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By Bolger Witthaus

"High performance production under challenging geological conditions is a product of several parameters. Main factors are the technical equipment and the knowledge in processing. The design and the support system of maingates must also be taken into consideration. The extraction of seams in German underground hard coal mines of RAG in a range of heights between approximately 1.1 and 4.0 m ( 4 and 13 ft) is done by plow and shearer technology. A complete system for quality management of production is required. The process examination of processes begins with procurement of the equipment. Education and training of the engineers and specialists, as well as for the miners are requirements for success as well.This report gives an overview of RAG measures for achieving high performance. Starting with examination of the processes accompanying procurement of equipment, the following important elements are described. Standards in education and training, technical design and elements of automated processing using stateof- the-art in control systems are critical to this system.The application of automation plays a decisive role in modem coal mines. The latest development in German coal mines is the fully-automated shearer system for longwalls. This helps to enhance safety, particularly in difficult ground conditions, by removing the shearer and shield operators from the immediate mining activity.INTRODUCTIONLongwall mining in carboniferous strata in Germany faces various challenges. Different longwall systems are used to mine up to 24 seams, each with a thickness between 1.1 m (3.6 ft) and 4.0 m (13 ft). High stresses due to a depth of up to 1,400 m (4,593 ft) requiring pillars between the panels, require special systems for roadway development, roadway support and reinforcemen.t as well as for longwall advance operations. Two main types of equipment for extraction are shearer and plow systems. While shearers are utilized for seam thicknesses above approximately 1.8 m (6 ft), plow systems can be used for smaller seam thicknesses between 1.8 m (6 ft) and approximately 1.0 m (3 ft)."

Jan 1, 2010

By Jeremy Theys

The Sacramento Regional County Sanitation District (SRCSD) is improving the sewer system in the Natomas area of Sacramento, CA. The Upper Northwest Interceptor (UNWI) system is an ongoing series of tunnel pipeline projects that will provide sewer conveyance for all residents of northeast Sacramento County to the New Natomas Pumping Station, located North of downtown Sacramento, also in the city of Sacramento limits. The New Natomas Pumping Station then carries wastewater flowsto the Sacramento Regional Wastewater Treatment Plant in Elk Grove via the existing Lower Northwest Interceptor. The Joint Venture of Traylor Brothers, inc. and J.F. Shea (TSJV) were awarded the UNWI 1&2 project in late September, 2007. The planned 5.8 km (3.65 mile) construction route for UNWI 1&2 begins at the New Natomas Pump Station and runs north along the East Drainage Canal to the begin-ning of UNWI 3&4 at Bridgecross Drive. UNWI 1&2 includes approximately 5.8 km (19,240 feet) of 3.65 m (144 inch) inside diameter pipe using Earth Pressure Balance Tunneling Methods. The project also includes 16.76 m (55 feet) of 3.04 m (120-inch) diameter interceptor pipeline, twenty three access manholes spaced approximately 305 m (1,000 feet) apart along the alignment, a 2.13 m (84-inch) by 3.65 m (144-inch) transition structure, a 3.65 m (144-inch) by 3.04 m (120-inch) transition structure, and the connection of three existing sanitary sewer lines to the interceptor.

By George H. Holmes

This paper is one of a series on mining methods and practices published by the Bureau of Mines. It describes quarrying methods and crushing practices of the U. S. Lime. Products Corp. at its Sloan and Apex quarries. Calcining and milling processes. at the Sloan and Henderson plants also are outlined. The Sloan dolomite and limestone quarries and plant are in secs. 12 and 13, T. 23 5., R. 60 E., and sec. 18, T. 23 S., R. 61 E., M.D.B. &M., Clark County, Nev., about. 17 miles by paved highway southwest of Las Vegas (fig. 1). The Apox limestone quarry, and crushing plant are in sec. 26, T. 18 S., R. 63 E., M.D.B.&M., Clark County, Nov., approximately 18 milos northeast of Las Vegas, adjacent to U. S. Highway 91 (fig. 1). The Henderson calcining plant is Inc., in sec. 13, T. 22 S., R. 62 E., 14 miles southeast of Las Vegas on U. in a reduction unit of Basic Magnesium, M.D.B. &M., Clark County, Nev., about S. Highway 93 (fig. 1).

