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By M. H. Pelkki, R. J. Sweigard, P. W. Conrad, J. M. Ringe, D. H. Graves

Research has shown that excessive soil compaction is detrimental to the establishment of trees in reclaimed soils. Dry bulk density and penetration resistance are two soil parameters that can be used to indicate the level of compaction in reclaimed soils. It is believed that reclamation practices which minimize soil compaction will produce better results when forestland is the intended post-mining use. The University of Kentucky has developed reforestation test cells on reclaimed soils at a surface mine in eastern Kentucky to test this hypothesis. Each test cell has been planted with various species of trees, and since 1997, data have been collected at these test cells for soil average dry bulk density, average penetration resistance, average depth to refusal, and average survival rate of trees. These data have been correlated to determine the relationship between soil compaction and tree survival rate. It is believed that the two soil compaction parameters may ultimately be used to predict tree survival rates on reclaimed surface mined lands.

Jan 1, 2001

By James A. Madonna

Several universities and colleges in Australia, Canada and the United States, recognized the need to establish or reorganize geoscience programs at the vocational and two-year technical levels. To accomplish this, a cooperative study with industry in one state or province from each country was initiated. These include Tasmania in Australia, British Columbia in Canada and Alaska in the United States. The objective of this work was to develop a diversified geoscience educational model at the two-year technical, vocational and continuing education level that provides the source from which any nation can draw the proper educational tools to create a new geoscience program, modernize an existing program or strengthen a weak program. More than 100 mineral industry organizations provided information regarding their specific training needs for support personnel. The data gathered have been used to identify training areas for semi-skilled, skilled and technical personnel in geochemical mineral exploration. These included geological, geochemical, environmental and computer skills. With the areas of major educational need defined for geochemical exploration, an educational field skills model was developed for the three levels of training. The international need for a study of this type was recognized as early as 1980, following numerous contacts from domestic and foreign geoscience educators regarding the structure and success of training within the University of Alaska Fairbanks Mining Extension program, located in Fairbanks. Contacts, correspondence and visits included educators from Australia, Brazil, Canada, Mozambique and the conterminous United States. The need was magnified between 1985 and 1989 at the University of Alaska Fairbanks, when an attempt failed at initiating and developing a two-year geoscience vocational-technology program leading to an Associate of Science degree.

Jan 1, 1994

By Shemek Oginski

The 5 km (16,194 lf) of 6.1m (20 ft) diameter TBM driven tunnel built by the J.F.Shea Co. will serve as sewage overflow storage. The TBM tunnel and four connecting tunnels excavated by drill and blast methods in hard granite will remain unlined. Based on the Geotechnical Baseline Report (GBR), the Contract documents provided a design with no required pattern of rock bolts, no probing, no pre-excavation or consolidation grouting, and lastly, no final lining of the tunnels was deemed necessary. This paper will provide a general overview of these topics with the emphasis on the actual ground conditions encountered versus the projected conditions provided in the GBR and how the GBR reflected the design of the project. Proposed modifications to the temporary and permanent ground support will be presented and the necessity of probing and pre-excavation grouting will be evaluated.

Jan 1, 2005

By G. Bulut

The recovery of copper and cobalt from ancient copper slags from the Küre region of Turkey was investigated. A fayalitic-type of ancient copper slag containing 1.24% Cu, 0.53% Co and 53.16% Fe were subjected to an experimental study. Two different routes were followed to recover the metal values. The first route was roasting the slag with pyrite followed by leaching. The second route was flotation of the slags, roasting of the flotation tailings with pyrite and leaching. The second route was found to be suitable for the treatment of the copper slag studied. In the flotation step, a copper concentrate containing approximately 11% Cu was produced with a 77% recovery while 93% of the cobalt stayed in the tailings. In the roasting experiments, the effects of roasting time, pyrite: slag ratio and roasting temperature on the dissolution efficiencies of Cu and Co were investigated, and the optimum conditions were determined. In the tests, 87% of the cobalt dissolution was observed after one hour of roasting at 500°C with a pyrite: slag ratio of 3:1. The copper dissolution in the tests was 31%. Based on the results of this study, a process flowsheet is proposed for the evaluation of these slags.

