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The Sandvik DE710 diamond core drill has become one of the most popular exploration drills available. It is operated on exploration and mining projects in locations around the world. The compact nature of the DE710 allows a small footprint on site. This has proven to be a great benefit in difficult access areas where site preparation can be limited. The DE710 drill is available in a range of con-figurations to suit global operational and environ-mental demands. It can be truck-mounted for greater maneuverability and ease of relocating the drill between sites. The drill can also be fitted to crawler tracks, powered from the rig hydraulic system, making it a small, self-propelled unit for areas with difficult site access. And the DE710 is available in a shack-mounted configuration for cold weather applications, ensuring a protected working environment for the drill crew. The DE710 is powered by a 6.7-liter engine that produces 141 kW (190 hp) at 2,200 rpm to power the rig hydraulic system. The engine complies with the Tier 3 emissions standards that are being introduced in many countries. The hydraulic system design for the Sandvik DE710 is based on the company?s philosophy of selecting the best available products to suit the application. The main hydraulic pumps and motors are the variable displacement design. This allows for greater fl exibility and precise control of critical drilling functions. The cooling system has also been designed for operation in ambient temperatures up to 50° C (122° F) to cope with varying global operational demands.

Jan 1, 2009

By Peter G. Chamberlain

The surge in gold and silver prices during the 1970's attracted many new mining operators and has rekindled interest among experienced ones. With its low capital investment requirements and fast payout, leaching has attracted many operators--particularly those with small or low- grade deposits. Although in certain situations leaching offers many advantages over conventional mining methods, many operators are uncertain how these relatively new techniques should be implemented. Consequently the Bureau of Mines has prepared this circular to disseminate information on gold and silver leaching practices, techniques, and problem. Engineering data gathered from 26 operations indicate that most ores are leached in heaps following crushing and distribution on pads. Metal values are recovered from cyanide leach solutions using either zinc precipitation or charcoal adsorption. Potential problems that my hamper on block development of a leaching operation are poor percolation characteristics of the ore, calcium salt buildup, low temperatures, and solution losses. An extensive bibliography on gold and silver leaching is appended.

Jan 1, 1984

By David E. Nicholson

The Bureau of Mines has constructed a special compression chamber for developing stress-strain relationships of mine backfill and simulating the high compressive ground pressures encountered in underground mines. A one-dimensional, or Earth-Pressure-at-Rest model is proposed for backfills subjected to these pressures. Results of the compression tests performed in the chamber indicate that reliable stress-strain relationships can be developed from these tests. The primary objective of the test series was to evaluate the effectiveness of compaction of common backfills in retarding ground closure. Analysis of the stress-strain data shows an eightfold reduction in relative strain in optimum well-graded, compacted backfills over poorly graded, loose backfill material. Analysis of the one-dimensional compression test data also indicates a significant retardation of transmitted stress in dense compacted samples of backfill and natural sands. Although not fully understood, a deformation mechanism is thought to occur at the intergranular contacts in cyclic fashion causing frictional heat loss and a reduction in transmitted stresses in the densely packed materials. The experimental values of transmitted stress are compared with theoretical values of effective stress calculated from the one-dimensional compression relationships developed by Hendron (4).3 These experimental values of transmitted stress compare very closely with the theoretical values of effective stress.