Sep 1, 1949

By Cotton R. E, Cox R

Digger Rocks is one of several nickel-bearing sulphide deposits localised within komaLiitic volcanics of the Forrestania Greenstone Bell within the Archaean YiIgam Block of Western Australia. Exploration and evaluation to April 1990 have delineated a resource, at a nominal cut-off grade of 0.5 per cent Ni, of 1.9 Mt at 1.1 per cent Ni within a lensoid sulphide body which strikes NW, dips 50°SW, and occurs at a depth of 30 - 110 m below surface. There has been no production from any of the Forrestania nickel deposits to-date. The deposit, comprising both disseminated and massive sulphides, has maximum dimensions of 400 m strike, 115 m vertically and 90 m true width. Generally the massive sulphides occur at the base of the disseminated sulphides, on or close to the base of the Digger Rocks dunite. The two main massive sulphide lenses have maximum dimensions of 120 m strike, 75 m vertically and 16 m true width (northern lens) and 60 m strike, 25 m vertically and 4 m true width (southern lens). The average grades of the massive and disseminated sulphides are 5.9 per cent i and 1.2 per cent i respectively. The massive sulphides generally have sharp hangingwall and fOOlwall contacts and display strong internal grade anisolropy.The massive sulphides are mainly within the supergene zone and consist of violarite, pyrite and marcasite. The disseminated sulphides are mainly within the primary zone and consist of pentlandite, pyrrholite, pyrite and minor chalcopyrite. Alteration zones, up to six in number, are complex in detail. The sulphides are capped by an average of 30 m of sub-economic, leached, siliceous, oxidised material.A genetic model for the deposit is described which places Digger Rocks in Category 2 (sulphides in large layered dunite bodies) of the latest classification of komatiitic nickel sulphide deposits in the Yilgam Block of Western Australia by Hill and Gole (1990).

Jan 1, 1992

By John Chadwick

In the drive for faster development advance rates, railbound systems can be advantageous, in the right application. Skanska BS, the contractor to mine-owner VSK Mining, is driving an adit access deep into the Volouské Urchy Mountain in Slovakia to access the Gemerská Poloma talc deposit, some 70 km (43 miles) west of Ko?ice, the economic and cultural center of eastern Slovakia. Skanska began driving the Stolna Elisabeth drift in February 2007, extending 25 kg/m (50 lb/yd) rails (with 750 mm or 30 in. between sleepers) into the mountain to provide the main haulage to bring talc ore and waste out of the new mine once it gets into production. It will also be used for personnel transport, and is now almost complete. As a second egress and for ventilation, a 2.5-m- (8.2-ft-) diameter shaft of about 300-m (984-ft) depth will be sunk from surface down into the orebody. Stolna Elisabeth is a 4.2-km- (2.6-mile-) long drift mined at a cross section of some 11 m2 (118 sq ft). The finished profile is 3.2-m- (10.5-ft-) wide by 3.6-m- (11.8-ft-) high. It has been developed on a very slight incline so that water runs down it to the portal. The portal is 450 m (1,476 ft) above sea level and the end of the tunnel is 475 m (1,558 ft) above sea level. Much of the rock through which it has been developed is magnesite with a compressive strength of 150 MPA. The drilling jumbo is an Atlas Copco twin-boom Boomer 282 Raildrill drilling 42- to 45-mm- (1.6- to 1.7-in.-) diameter holes to a maximum length of 2.5 m (8.2 ft), depending on ground conditions. This is a drill jumbo that Skanska has used successfully on a number of projects during its life. The condition of the rock through which this drift has been driven is extremely variable. Five classifications of A1 and A2 and B1, B2 and [ ] B3, where A1 is the most competent rock and B3 is the worst, have been established. The amount of support used, of which more later is determined by the classification of ground conditions. At the time of the author?s visit, the project it was about 40 days ahead of schedule, thanks to high advance rates. For instance 6 m/day (20 ft/d) was the plan for the advance to be achieved in B1 rock, but sometimes the crews have achieved 8 m/day (26 ft/d) ? which is the average for the whole project ? in those conditions. In A1 the best advance rate achieved is 12 m/day (39 ft/d). The loading unit is a Häggloader 9HR-B on crawler tracks with a pony track to run on the rails. This allows the unit to be towed up and down the drift along the rails. To load from the face, the Häggloader lowers the foot of its onboard conveyor system to the floor and the backhoe pushes material on to it. This digging arm-conveyor combination produces uninterrupted, optimal volume loading of the haulage unit, without spill ? a technique offering load capacities as high as 180 m3/h (6,356 cu ft/hour). Further-more, as an electrically powered machine, it offers great