Jan 1, 2007

By A. Adu-Acheampong

Stope instability at Ashanti Goldfields Company Ltd. (AGC) is a major concern. Using information from field mapping, major discontinuity sets were defined, Rock Quality Designation (RQD) was determined. The stability of wedges in the stope walls was assessed using the commercially available computer program UNWEDGE(Rock Science, 1992). Analytical results indicate that the presence of graphite infilling in joints contributes to instability, while steeply dipping joints control wedge slabbing and wedge toppling in the hanging wall. Stable stope widths and dimensions were determined to range from8.1 to 30 meters (m).

Jan 1, 2003

Minerals contribute to the U.S. gross domestic prod-uct (GDP) at several levels?mining, processing and manufacturing finished products. In 2008, the estimated growth rate for the real GDP of the United States was 1.1 percent and the nominal GDP was about $14.3 trillion (Table 1). Trends in various sectors of the domestic economy are often reflected in mineral production and consumption rates. For instance, continued declines in the U.S. housing market during 2008 were reflected in further reductions in the production and consumption of cement, clays, construction sand and gravel and crushed stone (mineral commodities that are used almost exclusively in construction), and those associated with the related manufacture of goods, such as ceramic tile, heating and air conditioning equipment, paint, plumbing and sanitaryware, roofing, wiring and wallboard ? all used by the housing industry. The widespread effects of the slowing domestic economy became more apparent in the last three to six months of 2008 when prices for many mineral commodities declined sharply after reaching record highs earlier in the year, indicating significantly lower demand for those materials or products containing them.

Jan 1, 2009

By Patrick J. Doig

In mid-1991, the design/build contractor set to work on the Pergau HEP in a remote corner of Malaysia, armed with only a slim volume of conceptual drawings. The next five years saw roads, power lines, buildings, dams, and over 30 kilometers of tunnel constructed, followed by two years of E&M installation and commissioning. Along the way, the project overcame geothermal groundwater, faults, land disputes, monsoon rains, marauding tigers, and those most fearsome of foes, politicians and the press. This paper is a testament to the ingenuity, commitment, and persistence of the construction industry professionals who brought this massive project to fruition.

Jan 1, 2001

By R. J. Pruett

The Georgia kaolin district is the world’s largest producer of kaolin. Georgia production tripled from 2.87 Mt (3.17 million st) in 1968 to8.35 Mt (9.2 million st) in 1998. Development of the Georgia kaolin industry can be traced to the identification of the mineral resource, the adoption of innovative process technologies and the creation of markets for kaolin. Georgia kaolin is an abundant and high-quality resource that shows a range of advantageous properties. Technical developments in wet processing, delamination and calcination allowed production of high-value products. The use of kaolin pigments in the paper industry provided a market that absorbed the rapid growth in kaolin production.

Jan 1, 2000

By C. Keefe, O&apos

Measurements of the quantity of entrained air within processes have challenged instrumentation developers for years. This is particularly true in the unique and demanding operating environment and applications seen in minerals processing. In minerals processing applications, entrained air plays a role in errors seen within nuclear density gauge measurements and in pumping efficiency. It is also the major component of flotation processes. The measurement of the quantity of entrained air, known as the gas void fraction, can now be performed from the outside of a process pipe through CiDRA?s patented passive sonar array-based technology. Within column cells, the need for a robust, no maintenance, accurate means of measuring the entrained air, known as gas holdup, has resulted in a submersible version of this technology. CiDRA?s meter technology calculates the gas void fraction by first measuring the speed at which naturally occurring sound in the process slurry propagates past an axial array of sensors. These sensors are incorporated in a band that is wrapped around the outside of the pipe, resulting in no process downtimes for installation and unprecedented reliability. A known relationship between this speed of sound and the entrained air captured in the form of bubbles within the slurry is then applied in CiDRA?s meter to calculate the gas void fraction. This paper will present the measurement technology, its operation, its calibration and its application in a variety of gas void fraction and gas holdup measurement situations such as pumping efficiencies, hydrocyclone operation, external sparging, and column cells.

