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By James H. Hance

Discussion of the paper of JAMES H. HANCE, presented at the New York meeting, .February, 1916, and printed in Bulletin No. 110, February, 1916, pp. 299 to 326. FREDERIC P. DEWEY, Washington, D. C. (communication to the Secretary*).-I wish to congratulate Professor Hance and to thank him for his extremely valuable addition to the fund of data upon this subject. It presents clearly conclusive and incontrovertible proof of the unreliability of drill samples of this class of metal for determining the value of such bullion. I am afraid, however, that his hope that some one in the mint service might continue such an investigation upon a more elaborate and exhaustive scale is, for the present at least, doomed to disappointment. As a practical matter, the mint service is not now particularly interested in segregation in crude bullion. So much trouble and friction arose, -both in the original purchase and in the transfer from one institution to another, in handling segregated bullion, that it became necessary to adopt drastic measures and now it is seldom attempted to determine the value of such bullion in the service. When bullion known or even suspected of being segregated is presented for purchase, it is at once strongly refined in the pot and the disturbing elements so far removed that concordant assays may certainly be obtained from the dip samples and a reasonably fair agreement obtained from drill samples. The mint service prefers that the owner should refine his bullion himself, but if he does not we will do it at his expense and return the slag to him when desired. Much work has been done in the mint service trying to arrive at the value of deposits of segregated bullion without strong refining. It was no unusual thing for such deposits to be melted three times at the purchasing office and to have 30 to 40 assays on the metal made before buying it. Yet when shipped to a mint it was found necessary to melt two or three times more and to make 15 to 25 assays before the mint would accept the bar, and there were considerable differences between the valuations of the two offices. In what was undoubtedly the worst case ever investigated, the purchasing office melted the deposit, 953 oz., three times, with a loss of 105.5 oz., and made 46 assays; the mint melted three times, with a loss of 6.6 oz., and made 28 assays. The mint allowed the purchasing office 1.178 oz. of fine gold more than the purchasing office claimed, while an extensive investigation of the mint samples by the Bureau, in which over 100 independent assays were made in various service laboratories, indicated that the metal carried 0.7 to 0.9 oz. more of fine gold than the mint allowed. Preliminary assays on the original metal

Jan 5, 1916

By R Fletcher

Pacific Hydro, a leading renewable energy company from Australia, is developing a series of cascading run-of-river hydro electricity projects in the Cachapoal Valley at the base of the Andes Mountains some 100 km south of Santiago in Chile.These projects involve extensive tunnelling and underground works. The first of these projects is called Chacayes (110 MW installed capacity) and is now under construction with Astaldi SpA of Italy and Fe Grande of Chile as the EPC contractor. Astaldi SpA also has an equity stake in the project. It involves 6 km of tunnels between 5 m and 7 m diameter, of which 2.3 km is by tunnel boring machine (TBM) (5 m diameter).This paper outlines the local and regional geology, the scope of the Chacayes project, actual tunnel excavation rates and support, and compares the decision making on Chacayes between a choice of traditional drill and blast excavation versus TBM.The paper also outlines the next project planned to be developed by Pacifi c Hydro in the valley – the 140+MW Nido de Aguila project involving a main tunnel drive of 11 km with two intermediate adits, plus a 388 m shaft and underground powerhouse. The main tunnel drives are likely to be excavated by two hard rock TBMs (5 m diameter) but will need to be able to withstand high overburden pressures of up to 1300 m.Subsequent projects in the planning stages include the downstream Coya II project with two tunnels totalling approximately 10 km, one of which may be by TBM.

Mar 8, 2011

Mine planners must select a single pit for design with the knowledge that the grades on which this selection is made are uncertain. This uncertainty is often ignored through the use of deterministic models of grade. This makes the mine planner?s job simpler but does not necessarily lead to the ?correct? or ?best? selection. Grade uncertainty is typically modeled with conditional simulation, which generates multiple possible realizations of grade that are equally probable (or improbable). M ethods for incorporating grade uncertainty into mine planning to date have either been perceived as being too complicated, not accessible or too time-consuming. This paper proposes a simple and efficient modeling approach to incorporating conditional simulation into pit optimization, which would be able to be applied in currently commercially available software. An example for R esolute M ining Ltd.?s Syama gold project in M ali shows the range of ore, grade and value outcomes associated with generating pits through a variety of methods.

