in the range 150 to 730°C

Jan 1, 1967

By Evans B. Mayo

This study examines the structural framework of the Southwest for evidence of the four principal trends of lineament tectonics. It attempts to classify their intersections and compares the positions of those trends that appear most favorable with the positions of the presently known mining districts. This is a controversial topic. The author, in presenting his analysis, is aware that the study of lineament tectonics and relation of ore districts to regional structure is complicated by insufficient data and, unavoidably, by personal bias. The development of lineament tectonics has been summarized by Umbgrove.' Early attempts to fit ore districts into the Cordilleran framework were made by Billingsley and Locke, who do not refer to lineament tectonics, although their appoach is similar; they observe that heat and fluids, including the ore-depositing fluids, are most likely to rise at or near intersections of major structures where the crust is fractured, or weakened, to great depth. As a result of studies distributed over the earth-including ocean basins as well as continents— some tectonists recognize four dominant structural trends: 1) northwest; 2) northeast; 3) nearly east-west, or equatorial; and 4) nearly north-south, or meridional. Baker' proposed a theory to account for these trends and Sonder" called their world-wide arrangement the regmatic shear pattern. Moody and Hill proposed a much more complicated shear network which, although fascinating and perhaps ultimately useful, will not be followed here. In a recent review of deformation within the Cordillera, Wisser mentioned the four fundamental directions. The fact that many geologists deny the existence of the regmatic shear pattern implies that the fundamental structures are far from obvious. It may mean, also, that some geologists are not accustomed to examine regional and world maps analytically. The maps require much study, and certain features should be isolated on overlays. Even so, with the present limited knowledge, uncertainties remain. The following analysis is a qualitative experiment, subject to change as information accumulates, and should be supplemented by the western sheet of the Tectonic Map of the United States.' Many have probably gained the impression that the Cordillera of the West is oriented northwest-southeast and from this, unless experience rules otherwise, it is natural to assume that the structure likewise trends northwest-southeast. To an important extent this is true, yet anyone tracing off the western sheet of the Tectonic Map all the recorded north west-southeast structures may be surprised to see what a small part of the entire area they occupy. In southwestern U. S. (Fig. 1) the northwest-southeast structures are mostly restricted to the eastern, southern, and western parts. The eastern margin of the Cordillera from northern Colorado to the International Boundary near El Paso, Tex., is obviously determined by some structure other than northwesterly ones. In eastern Nevada and far southward toward the Gulf of California many mountain ranges, valleys, and faults are meridional. In the Rocky Mts. of Colorado, and at many places in southern Arizona, the crystalline Pre-Cambrian is foliated northeast-southwest. The Uinta Mts. of Utah trend approximately east-west, in much the same way as a broad belt of transverse, west-northwest structures—the Texas lineament of Hill",' and Ransome10 in southern Arizona, southwestern New Mexico, and southern California. It seems, then that there are four regmatic shear directions in the Southwest, but at many places they are discontinuous, and their projections must be inferred. To clarify these trends the four sets have been isolated into two systems: 1) northwest-northeast and 2) east-west-north-south (Figs. 2 and 3). Northwest-Northeast System: A number of prominent northwest-trending zones of structure are easily recognized. They are designated by circled Roman numerals, the northeast-trending structures by circled capital letters. Perhaps no two geologists would agree completely on the positions of all these belts. Names given below are for convenience only, and may be discarded where other names have priority. (I) The Sierra Nevada-Lower California belt contains the Jurassic-Cretaceous granitic massifs of the Sierra Nevada and Lower California. These