Jan 1, 1968

By P. D. White, J. T. Moss, J. S. McNiel

Results of experimental work on the in situ combustion process were first published in this country in 1953' when Kuhn and Koch described results of a three-well test in Jefferson County, Okla. Shortly there-after Grant and Szasz2 described field studies in Nowata County, Okla., in which several modifications of the basic combustion process were investigated by injection of various oxidizing gas mixtures other than air. In the past five years speculation concerning this method of oil recovery has continued, with a number of interesting papers on various aspects of in situ combustion appearing in the literature3,4,5,6,7. However, only one of these' provided actual engineering information on pattern-type field experiments. The primary objectives of the Oklahoma three-well test were to check operation of specially designed electrical ignition equipment and evaluate problems associated with initiation of combustion in a natural oil sand reservoir. Immediately following successful completion of this test in the Pontotoc sand, plans were activated for a second field experiment to be conducted in the same sand a short distance away. The objective of this second field test1 as to establish a relationship between field observations and information derived from laboratory experimentation and computations. Secondary objectives of the test were to develop operating tech- niques and define operating problems, and to obtain information that would assist in the economic and engineering appraisals of the process. Careful planning was required since this test was designed with the hope of providing basic information on the characteristics and operation of a combustion drive. It was necessary to include test facilities which would not only permit precise control and measurement of the process variables but would also provide sufficient flexibility of control to insure that variables in the combustion process could be studied during the course of the test under a wide range of conditions. In order to accomplish these objectives elaborate instrumentation and several control wells were included in the design of the test installation. This type installation was felt to be well suited for an experimental study, but was in no respect intended as a counterpart of a commercial-scale operation. PRELIMINARY LABORATORY WORK An important objective of the field experiment, as stated previously, has been to obtain reliable data on operating variables which might be compared with similar data obtained in laboratory combustion tests. The value of successful correlations between quantities that can be measured in the laboratory and those obtained in a full-scale commercial venture has been well established for other industrial processes. The difficulty of following a process underground should make such correlations even more essential here. Also. approaching the field test phase of development of in situ combustion, it was considered desirable to precede design of test equipment with laboratory combustion studies under a variety of conditions to establish likely ranges of variables. These included temperature, air injection rate, frontal advance rate and produced gas composition.

By M. M. Gilkey

This report on water requirements for the New Mexico mineral industry includes data on industrial operations closely related to mineral production. It contains general information concerning public water supplies at five principal cities and a short discussion of water needs for agriculture. Background data are presented regarding the geography, hydrography, and water laws of New Mexico, as well as precipitation and evaporation characteristics of the State. Most of the information on source, quantity, distribution, cost, and treatment of water was obtained in interviews with company representatives. Water systems at 46 operations are illustrated by schematic waterflow diagrams. The study revealed that the New Mexico mineral industry in 1962 used about 16 billion gallons of new water and reused 152 billion gallons, a total usage of 168 billion gallons. Consumption amounted to 7.6 billion gallons. Of the 16 billion gallons of new water, 11.9 billion gallons was "self- supplied" from ground water sources, and 2.6 billion gallons was self supplied from surface sources. Approximately 1.5 billion gallons was purchased. Some of the water from company-owned wells is piped as far as 30 miles. Reported costs of the self-supplied new water range from 1 to 20 cents per 1,000 gallons and average 8 cents per 1,000 gallons for power and maintenance. Inadequacy of the supply of new water at many operations necessitates large-scale recirculation. Reported costs for recirculating water (power and maintenance) average 1.8 cents per 1,000 gallons. For the entire New Mexico mineral industry in 1962, the value of product was $42 per 1,000 gallons of new water intake and $88 per 1,000 gallons consumed. Analyses of 40 water samples, listed in the appendix, show the wide range in water quality at plants throughout the State. Projection of the 1962 water requirements of the mineral industry indicate that the total demand for new water will increase from the 16 billion gallons used in 1962 to 24 billion gallons in 1980, a 50 percent increase; and to 36 billion gallons in 2000, which is 125 percent more than the 1962 intake of new water.