Jan 1, 2009

By J. Noll, D. Yantek, B. Kendall, J. A. Organiscak

"A self-tramming remotely controlled mobile Dry Scrubber (DS) was developed by J. H. Fletcher and Co. under a contract with the National Institute for Occupational Safety and Health’s Office of Mine Safety and Health Research (NIOSH OMSHR) to reduce exposure of miners to airborne dust. NIOSH OMSHR laboratory testing showed that the DS averaged greater than 95% dust removal efficiency with the disposable filters and averaged 88% and 90%, respectively, for the optional washable filters in their pre-washed and post-washed test conditions. Although the washable filters could be reused, washing them generated personal and downstream respirable dust concentrations of 1.2 and 8.3 mg/m3, respectively, for a 10-minute washing period. The DS’s velocity pressure regulated variable frequency drive (VFD) fan maintained a relatively consistent airflow near the targeted 1.43 and 4.25 m3/s (3000 and 9000 ft3/min) airflow rates during most of the laboratory dust testing until reaching its maximum 60 Hz fan motor frequency or horsepower rating at 2610 Pa (10.5 in. w.g.) of filter differential pressure and 3.97 m3/s (8420 ft3/min) of scrubber airflow quantity. Laboratory sound level measurements of the DS also showed that the outlet side of the scrubber was noisier and the loaded filters increased sound levels as compared to clean filters at the same airflow quantities. With loaded filters the scrubber reached a 90 dB(A) sound level at 2.83 m3/s (6000 ft3/min) of scrubber airflow, indicating that miners should not be overexposed in relation to MSHA’s permissible exposure level (PEL) of 90 dB(A) at or below this airflow quantity (1). The DS’s washable filters were not used during field testing because of their lower respirable dust removal efficiency and the airborne dust generated from filter washing. Field testing the DS with the disposable filters at two underground coal mine sections showed that it can clean a portion of the section return air and provide about a 50% dust reduction at the face area downstream of the continuous miner (CM) operation. INTRODUCTION Coal mine worker overexposure to coal and quartz dust continues to be a problem at underground coal mining operations in the U.S. The U.S. Mine Safety and Health Administration (MSHA) recently changed the standard for coal mine worker respirable dust exposure from 2.0 mg/m3 average during an eight-hour shift to 1.5 mg/m3 average over the actual working shift, effective as of August 1, 2016 (1). If more than 0.1 mg/m3 quartz mass is found in the coal mine worker dust sample, the applicable respirable dust standard is further reduced to the quotient of 10 divided by the percentage of the quartz in the sample (1). Over 90% of the mechanized mining units operating in U.S. underground coal mines are continuous mining machines (2). MSHA inspector dust samples from 2009 to 2012 showed that 3.7% of continuous miner operators exceeded the 2.0 mg/m3 dust standard, while 8.8% of these dust samples exceeded the new 1.5 mg/m3 dust standard and 9.7% of these samples exceeded the reduced quartz standard (3). Additionally, roof bolter dust samples at these mechanized mining units during this same period have exceeded the 2.0 mg/m3, 1.5 mg/m3, and reduced quartz levels at 1.1%, 3.7%, and 10.6%, respectively (3)."

Jan 1, 2016

By Rodney Pinksen, Remi Proulx

The IOC Mining and Operations Departments expressed an interest in reducing the amount of oversize material (>1.0 m x 1.0 m x 1.0 m; 39 inches x 39 inches x 39 inches) being produced and its subsequent downstream effects including in-pit sorting and reduction, damage to mobile equipment, and significant delays in the loading pockets and the primary crushers. To achieve this goal an oversize reduction project was undertaken by the mine and to date has significantly improved upon baseline measures.

Jan 1, 2011

By T. Iriyama, Y. Fujita, R. Omatsuzawa, T. Nishio, N. Okochi

Sm-Fe-N hard magnetic materials are good candidates for high performance bonded magnets. Rapid-quenching method should be highly favorable for the fabrication of isotropic magnet powder with high coercivity. We report here the magnetic properties of Sm-Fe-N powders produced by a few kinds of rapid-quenching processes. Sm2Fe17 powders were prepared by gas-atomization (GA), rotating-disk atomization (RDA), or melt-spinning (MS) method. The Sm2Fe17 GA powder had dendritic solidification structure, while the RDA powder had micro crystalline structure with size of 3-5 µm. Coercivities (HcI) of their nitrides were 56 (for GA) and 217 ka/m (for RDA). The H cI of nitrided Sm2Fe17 MS powder with wheel velocity (V5) of 5-20 m/s, which had also microcrystalline structure, was 53 to 509 Wm. These coercivity values are not sufficient for bonded magnet applications. From the study of additive effects, it was found that C, C-Si, C-Ga, or C-Al addition decreased the crystal size of quenched Sm2Fe17 powder. Consequently, high coercivity was obtained; the HcI of nitrided Sm2Fe1C1Si1 MS powder with V5 of 10 m/s was 898 kA/m, and 708 kA/m for the nitrided Sm2Fe15C1Si1 RDA powder. The compression-molded bonded magnet made from Sm2Fe18Si0.1Nx MS powder showed a high (BH)max of 82.7 kJ/m3.