Jan 1, 1959

By C. S. Matthews, M. J. F. Rosenbaum

The purpose of thir work wax to lcarn it~lzut infori~lation could he obtained from various typs of pilot water floods and to attempt to find the optunum pilot patter11, for a revervoir which had previously been depleted by a solution gas drive. The study was made in the laboratory with mathemetical methods a dynamic analog and a potentiotnetric analog. Results werp tested against the field llistorics of a nrrnlber of pilot water floods. At a reasonable valrre of currzulative injection, the total production rate for the one-injector five-spot should reach about 6.5 per cent of injection rate, and for a four-injector five-spot, about 9 per cent. Accurate estimates of ultimate recovery cannot be made on the basis of such snzall prorluction rates. However, with a pilot composed of nine ir1jector.s and 16 producers the production rate is approximately 50 pcr cent of injection rate at a reasonable value of camulative injection. Sonle inforn~ation for extended performance predictions might he obtained from such a large pilot. These conclusions were drawn on the basis of results obtained for unit mobility ratio, and a sturly using tlre potentiometric analog was made of the effect of other mobility ratios to determine the range of applicability of these predictions. For the four-injector, five-spot pilot with the ratio of production to injection rate (before water breakthrough) is about twice that for with it is about two-thirds; and with M0= 10, it is about one-third For high mobility ratios, it was found that the production rate increased considerably as water-cut increased. These result can be used to modify, qualitatively, the inter.pretntions based on curves for the unit rnobilit\. ratio CaSeS. It was found that the maximum ratio of production rate to injection rate obseriled in field pilot floods was of rhe scime order as that prerdicted by these methods. The time required to reach thisr maximum did not generally agree with the time predicted for a homogeti~orir reservoir. The differcrlce between predicted and observed time of response gives an indication of the permeability profile and of the condition of the producin,g wells. Pilot water floods of the pattern type are generally carried out in reservoirs which have been depleted by solution gas drive and are at low pressure. Under these conditions, oil and water can be considered incompressible. It is assumed that, as the water is injected, an oil bank forms ahead of it and that there is a distinct interface between the water zone (or bank) and the oil zone (or bank) and between the oil zone and the region ahead of the oil zone. It is further assumed that only gas is mobile in the unflooded (gas) region, only oil is mobile in the oil bank and only water is mobile in the water bank. The saturations and the mobilities associated with each zone are assumed uniform. We idealize our reservoir to be homogeneous, horizontal and of constant thickness. Effects of gravity within the producing layer are assumed negligible. If the actual time-dcpendent flow problem is approximated by a acries of steady-state problems. the potential and stream function in the oil bank and water hank satisfy Laplace&apos;s equation in two dimensions. We can therefore use a poteiitiometric analog of this system. Potentiometric models have yielded uscful results in this laboratory&apos; and clsewhere in the study of a variety of secondary recovery problems. For the case where M = I, we generally prefer to use theoretical mcthods as well as a simpler dynamic analog. Except where otherwise noted, the ratio, side of five-spot/wellbore radius. is taken to be 3,600. a figure which corrcsponds to a normal-size wellbore in a 10-acre well spacing. THEORETICAL EVALUATION OF VARIOUS PILOT PATTERNS, Mw0 = 1 <&apos;he theoretical models which we used to examine the performance of various pilots are shown in Fig. 1. Image theory was used to determine the ratio of production rate to injection rate as a function of the volumc of the flood. The ratio of production rate to injection rate was chosen because this is an easily measurable quantity which is characteristic of a pilot

