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By Tobias W. Goodman

The U.S. Bureau of Mines used a contact drilling strategy with short-collared assemblies in order to rotary drill long horizontal methane drainage holes in the Beckley Coalbed near Glen Daniel, WV. By decreasing the thrust, and increasing the rotation of the drill bit when in contact with the roof and floor rock, assemblies with 1-ft (0.3-m), 10-ft (3- m), and 144 (43-m) long collars were made in deflect and stay in the coalbed. Successful application of this procedure resulted in holes drilled full length in coal with the original assembly including bit. The test assemblies deflected away from the original bearings of the holes during drilling and arced. Four holes were drilled to depths ranging from 478 ft (145.7 m) before the strategy was developed, to 4,034 ft (1,230 m) when the strategy was applied. Total length of the holes in coal was 8,590 ft (2,618 m). After 577 days the total methane captured by the holes was 221,600,000 ft? (6,276,000 m3). The methane flow rate peaked at 805,000 ft3/d (22,798 m3/d), then decreased to stabilize at 670,000 ft?/d (18,974 m3/d) after 6 months.

Jan 1, 1989

By R. A. Jankowski, J. D. Potts

Four underground respirable dust surveys were conducted to determine factors affecting belt entry dust levels and how using belt air to ventilate work areas affected dust exposures. Belt entry dust levels on the surveyed longwall and continuous miner section averaged 0.59 and 0.26 mg/m3, respectively. The stageloader/crusher contributed an additional 0.5 to 0.9 mg/m3 of dust to belt air, while the feederbreaker contributed 0 to 0.2 mg/m3 of dust. A 1000 ft (305 m) increase in belt entry length or a 200 to 500 tons/shift (181,000 to 454,000 kg/shift) increase in production resulted in roughly a 0.1 mg/m3 increase in dust. Using the belt entry as an intake entry on the continuous miner section appeared to reduce dust levels by 0.1 to 0.3 mg/m3 during cutting. Belt air was not used to ventilate the face on the longwall section.

Jan 1, 1992

By B M. Washington, R R. Lovel, G J. Sparrow, M J. Fisher-White

"Work in CSIRO with several Australian ores has shown that?by heating a high-phosphorus goethitic iron ore and then leaching with caustic or acid, the phosphorus levels in the ore can be lowered to a commercially acceptable value of around 0.08 weight per cent phosphorus (P). With increasing heating temperatures up to 1300¦C, the extent of phosphorus removal increased. This is considered to be associated with mineralogical changes in the iron oxide phases in the ore during the heating.Little phosphorus was removed from four high-phosphorus goethitic iron ores containing 0.12 - 0.16 weight per cent P in leaches with boiling water. However, leaches with sulfuric acid removed ten to 20 weight per cent of the phosphorus and leaching with sodium hydroxide removed even more phosphorus to produce leach products containing between 0.08 and 0.12 weight per cent P from three of the ores.After heating at 300¦C, leaches at a low pulp density of 0.5 weight per cent solids with 1 - 5 M NaOH and 0.1 - 1 M H2SO4 gave leach products with 0.06 - 0.09 weight per cent P. The lixiviant concentration required and extent of phosphorus removal depended on the ore. At a higher pulp density of 25 weight per cent solids, leach residues with between 0.07 - 0.14 weight per cent P were obtained.More phosphorus was removed from ores heated at 900¦C. Leaches at both the low and high pulp densities, with 1 - 5 M NaOH and 0.1 - 1 M H2SO4, gave leach products with 0.05 - 0.09 weight per cent P. However, the lowest phosphorus levels were obtained after heating the ores at 1200 or 1300¦C and leaching with 1 M NaOH or 0.1 M H2SO4. Leach residues with 0.02 - 0.06 weight per cent P were obtained, even from leaches at the higher pulp density of 25 weight per cent solids. Leaching tests on heated samples under heap leaching conditions at ambient temperatures have shown that phosphorus levels can be reduced to similar levels to that achieved in an agitation leach within ten days.CITATION:Fisher-White, M J, Lovel, R R, Sparrow, G J and Washington, B M, 2013. Effect of temperature on removal of phosphorus from Australian Goethitic iron ores, in Proceedings Iron Ore 2013, pp 373-380 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Aug 12, 2013