Jan 1, 2000

By Warder A. Shaver

INTRODUCTION Mine No. 95 of the Consolidation Coal Co., a divi¬sion of Continental Oil Co., is located in Shinnston, WV. Mine No. 95 was opened in January 1967 and has been operating continuously since that time. The mine has 544 hourly and salaried employees. In 1974 this mine was the largest coal producer in West Virginia. Opera¬tions include two longwall sections and eight Marietta Borer continuous miners. Originally there were also two Jeffrey Heliminer continuous miners, but use of these was discontinued in June 1978. Production averages 12 727 t/d (14,000 stpd). Two portals now service the mine with a third portal proposed for 1978. GEOLOGY Mining takes place in the Pittsburgh coal seam which dips 0.04 rad (21h°) northwest. The working seam is steam coal with an average height of 2.2 m (86 in.). Immediately above the seam, the overburden is com¬posed of 6.1 m (20 ft) of shales followed by 6.1 m (20 ft) of weak sandstone and 51.8 (170 ft) of shaley limestone. The shale weathers when exposed to mois¬ture. Surface water has been a problem due to a fire¬clay floor which turns to paste when mixed with water. GENERAL DESCRIPTION The present mining plan for Consol's No. 95 mine has a proposed life of 35 to 40 years. One set of main headings has been driven in the north direction and one set in a western direction. These are the areas in which the majority of the virgin coal lies. The panels are driven off of the mains at 2134 to 3048-m (7,000 to 10,000-ft) intervals. The continuous mining sections are set up 133 x 610 m (600 x 2000 ft). Longwall sections are 259 x 1220 m (850 x 7000 ft). All of these mining sections are retreating uphill. Development is with continuous miners. A Marietta Borer miner is used for the main development work. These miners cut 4.1 m (13 ft 6 in.) wide and 2.1 m (7 ft) high, leaving an arch cut. Along with the arch, a 152.4-mm (6-in.) thick layer of roof coal is left in place to protect the roof from weathering. The combi¬nation of the self-supporting arch and the roof coal al¬lows the entries to be driven 30.5 m (100 ft) without bolting. Full bolting, however, is carried outby the last open crosscut. This bolting is done on 1.5 x 1.5-m (5 x 5-ft) centers using 1.5-m (5-ft) long roof bolts with tandem anchors and 152.4-mm (6-in.) steel washer plates. When needed, crib blocks with a 50.4-mm (2-in.) drill hole are used as supplemental supports. Three-hole steel channel with the ends curved down¬ward to conform to the borer arch may also be used. In the mining sections, a 14 BU loader is used be¬hind the Marietta miner to load two Torkar shuttle cars. These haul to Long-Airdox feeder breakers which load onto a 1.07-m (42-in.) wide conveyor belt. With the exception of the shuttle cars, which use 275-V d-c power, the face equipment operates on 550-V a-c power. The main power line is 7200 V ac. The same loading and hauling equipment is used with the Marietta Borer miner for driving both panel and development entries. In the retreat sections, the same equipment with the exception of the miner is utilized. Because of the advantage of a variable cutting hight, Jeffrey Heliminer ripper-type continuous miners were employed in these retreat sections. For longwall mining the first unit uses an Eickoff EDW 300-L double-drum ranging shearer. Drum di¬ameter is 1.69 m (66 in.) with a drum rotation speed of 45 rpm. The drum is equipped with 48 Carmet 34-L-1 bits. The 300-kW (400-hp) motor operates on 1000 V. Tramming of the shearer head is done through a hydraulically driven 26-mm wide chain. Maximum cutting speed is 7.93 m/min (26 fpm), and maximum tram speed is 12.8 m/min (42 fpm) with an average cutting speed of 3.66 m/min (12 fpm). Web cutting depth is 762 mm (30 in.) and the cutting height is 2.13 m (7 ft). The conveyor width is 686 mm (27 in.). It uses a single chain, 26 mm thick, with flights on 0.91-m (3-ft) spacing. The face conveyor speed is 64.3 m/min (211 fpm) when driven by two 112-kW (150-hp) 1000-V Louis Allis motors through torque convertors. The roof support used is Gullick Dobson chocks [6-legged 77.3 t (85 st) per leg, 463.6 t (510 st) total capacity] with double-acting leg jacks. The hydraulic pressure for the legs is 13.8 MPa (2000 psi), and the hydraulic pressure for rams is 6.8 MPa (1000 psi). The yield valves are set at 57.8 MPa (8500 psi). Each chock weighs 4.09 t (4.5 st). The stage loader is an Eickoff drive, with a 68-kW (90-hp), (00-V Reliance motor. A 26-mm thick single strand chain with flights on 0.6-m (2-ft) centers runs at 75.89 m/min (249 fpm). The stage loader is self-advancing and is connected to the belt tail piece with a special fifth wheel attachment so that the stage loader can advance and guide the tail piece. The hydraulic pressure for the roof support is supplied by two 45-kW (60-hp) Reliance motors op¬erating on 699 V, with 1.26 L/s (20 gpm) constant volume Gullick hydraulic pumps. Each pump services alternate chocks, with the two pressure lines connected at the tailgate by 25.4-mm (1-in.) hydraulic hoses, rated at 27.21 MPa (4000 psi). The 6.8 MPa (1000 psi) pressure rams advance is supplied by a 45-kW (60-hp), 600-V Reliance motor and a 1.26-L/s (20-gpm) con¬stant volume Gullick pump. Return lines are 35.5-mm (1.4-in.) diam hydraulic hoses with rated capacity of 27.21 MPa (4000 psi). The fluid used in the hydraulic system consists of 95% water and 5% Hy-chock fire resistant water soluble oil. Water is stored and mixed in a 6096-L (2000-gal) track-mounted tank and goes to a 655-L (180-gal) reservoir at the pump. The belt conveyor structure is 1.07 m (42 in.) wide with 0.47 rad (27°) idlers, bottom stands located on 3.05-m (10-ft) centers, top stands on 1.5-m (5-ft) centers, and is supported on a 19-mm (3/4-in.) diam