Jan 1, 2008

By P. C. Holloway

La Oroya zinc ferrite (19.5% Zn, 26.6% Fe, 0.052% Ga, 0.075% In) was roasted with coal under conditions more oxidizing than normally used in a Waelz kiln, and the roasted residues were leached with 200 g/L H2SO4. Overall zinc recoveries of up to 97% were achieved, with 35% of the zinc reporting to the fume after roasting at 1,100°C with a bituminous coal addition of 20%. Under these conditions, 93.2% of the Cd, 13.4% of the Cu, 33% of the Ga, 61% of the In and 18.2% of the Pb were recovered either in the fume or in the leach solution. At 2006 metal prices, the value of the recovered metals is comparable to that obtained from commercial Waelz kiln processing of this residue. Zinc in the franklinite (ZnFe2O4) mineral is reduced to zinc vapor, and, after oxidation, some of the zinc is found as ZnO or as Zn2SiO4 in the roasted residue. Iron reacts to form either hematite (Fe2O3) or magnetite (Fe3O4), with no metallic iron detected in the roasted residue, and less than 2% of the iron is solubilized during leaching. The residue produced after roasting and leaching is, therefore, much more stable than the residue from Waelz kiln processing and should be suitable for disposal in a landfill.

Jan 1, 2008

By Maurice Grieves

GC. Waterman 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. W.E. Hawes The author, in mentioning the South African developments of the cactus grab, ignores the parallel development of the Cryderman mucker in this hemisphere. Perhaps a slightly longer history of blind hold drilling would have been in order. Blind shaft boring got a major boost at the Nevada Test Site, as part of the nuclear weapons development, when it was essential to be able to quickly assess depth for weapons testing. One of the early civilian attempts to utilize this technology occurred when Kerr McGee drilled the shaft at the Section 19 Mine at Ambrosia Lake, NM. This twin masted drill rig belonged to a subsidiary of Kerr McGee, not Shaft Drillers. This was a difficult hole, due to bentonitic shales that decomposed. Later advances in drilling fluids eliminated this problem. The Conoco project referred to is not entirely correct, and needs minor amplification: The depth of the development (not pilot) shaft drilled was 684 m, at a diameter of 3.05 m, with a 2.16 m ID hydratatic casing being installed to a depth of 669 m. Conoco Inc., not Challenger, executed the drilling, however Conoco did use Challenger Drilling Company's Rig 14-S. Rig 14-S is unique in that it is specifically designed for shaft drilling, rather than being an oil field rig used for this service. All work was planned and directed by Conoco personnel. Deviation of the aforementioned shaft from vertical was 31 cm. Two smaller shafts (pilot shafts) were drilled with the same rig. These smaller ones were drilled 1.8 m in diameter, to depth of 665 m and cased with .91-m-diam casing, to depth of 650 m to be used as pilot holes for 5.5 m diameter shafts. The total elapsed time for all drilling, casing, cementing, and rig moves was 50 weeks. The constraints mentioned of vertical accuracy, torque, wall collapse, rig set up time, lining time etc., pumping out, etc., may not be as serious as the author implies. First, vertical accuracy: The record achieved by Conoco indicates that with reasonable care, good accuracy can be obtained, primarily by having a fair amount of weight in tension, rather than on the drill bit. Second, the amount of torque that can be applied to the drill stem has been greatly increased by the introduction of larger diameter drill stems, such as the new Hughes rig uses. Third, risk of collapse of shaft walls is minimal with the proper mud program. Basically, the walls are supported by the weight of the mud and it is not removed until some type of casing is installed and grouted in. Here in the real challenge of blind drilling shafts to devise a more economical method of lining bored shafts than using steel casing. The remaining issues were minor at Conoco, provided adequate planning and scheduling takes place. An excellent paper, "Shaft Drilling--Crownpoint Project" by Hassell H. Hunter, presented at the Fifth Uranium Symposium in Albuquerque in Sept. 1981 contains details of this project. A combination of blind boring, with enlargement by either mechanical means (boring machine, or tunneling type shield) or slabbing down, with muck removal through a larger bored shaft would seem to be the future trend in shaft development. The economics, especially for wet conditions, favor this concept over conventional sinking, as has been demonstrated at Kerr McGee --Churchrock, UNC--Churchrock and Conoco at Crownpoint.