By Q. Zhang

The necessary capability of the gravity transportation pipeline system for cemented coal gangue backfilling at the Suncun Coal Mine, China, is about 74.4 m3?h-1 according to mining production and working times in the mine. It is estimated that the working velocity and maximum pipe length-backfill depth ratio of the backfill system are 3.82m?s-1 and 5.4, respectively and that the probable backfill capability of the system is 110.18m3?h-1. Results of hydraulic calculations for the backfill transportation system show that the designed backfill system is reliable in flowability and sufficient for backfill capability needs.

May 1, 2007

By D. J. Fray, F. Tailoka

Microporous ceramic tubes (15 micrometre pore diameter) and PVC membranes (10 micrometre diameter) have been used as spargers in the electrowinning of copper from concentrated and dilute chloride electrolytes. At a current density of 600 A/m2 and a nitrogen gas sparging flow rate (at NTP) of 1.53 cm3/s per unit electrode length, it was possible to electrowin planar rather than powdery deposits with copper concentrations as low as 0.157 M. To obtain a planar deposit at higher concentrations of cupric chloride the most realistic optimum current density appears to be between 1 000 and 1 500 A/m2.

Jan 4, 1993

By J. A. Lombardi

Innovative hard rock narrow-vein, longwall, and stope leach mining systems concepts are described and evaluated economically. The principal tool proposed for use in these systems for both mine development and production is the U.S. Bureau of Mines radial-axial splitter. The mining systems concepts are prefeasibility level costed and projected to be cost competitive for steeply dipping veins, 0.6-1.2 m wide, for horizontal tabular ores 2 m -5 m thick, and for leachable ores with greater than a 1.7:1 stripping ratio. In all of the systems the orebody, less swell, is retained underground for use as backfill.

Jan 1, 1994

By Yoginder P. Chugh

The control of soils has been found to be one of the highest priority factors influencing the return of reclaimed prime farmland to production. The compaction of soil is primarily a result of the movement of high ground pressure wheel scrapers and dozers over soils during the soils replacement and regarding process and occurs to a maximum depth of about 1.2 m (4 ft) The deleterious effects of soil compaction can be controlled by deep soil loosening to a depth of 1.2 m (4 ft). Most farm tillage equipment can only loosen soils to a maximum depth of 0.33 m (1.0 ft). Deeper soil loosening equipment, in the 0.33-1.2 m (1.0-4.0 ft) range is currently avail- able in the U.S. and abroad which may have potential for application on reclaimed prime farmland. This study has appraised such available equipment and the results indicate that a few manufacturers in the U.S., West Germany, and England market static and vibratory soil loosening equipment which have potential for deeper soil loosening of reclaimed prime farmland.