Jan 1, 1982

The oxidation of ferrous sulphate to ferric sulphate with air and oxygen in a gas lift percolator was investigated with initial Fe2+ concentrations in the range 0·01 M to 1·0 M,temperatures in the range 20-80°C, and pH in the range oto 2.The rate of formation of Fe3+ was found to be well described by the rate expressionEquation 1and a consistent mechanism is proposed.The presence of Cu2+ was found to cause an increased rate of oxidation, the effect being greater with Cu2+ concentrations increasing in the range 0·001 M - 0·01 M in 0·2 M Fe2+. The effect of other solutes containing Ni, Zn, Mn, Co, Hg, Cl, Mo, NH4, As, Cr, or Na at a concentration of 0·001 M in 0·2 M Fe2+ was negligible.

Jan 1, 1972

By B. Whistler

The Alpala porphyry copper-gold deposit is a recent discovery located in the Imbabura Province of northwestern Ecuador. The drilling of seventeen diamond core holes over a 900 m by 400 m area has defined a northwesterly-trending, steeply northeast-dipping dike-stock complex of diorite to quartz diorite that exceeds 1300 m in height. This intrusive complex is hosted by a sequence of andesitic volcaniclastic rocks and lavas. The host-rocks are interpreted to be part of the Paleocene to Late Eocene Macuchi Formation (BGS-CODIGEM, 1997; Cruz, 2007). The best drill intercept to date is 1312 m at 0.67 % Cu and 0.63 g/t Au from 128 m depth in CSD-15-012, which includes 576 m at 1.03 % Cu and 1.19 g/t Au.