Jan 3, 1983

By Brian Heer, David Young, Randall Essex

The Manapouri Power Station is New Zealand’s largest hydroelectric facility located on the South Island of New Zealand. Constructed in the 1960’s, it includes a 700 MW under-ground power station and a 9.8 km (6.1 mile), 9.2 m (30 ft) diameter tailrace tunnel. The output of the power station is currently limited due to excessive head losses in the tailrace tunnel. To remedy this situation, the Electricity Corporation of New Zealand (ECNZ) retained Woodward-Clyde to design a second tailrace tunnel and appurtenant facilities and prepare Contract Documents for construction. The Project will involve the construction of a 9.8 km(6.1 miles), 10 m (33 ft) diameter, TBM-mined tunnel, hydraulic control structures at the upstream and downstream ends of the tunnel, an underground dewatering shaft and station, and tailrace channel improvements. This paper summarizes the key design and construction issues addressed during the planning and design phases, outlines the risk management provisions incorporated in the Contract, and overviews the Interactive Tender Process (ITP)that was implemented during the pre-tendering and tendering periods.

Jan 1, 1997

By K. A. Lewandowski

The efficiency of heap leaching is reduced by poor solution flow characteristics (i.e., channeling and ponding) that result from fines migration. Agglomeration of the ore can immobilize fines, but for agglomeration to be successful in copper leaching, an acid-resistant binder is needed to stabilize the agglomerates. Until now, there has not been a satisfactory method to evaluate whether a binder was effective, other than squeezing a handful to see if it ?feels right.? This paper reports on several procedures that have been developed to quantitatively determine whether a binder gives good performance in an acidic solution. These procedures have been used to identify binders that would be suitable for copper heap leaching.

Jan 1, 2008

By Richard D. Lang

Introduction Argyle is located in the Kimberley region in northwest Australia (Fig. 1). It is a significant diamond discovery by world standards and is the first major diamond find in Australia. The deposit consists of a pipe and associated alluvial deposits. Before the 1970s, diamonds were known to exist only sporadically in Australia. They have been found occasionally in association with gold deposits in Western Australia. There is a small deposit near Copeton in New South Wales. Some alluvial occurrences have also been found in Victoria, South Australia, and New South Wales. The first indicator minerals in the Kimberley region of Western Australia were discovered in January 1973. Production from the alluvial deposits at Argyle began in 1983. A $430-million construction program to develop a mine and treatment plant capable of producing 25 million carats per year was completed in December (1 carat = 0.2 grams). Ownership and management Participants in the project are known as the Argyle Diamond Mines Joint Venture. The group consists of CRA Ltd., 56.8%; Ashton Mining Ltd., 38.2%; and the Western Australian Diamond Trust (formerly Northern Mining Corp.), 5%. The project is managed by Argyle Diamond Mines Pty. Ltd., a wholly-owned subsidiary of CRA. Marketing is undertaken separately by the Western Australian Diamond Trust and by Argyle Diamond Sales Ltd. (for CRA and Ashton Mining groups). History Ownership of the Exploration Joint Venture has changed over the last 12 years as various parties changed their shareholdings. The original Kalumburu Joint Venture was formed in 1972 with five participants, each holding a 20% interest. By 1983, a number of ownership changes had occurred. The end result is that two original participants have sold out. CRA now holds a 56.8% share and the Ashton Mining Group holds 38.2%. CRA also assumed project management in January 1977. Table 1 demonstrates the long lead times involved in developing a project.

Jan 1, 1986

By D. Pappas

Over the past decade the number of contractors working in underground coal mines has nearly doubled; however, the effect this has had on the industry?s injury rates has been the subject of much speculation. The confusion exists because the U.S. Mine Safety and Health Administration (MSHA) database does not assign contractor hours to the individual mining operations where they worked. To better understand the safety record of the contractor segment of the industry, the National Institute for Occupational Safety and Health (NIOSH) undertook a comprehensive analysis of the data contained in the MSHA database. The first part of the study compared overall contractor and operator trends from 1983 to 2009 related to employment, hours worked, injury rate, and days lost. One unexpected finding was that the larger contractors tended to have higher injury rates than the smaller ones. The second part of the study directly addressed the role of contractors on the safety record of 10 large underground coal mines. A detailed analysis provided both injuries and estimated contractor hours worked at these mines for the period 1992?2007. Comparison between the estimated contractor injury rates and the reported operator injury rates at these mines indicated that the contractor injury rates were significantly higher at most of the operations studied. However, the industry-wide data indicates that contractor and operator injury rates have converged since 2005 and are now equivalent.