Jan 1, 1965

By M. M. Gilkey

This report on water requirements for the New Mexico mineral industry includes data on industrial operations closely related to mineral production. It contains general information concerning public water supplies at five principal cities and a short discussion of water needs for agriculture. Background data are presented regarding the geography, hydrography, and water laws of New Mexico, as well as precipitatton and evaporation characteristics of the State. Most of the information on source, quantity, distribution, cost, and treatment of water was obtained in interviews with company representatives. Water systems at 46 operations are illustrated by schematic waterflow diagrams. The study revealed that the New Mexico mineral industry in 1962 used about 16 billion gallons of new water and reused 152 billion gallons, a total usage of 168 billion gallons. Consumption amounted to 7.6 billion gallons. Of the 16 billion gallons of new water, 11.9 billion gallons was "self- supplied" from ground water sources, and 2.6 billion gallons was self supplied from surface sources. Approximately 1.5 billion gallons was purchased. Some of the water from company-owned wells is piped as far as 30 miles. Reported costs of the self-supplied new water range from 1 to 20 cents per 1,000 gallons and average 8 cents per 1,000 gallons for power and maintenance. Inadequacy of the supply of new water at many operations necessitates large-scale recirculation. Reported costs for recirculating water (power and maintenance) average 1.8 cents per 1,000 gallons. For the entire New Mexico mineral industry in 1962, the value of product was $42 per 1,000 gallons of new water intake and $88 per 1,000 gallons consumed. Analyses of 40 water samples, listed in the appendix, show the wide range in water quality at plants throughout the State. Projection of the 1962 water requirements of the mineral industry indicate that the total demand for new water will increase from the 16 billion gallons used in 1962 to 24 billion gallons in 1980, a 50 percent increase; and to 36 billion gallons in 2000, which is 125 percent more than the 1962 intake of new water.

Jan 1, 1965

By R. G. Warncke, J. H. East, D. Harrington

The mining industry of the United States has long been severely criti- cized because of its high rate of accident occurrence, not only as compared with other industries of the United States but also with the mining industries of other countries, particularly Europe. Statistics, often not strictly com- parable, indicate that the accident-frequency rate in all mining is currently about three times and the accident-severity rate is about seven times as great as the average for general industrial work in the United States. These fig- ures are by no means comforting to humanitarian mining men, irrespective of their status in the industry, and they constitute a challenge that certainly cannot continue to be ignored much longer. The picture becomes much darker when the above statistics are augmented with data indicating that on an exposure basis our mines (more particularly our coal mines) have accident rates two, three, or even more times as great as more or less similar mines in Europe, particularly England, Wales, Belgium, France, and Germany. The sting is alleviated to some extent by the knowledge that our coal mines produce about 4 to 5 times as much coal per man day as is produced in at least some European mines. However, all of these foreign acci- dent statistics should be analyzed carefully before we become very much con- cerned about them, because they are not strictly comparable, for several reasons, not the least of which is the fact that foreign methods of compiling or recording accident statistics often differ materially from ours. Moreover, foreign mining and other conditions are definitely dissimilar to ours in many phases, and numerous other factors affect the comparability of foreign mine- accident statistics with those of the United States. Unquestionably, it should also be taken into consideration that the accident rate of the 140,000,000 or more constituting the populace of the United States as a whole is much higher than that of any other major country in the world and is two, three, or more times the rate of the European countries that have more favor- able mine-accident rates than those of the United States (table 1). Moreover, it is significant that the mine-accident-severity rate of British mines is about six times the average accident-severity rate of the major industries of the British Isles, or but little more favorable than the comparable figures in the United States; namely, that mining has an accident severity rate about seven times the average of about 30 of the major industries of this country.

Apr 1, 1949

By N. Shariatmadari

In recent years, determining bearing capacity of piles from in-situ testing data as a complement of static and dynamic analysis has been used by geotechnical engineers. In this paper, different approaches for estimating base bearing capacity of piles from SPT data has been compared and evaluated. A new method for pile base bearing capacity in sand is proposed using SPT records. Data averaging, failure zone extension, and plunging failure of piles have been addressed in the proposed approach. A data bank comprising 10 full scale static load tests on instrumented piles and 26 dynamic tests, analyzed with signal matching technique by CAPWAP software, was used for validation. The SPT data that was performed close to pile locations is also included in the data base. Comparison of current methods by error investigation with Log-Normal approach demonstrates that the proposed method predicts pile base capacity with more accuracy and less scatter in comparison to other methods.