May 1, 2008

By G. B. Misra

Attempts have been made during the last three decades to establish a correlation between respirable dust dosage and the degree of progression of pneumoconiosis or silicosis amongst mine workers. Fair correlation has been possible only after the Johannesburg Pneumoconiosis Conference established mass concentration of dust in the respirable size range as the true measure of the incidence of pneumoconiosis. The author had established a good correlation between cumulative total respirable dust dosage as well as cumulative silica dosage and the category of pneumoconiosis according to ILO classification at Mufulira Copper Mine. Studies at Hutti also show a linear relationship between the average cumulative dust and silica dosage md the category of pneumoconiosis. However the maximum tolerable cumulative dust dosage for - prolonged exposure is 182 mg month/m3 - at Hutti as compared to 275 mg month/m3 at Mufulira. Similarly the maximum tolerable silica dosage for - prolonged exposure is Only 35 mg month/m3 at Hutti as compared to 55 mg month/m3 at Mufulira. The TLV for dust concentration at Hutti works out at 0.6 mg/m3 compared to 1.1 mg/m3 at Mufulira and the TLV of silica concentration at Hutti is 0.117 mg/m3 compared to 0.25 mg/m3 at Mufulira. The Hutti figure is however, closer to the TLV of 0.1 mg/m 3 recommended by ACGIH.

Jan 1, 1987

By Tore M. Larsen

The Rødsand Gruber mining company is a subsidiary of the Norwegian industrial corporation Elkem AS, and its mine is situated in the northwestern part of southern Norway. The ore from this mine is a relatively low grade, magnetite ore with a small amount of vanadium. The ore bodies are very irregular in shape, resulting in a high amount of waste rock intrusion in the stopes. The annual underground production in the mines is approximately 1 million t. Because of varying dimensions of the ore bodies, both transverse and longitudinal stopes are employed. The inclination of an ore body is usually 1.04 to 1.59 rad (60 to 90°), and the horizontal width varies up to 100 m. Sublevel stoping has been the traditional mining methods in this mine. Until a few years ago, the ore was primarily blasted using horizontal blasthole fans drilled from Alimak raises. Recently, however there has been an increasing demand for a les expensive mining method. Alternate systems were considered, and finally a method based upon large blasthole stoping was found to be the most suitable. GENERAL LAYOUT The general layout of a stope is shown in Fig. 1 As with all the figures in this chapter, only the prin¬ciples of the method are emphasized and the details may not be valid for any particular site in the mine. In transverse stopes, the horizontal width of the ore determines the length of the stopes. The average length is approximately 50 m, the width of the stopes is 20 m, and there is a 10-m wide vertical pillar between. The distance between the main levels is 60 m. Usually a horizontal (sill) pillar 10 m thick is left below the mucking level. Each stope is first undercut along the entire stope length. This undercut is made with 50.8-mm (2-in.) diam drill holes in a fan pattern. At the moment this is being done by a trackless Atlas Copco Promec jumbo equipped with two Cop 130 pneumatic rock drills. The drill drift is 4 X 4 m, and the crater is 20 m high. This leaves a 36-m stope height to mine with large blastholes. To achieve a fair portion of vertical drilling withlarge blastholes, there are two parallel drill drifts at the top of each stope. These drifts are 4 m high and 9 m wide, separated by a 4-m wide pillar in the middle of the stope. To ease the drilling of perimeter holes, the drifts extend 1 m into the vertical pillar on each side of the stope. The pillar will thus be only 8 m wide on the drill drift level, but 10 m wide down the rest of the stope.

Jan 1, 1982

By Robert B. Bittner

The Sutong Bridge across the lower Yangtze River in China will have, when completed, the longest span (1,088 m) and the highest towers (306 m) of any cable stayed span in the world. The foundations for this record setting span will also be record setting. They are located in water depths exceeding 30 m with maximum flows exceeding 3 m/sec. The soil at the site of the main pylons consists of layers of silty sands and silty clays extending to bed rock at 270 m below river elevation. This paper describes the innovative methods used by the team of foundation designers and constructors to support this record setting bridge at this very challenging site. This paper covers three specific topics related to the foundations of the Sutong Bridge: ? Design and construction of the 131 drilled shaft (2.8/2.5 m diameter and 114/117 m long) under each of the two main pylons, ? Design and construction of the scour protection for the two main pylons, ? Construction methods used to construct the waterline pile caps (113.8 m by 48.1 m by 13.3 m deep) under each pylon.