By J. J. Forbes

&apos;"THE Federal Coal Mine Safety Act (public Law T. 552. 82nd Congress) was approved oil July 16, 1952. It incorporates, as Title I, the Coal Mine Inspectio1.1 and Investigation Act of May 7. 1941 (Public Law 49, 77th Congress), which gave Federal inspectors only the right to enter. coal mines for inspection and investigation purposes but no power to require compliance with their recommendations. Title 11 contains the enforcement provisions of the act; its purpose is to prevent major disasters in coal mines from explosions, fires. inundations. and man-trip 01. man-hoist accidents. At this point a brief account of events that preceded the enactment of the Federal Coal Mine Safety Act seems appropriate. The hazardous nature of coal mining was recognized by the Federal Govermment as long ago as 1865, when a bill to create a Federal Mining Bureau was introduced in Congress. Little was done, however, until a series of appalling coalmine disasters during the first decade of this century provoked a demand for Federal action. As a result an act of Congress established a Bureau of Mines in the Department of the Interior on July 1, 1910. The act made it clear that one of the foremost activities of the Bureau should be to improve health and safety in the mineral industries. One of the first projects selected by the small folce of engineers and technicians then employed was to determine the causes of coal-mine explosions and the means to prevent them. By investigations aftel mine disasters the fundamental causes and means of prevention were soon discovered, and the coal mining industry was informed accordingly. However, despite this knowledge and the enactment of State laws and the Federal Coal Mine Inspection and Investigation Act of 1941, mine disasters continued to occur with disheartening frequency and staggering loss of life. The devastating explosion at the Orient No. 2 mine on December 21, 1951, resulted in the death of 119 men. The Orient disaster rekindled the memory of the Centralia. Ill., disaster of March 25. 1947, which caused the death of 111 coal miners. These two tragedies ultimately brought about enactment of the Federal Coal Mine Safety Act. The act is a compromise measure. Senator Matthew M. Neely of West Virginia and Congressman Melvin Priec of Illinois introduced almost identical versions in the 82nd Congress, but they were considered too drastic. The final version was introduced by Congressman Samuel K. McConnel, Jr., of Pennsylvania, after considerable discussion and amendment in committee hearings. It was passed by the Congress and became effective when signed by the President on July 16, 1952. The act is somewhat limited in scope because it applies only to approximately 2000 coal mines in the United States and Alaska that employ regularly 15 or more individuals underground. It exempts approximately 5300 mines employing regularly fewer than 15 individuals underground and all strip mines, of which there are about 800. Moreover, it covers only conditions and practices that may lead to major disasters from explosion, fire, inundation, or man-trip or man-hoist accidents. According to Bureau records, such accidents have resulted in less than 10 pct of all the fatalities in coal mines. It is important to mention that the law is not designed to prevent the day-to-day type of accidents that have caused the remaining 90 pct or more of the fatalities, because it was the specific intention of the Congress to reserve the hazards which caused them to the jurisdiction of the coal-producing states. Many who opposed any Federal legislation that would give the Federal inspectors authority to require compliance with mine safety regulations claimed that such legislation would usurp or infringe upon States&apos; rights. To assure that the principle of States&apos; rights would be preserved, the act provides for joint Federal-State inspections when a state desires to cooperate in such activities. The Director of the Bureau of Mines is required by the act to cooperate with the official mine-inspection or safety agencies of the coal-producing states. The act provides further that any state desiring to cooperate in making joint inspections may submit a State plan for carrying out the purposes of this part of the act. Certain requirements are listed: these must be met by a state before the plan can be accepted. The Director of the Bureau of Mines, however, is required to approve any State plan which complies with the specified provisions. The Director may withdraw his approval and declare such a plan inoperative if he finds that the State agency is not complying with the spirit and intent of any provision of the State plan. When this paper was prepared, agreements for joint Federal-State inspections had been entered into with Wyoming and Washington. A few other states have indicated their desire to submit a State plan and negotiations toward that end are now under way. Reluctance to enter into such agreements may be due to the mine operators&apos; knowledge that in the states that adopt a cooperative plan they are prohibited from applying to the Director of the Bureau of Mines for annulment or revision of an order issued by a Federal inspector and must appeal directly to the Federal Coal Mine Safety Board of Review for such action. Experience has proved that review by the Director as provided in the act is a less expensive and time-consuming procedure to all concerned than applying to the Board. Reluctance also may stem from the fact that joint Federal-State inspections somewhat restrict the movements of the State mine inspectors and tend to reduce the number of inspections of mines. Where a State plan is not adopted, the Federal coal mine inspector is responsible under the law to take one of two courses of action if he finds certain hazardous conditions during his inspections. The first action involves imminent danger. If a Federal inspector finds danger that a mine explosion, mine fire, mine inundation, or man-trip or man-hoist accident will occur in a mine immediately or before the imminence of such danger can be elim-

Jan 1, 1955

By J S. Dunlop

At its most basic level, minerals project management may be broken down into a series of logical chronological phases, described in this paper as: exploration, discovery; scoping and pre-feasibility; feasibility and approval; project financing and commissioning; operation; expansion and closure. Each phase has its own peculiarities and challenges, any one of which can potentially upset the forward progress of a minerals project.Reference is made to each of these project phases based on the authorÆs many years of first-hand experience, in an attempt to identify issues within the project management spectrum where additional caution may be warranted. Left un-addressed, these issues can cause undue delays in the project timeline, ultimately delaying approval at significant project cost.Finally, reference is made to the Mine ManagerÆs Handbook recently published by the AusIMM which is an excellent project management resource, considered to be very relevant for minerals treatment plant managers, whether the context be constructing, operating, expanding or shutting down.CITATION:Dunlop, J S, 2013. Mineral project management - A perspective from four decades in the industry, in Proceedings MetPlant 2013 , pp 13-27 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 15, 2013