Jan 1, 2006

By R. W. Nice

"The Similkameen Mining Company property is situated about 9 1/2 miles south and west of Princeton B.C. Owned and operated by Newmont Mining Corporation, mining is presently taking place in the Ingerbelle pit. At 15,000 tpd of 0.50% Cu ore the Ingerbelle pit is sufficient for about 8 years. Another 8 years ore supply is available from the old Granby workings of the Copper Mountain property which operated from 1927 to 1957.The pit operations commenced in late 1970 with the concentrator starting in March 1972. Designed and built by Bechtel Corporation under the supervision of Newmont the concentrator's design was based on single-stage crushing and single-stage modified autogenous grinding. The modification was installation of a crusher to crush the critical size build-up. The description of the operation follows.CRUSHERThe primary crusher is an open pocket Allis-Chalmers. 54"" x 74"" gyratory. The basic closed side setting is 8"". The crusher is fed from one side with 100 ton Electra-haul trucks. The crusher pocket discharges to a 96"" wide conveyor belt, then to a 48"" conveyor system including a stacker belt. The crushed product is piled on a stockpile with a capacity of about 30,000 live tons. This stockpile feeds the concentrator."

Jan 1, 1975

By Cathy Aimone-Martin, Caitlin Woznick, Brent Meins

This study is a continuation of previous work in which the responses of tall structures to close-in construction blasting at two rock excavation sites are evaluated. A total of eight buildings were instrumented to calculate blast-induced wall strains. Of these structures, five high-rise buildings measuring 166 ft (51 m) to 294 ft (90 m) in height (15 to 24 stories), were included for this study and compared with the results of previous studies that included one to three story residential structures and mid-rise structures (four to six stories). All five structures were built between 1908 and 1940 with interior walls comprising aged plaster. Three of the structures are designated Landmark, subject to a 0.5 in/s (12.7 mm/s) regulated peak particle velocity (PPV) limit. The regulated PPV limit for the two other structures was 2.0 in/s (50.8 mm/s). To date, a total of 450 close-in construction blasts have taken place over six months at the two adjacent projects, blasting three to six times per day at each site. Seismograph instrumentation included tri-axial geophones in the ground and bi-axial geophones (horizontal components only) mounted on structure exteriors at the corners closest to blasting. Peak global wall tensile strains were calculated from time-correlated differential structure displacements in two horizontal directions as measured by upper and lower geophones mounted near the roof and at street levels. Wall strains were correlated with ground excitations to establish a stiffness factor for each structure defined as the upper bound linear regression slope for strain as a function of ground displacement. In addition, a structure-specific regression best-fit function was established to relate horizontal components of displacement to velocity. These two functions were used to determine the maximum ground motion velocity for a desired limit placed on wall strain. A 25 ?-strain limit was suggested for both projects to protect interior plaster walls. To date, global wall strains have not exceeded 0.53 ?-strains (or 47 times the project limit) for peak particle velocities up to 3.49 in/s (88.6 mm/s). This demonstrates the low strain response of tall structures to ground velocities well above regulated limits. The stiffness factor computed for the most flexible high-rise building was 688 in-1 (17475 mm-1). In comparison, stiffness factors for the most flexible residential and mid-rise structures previously studied were 8651 in-1 (219735 mm-1) and 4387 in-1 (111430 mm-1), respectively. Based on a large number of measurements made at these study structures, the maximum horizontal ground velocity needed to achieve 25 ?-strains in walls ranged from 36 in/s (914 mm/s) to 192 in/s (4877 mm/s) because of the high stiffness of these buildings. Such high amplitudes of ground vibrations can never be safely achieved during close-in blasting in urban environments. These excessive values stand to show how unfounded the current limits of 0.5 in/s (12.7 mm/s) and 2.0 in/s (50.8 mm/s) are when applied to massive, tall structures. This study demonstrated that upper structure motions are highly attenuated from the lower (street-level) motions because of the long travel distances up the buildings. High-rise structures are far stiffer than mid-rise and residential structures and can withstand ground vibration amplitudes far greater than the currently enforced blasting limits in New York City.