Jan 1, 1985

By P. G. Carpenter, R. G. Nisle, F. W. Bubb

It is the purpose of this paper to describe an electrical computer which has been constructed to solve the equations for vapor-liquid equilibrium in multi-component systems. The instrument consists of seven component-computing units each with proper indicating means and power supplies. Each unit is a resistance network with a voltage matching servomechanism, and each provides an output voltage proportional to the mol-fractions for vapor and liquid phases. These voltages are summed and matched with a reference voltage to provide the solutions. Any reasonable number of such units may be put together to make a computer. The theory and operation of the computer is discussed. A number of applications and examples of computer results are given. The computer yields the over-all vapor or liquid fraction to a probable error of 0.002. An interpolation method is described which reduces the probable error to 0.0002. INTRODUCTION The process known as Flash Vaporization may be described as follows: A mixture of known composition of relatively volatile components is allowed to come to thermodynamic equilibrium at some given temperature and pressure by any path whatsoever. In general, a vapor and a liquid phase will be present at equilbrium. The problem is to determine the fractions of the mixture in the vapor and liquid phases, and to determine the mol fractions of the various components in each phase. The relations that exist between these various quantities at equilibrium are well known and will be given later. These equations are difficult to solve, yielding only to trial-and-error methods. The virtues of the analog computer are its speed and the automatic character of its calculations. The device is operated by turning a crank which varies the value of v, the total vapor fraction. This actuates a number of servomechanisms which perform automatically all computations. This computer facilitates the solution of such problems as (to mention a few] : Analysis of separator operation, studies of changes in composition of reservoir fluids with pressure decline, and analysis of natural gasoline plant operation. Examples of some of these problems are given in detail later. STATEMENT OF THE PROBLEM The essential equations are xm = 1+(km-l)v . . (1) Kmzm ym =------------------------. . (2) 1 + (Km-l)v ; n n S S xm= --------------------=1 1 + (Km-l)v m = 1 m=1 ......(3) n n S = S Km Zm =1 S ym 1 + (Km-1)v m=1 m = l ......(4) where ym = Vm/V = Mol fraction of m-th component in vapor phase. xm = Lm/L = Mol fraction of m-th component in liquid phase. Zm = Fm/F = Mol fraction of m-th component in mixture. KM = ym/xm = Equilibrium constant for m-th component at the given temperature and pressure. v = V/F = Mol fraction of vapor in the mixture. F = Total mols of a mixture of n components. V - Total mols of vapor in the mixture. L = Total mols of liquid in the mixture. Fm = Total mols of m-th component. Lm = Mols of m-th component in liquid phase. Vm = Mols of m-th component in vapor phase. Assuming the total mols of the mixture F and its composition (all the Fm) to be known — either from a quantitative analysis or from the amounts put together to make the mixture — the zm = Fm/F may be regarded as known quantities. From the given temperature and pressure of the mixture the K., are known (principally from experimental data). The primary problem is then: having given all the zm and Km to calculate v and all the fractions xm and ym. Having calculated these ratios, it is a simple matter to calculate V. L. Vm, Lm from their defining equations. SOLUTION OF THE PROBLEM Basic Computer Unit The computer consists of n units, each of which provides voltages xEo and YE, proportional to x and y as in

Jan 1, 1950

By Bernhard Bruggaier

Outside the oil and gas industry there is an increasing demand to install big piles for the foundation of ports, bridges, and wind-energy plants. This article introduces the design and techniques in using large MENCK pile driving hammers to install big foundation piles. Piles with a length up to 120 m, an outside diameter of 1.0 m to 3.15 m and weighing between 25t and 400t are being driven for these types of foundations. The required energy lies between 200 and 1700 kNm. Pile driving hammers that bring this amount of energy weigh from 38 to 300t and have a length between 10 and 20 m. Handling these big pile driving hammers requires a crawler crane on land or cranes on floating pontoons or jack-ups when working in water. Possible are either vertical or inclined piles, using a leader for inclinations up to 1:1.7 or up to 1:3 in a free riding version. Project descriptions demonstrate the versatility and practical use of the equipment under diverse operating conditions.

Jan 1, 2002

By Richard E. Sabaski

The demand placed on the mining industry in the past 15 years for increased production, coupled with lower grade ore and higher cost capital, has resulted in a noticeable increase in grinding mill sizes. The pressure to increase mill size continues. During the mid sixty's and early seventy's, ?larger? conventional ball mills were in the range of 13 - 16-1/2 f t. diameter (4 - 5 m) increasing by the mid 70's to 18 ft. (5.5 m) diameter. At the same time, autogenous mills expanded from 24 - 32 ft. (7.3 - 9.1 m) to the present 34 and 36 ft. (10.b and 11.0 m) diameters. These size pioneers of 16-1/2, 18, and 36 ft. diameter mills from Allis-Chalmers have been free from structural problems, because they were analyzed, modeled, designed, and field tested to meet the challenge.

Jan 1, 1981

By Ralph Musick, Brian Briggs

The U1h Shaft Project is a design/build subcontract to construct one 6.1 meter(m) (20 feet [ft.]) finished diameter shaft to a depth of 321.6 meters (m) (1055 ft.) at the Nevada Test Site (NTS). Atkinson Construction was subcontracted by Bechtel Nevada(BN) to construct the U1h Shaft for the Department of Energy (DOE). The project consists of furnishing and installing the sinking plant, construction of the 321.6 m(1055 ft.) of concrete lined shaft, development of a shaft station at a depth of 297.5 m(976 ft.), and construction of a loading pocket at the station. The outfitting of the shaft and installation of a new hoist may be incorporated into the project at a later date. This paper will describe the design phase, the excavation and lining operation, shaft station construction and the contractual challenges encountered on this project.