By Bryan M. Maybee

The scheduling of underground mining activities is a complex task, in particular because of the requirement to satisfy a large number of constraints. A Schedule Optimization Tool (SOT) for underground mine planning is described, and its use for investigating strategic decisions is illustrated by means of a case study. The main components of SOT include: a genetic algorithm, which searches for optimal schedules; guidance, the application of heuristic rules to guide the genetic algorithm; and sliding, which implements just-in-time development. For the case study, SOT is employed to generate schedules at various ore capacities, accounting for the appropriate financials in each schedule evaluation. The case study finds that preliminary scheduling performed before invoking SOT, intended to provide a good starting point for the search, does not ultimately lead to higher value schedules. Two key uses for SOT are identified. A mine planner can save time by using SOT to produce high-value scheduling alternatives. This process ensures a broader investigation of the value that the deposit can yield. Also, SOT can be used to validate strategic and tactical decisions, by scheduling under alternative scenarios. Both uses produce high value schedules conforming to all imposed resource constraints.

May 1, 2009

By S. R. Seagle, R. Schuhmann, N. A. Parlee

Rates of CO evolution and CO absorption were measured for liquid-iron alloys containing from 0.15 to 4.4 pet C, using a modified Sieverts apparatus. The alloys were held in alumina crucibles, so that both crucible-metal and gas-metal reactions occurred simultaneously. The data are interpreted on the hypothesis that the C-O reaction rate is controlled by oxygen diffusion across the boundary layers at the gas-metal and crucible-metal surfaces. CARBON-monoxide evolution from liquid Fe-C-0 alloys is a key reaction in steelmaking processes, both in the steelmaking furnace and in the ingot mold. Also, this reaction probably is an essential step in the desulfurization of iron under blast-furnace conditions, for which carbon is the principal reducing agent or deoxidizer present in the iron. The thermodynamic properties of the solutions of C and 0 in Fe and the equilibria of these solutions with gaseous CO and CO, have been investigated in some detail and are reasonably well understood. However, while these equilibrium data have accumulated in the laboratory, other data have accumulated to show that equilibrium conditions are often not achieved under plant operating conditions. Thus, an understanding of the rate and mechanism of the carbon-oxygen reaction not only has theoretical interest, but ultimately may assume considerable practical importance. Reaction Mechanisms and Rate Equations The over-all reaction under consideration is C (in liquid Fe) + 0 (in liquid Fe) + CO (gas) [I] Although this appears to be a straightforward heterogeneous reaction, involving just two phases (liquid metal and gas), a wide variety of reaction mechanisms has been proposed. Different authors have expressed divergent views as to the nature of the rate-controlling steps. Accordingly, a brief discussion of the possible reaction steps will be given. 1) Homogeneous Reaction Within Liquid Iron— Feild&apos; and Jette2 have applied chemical reaction-rate theory as if the C-0 reaction were a homogeneous, second-order reaction occurring within the liquid-metal phase. This mechanism necessarily produces CO "molecules" dissolved in liquid iron. In order to yield CO gas, the dissolved CO must nucleate CO gas bubbles, and the dissolved CO also must be transported to the bubble surface where it enters the gas phase. However, this mechanism is inconsistent with present concepts of the structure of liquid-metal solutions. 2) Reaction at Gas-Metal Interface—At the surface of contact between gas and metal, a second-order reaction between dissolved C and 0 may occur to produce a CO molecule which enters the gas phase. This reaction presumably involves an "activated complex" structure in the interface. Some such surface mechanism appears essential because the carbon and oxygen do not occur in the same species in the two phases and therefore must react in some way at the surface. At liquid-iron temperatures, calculations based on reaction-rate theory indicate that the surface reaction must proceed extremely rapidly and thus cannot be a rate-determining step in the C-O reaction under ordinary conditions.3 ccordingly when the over-all reaction is measurably slow, the surface reaction may be considered at equilibrium; that is Cc* Co* = m&apos;Pco [2] in which Co* and Co* are the concentrations of carbon and oxygen (weight per unit volume of metal), respectively, in the liquid-metal phase at the interface, Poo* is partial pressure of CO in the gas phase at the surface (atmospheres), and m&apos; is the mass-action constant for Reaction [I]. Since activities of C and 0 are not strictly proportional to concentrations, m&apos; is not a true thermodynamic equilibrium constant but varies with carbon content.&apos; 3) Mass Transport—If the reaction is to proceed at the gas-metal surface, as described above, dissolved carbon and oxygen in the liquid iron must come to the surface and gaseous carbon monoxide must move away from the surface. Details of the mass-transport mechanisms depend on the kind of system under consideration. For a stirred metal bath in contact with a CO gas phase, the transport of C and 0 to the surface can be described by 1 Do Mols CO evolved per sec = — 1/16— Do/do Agm (CO-Co*) = 1/12 Do/do Agm *(Co-Co*) [3] In this equation, Do and D, are the diffusion constants for oxygen and carbon in liquid iron, 6, and 6,. are the boundary-layer thicknesses, A,,,, is the