Jan 1, 2013

By R. R. Bassermann

THE property of Howey Gold Mines, Limited is situated at Red Lake, Ontario, one hundred and seventy-five miles east of Winnipeg and eighty miles north of the Canadian National Railway line. The mine has been a continuous producer since April, 1930, and at present is hoisting 1,350 tons of ore per day. The shaft consists of three compartments (see Figure 1, sketch B). Hoisting is done by two skips in balance, one of which is interchangeable with a cage. Sinking had been carried out at three different periods. From the surface to the 550-foot level, the sets had been spaced at seven-foot intervals, from the 550-to the 1,060-foot level at six feet, and from the 1,060-to the 1,500-foot level at five feet. During the five years of operation, the shaft had become so badly worn, due to spill from the tipples and, to a certain extent, from the loading pockets, that re-timbering was considered necessary. It should be explained that no crushing is done underground because of the ad verse effect this would have on the efficiency of the sorting. The grizzly rail clearance above the ore pockets is fourteen inches and consequently the ore hoisted is extremely coarse. More than average spill from the loading pockets and tipples under such conditions is inevitable, and it is debatable whether underground crushing with consequent increased sorting costs would not be justified. The cost of complete re-timbering according to standard practice would have been prohibitive both directly and indirectly: A lengthy shut-down of the shaft, with the consequent disorganization of the staff and workmen, could not be considered. Accordingly, it became necessary to devise some method of repair that could be carried to completion in a very limited period of time. The shaft timber for the first 250 feet was not in need of repair. From the 250-foot to the 1,060-foot level the wall plates, posts, and blocking were in good condition. These had been protected by a one-inch plank lining put in at various periods during the past three years without interfering with hoisting operations. The end plates and dividers, however, could not be protected in this manner because of the lack of skip clearance. It was decided that, if the dividers in this section could be replaced, the shaft would be as satisfactory as if completely re-timbered. Several factors had to be taken into consideration in deciding on a suit-able method of repair. In the first place, the tipple plates had to be re-designed. To complete the job in a reasonable time, not only had work to be done on the tipple, but simultaneously from several points in the shaft itself. Since all power cables, pipe lines, etc., were in the manway compartment of the shaft, this compartment had to be left undisturbed.

Jan 1, 1935

By Daniel Augusto Godinho de Carvalho

Since the beginning of BOF process, slag formation has been subjected to extensive development. During the initial stages of the blow, fast slag formation is primordial for the process performance. This is mainly to allow oxidizing reactions, and also to protect the refractory. A modeling of BOF process for the purpose of process optimization requires precise knowledge concerning the dissolution of CaO and MgO based materials. Using this knowledge, the optimal process conditions can be ensured by a controlling of the slag formation during the converter steelmaking process. Taking into account an application of large amount of different raw materials, as well as location-specific process conditions, the characterization of the dissolution behavior of different additives is an important key to improve the technological process. Thereby, in the frames of this project, an appropriate model for slag formation is established and compared to operational results for a 330t BOF, operating at ThyssenKrupp CSA, Santa Cruz, Brazil.

Aug 16, 2017

By J. P. Dunlop

The fact is not.ablc though probably little known that thc Unitcd Statcls is the only nation obtaining and distributing through its Govcrn-ment burcaus any data pertaining to waste metals and drosses. Some foreign goverrlnlents do issuc reports relating to rnine or smelter production of prirrlar~ metals, but none gives in scope and detail t'he information published now by the U. S. Bureau of Mines and formerly by the U. S. Geological Survey; and this is despite the fact that the funds annually allotted for those purposes are insignificant in comparison with allotments for the preparation of data relating to other industries of like importance. The earliest data on the production of metals from secondary sources in the United States were contained in a press bulletin issued by the U. S. Geological Survey on Jan. 4, 1909. The information published was that for thc year 1907 obtained ant1 compiled by C. E. Siebenthal, who was one of the first to visualize the growing importance of the scrap-letal indust,ry. Thc Nationd Conscrvation Commission had emphasized its opinion that the known domestic supply of certain kinds of the morc important irietals was being rapidly marketed and that while there was no immediate cause for concern it was best to conserve such resources. Not only had there been a lanlentable lack of any data regarding the quantity of scrap inetals available or used but also gross carelessness and indifference in the segregation, collection, and effective use of scrap material. Siebenthal realized that even approximately correct statistics of recoveries from the more important base-metal scrap and drosses wcrc nearly as essential t,o both producers and consumers as the stat'istics of the output of primary metals. There were not then nor are there now sufficient funds for complete collection of data, nor was it considered the province of the Geological Survey t'o do more than determine approximately the recovery of secondary metals for use in supplementing those given for the primary metals. Undoubtedly the first bulletins issued did not give the full output, for it takes time to establish contacts for an inquiry covering so