Jan 1, 2005

By Peter Makuvise, Alain M. Bantshi

"The Mazowe Gold Mine is managed by Metallon Gold (Zimbabwe) and is situated in the west-central part of the Harare greenstone belt. Orebodies here generally comprise shear zones which are in-filled with gold-bearing sulphides and quartz. The mineralised zones are up to 1 m in width, have average grades of 4 to 5 g/t, and mostly dip between 10º and 60º to the north.Mazowe Mine is one of the oldest mines in Zimbabwe, and exploration and development in this region dates back to 1890, with over 1.4 million ounces of gold produced to date. Mazowe mine comprises two underground operations, Mazowe and the BSV sections. Ore is processed in a single plant which consists of conventional crushing and milling and carbon-in-leach facility.Baldmin Projects was awarded the contract for the design, construction and commissioning of the 60,000 tonnes per month plant at Mazowe mine in 2014 for the processing of sand and slimes tailings from previous operations at Mazowe mine. The average grade of the tailings is 1.1 g/t. The plant consists of feed preparation, grinding, flotation, thickening of flotation concentrate, and carbon in leach (CIL). A pump station for hydraulic mining of the slimes tailings dam also part of the plant.Key features of the Mazowe sands Plant are described and include:• Hydraulic mining of gold tailings dump.• Feed preparation section that includes a (W x L) 2.4 m wide x 4.8 m trash screen, 250 m3 surge tank, a cluster cyclones with 2 x 500 mm diam. Cyclone.• Comminution circuit designed for a bond work index range of 10-12 kwh/tonne. comminution circuit is constituted with 3 ball mills, 1.83 m diam. X 6.71 m Long, 280 kW.• Flotation circuit comprising rougher bank, scavenger bank and cleaner bank. Rougher Scavenger bank is constituted of 6 x 6m3 flotation cells and 3 x 30 m3 flotation cells and the cleaner is constituted of 3 x 6 m3 flotation cells.• Concentrate Thickening with 6 m diam. Thickener.• Concentrate CIL (carbon in leaching): CIL circuit is comprised of 7 tanks with 2 leach tanks and 5 Carbon in leach."

Jan 1, 2017

By G. C. Waterman, W. E. Hawes

Mr. Grieves' paper on "Shaft Sinking Today --A Boring Business Tomorrow" in the Dec. 1981 issue of MINING ENGINEERING is an excellent description of recent improvements in speed and costs of shaft sinking. However, shaft boring techniques are not a recent development. In 1935-36, the Idaho Maryland Co. (gold), Grass Valley, CA, bored the vertical 1.5-m-diam Idaho No. 2 ventilation and supply shaft to a depth of 346 m. The equipment was designed and perfected by Branner Newsome and was, I believe, described in "Transactions." Pickands Mather later bored a 1.8-m-diam shaft to a depth of about 305 m using Newsome's design. The Idaho No. 2 shaft cut through hard gabbro, diabase dikes, soft serpentine, very incompetent ankeritized serpentine and a strong fault zone. The completed shaft did not require timber support; the fault zone was cemented. Equipment consisted of a rotating 1.5-m or 3-m core barrel with a slotted bottom which cut a 76-mm kerf. The cutting agent was chilled steel shot introduced into the slot as needed. The driving mechanism and core barrel were lowered into the shaft, the former secured to the wall and the inshaft "engineer" operated the motors which controlled core barrel rpm and advance. After an advance of 1.5-3 m the equipment was lifted out of the shaft, a half(?) stick of powder cut off the drilled core, and it was hoisted to the surface with a cable attached to an eye bolt at the tope of the core. A "shift" consisted of three men: hoistman, equipment operator (down the shaft), and a surface laborer. Advance was variable as equipment and techniques were perfected. Near the end of the job advance was, as I remember, 1.5-3 m per shift. Costs were, as I recall, about $115/m. Up to 4 m cores were lifted out in one piece and swung to the dump by a stiff leg derrick. The Newsome equipment was relatively inexpensive to build and operate and his method should be (more than?) competetive with the methods described by Grieves.

Jan 1, 1983

By David Belknap, Clay Haynes, David Egger

Rehabilitation challenges were met nearly two months ahead of contract schedule for the Design-Build, Fast Track Scituate Aqueduct & Tunnel project in Cranston, Rhode Island. The aqueduct and tunnel were built in the 1920s and consisted of three sections—a 2.3 m diameter reinforced concrete section that extends approximately305 m before it transitions to a horseshoe-shaped 2.3 m wide by 2.1 m high cast-in-place concrete tunnel that continues for approximately 5.3 km until it transitions back to a 2.3 m circular section that extends approximately 1.9 km to the end of this project. Major repairs were required to assure sustained operation of this water supply tunnel.