Jan 1, 2001

By Cheowchan Leelasukseree

Mae Moh mine, operated by Electricity Generation Authority of Thailand (EGAT), is the largest open pit mine in Southeast Asia. Its production is approximately 15 million lignite tons and 90 million m3 of waste, annually. The bottom of the final pit limit is designed to be approximately 500 m in depth. Presently, the pit depth is almost 250 m below original surface. The complexity of geological structures, such as faults, has caused a number of geotechnical difficulties. In a challenging lowwall area, called Area 4.1, an inclined weak bedding plane, G1, will be daylighted along the lowwall as following by mine plans. The dimensions of Area 4.1 are nearly 300 m wide and 90 m high. The friction and dip angles of the G1 interface are reportedly 12° and 15°, respectively. A three million m3 claystone block, resting above G1, would potentially slide down, and possibly cause a massive failure. EGAT engineers are acutely aware of the potential gigantic sliding failure and are modifying the mine plan in Area 4.1 to improve mine safety and minimize the displacement of the potential sliding block, reduce the possibility of failure and lower the additional mining cost as much as possible. The cut and fill mining technique was selected and studied to achieve those goals for mining in Area 4.1. With this mining technique, G1 will be partially daylighted. Therefore, an arch formation is induced with a sliding block above the mining area, which could perform as a temporary support undercutting a certain span. Initial studies of the behavior of the partially undercut slope are presented in this paper, along with a review of the similar failure history in Mae Moh mine and a 3-dimensional numerical model of the mine. The failure history review provides an estimation of undercut span without failure. The three-dimensional numerical model results, using 3DEC, gave an initial approximation of the displacement of the sliding block in Area 4.1. The next exciting study of the behavior of partially undercut slope is a trial undercut slope field experiment in Area 4.1. Currently, a trial undercut slope field experiment in Area 4.1 is under construction. In the field experiment, a number of monitoring programs will be installed in Area 4.1 to monitor and obtain information to allow a better understanding of the behavior.

Jun 1, 2012

By P. E. Weidmark

"PART I - Published PapersBallivy, G., and Dupuis, M., ""Laboratory testing for the design of grouted rock anchors""; Proc 6th Pan Am Conf Soil Mech Found Eng; Lima; 1:277-291; 1979.Ballivy, G., Ladanyi, B., and Gill, D., ""The effects of moisture content on the strength of fissured rocks""; Proc 1// 1ntnl Congr 1.A.E.G.; Madrid; //:2:7-16; 1978.Baumgartner, P., and Stimpson, B., ""Development of a tiltable base friction frame for kinematic studies of caving at various depths""; Int J Rock Mech Min Sci Geomech Abstr; 16:4:265-267; 1979.Bawden, W.F., and Roegiers, J.-c., ""Discussion and interpretation of the mathematics governing radial flow in a single horizontal rock fracture""; ISBN 07727-7002-6; Dept Civil Engng publication 79-11; Univ Toronto: 134p;1979.Chatterji, P.K., Smith, L.B., and Insley, A.E., ""Construction of Saline Creek Tunnel in Athabasca oil sand""; Can Geotech J; 16:1:90-107: 1979.Coates, D.F., and Gyenge, M., ""A case history of pit slope design""; Proc 4th ISRM Congr; Montreux, Switzerland; Sept 1979; 1:Theme IV:591-596; 1979.Crawford, A.M., and Curran, J.H., ""A dynamic direct shear machine for determining the time-dependent behaviour of rock joints"": Dept Civil Engng publication 79-08; Univ Toronto; 1979.Crawford, A.M., and Curran, J.H., ""A displacement discontinuity approach to modelling the creep behavior of rock and its discontinuities""; Dept Civil Engng publication 79-09; Univ Toronto; 1979.Crawford, A.M., and Curran, J.H., ""A higher-order displacement discontinuity element""; Dept Civil Engng publication 79-10; Univ Toronto; 1979.Cruden, D.M., ""A computer-based system for rock mass description""; Proc 1// Intnl Congr I.A.E.G.; Madrid; //:2:155-159; 1978.Cruikshank, D.J., and Curran, J.H., ""A.B.LE. approach to modelling of fluid flow in fractured rock masses""; Proc 7th Cdn Congr Appl Mech; May 1979; Sherbrooke, Que; 887-888; 1979."