By J. Keith Haddock

"Introduction The Boundary Dam Mine is located 5 km south of the City of Estevan, Saskatchewan and is operated by Manitoba and Saskatchewan Coal Company Ltd., one of the Luscar companies. The entire production of 1.7 million tonnes per year is delivered to the Boundary Dam Generating Station of Saskatchewan Power Corporation. Geology There are basically two seams mined on the property, the main Estevan seam ranging in thickness from 3.0 m to 4.5 m with partings and the upper Roche Percee seam (roughly 15 m to 23 m above the Estevan) which ranges from 0.6 m to 2.0 m in thickness. Most of the reserves are concentrated in the Estevan seam, the Roche Percee only being mined whenever its thickness makes it economic to recover. The coal measures belong to the Ravenscrag Formation of Tertiary (Paleocene) Age. This formation is widespread in North Dakota and Montana and contains the lignite fields of North Dakota and subbituminous coals of Montana. The formation is gently tilted to the southeast at Boundary (0.5 degrees) and consists of weak sandstones, siltstones and silty mudstones. Reserves and Coal Quality The present minesite encompasses a coal reserve of approximately 50 million tonnes calculated to a maximum stripping depth of 37 metres or 20:1 mining ratio. The reserves form only a part of the total reserves in the Estevan area which are owned by the Saskatchewan Power Corporation. The coal being mined is classed as Lignite ""A"" and has the following specifications:"

Jan 1, 1985

By G. A. Koehler

The Hiawatha mine, a producer of iron ore, is near the City of Iron River, in Iron County, Michigan. This mine, operated by The M. A. Hanna Co., has been chosen by the author as a typical iron mine in this district that is using the cut-and-fill mining system. General Geology The Hiawatha ore body lies along the steeply dipping north limb of a syncline that closes around on the east end to form a bathtub-shaped structure. Throughout this major structure there are very many smaller wrinkles or folds, which vary the direction as well as the thickness of the ore body. The total length of the ore formation with minor interruptions so far developed is about 3600 ft. The ore body throughout the entire present depth of 2000 ft. has averaged approximately 40 ft. in thickness; it is somewhat narrower at the upper levels and shows a marked increase in thickness on the bottom levels. This increase in thickness probably indicates that the bottom of the trough is near or that a major monoclinal fold has been encountered, which flattens the ore and gives apparent horizontal thickening. The footwall of this ore body is generally a black, graphitic slate, very rich in iron sulphides, quite combustible when exposed to air, and very treacherous to handle during mining operations. The hanging wall, in most places, is a cherty, fairly hard iron formation, and in general stands quite well ill open stopes. There are places, however, where black slate is found in both the hanging wall and the footwall, which makes mining difficult, as the black slate tends to slough off after being exposed to air. In these places it becomes necessary to leave a scale of ore 5 to 6 ft. thick on both hanging wall and footwall, which is not recovered. Fig. I shows a typical vertical cross section through the Hiawatha ore body. Type of Stoping Because of the size and shape of the ore body and the hard, dense structure of the ore, sublevel stoping has been an efficient and economical method of extracting the ore. This method, with various minor changes, has been used since the mine was opened in 1893. There have never been any stope-filling operations, however, until recent years, when crushing of pillars and black-slate fires forced the issue and methods were arrived at to fill all stopes. Development of Deeper Levels Shaft sinking to deeper levels in development operations has been done successfully by two methods without interference or hindrance to production. These systems are: I. Sinking an auxiliary winze a short distance from the shaft, then drifting back to the shaft location and putting up a pilot raise through to the bottom of the shaft. This raise is always in the manway compartment of the shaft, so that no work is ever done under a cage or skip. A 20-ft.