Jan 1, 2005

By Joel K. Waterland

GENERAL DESCRIPTION Homestake Mining Co.'s operations are at Lead, SD. Production from the various methods used are: 1 155 753 t/a (1,274,000 stpy) by open cut-and-fill, 68 039 t/a (75,000 stpy) by square set cut-and-fill, 144242 t/a (159,000 stpy) by shrinkage stoping, 107 048 t/ a (118,000 stpy) by blasthole stoping, and 67 132 t/a (74,000 stpy) in development work for a total of 1 542 214 t/a (1,700,000 stpy). Mining takes place from the 616-m (2000-ft) level to the 2073-m (6800 ft) level (Fig. 1). The average travel time to the stope is 45 min. GENERAL ORE BODY REQUIREMENTS AND LIMITATIONS Size, Shape, and Dip of Ore Body Six major ledges varying from 914 m (3000 ft) long by 228.6 m (750 ft) wide to 305 m (1000 ft) long by 122 m (400 ft) wide are mined. Within these ledges are irregular-shaped areas of varying sizes. These ledges dip at approximately 1.2 rad (70°) to the east and plunge to the south from 0.26 to 1.57 rad (15° to 90°). Ground Conditions The ore is cummingtonite (chlorite schist) which is very competent except for major bedding planes and fracture patterns. The footwall is a slate but because of the dip presents very few problems. The hanging wall is also a slate which because of the dip and bedding is very weak and requires considerable rockbolting. No problem has been encountered with rock bursts to a depth of 2438 m (8000 ft). Ore Value The ore values vary somewhat from ledge to ledge, but the average ore delivered to the mill is approxi¬mately 6.2 g (0.2 oz) of gold per ton. A small amount of silver is the only byproduct of this operation. Annual production of gold is approximately 10 575 183 g (340,000 troy oz). Production Requirements The mine has a mill capacity of approximately 1 723 651 t (1,900,000 st) but because of several fac¬tors (manpower, depth, etc.) this mill capacity has not been attainable for the past several years.

Jan 1, 1982

By James M. Spear

Sample results suggest the only mineral resources in the study areas are placer gold deposits along the South Fork Kern River in the Woodpecker RARE II unit (fig. 1). These deposits, in Rockhouse Basin, total 4.5 Figure 1.--NEAR HERE. million cu yds (3.4 million cu m) of indicated subeconomic gold resources in four deposits of "older" gravel. Gold value in samples averages $0.36 per cu yd ($0.47 par cu m) and is as much as $4.44 per cu yd 1/. Cost of mining the gravel and recovering the placer gold is estimated to be more than $2 per cu yd ($2.62 per cu m) for a deposit of this size and type. Additional inferred subeconomic placer gold resources with similar grade probably occur in four other deposits of "older" gravel on which no mining claims were located. Placer gold occurrences on upper Trout Creek contain gold worth as much as $4.50 per cu yd ($5.88 per cu m) 1/. Average grade is much lower, and only about 1,000 cu yds remain (760 cu m). These occurrences do not constitute a resource, but may attract hobbyists not dependent upon making a profit.

Jan 1, 1982

By J. C. Motz

With ore grades continually decreasing and finer mill feed requirements, it becomes increasingly important to have the capability to handle the larger tonnages economically. With the 10 ft. Standard capable of reducing 1,633 metric tons per hour (1,800 short tons per hour) to 80 percent passing 50 MM (2 inches) and the 10 ft. Short Head following to produce a mill feed of all minus 12 MM (1/2 inch) at a rate of 608 metric tons per hour (670 short tons per hour), we feel significant savings will be realized in installation and operating costs. The largest concentration of cost in the secondary and tertiary crushing stations is the building itself that houses the process equipment. An example of the savings to be attained can be seen from an examination of a 40,820 metric ton (45,000 short ton) per day crushing plant. A total of nine 7 ft. Symons Cone crushers is used to generate 12 MM (1/2 inch) ball mill feed. The average center distance between units is 6.1 M (20 feet) and allowing a 6.1 M (20 feet) nominal distance from the centerline to the wails, the building becomes 61 M (200 feet) long. For this production rate, four 10 ft. Symons Cone crushers would be required. With the 10 ft. crushing plant requiring a center distance spacing of 7.6 M (25 feet), the building length totals approximately 38 M (125 feet). Construction costs vary depending on location, but a reduction in building length from 61 M (200 feet) to 38 M (125 feet) will result in estimated savings of between $2,500,000 and S4,000,000. While the 10 ft. Symons Cone has been involved in a number of feasibility studies for some time, the Process Machinery Division of Rexnord commenced detail design work on the 10 ft. Cone crusher in 1972. For comparison purposes, the 10 ft. and 7 ft. capacities are given in Table 1.

Jan 1, 1978