Jan 1, 1980

By Peter van Dijk, Gary Almeraris, Philip Rice

This paper describes the Contractor’s value engineering design and the construction of the Interstate I-90 highway tunnel sections under the South Station railroad yard in Boston, MA for the Central Artery/Tunnel Project, using tunnel jacking and ground freezing techniques. Three tunnel box sections were constructed and installed, with typical cross-sectional dimensions of 11.6 m x 23.8 m, and respective lengths of 50.9 m, 78.7 m and 115.6 m. A 12-compartment concrete shield was built as part of the front of each multi-sectional tunnel box structure. The jacking systems employed consisted of rear and intermediate jacking stations and up to 87 individual jacks, providing a maximum total thrust capacity of 42,300 t. The tunnel alignments pass through a geologic profile consisting of historic fill, which contains remnants of old waterfront structures, building foundations and a forest of 19th century vintage timber piles, underlain by organic sediments and marine clay. Ground cover over the tunnels is in the range of 1.8–7.6 m. The groundwater table is at depths of 1.5–3 m below surface grade. The paper discusses the techniques developed and equipment used to successfully build and install these tunnel box structures with minimal impact on the commuter and long distance railroad operations directly above the tunnel alignments. Figure 1 shows the track area at the I-90 EB and WB tunnels.

Jan 1, 2001

By Erik Bengtsson

Phenol was discovered in coal tar by Runge (1)1/ in 1834 and was isolated (2) in 1841 as pure phenol by distilling the tar, extracting with caustic soda, acidifying the extract, and fractionating the tar acids. Later, o-cresol was isolated (3) through careful fractional distillation of the tar-acid mixture. The m-and p-cresols distilled within a boiling range of about 1.50 c. and had to be separated by other means. Raschig (4) fractionated the tar-acid mixture to obtain pure o-cresol, sulfonated the mixture of m-and p-cresols, and treated the cresol-sulfonic acids with superheated steam at 1200 to 1300 c. to hydrolyze the m-cresol sulfonic acid, the p-isomer was not affected at that temperature. He also attempted to make use of the different crystallizabilities of the m-and p-cresol sulfonic acids (5). When treated with sodium bisulfate and sodium pyro-sulfate at 100° to 110° C. or with sulfuric acid (90/0) at 20° C., only the m-cresol was sulfonated (6). Hoffman-La Roche (7) treated the meta-para mixture with concentrated sulfuric acid at 40° to 100° C., when m-cresol and part of the p-cresol was sulfonated, and precipitated the m-cresol sulfonic add by diluting with water. The same sort of separation may be carried out by means of 80 percent sulfuric acid at 35° C. and 35 mm pressure (8).