Jan 1, 1959

By R. H. Jacoby, V. J. Berry

When dry gas is injected into a reservoir containing a volatile crude oil, a significant amount of the reservoir liquid phase may become vaporized. The resultant rich gas phase, when subsequently produced, con-tributes to tank oil production. This contribution assumes greater importance, the more volatile the oil in-olved. Oil recovery may be sub-stantially greater than that predicted by conventional frontal-drive methods, which do not consider the vaporization equilibrium between the reservoir oil phase and the injected gas. A calculation method has been (developed to account for vaporiza-. tion of the reservoir liquid phase during gas-injectiorz operations, and for the tank oil production which results from this factor. Recovery performance calculations are presented for a reservoir containing a highly volatile oil. Tank oil recovery is calculated to be about twice that predicted by the use of the conven-rional frontal-drive equations. In contrastto usual pressure maintenance performance results, in which the gas-oil ratio rises at an increasing rate after gas breakthrough, the pre-dicated gas-oil ratio rises rapidly to about 12,000 scf/bhl and then rises much less rapidly. During gas inje-tion, most of the reservoir liquid phase contacted is evaporated by the dry injection gar. The gas-oil ratio during this period is dependent upon reservoir pressure. The higher the operating pressure, the lower the gas-oil ratio. The predicted behavior i., in accordance with laboratory PVT tests made to .simulate the vaporization behavior. In addition to recovery performance predictions, results of the calculation procedure include complete wellstrearn composition data of value in the design of gasoline plant facilities often used in con-rrction with gas-injection operations. INTRODUCTION In the cycling of gas-condensate reservoirs, dry gas is injected to maintain reservoir pressure during wet gas production and to thereby eliminate or reduce ultimate loss of liquids due to retrograde condensation within the reservoir. Gas injected into crude oil reservoirs has a dual function. It displaces oil to the producing wells and at the same time serves to partially or fully maintain reservoir pressure. Oil shrinkage which would occur upon pressure reduction is thereby minimized or eliminated. Accepted calculation methods are available&apos;= for predicting recovery performance of either gas-condensate reservoirs or crude oil reservoirs which are being subjected to gas injection. In gas-condensate reservoirs, any retrograde liquid formed does not flow, and it is necessary to account only for the vaporization equilibrium between this liquid and the injected gas. Conversely, in normal crude oil reservoirs, both the oil and gas phases flow, but it has not been considered necessary to account for any vaporization of the reservoir liquid which might occur upon contact with the dry injection gas. Recently, high shrinkage reservoir fluids known as "volatile oils" have been found in increasing amounts. These oils are characterized by tank oil gravities above 4.5" API, solution gas-oil ratios above 1,000 scf/bbl, and reservoir volume factors above two. Special techniques have been devised for predicting depletion performance of reservoirs containing such oils.74.5.1; One of the characteristics of reservoirs producing volatile oils is that the reservoir gas phase carries a significant amount of oil which is recoverable as stock-tank liquid. This unusual vaporization be-havior implies that an appreciable amount of reservoir liquid would be vaporized upon contact of the oil with dry injection gas. In a gas-injection operation, tank oil recovery would he obtained not only through frontal displacement of the reservoir liquid by the injected gas, but also through production of the rich gas phase. This means that not only are improved methods needed to predict recovery of such oils from reservoirs undergoing gas injection, but it would also be expected that high oil recoveries might be obtained by such operations. When relatively dry gas is injected in to a volatile oil reservoir, phase equilibrium between the injected gas and the reservoir oil will tend to bc established. Initially the most volatile components, such as propane, the butanes and the pentanes, will account for most of the material transferred from the oil phase to the gas phase. As the partially stripped oil phase is contacted with additional dry injection gas, the heavier intermediate components, such as the hexanes, the heptanes and the octanes, will gradually transfer to the gas phase in increasing amounts. This is because the supply of lighter components in the oil phase dwindles due to the stripping action of the injected gas. This stripping action will usually continue to be effective down