Jan 1, 2015

By Wolfgang G. Brunner

The closing of the gap in the European waterway- net is one of the big and challenging traffic projects following the German reunification. The water canals in the North German plain connect the Rhine river, the border between Germany and France, with the Oder, the German border river to Poland. As early as in the 30's of the 19th century the planning and construction of a ship- lift and a canal bridge started, in order to close this last big gap in the system between the Mittelland Canal and the Elbe- Havel Canal near Magdeburg. This main waterway connects the West of Germany, known for its coal mines and steel mills, with the capital Berlin, and from there it goes on into Poland. The building activities started in 1934 but came to a complete stop in 1942 because of the war. The communist government in East Germany was not interested in continuing this huge project, including the ship lock Rothensee, the canal bridge crossing the river Elbe, the former border river between West and East Germany, and the lock Hohenwarthe due to lack of political backing along with low construction capacity. Now this important project is in progress and this report describes the special foundation techniques used on the great ship lock Hohenwarthe. Hohenwarthe is a double- bay ship lock of considerable scale, which is notable for recirculating water rather than letting it flow. It includes 44 800 m2 ( 481 720 sq. ft) of up to 55 m ( 180 ft) deep cut- off walls, ground water control, 1248 bored piles with a depth of 22 m ( 72 ft ) serving as foundation and 28 000 m ( 91 864 ft) of soil densification with vibroflotation equipment for ground improvement. In the year 2003 the project shall be completed and all ships can travel from the West to the East through Germany without expensive and time- consuming detours through other canals and locks.

Jan 1, 2000

By James R. Wait

The present final report constitutes the total productivity of a series of contracts from the U.S. Bureau of Mines and the Institute for Telecommunication Sciences over the period 1971-1979. The emphasis on the research has been analytical. In this "executive summary" we give a succinct description of each phase of the investigation. In the first phase of the investigation that was stimulated by the National Academy of Engineering, we concentrated on the analytical studies of techniques relevant to mine rescue. Several specific procedures for locating a buried source were analyzed. In the next, but overlapping phase of the investigation, we dealt with the fundemental character of how electromagnetic waves propagate in mine tunnels. At the outset we allowed for the finite conductivity of the tunnel walls and the presence of axial conductors such as cables, pipes, and tracks. The third phase of the investigation was devoted to the analyses of various specific geophysical engineering problems. The related fourth phase dealt with the general subject of non-destructive testing of wire ropes such as used for mine hoists.

Jan 1, 1980

By P. M. Lin

A subsidence monitoring program over a longwall panel was established (1) to explore the impacts of dynamic subsidence on the ground surface and structures and (2) to correlate the movements between the structures and their corresponding points on the ground. The program consisted of a transverse line survey and a series of survey points around the structures on the ground surface and their corresponding survey points on the exterior walls of the structures. The vertical displacement of each point can be expressed as: where Si is subsidence at i point; Sif is the final subsidence at i point; F is face location in terms of distance from the i point and SD is seam depth. The development of subsidence can be divided into five stages. Damage to the structures starts at the beginning of the second stage, The maximum damage developed in the middle of the third stage. The relationship of the vertical displacement between ground surface and structures can be expressed as: G = aH + b where G is ground displacement, H is structures displacement, and a and b are constants. The relationship is linear. However, the strain is non-linear and more complicated because of the interaction of the ground surface and structure.

Jan 1, 1989

By Jon H. Gould

This paper describes the following at the new Advanced Solid Rocket Motor (ASRM) nozzle manufacturing facility, which is being constructed inside an existing manufacturing facility: 1) generalized subsurface conditions, 2) pile load test program and evaluation, 3) pile selection, 4) problems associated with pile driving operations, and 5) production pile installation. The poor subsurface conditions, known and unknown underground obstructions, and work environment inside the existing building, used to manufacture the Space Shuttle external fuel tanks, required special pile installation equipment, procedures, and work schedules to install the economical pile foundation system ahead of schedule and without a lost time accident. GEOTECHNICAL STUDIES In April and May 1990, 16 soil test borings were drilled inside the existing building with a truck-mounted drill rig. The borings extended to depths ranging from 60 to 65 ft (18.3 to 19.8 m) below the floor slab. Typically, the silty and clayey soils were sampled by pushing "undisturbed" Shelby tubes, and the more sandy soils were sampled by driving a Standard Penetration Test (SPT) split-spoon sampler. The unconfined compressive strength, water content, and unit weight of the Shelby tube samples were measured in the laboratory. A geotechnical report was issued in May 1990 presenting the results of the field and laboratory testing and foundation recommendations (Gilbert, 1990).