Jan 1, 1950

By Gordon M. Pugh, David G. Rasmussen

INTRODUCTION The two unit operations in cut-and-fill stoping that would benefit most from increased mechanization are drilling blastholes and moving broken ore to the ore¬passes. Drilling in current practice is done with hand¬held jackleg and stoper drills. Broken ore is commonly moved from the face to the orepass by a small slusher. This type of equipment can be easily moved through the raises. Hand-held drills have a slow drilling rate and require one operator for each machine. Drilling can be done much faster by small jumbo-mounted machines as two or more drills can be operated by one miner. Load-haul-dump (LHD) units move broken ore from the face to one or more orepasses at a much faster rate with less loss and dilution than slushers. These units are mobile and may be used at more than one place during a shift. The problem which must be overcome in applying this higher degree of mechanization is the restricted means of access inherent in current practices. Cut-and¬fill methods are usually applied to mining veins that are narrow, irregular in dip and strike, or where ground con¬ditions require the support of sand fill. The mechanized equipment which is needed to accomplish high produc¬tion cannot be raised or lowered through raises without being disassembled. A jumbo which will give the desired performance would be mounted on a rubber-tired carriage with a diesel engine for propulsion. Twin hydraulic booms run by a small air motor would carry pneumatic drills. Over¬all dimensions of the unit would be about 8 m (26 ft) in length, 1.6 m (5'/z ft) in width, and 1.6 m (5'/z ft) in height. An LHD of the desired capacity would be rubber¬tired and either diesel or electric-powered. The overall dimensions of a 2.3-m (3-cu yd) unit will be about 8 m (26 ft) in length, 1.8 m (6 ft) in width, and 1.5 m (5 ft) in height. It will weigh about 12 700 kg (28,000 1b). It can readily be seen that conventional raises can¬not handle this size equipment. The most desirable means of access enables equipment to be driven into the stoping areas. Ramps can enter the stope by being driven upward in the vein walls or by being developed in the vein on top of the fill as the stope advances. This ramp system not only permits the use of rubber-tired equipment in multiple working places but also enables its removal for repairs without component breakdown. In the proposed method, as in conventional cut-and¬fill stoping, sand fill will replace the mined-out ore. By constructing an inclined timber floor at one end of the stope, a ramp can be maintained on top of the fill and advanced as the stope is advanced cut by cut (Fig. 1). The ramp is, therefore, developed during the stoping process and is only driven as a tunnel when rising over and around the sill and crown pillars. A stoping section consists of three blocks of different geometric shapes. The first and last blocks to be mined will be triangular in the plane of the vein, and the blocks between will have the shape of parallelograms (Fig. 1). The first block will develop the initial ramp by breast stoping and filling to the ramp floor (Fig. 2). The ramp will need to be completed through to the next level be¬fore stoping of the subsequent block can begin. The subsequent blocks are mined by back stoping with the next ramp being developed at the other end of that block (Fig. 3). The length of each subsequent block can be varied without changing the slope of the ramp. As the new ramp is developed, the previous ramp is filled along with the stope cut. Access up and down is maintained through the stope. The final block is again breast stoped as no new ramp will be developed at the end line for jumbo egress. Boreholes provide ventilation between levels as well as utility access. They can also provide a secondary escapeway for personnel during periods of development. These boreholes would be located about 3 to 4.5 m (10 to 15 ft) in the footwall and spaced about 183 m (600 ft) apart along the vein. The orepasses are spaced about 61 m (200 ft) apart in the vein. This spacing allows not only a short tram for the LHD unit but also provides surge capacity for the ore between haulage trains. These passes are raised from the haulage levels. As development progresses, the raises are driven through the sill pillar, and then car¬ried in the fill as stoping progresses. Ventilation and utilities can be maintained through an orepass when it is not needed for ore. The end line orepasses are intended for this purpose as well as for personnel emergency egress. Overall dimensions and criteria assumed to illustrate the ramp method of access in a highly mechanized cut¬and-fill stope are: haulage level spacing of 61 m (200 ft); ramp gradient of 17%; average dip of approximately 1.57 rad (90°); initial stope length of 358 m (1177 ft); subsequent stope length of 179 m (588 ft); average stoping width of 3 m (10 ft) ; sill pillar thickness of 3.6 m (12 ft); crown pillar thickness of 1.8 m (6 ft); bored raises of 1.5-m (5-ft) diam; and cribbed orepasses of 1.2 x 1.2 m (4 x 4 ft) through the sill pillar and 1.2 m (4 ft) square inside the cribbed portion. DEVELOPMENT The ramp access system lends itself well to conduct¬ing the several phases of development concurrently. Haulage Drifts and Raises The first development step will be to drive the 3 x 3-m (10 x 10-ft) haulage drifts above and below