Jan 1, 1946

By Howard C. Pettibone

Mine safety personnel reported occurrences of excessive "glove-fingering" and other problems with resin-grouted roof bolt installa tions. To study these problems, the Bureau of Mines designed a test plan to (1) obtain baseline information on several resin cartridges and (2) investigate the variables involved in the installation of a resin bolt. Three brands of resin representative of the industry in January 1983 were selected for cartridge evaluation. The installation variables selected for study were the method of installation, the hole depth, and the hole annulus. These variables were studied by installing a series of bolts in concrete blocks using different values for each variable. Results of the research show overspinning the resin may cause problems, but they can be avoided by using the spin times recommended by manufac-turers. Excessive glove-fingering is not a problem if the manufacturer-recommended installation procedures are followed.

Jan 1, 1987

By J. E. Gebhardt, D. McBride, T. N. Croft, M. Cross

"Heap leaching is well-established as a relatively low-cost and low-energy method of extracting metals from low grade ores. However, the design methodology, attention to the engineering details during heap construction and the long time scale in metal recovery makes the process susceptible to significant variation in the success of a heap leach operation. Often the laboratory predicted recovery rates are not achieved on site due to poor understanding of the complex physics of flow and coupled solid-liquid-thermal reactions within the multi-dimensional heterogeneous ore body. Given the large scale of commercial operations, small changes in metal recovery can substantially change the operating profitability. While there are a variety of approaches to heap leach modeling, computational fluid dynamics (CFD) modeling coupled with leach kinetic models offers a methodology to describe and understand the complexities of heap leaching. The CFD model uses leach kinetics specific for the mineralogy of the ore, including gangue chemistry, and provides an energy balance to calculate and predict temperature. Coupled with the CFD model’s inherent capability for modeling complex fluid flow, the physics and the chemistry of the heap can be simulated. This provides a tool for engineering design, as well as forecasting heap production under different scenarios. The calibrated model is used to simulate a three-dimensional test heap under fluctuating meteorological data, such as precipitation events, periodic solution application, and notably temperature. These simulations provide support for a methodology to design and engineer heap leaching of low grade copper sulfide ore. INTRODUCTION Worldwide copper ore reserves are faced with decreasing head grades with copper sulfides, particularly chalcopyrite being the predominant mineralogy (Nesse, 2000). Heap leaching offers an economical way to recover copper from low grade ore. The ore is stacked and a leach solution is applied and allowed to trickle though the porous bed of ore to extract target metals. The heap can be built with run of mine material or crushed ore and is constructed in order to achieve percolation of the leach solution to all regions. Bartlett (1998) provides an authoritative overview of the solution mining process where the leach solution interacts with the solid, the lixiviant dissolves the target minerals from the solid into the solution, and the dissolved species are transported out of the heap for subsequent recovery. Petersen (2016) provides a recent review of heap leach technology and emerging and potential future applications, and Ghorbani et al. (2016) summarizes the current state and challenges of the technology. Early applications of heap leaching offered a low cost technique to achieve high recoveries from oxide ores, typically greater than 70% depending upon the kinetics of reaction. As world copper resources transition to mainly copper sulfide mineralogy, such as chalcopyrite based ores, achieving high recovery from these copper sulfide deposits is proving more difficult due to the complexity of the leach kinetics, which leads to low leaching efficiency."

Jan 1, 2018

By H. P. Carlisle

The significance of the mineralogical and surface characteristics of gold grains in the recovery process by C. E. FEATHER and G. M. KOEN Published in the Journal. February 1973. DISCUSSION H. P. Carlisle. B.Sc. (Eng.) Rand (Fellow) The authors are to be congratulated on drawing attention to the importance of surface characteristics in the treatment of gold grains. I must, however, confess to considerable doubts about the validity of the distribution of gold in "free" and "locked" form set out in Tables I to IV.

By Robert S. Young

During the last quarter-century, exploration for metallic deposits in Maine has been sporadic with peaks generally coinciding with periods of high metal prices. Known cases of regional or semi-regional evaluations are few; most efforts have been local and usually intensive. On the basis of broad geologic groups, the State has been divided into five fundamental environments: ( 1) coastal metamorphic belt; (2) southern volcanic belt; (3) Silurian slate belt; ( 4) northern volcanic belt; and ( 5) Devonian slate belt. Although not restricted to these areas, exploration in recent years has been concentrated in the northern and southern volcanic belts. Notable examples of recent exploration activity, directed to sulfide minerals, include: ( 1) the Knox County nickel-copper-bearing ultramafic rocks; (2) the Harborside (Cape Rosier) copper-zinc deposit; (3) the Second Pond (Black Hawk) zinc-copper-lead-silver deposit; ( 4) the Big Hill-Barrett lead-zincsilver- copper-gold zones; (5) the Parmachenee copper-zinc belt; ( 6) the Moxie Lake nickeliferous peridotite; and (7) the Attean quartzmonzonite molybdenum-copper occurrences. The Second Pond and Harborside prospects appear to be close to production. Systematic investigations for non-metallic deposits appear to have been aimed principally toward limestone, asbestos, diatomaceous earth, and spodumene-bearing pegmatites. Advances in prospecting methods, high metal prices, a favorable business outlook, increasing cooperation from large landholders, and some discoveries support the probability of increased exploration activity in Maine during the coming decade.