By I. S. Thompson

A program of geochemical till prospecting, carried out by Derry, Michener & Booth in 1974, was instrumental in t~ e discovery of a copper deposit in Currie Township, Ontario, beneath thick lacustrine deposits of the Abitibi Clay Belt. The prospecting system involved the continuous collection of glacial sands and gravels beneath the clays, using a track-mounted, dual-tube rotary drill, and the separation and geochemical analysis of the heavy mineral content for Cu, Zn, Pb, Ag and Au. In one of 22 holes drilled down-ice from nine Canadian Nickel airborne EM conductors, fragments of chalcopyrite-bearing graphitic argillite, assaying 12% Cu, were recovered from a 14-foot gravel section beneath 83 feet of clay. Ultimately, mineralization was identified in 10 of 11 additional holes drilled up to 3,000 feet down-ice from the Currie deposit. The glacial fan, when compared to the 1,200- by 165-ft area of the subcropping mineralization, is at least 20 times the area of the deposit and is the result of glacio-fluvial reworking of the mineralized ground moraine. The Currie deposit is contained within a Keewatin succession of graphitic argillite, dacite tuff and graywacke, which is intruded by dacite and feldspar porphyry sills and dykes and truncated by a Keweenawan diabase dyke. Grade and thickness of the mineralization, which lies chiefly within the argillite, is related to felsic intrusives and hydrothermal alteration, and to fold hinges.

Jan 1, 1979

CONTENTS PAGE HESSE, A. W,-Safeguarding Coal-mining Operations against Danger from Oil and Gas Wells. Discussed by A. W. Hesse, T. G. Fear, George H. Ashley, George S. Rice, W. E. Fohl, R. V. Norris, S. A. Taylor, F. B. Tough, C. W. Gibbs, A. Hurlburt, S. W. Meals 1 DAWSON, THOS. W.-Belt Conveying of Coal at H. C. Frick Coke Company Mines. Discussed by F. F. Jorgensen, Thos. W. Dawson, Graham Bright, A. W. Hesse 13 OTTO, HENRY H.-Ultimate Recovery from Anthracite Coal Beds. Discussed by R. V. Norris, S. A. Taylor, Henry H. Otto, Douglas Bunting, Graham Bright, Howard N. Eavenson, J. B. Warriner 15 HOSLER, Rush N.-Schedule Rating Coal Mines in Pennsylvania for Compensation Insurance Rates. Discussed by E. A. Holbrook, W. W. Adams 17 ARCHBALD, HUGH.-Application of Gaussian Curve to Mining Industry. Discussed by Donald A. Laird, T. T. Read 19 ASH, SIMON H.-System of Coal Mining in Western Washington. Discussed by Eli T. Conner, Simon H. Ash 21 MILLER, J. S.-Method of Mining a Steeply Pitching Anthracite Vein by Successive Skips. Discussed by Eli T. Conner, J. S. Miller, J. B. Warriner 22 HARRINGTON, GEORGE B.-New Orient, an Unusual Coal Mine. Discussed" by George S. Rice, Eli T. Conner, John A. Garcia, George B. Harrington, F. F. Jorgensen, Graham Bright, W. M. Hoen, George N. Simpson, W. C. Adams, R. W. McNeill, Chas. C. Whaley, Carl Scholz, E. T. Gott, Andrews Allen 23 PERROTT, G. ST. J.-Properties of Liquid-oxygen Explosives. Discussed by George S. Rice 48 Safeguarding Coal-mining Operations against Danger from Oil and Gas Wells Discussion of the paper of A. W. HESSE, presented at the New York Meeting, Feb-ruary, 1925, and issued, as Paper No. 1412-F, with MINING AND METALLURGY, February, 1925. A. W. HESSE.-The well under investigation, No. 101, Fig. 2, was started about Sept. 5, 1921, and drilled-in in November. The first hole was started with a 13-in. bit and reached the Pittsburgh coal at a depth of 563 ft.; at 602 ft. the 10-in. casing was run in and the well was continued with a 10-in. bit and 1400 ft. of 8-in. casing was put in. It was then continued to a depth of 1770 ft. and 6-in. casing put in; then continued to a depth of 2555 ft. where the gas was struck. As the well produced only about 100,000 cu. ft. per day. 2-in. tubing was used to conduct the