Jan 1, 1992

By J E. Everett, D Linden, P Maney

In response to the implementation of a new schedule for iron ore fines in the International Maritime OrganizationÆs (IMO) International Maritime Solid Bulk Cargoes (IMSBC) Code, improved measures for the management of moisture have been developed. In particular, prediction of cargo moisture allows management of iron ore fines within the supply?chain to facilitate the safe?shipping of iron ore fines.An ARIMA model, augmented to allow for shipment tonnage, has been developed to predict the moisture level from previous shipments of the same product. The model explains about 60 per cent of the moisture variance. It was found that inclusion of other available information, such as ore composition, train/rake assays and recent rainfall at port and mine, did not add to the predictive power of the model.CITATION:Everett, J E, Linden, D and Maney, P, 2013. Predicting iron ore fines shipment moisture, in Proceedings Iron Ore 2013 , pp 17-26 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 12, 2013

By William J. Purdy

Successful design, operation and completion of a surface waste management facility for a diamond mine located in Canada?s northern environment requires careful planning to minimize short and long term environmental impacts. As part of the mine development application process, best practice standards were applied using conventional engineering to reduce the environmental effects of the mine waste facility at the Snap Lake Diamond Project. This paper presents the design criteria, site conditions, material characteristics and engineering management objectives considered for optimization of the development plan. Insight into the engineering analyses, development alternatives, monitoring program and reclamation strategy plan to be undertaken during mine operations to provide opportunity for adaptive management and to minimize environmental impacts is provided.

May 1, 2004

The central question concerning the health risks associated with mining and processing of taconite ores in the Eastern Mesabi Range is ??does asbestos, of any variety, occur within the gangue mineral population?? Free silica in the form of quartz does exist in the taconite, but silicosis has been relegated to a secondary issue. The focus has shifted to the asbestos minerals and the malignancies that might occur following low-level exposure to dust. Although mineral fibers do exist in the Mesabi taconite ores, Ross, Nolan and Nord conclude ??no asbestos?, as explained in their contribution ??The Search for Asbestos within the Peter Mitchell Taconite Iron Ore Mine, near Babbitt, Minnesota?. Their area of study is located in the eastern Mesabi Range. It is here that the tectonic and metamorphic effects associated with the intrusion of the Duluth Gabbro have been profound. A variety of amphibole minerals have developed as a result of these geological events, some with names recognized within the US Asbestos Standard. The issue tackled by the authors is whether or not these amphiboles are present in the mine and if so whether or not they are asbestiform in habit.

Jan 1, 2008

By D. Mohapatra

The possibility of using 2,2,4-trimethyl-1,3-pentanediol dissolved in tetrachloromethane as extractant for boron recovery is discussed. The parameters optimized for this purpose are contact time, boron concentration in the aqueous phase, extractant concentration, equilibrium pH and phase ratio. NaOH is used as stripping agent for boron recovery from the loaded organic phase. Boron extraction is found to be a slow process (equilibrium time 20 minutes) compared to boron stripping, which is reasonably fast (equilibrium time 5 min). Maximum boron extraction is possible at a broad pH range from 1.0 to 6.0. Using 1.0M extractant concentration, a 1:1 phase ratio (organic: aqueous) and two stages, ~99% boron extraction is achieved. More than 97% boron stripping is possible using 0.5M NaOH at a 1:1 (organic: aqueous) phase ratio in two stages. Overall, a 95.6% boron recovery was achieved. Loss of solvent into the aqueous phase and co-extraction of Ca, Na and K into the organic phase at higher equilibrium pH are two major concerns in his process.