Jan 1, 1968

By John L. Miles

John L. Miles (Arthur D. Little, 1nc.)—Pollack and orris" have reported measurements on electron tunneling through A1-A12O3-A1 sandwiches in which the oxide was formed by gaseous oxidation in a glow discharge. From these measurements they deduced the asymmetry of the barrier and, since this is small, conclude that the mechanism suggested by Mott19 for the growth of oxide in thin A12O3 films is inapplicable. In earlier papers20 Pollack and Morris report similar work for oxide films grown thermally. In this case they find a greater asymmetry and conclude that the Mott mechanism is valid. I would like to point out that both these conclusions are quite unjustified. Mott suggests that the growth of the oxide film on aluminum results from the passage of ions through the already present film of oxide under the action of an electric field. This field results from a constant voltage which is in effect a contact potential between metal on one side of the barrier and adsorbed oxygen ions on the other side of the barrier. The theory does not require that the oxide grown is nonuniform either in stoichiometry or structure. It does however specifically assume that the partial layer of ionized oxygen on the surface remains adsorbed on the surface of the growing oxide. In other words, the so-called "built-in field" remains in the oxide only as long as the ionized oxygen is present. When a counter electrode of aluminum is deposited on the oxide, it will react with the adsorbed oxygen on the surface of the oxide, thus forming a small additional amount of oxide. It is clear, then, that there is no requirement in the Mott theory of oxide growth which would necessitate tunneling currents through an Al-A1203-A1 sample to be different when the polarity is reversed. Neither does the theory eliminate the possibility that some additional mechanism could cause the tunneling barrier to be asymmetric and hence tunneling currents to be a function of polarity in such a sandwich. Thus these tunneling-currents measurements are not germane to the question of whether the Mott mechanism is the true method of growth of aluminum oxide films. In fact, it is not surprising that there should be a difference between the oxide properties at the two interfaces (with resulting asymmetry in the tunneling barrier) since the growth conditions and growth rates must have been quite different at these two positions. S. R. Pollack and C. E. Morris (authors' reply)— The point raised by Miles above is one has caused some confusion in the past. The following is an attempt to clarify this point. The built-in field which is responsible for the growth of the thermal oxide at low temperatures arises, according to Mott, because of the passage of electrons from the Fermi surface of the oxidizing metal to surface states introduced by the adsorbed oxygen. It is assumed that the energy of these surface states lies below the Fermi energy of the metal. Electrons therefore continue to flow from the metal to the surface until the built-in electric field raises the potential energy of the surface states to the value of the Fermi energy in the metal, at which time equilibrium is obtained between the surface states and the metal. That is in equilibrium as many excess electrons pass from the metal to the surface per unit time as vice versa. The surface of the oxide prior to deposition of a metallic counterelectrode can then be pictured as follows. The Fermi energy lies in the energy gap of the oxide and is essentially pinned at the energy of the oxygen surface states. The vacuum work function of the oxide is then given by the sum of the electron affinity of the oxide (i.e., the difference in energy between the vacuum and the conduction-band minimum) plus the energy difference between the conduction-band minimum and the Fermi energy. The deposition of a metal onto the surface of the oxide can result in a transfer of electrons across the extremely thin oxide only if there is a contact potential difference between the deposited metal and the parent metal or oxide. That is if the vacuum work function of the deposited metal differs from that of the parent metal, then charge can be redistributed across the oxide in order to equilibriate the Fermi energy across the structure. (It should be