Jan 1, 1982

By K. J. Reid, I. Iwasaki, T. Sabelin

Introduction Occurrences of platinum group minerals (PGM) in mafic and ultramafic rocks in association with Cu-Ni mineralization are well known. Yet for the Duluth Complex, a large body of dominantly mafic igneous rocks in northeastern Minnesota with significant Cu-Ni mineralization (Bonnichsen, 1972; Boucher, 1975; Watowich, 1978; Weiblen and Morey, 1980), little has been published concerning its platinum group element (PGE) contents and mineralogy. In a previous study of massive sulfide rock, sperrylite (PtAs2) was found in nickel arsenide and recovery of arsenide minerals by flotation has been investigated (Iwasaki et al., 1986; Ryan and Weiblen, 1984). Recently, an announcement was made of the discovery of significant quantities of platinum and palladium [peak assay values of 9 g/t (0.27 oz per st)] in an interval of a drill core with disseminated sulfide mineralization 16 km (10 miles) southeast of Ely, MN (Anon, 1985; Sabelin, 1985). This paper describes PGM from the Pt-Pd-enriched zone and evaluates their potential for concentration. Geologic setting The Pt-Pd zone occurs in the South Kawishiwi in¬trusion near its basal contact at 734 to 737 m (2408 to 2418 ft) depth. The South Kawishiwi is one of several minor mafic intrusions that constitute the western margin of the Duluth Complex near Ely, MN (Fig. 1). It is com¬prised of a 70.1- to 350.5-m (230- to 1150-ft) thick sul¬fide-bearing basal unit of troctolite with minor amounts of anorthosite, norite, picrite, and oxide layers overlain by a troctolite unit with interlayered anorthosite (Bonnichsen et al., 1980; Foose, 1984; and Foose and Weiblen, 1986). The basal unit, in which the Pt-Pd zone is found, is markedly heterogeneous in rock type, lacking any coherent internal stratigraphy, and shows evidence of both pre- and post-consolidation faulting (Foose and Weiblen, 1986). Geochemistry Chemical analyses were made on 0.3-m (1-ft) intervals of crushed drill core material (1/4 core, BX size) at 731.5 to 740.7 m (2400 to 2430 ft) depth and the results are presented in Fig. 2. The analytical techniques used were atomic absorption spectrometry for silver, chromium, copper, and nickel; a coulometric titration method for sulfur; and a combined fire assay-DC plasma technique for gold, palladium, and platinum. Figure 2 shows the significant enrichment of platinum and palladium at 734 to 737 m (2408 to 2418 ft) depth; it corresponds to a stratigraphic interval rich in oxide minerals. Average values for platinum and palladium contents over the 3-m (10-ft) interval are 2520 and 2580 ppb, respectively, or 5.1 g/t (0.15 oz per st) platinum plus palladium. Platinum plus palladium contents from individual 0.3-m (1-ft) intervals are as high as 9120 ppb or 9.1 g/t (0.27 oz per st) and average 7.4 g/t (0.22 oz per st) in a 1.5-m (5-ft) interval in the richest part of the zone. The Pt-to-Pd ratio is nearly 1:1 over the entire 9.1-m (30-ft) interval and the platinum and palladium correlate closely with chromium, which ranges from 1.9% to 5.3%. Copper, nickel, and sulfur contents for the drill core interval are also given in Fig. 2. In¬terestingly, the Pt-Pd high corresponds to a depletion in copper, nickel, and sulfur. Mineralogy of the Pt-Pd zone The Pt-Pd zone occurs within an oxide-rich zone or layer that consists essentially of oxide minerals, plagioclase, and olivine. The oxides may constitute as