Jan 1, 2011

By R. J. Boyd

SUMMARY The productivity of drill and blast methods of mining in hard rock is limited by the cyclic nature of the operation, and a plateau of performance has been reached, even with the most advanced equipment developments. Higher mine operating costs and lower metal prices have more than ever generated the need for breaking through these productivity barriers. After initial field testing The Robbins Company has designed and manufactured, on a joint venture basis with Mount Isa Mines, Ltd., Australia, a prototype "Mobile Miner" for continuous cutting of hard rock. This paper describes the cutting concept, the initial application of a completed 1,156 m (3792 ft.) long 14% decline, actual performance and delay analysis, developmental modifications, anticipated performances of a fully developed machine, and possible future applications. Sincere appreciation is expressed to the management of Mount Isa Mines, Ltd., for allowing access to the data used in this paper. The views expressed here are those of the author and not necessarily those of Mount Isa Mines, Ltd. INTRODUCTION A paper titled, "Design and Development of a Mobile Mining Machine", was presented by J.P. Kindsvater at the RETC, New York, in 1985. This paper described the evaluation of a test fixture in an andesite quarry near Seattle, followed by design, construction and

Jan 1, 1987

By C. Wen

In Huoqiu region of Anhui Province, Fe2O3 occurs mainly in the form of specularite. Besides quartz, the gangue minerals include iron-bearing silicates such as chlorite and sericite which account for about 10% of the total gangue. The separation effect of conventional collectors is poor in the flotation of the minerals, that is, the average grade of iron concentrates is only 60-62%. Using the developed amphoteric collector can successfully realize not only the separation of specularite from various gangue minerals with results of iron concentrate grade above 65% and iron recovery of above 80% but also low temperature flotation at 15°C.

Jan 1, 2012

By Antonio De Biase

In October 2006, a 7m Double Shield TBM, boring through very poor volcanic formations, was pushed back by fluid mud, which presence had not been detected during previous field investigations, due to the high covers characterizing the majority of the tunnel axis. The present work describes the investigations made to gain better knowledge of the conditions in front and all around the TBM, the special measures implemented, the exploratory and bypass tunnels excavated, the extraordinary occurrences came true during the execution of these tunnels, the drainage campaigns, the design of the chamber excavated to free the TBM, the dismounting and restarting of the TBM.

By Harry R. Price, Bharat Doshi, Awni Qaqish, Abdul-Ghani Mekkaoui, Thaddeus Znoy, Ramesh Shukla

In 1992, a detailed inspection of the Southfield Sewer revealed a reach of approximately 0.8 kilometers in length, located directly below the freeway lanes, which was severely distressed. Failure of this section of the sewer would result in the closing of the busiest roadway in the state as well as the potential backup of sewage into homes and spilling into nearby surface waters. As such, repair of the distressed section required development of a procedure that would not jeopardize the conveyance of flow through the existing sewer nor impede travel on the overlying freeway. Based on these constraints, the DWSD and NTH team developed a repair concept that entailed the mining of a new, 3.05-meter (10-foot) diameter bypass away from the traffic lanes. The new bypass section was connected to the existing sewer, up stream and down stream of the distressed area, using a “top down” construction approach, that required only minimal freeway closures during weekend periods. This paper, prepared jointly by the Owner, Consultant and Contractor will describe existing conditions, design considerations and construction methods.

Jan 1, 2001

By J. Evans

Including: Tailings thickeners, CCD?s, paste thickeners, HI-rate thickeners, lamella thickeners, high compression thickeners, clarifiers etc. Flocculant dosing control for mining thickeners and clarifiers is influenced by a number of process factors. 1. Flow rate into the thickener or clarifier a) As the flow rate increases into the thickener or clarifier ?retention time? decreases which causes a greater affect hydraulically on the settling solids. So as the flow rate increases it will generally mean an increase in flocculant dosing rate. b) As the flow rate decreases into the thickener or clarifier the retention time increases which reduces the hydraulic affect on settling, improving settling rates. This will generally mean that as the flow rate decreases, we will decrease the flocculant dose rate.

Jan 1, 2011