Jan 1, 1965

During the period January 1, 1977 through July 31, 1977, 5 percent (264,000 tons) of U.S. capacity was closed due to power curtailments in the Pacific Northwest because of the shortage of water. During July, one plant in Tennessee cut production by 50,000 tons annually due to a power reduction by the Tennessee Valley Authority. Full capacity production was resumed at a West Virginia smelter during September. If demand for aluminum is strong it seems likely that aluminum supply during the remainder of 1977 and in 1978 will be tight. The price was increased 3¢ per pound to 51¢ per pound effective March 28, 1977, and by an additional 2¢ per pound to 53¢ per pound, effective July 5, 1977. New projects and expansion of existing facilities are planned by Australian aluminum companies over the next 2 to 3 years. The companies listed have committed themselves to spending approximately A$801 million: Alcoa of Australia Ltd. -New alumina plant at Wagerup, south of Perth. Darling Ranges bauxite is located near the proposed project. A$650 million. Coma 1 co Industries Pty. Ltd. -Increase bauxite mining capacity and upgrade or replace other facilities at its Weipa mines on Cape York Peninsula region of Northern Queensland. A$81 million. Alcan Australia Ltd. -Expansion program that will increase capacity of the Kurri Kurri aluminum smelter north of Sydney about 50 percent, and improvement of other facilities at the plant. A$70 million.

Jan 1, 1977

By K. M. Kim, J. B. Kemeny

A case study was conducted with 5 shots with varying pre-blast block sizes and explosive energies. From this case study a site-specific predictive fragmentation model was developed. There are two main innovations in this study. One is using an image analysis program (Split-Desktop 3.0) on the bench face to obtain the pre-blast block size (F80) quickly and consistently. Another innovation is to utilize the Schmidt Hammer Hardness (SHH) to estimate intact rock strength.

Jan 1, 2011

By T. G. Digges, M. R. Achter

ALTHOUGH zone melting has found favor in recent years because of its convenience and its faster rate of production of single crystals, the older technique of strain annealing still has a number of advantages. It has been reported1 that the lattice of a crystal prepared in this manner has fewer defects than one grown from the melt and, of course, the control of the external geometry is easier. Also, the diameter of the crystal which can be grown by zone melting has a theoretical limit determined by the surface tension, thermal conductivity, and specific gravity of the material, while the size of that which can be grown in the solid state is limited only by practical considerations. In a program for the study of the structure-sensitive properties of niobium and its alloys it was desirable, for ease of handling, to have available single crystals of large diameter. This requirement dictated the choice of the strain-anneal technique. Induction heating was chosen in the present work in preference to other methods because of its deep penetration of large-diameter sections. A five-turn coil, powered by a motor generator operating at 10,000 cps, is contained in a vacuum chamber. The work is lowered through the coil for both the recrystallization and the crystal-growth steps. Starting with electron-beam melted material, the bars are cold-swaged, recrystallized, strained in a tensile machine, and then subjected to the growth step. Single crystals have been grown under the following conditions: Reduction in area during swaging— 79 to 99 pct, recrystallization temperature—1500o to 1667°C at the rate of 1.25 in. per hr, strain—1.0 to 2.5 pct, growth temperature—1600o to 2400°C at rates from 0.15 to 1.25 in. per hr. Starting grain sizes achieved after the recrystallization step varied from 0.6 to 1.55 mm. The starting grain size was found to be critical; if it is too fine and has any residual cold work, as shown by the presence of a fiberlike texture, a single crystal cannot be grown. Using this technique niobium single crystals have been grown from 0.250 in. diameter and 12 in. long to 1.00 in. diameter and 5 in. long as shown in Fig. 1. The 1-in. crystal is the largest to be reported for a refractory metal as compared to 0.500 in. for molybdenum grown by zone melting.2 As with zone melting, control of orientation is possible by the use of special procedures. Other investigators, see for example Williamson and Smallman,3 have reported that orientation control may be achieved by a bending technique. In the present work the strained rod is partially lowered through the coil to start the growth of the crystal. Then it is removed and bent at a point ahead of the interface between the single and polycrystalline portions. Finally, it is returned to the chamber and growth is continued "around the corner". The amount of bending which can be imparted to the rod is limited by coil geometry which up to now has been 10 deg. However, by repeating the bending and growing operations it should be possible to attain any desired orientation. The perfection of these single-crystal specimens as a function of process variables is being measured by X-ray and metallographic techniques and the results will be reported in a forthcoming paper. This work was partially supported by the Advanced Research Projects Agency.