Jan 6, 1925

By Peter Stacey

This presentation provides comments on the increasing challenges associated with pit slope designs from the perspective of a design reviewer. Besides the technical issues related to the significant increase in current and proposed slope heights since the primary methodology in current use for pit slope designs was developed, other factors are coming into play. These include changes in mining equipment and associated operating practices, as well as a growing awareness on the part of mining executives and other stakeholders of the significance of stable, or at least well-managed, slopes. In the latter area, the associated clarification of responsibility has led to the increased use of either in-house or independent review consultants working on the behalf of management. From a technical perspective, since the mid-1970?s there have been significant improvements in the tools, particularly analytical methods, that are available to slope designers. There is also an increased understanding of the importance of a detailed geological model, with strong emphasis on alteration and/or structure to form the platform for the slope designs. However, the basic technology for determining the rock strength aspects has changed little, although an expanding body of experience is being accumulated in its use and there is a reasonable degree of comfort in its application for slope heights of up to at least 500 m. In addition, slope management is now viewed as a critical part of the implementation of slope designs and this has been supported by continuing improvements in slope monitoring systems. In recent years safety has become a primary concern, in part as a result of some major slope failures that have attracted the attention of regulators and the public, as well as mining executives, who are being held more responsible for unsafe conditions and associated events. There is therefore a growing demand for risk assessments to replace the deterministic design approaches upon which the current design technology is largely based. The focus of mine executives on safety has undoubtedly in part facilitated the recent long overdue revitalization of research into the design and stability of large pit slopes, the requirement for which has been recognized for many years by practitioners. This research, as well as the CSIRO Large Open Pit Study, will almost certainly provide advances beyond the empirical strength determination methods such as the Hoek-Brown failure criterion, which form the basis for current rockmass strength determination. At the same time, as mentioned above, there is a growing awareness of the requirement for a detailed structural model as a major component of every large slope design. In addition, other areas requiring further research include the impact of groundwater pressures on rock mass strength, particularly in rocks with low permeabilities, and the role of stress in high open pit slopes. From the perspective of design implementation, even with the recent significant advances in metal prices, mine operators remain under pressure to minimize mining costs. To address these constraints, mining equipment of ever increasing size is being introduced. There are, however, some disadvantages to this trend in the area of slope design. For example, the large electric shovels are not well designed for scaling bench faces, nor are they cost effective in this mode. As a result, specialized equipment may be needed in the mining cycle to perform the clean-up duties, which in turn increases operating costs. Further, where the large equipment is used in minimum width pushbacks to reduce the instantaneous stripping ratio, the advance rates can be high, but at the same time the associated changes in operating procedures are often not conducive to the concurrent use of such measures as controlled blasting, careful scaling and drain hole installation that generally improve stability. The resulting conflicts which may arise between the interests of production and those of slope stability are often exacerbated by the fact that these stabilization techniques actually increase the operating costs on which the operations manager is frequently judged, even though there is an overall increase in profit. Meeting the objective of developing slope designs which are practicable , i.e., achievable in terms of every aspect of the operating constraints in the specific pit, requires interaction and compromise between the geotechnical engineer, mine planners and operating staff during the formulation of the design criteria. In the current environment, it is often no longer sufficient to present slope designs in deterministic terms to a mine planner who accepts them almost without question. Increasingly, the requirement is that they be proposed within the framework of risk levels, related both to safety and to economic outcomes, to a decision maker who may not be a technical expert in the mining field. In this context, the mine executives must have sufficient information and understanding to be able to establish acceptable levels of risk for the company and other stakeholders; in this process the slope designers must play a major role. These changing requirements for presentation of slope designs necessitate clear communication of the basis for the design, which may be very complex, so that the implications of the designs may be thoroughly understood by all concerned. To ensure clarity at this stage, the slope designer must not only quantify the uncertainty in the input parameters, including the geological model, but must also be able to communicate and defend the resulting design recommendations. Where particular designs may have high associated risks, simply because of the degree of uncertainty in the data, this must also be recognized and proposed methods of reducing the uncertainty included in the presentation. In summary, from a reviewer?s perspective slope designs must not only be technically sound, but must also address the broader context of the mining operation as a whole, taking into account such factors as safety aspects, the available equipment to implement the designs, and the acceptable risk levels for the company. In addition, the designs must be presented in a way that will allow the mine executives, who are ultimately responsible, and the operators, who implement the designs, to fully understand the basis and short-comings of the designs and the risks associated with deviation from any constraints defined by the designer. It goes without saying that there must also be a well-defined monitoring system to confirm stability and detect and manage any variations in the design model or unexpected instability.