Jan 1, 1987

The Ortiz Mountain property is located approximately 40 km (25 miles) southeast of Santa Fe and 13 km (8 miles) from the village of Cerrillos. The property embodies a Spanish mineral grant from which Gold Fields acquired about 2, 200 hectares (900 acres of surface and mineral rights. Following a period of exploration, metallurgical testing and development of a feasibility study in 1974, the decision was made, because of the higher gold price in the fall of 1978, to place the property into production. the gold is confined in a quartzite breccia; it is an unusual occurrence because the gold in this host rock is leachable at minus 19 mm (3/4 in. ). Total ore reserves are estimated at approximately 3.6 million mt (4 million st) averaging 1.89 g per mt (0.055 oz per st) based on a cut-off of 0.7 g (0.02 oz). The ore will be recovered by open pit mining at a stripping ratio of 1. 2:l. Production is scheduled for 13,600 mt (15,000 st) of ore per week or approximately 680,000 mt (750,000 st) per year. Primary crushing of pit-run material is by a 1,070 by 1,015 mm (42 in. by 40 in.) jaw crusher. Crushed ore is conveyed to stockpile and then to a 2.13 m (7 ft) standard crusher with the screened over- size crushed in a 1.5 m (5 ft) tertiary cone crusher. The screen undersize, the final product for leaching, is minus 19 mm (3/4 in. ). A 1, 360 kg (3,000 lb) sample is cut from the crushed ore stream, further crushed in a small jaw crusher and then sampled in two more stages by Vezin samplers to produce a final 13.6 kg (30 lb) sample for assaying. Pebble lime is then added to the ore on the belt as it is sent to the conveyor gantry spanning the leach pads. A tripper on this moving gantry stacks the leach ore without compaction. The entire crushing plant and conveying system is of open-air construction. Leach pads have a total length of 515 m by 60 m (1,700 ft by 200 ft) wide with usable dimensions of 490 m by 58 m (1,600 ft by 190 ft). The directional length slope is 3. 870 with 370 on the width. The pad is constructed of 175 mm (7 in. ) thick asphalt containing a rubberized asphalt membrane. It is divided by a 300 mm (1 ft) high curb into eight segments. Each segment is loaded with 13,600 mt (15,000 st) of ore, which is one week's production. Optimum working height of a pad is considered to be 3. 0 m (10 ft) with a maximum height of 4. 0 m (14 ft).

Jan 1, 1981

By E. B. Johnson, G. Kjerbol, L. J. Holloway

The Mount Baker Ridge Tunnel is a major improvement of Interstate Route 90 in the State of Washington. The total length of planned underground construction is about 3,400 feet (1,040 m) of which 1,332 feet (406 m) will be constructed by a stacked-drift tunneling method. Sized for vehicular traffic on two levels, the net interior diameter will be 63 feet (19.2 m) requiring an exterior diameter of about 80 feet (24.4 m). The material to be excavated is predominantly stiff clay, and this will be the largest-diameter tunnel ever attempted in soft ground. A construction contract for the stacked-drift tunnel was awarded in late 1982.

Jan 1, 1983

By Sven Petersen

In April 2010 three REMUS6000 series autonomous underwater vehicles (AUVs) have been used to simultaneously map large areas near the Mid-Atlantic Ridge (3°N) axis in water depths ranging from 1,000m to 4,000m. One vehicle was provided by the Leibniz-Institute for Marine Science, IFM-GEOMAR (Kiel, Germany) the other two were provided by the Waitt Institute for Discovery (La Jolla, USA) and managed by the Woods Hole Oceanographic Institute (USA). The REMUS6000 vehicles, manufactured by Hydroid Inc. (USA), are "cigar-shaped" vehicles (4m long and weighing approx. 900kg) and consist of a tapered forward section, a cylindrical midsection and a tapered tail section. An internal titanium strongback, which extends through much of the vehicle length, provides the structural integrity and acts as a mounting platform for syntactic foam, equipment housings, various sensors and release mechanisms. The AUVs can be used in various payload configurations such as multibeam bathymetry, electronic still camera, or subbottom profiler. For these missions the electronic still camera configurations was installed. An EDGETECH 2200M dual frequency (120/410 kHz) side scan sonar was used to map the seafloor, but other frequencies are readily available for more options in resolution and range. The 120 kHz side scan transmitter was used for most missions in order to map large areas (increased range) and to be able to fly at higher altitudes (between 70m and 90m) above the rough seabed encountered in this area. The maximum slant range was between 600m and 700m on both sides resulting in a swath width of up to 1.2 km. Line spacing ranged from nearly 1,000m in flat areas down to 400m in steep terrain. Side-scan mosaics were generated with a 1m resolution. Selected areas were targeted by the higher resolution 410 kHz transmitter as well as by bottom photographs flying as low as a few meters above the seabed. The vehicles navigated autonomously using a combination of inertial navigation (Kearfott T-24 or T-16 INS) and long baseline acoustic navigation by computing their range to two moored acoustic transponders. Working efficiently in this rough topography at speeds above 3 knots was only possible due to the obstacle avoidance capabilities of these AUVs.

Jan 1, 2011