Jan 1, 2008

By J. E. Campbell, A. W. Wesner, M. Pobereskin

This paper summarizes the results of an investigation sponsored by The Coates Steel Products Company. The objective was to determine the relative effects of certain chemical and physical properties on the wear resistance of grinding balls when grinding cement. In making this study, a number of types of balls were used in both dry and wet grinding in commercial ball mills. The radioactive-tracer technique was used in separating the test balls from the rest of the mill charge and in sorting the different types of halls that were all tested at the same time for each of the grinding operations. PROCESSING OF GRINDING BALLS In order to determine whether the properties of grinding balls, namely, composition, microstructure and hardness, can be related to the rate of wear, the halls used in this study were selected to represent various combinations of these properties. It was recognized that the mechanism of wear was not the same in the wet mill as in the dry mill. Therefore, the types of balls tested in each of the two environments were not necessarily the same. However, all of the types of balls evaluated in the dry mill were tested simultaneously for one time period, while those evaluated in the wet mill were tested simultaneously for three different time periods. All of the test balls were separated from the charge when the mill was dumped. All of the balls were of 0.940-inch diameter when charged, and all but the regular production balls were finish ground to this size to eliminate the decarburized surface layers. The steels and processing methods used, as well as the micro-structures and hardness levels of the balls, are described as follows: Regular production forged SAE 1095 steel balls were used in both the wet and dry mills. They were forged to a nominal diameter of 0.940 inch and quenched in oil from the forging temperature. Each retained a partially decarburized surface layer from the forging bar stock; the maximum depth of this layer was about 0.030 inch and the decarburization was negligible where the ball was sheared from the bar. The hardness at the surface was 40 to 46 Rock-well C in the partially decarburized area and 57 to 62 Rockwell C in the areas not decarburized. The hardened zone extended to a depth of about one-third of the radius. Below the decarburized layer, the hardened zone consisted primarily of hard as-quenched martensite with varying amounts of retained austenite. This structure changed to fine pearlite at a depth of about one-third of the radius. In order to eliminate the effect of the decarburized layer, a set of production balls of SAE 1095 steel was forged to about a 1-inch diameter, quenched in oil from the forging temperature, and ground on ball-bearing grinding equipment to a 0.940-inch diameter to remove the decarburized layer. The balls' hardened zones were similar to those in the previous set. The hardness of the finished surfaces was 52 to 60 Rockwell C. A slight amount of decarburization apparently remained after the removal of about 0.030 inch from the surface. Balls of this type were tested in the dry mill only. Another set of forged SAE 1095 steel balls of 1-inch diameter was heated to 1500 F and quenched

Jan 1, 1961

By E. D. Gardner

Open-cut mining of copper ores in the western United States is a relatively new industry. At first the mining methods were mainly adapted from those used at open-cut iron mines and at quarries, but now efficient practices that are more suitable for the copper deposits have been developed. The pioneer in this field is the Utah Copper Co., of Bingham, Utah, which deserves great credit for working out much of the early procedure. Stripping began at that mine in 1906, and the large Garfield concentrator was put in commission in 1908. Next in order of development were the Nevada Consolidated Copper Co., at Ely, Nev. ; the Chino mine of the Ray Consolidated Copper Co., at Santa Rita, N. Mex. ; the New Cornelia Copper Co., at Ajo, Ariz. ; the Sacramento Hill mine of the Copper Queen Branch, Phelps Dodge Corporation, at Bisbee, Ariz.; and the open cut of the United Verde Copper Co., at Jerome, Ariz. The Sacramento Hill mine began to produce in 1923, and the United Verde open-cut mine has produced ore mainly in connection with stripping. The writer visited the Utah Copper mine in September, 1923; the Nevada Consolidated in August, 1923; the Chino in July, 1924; and the New Cornelia, the Sacramento Hill, and the United Verde open cut at various times during 1922, 1923, 1924, and 1925, mostly in connection with other work.

Jan 1, 1927