Jan 1, 1964

By E. M. Hill, D. E. Walsh, J. T. Hollow, H. K. Lin

Fairbanks Gold Mining, Inc. (FGMI) began blending and processing True North ore at their Fort Knox mill on April 12, 2001.Gold extraction subsequently dropped from 88% to 71% during the last week of April 2001. 25% of the gold values remaining in the leach tailings were gravity recoverable free gold. Significant quantities of stibnite and arsenopyrite were also identified in the tailings. In laboratory bottle roll tests, the gold extraction from a composite of leach circuit feed collected during the same period was 72.6%, when leached at an average NaCN concentration of 0.16 g/kg. The extraction increased to 82.4% and 91.7% with 0.5 g/kg NaCN, and 0.5g/kg NaCN plus 0.075 g/kg Pb(NO3)2, respectively. A passivating layer surrounding liberated gold particles and resulting from sulfide mineral dissolution, likely contributes to reduced gold recovery, which can be substantially increased by the addition of lead nitrate. Lead species remain in the solid residues after leaching and were not mobilized using Meteoric Water Mobility Procedure.

Jan 1, 2003

By F. M. Benavides

The mining complex being analyzed consists of four mines and two processing plants located in an area where there are several existing modes of transportation. Figure 1 shows a map of the general layout of the four proposed mines and the two existing processing plants. The mines are identified by the numerals "I" through "IV" and the plants by the letters "A" and "B". A proposed overland conveyor would collect the product from all four mines and convey it to Processing Plant A. Alternate transportation systems are as follows: Railroad transport from Mines I and II to Plant B; Highway trucking from Mines I and II to Plant A; Off-the-road truck haulage from Mine III to Plant B and from Mine IV to Plant A. River transportation from the mines to the processing plants was not considered in this study under the assumption that the high capital investments for the loading and unloading facilities, in addition to difficult traffic problems, would make this mode of transportation unattractive.

Jan 1, 1980

By R. C. Adams

This report describes the feasibility of mining the oil-shale deposits of the Piceance Creek basin in Colorado by an integrated open pit system. The study is based upon a comprehensive analysis of pertinent environmental, economic, and engineering factors. Ten mining cases are investigated, and the mine design and cost estimates for each case are presented in this report. Nine of the cases represent three cutoff grades, defining oil shale ore as equal to and above 15 gallons per ton (gpt), 20 gpt, and 25 gpt, at three ore production levels of 500,000 tons per day (TPD), 1,250,000 TPD, and 2,500,000 TPD. The tenth case is an alternate design of the 500,000 TPD production rate at the above 15 gpt cutoff grade. The report consists of Volume I, a technical and cost comparison of the ten cases; Volume II-Part I, a detailed explanation of the estimating methods used; and Volume II-Part 2, a more detailed discussion of the technical and cost features of the ten cases than is presented in Volume I.

Jan 1, 1976

By J. S. Diller

In 1827, chromite was discovered near Baltimore by Isaac Tyson, Jr., who initiated the mining of chrome ore an:! later (1845) .the manufacture of chromium compounds in this country. From 1828 to about 1860 the United States, supplanting Russia, supplied chromite for most of the world. Chrome deposits were found also in California, and the Tyson Company shipped ore to Philadelphia by way of Cape Horn. In 1848, J. Lawrence Smith, an American in the employ of the Porte, discovered large bodies of chromite in Asia Minor. They were developed in 1865 and became the chief source of the world's supply for a number of years. The discovery of the large high-grade bodies in New Caledonia, and later in Rhodesia, with other foreign sources, gives at present an abundant supply for all demands. For a number of years before the great war, the United States produced only a few hundred tons of chromite annually, for local use, but during the war the demand for domestic ore resulted in a production of more than 82,000 long tons of crude ore in 1918, and it was demonstrated that the United States has reserve deposits equal to a war demand of several years' duration. Now that the war is over, the country is conserving its domestic supplies by employing higher-grade but cheaper ore from foreign countries. ChromiUm-bearing Minerals Oxides: Chromite-—chromic iron ore................ FeO.Cr2O3 Picotite—chromic iron ore................. (MgFe)O.(Al,Cr) Chromates: Crocoite—lead chromate................... PbCrO4 Phoenicochroite—basic lead ohromite........ 3PbO.2CrO3 Vauquelinite—phospho-chromate of lead..... 2(PbCu)CrOd.(PbCu)3P2O8 Sulfates: Redingtonite—hydrous chromium sulfate.... Knoxvillitel—hydrous basic sulfate of chromium.............................. Daubreelite-—iron-chromium sulfide (known in meteorites only)...................... FeS.Cr2S3

Jan 1, 1920