Jan 1, 2006

By Tom Watts

The practice of pumping bulk blasting agents is regulated by many agencies around the world. The explosives manufacturers have done an excellent job of hazard and operability studies allowing criteria to be established for the safe pumping of water based blasting agents. The challenge of a Bulk Trucks or Mobile Manufacturing Units (MMUs) manufacturer is to take the objectives of the regulatory agencies and the explosives manufacturers and responsibly design and build MMUs that are safe, compliant, and cost effective without compromising serviceability.

Jan 1, 2011

By Terrence D. Chatwin, James Hendrix

Based on the findings within their report to Congress, the US Environmental Protection Agency (EPA) announced that it had decided not to regulate mining wastes under the hazardous waste regulatory program. However, they found that some mining wastes are hazardous and may be causing environmental damage. Cyanide wastes from heap leach operations were included in this group. In evaluating the effects of cyanide use in mining one finds a significant lack of information on the fate of cyanide in soils. Cyanide attenuation and degradation of soils is important, because ground water is a major potential pathway for cyanide contamination. There is a strong potential for cyanide to be removed from solutions as it migrates through soils· by volatilization, chemical reaction, adsorption, oxidation and biological processes, but no definitive investigation has defined these mechanisms adequately to be able to predict the fate of cyanide in soils.

Jan 1, 1998

By L. L. Wyman

IN order to clarify and amplify the existing data concerning the action of the cementing material in cemented tungsten carbide alloys, the authors have initiated this investigation of the entire range of cobalt-tungsten carbide alloys. Inasmuch as the ultimate objective is relative to what actually goes on during the sintering of cemented tungsten carbide materials, this work was necessarily restricted to heat treatments similar to those used in actual production of these materials. In the course of numerous experiments, the authors have noted several conditions that indicated that there was a solubility to be considered. Among these factors are the following: 1. Many of the alloys showed a much larger amount of binding constituent to be present than could possibly be accounted for by the cobalt content. 2. In many areas, grains of the carbide constituent are much larger than the particles of carbide originally added. In addition, these grains are of very regular contour. 2a. In samples of cemented tungsten carbide which had been fused in the atomic hydrogen torch in the presence of excess hinder constituent, immense grains are formed, and their shapes are very regular. This is also true when the cemented tungsten carbide of 13 per cent. Co content is fused alone in the atomic hydrogen torch. Contrary to general expectation, chemical analysis of this material, after fusion in the atomic hydrogen torch, checks the analysis of unfused material. 3. In making the cemented tungsten carbide materials by the process of exerting the pressure at the time of heating a certain portion of the contents squeezed out of the mold. Chemical analysis has shown that this material contains approximately 12 to 20 per cent. of tungsten. The microstructure shows a cored dendritic structure interlaced with eutectic network, and some graphite, as shown in Fig. 1. 4. Thermal analysis of these materials has consistently indicated an arrest point close to 1350° C.

Jan 1, 1930