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By Charles R. Nelson

INTRODUCTION The support of openings in weak sandstones can be achieved by spraying on a liquid grout which will soak in and harden to form a shell. In the St. Peter sandstone of the Minneapolis-St. Paul area of Minnesota compressive strength is increased from about 0.7 MPa (100 psi) to over 7 MPa (1000 psi) to depths of 152 mm (6 in.). A shell of this strength and thickness can pro¬vide support much as shotcrete does but at lower cost. If the spray-grouted shell is inadequate for the support needed, it provides a good base for shotcrete application, which is often difficult in weak or friable sand¬stones because the fresh shotcrete peels off along with a thin layer of sandstone. The spray-grouted technique was initially developed at the Civil and Mineral Engineering Dept. of the University of Minnesota using funds provided by RANN Div. of the National Science Foundation. It was used as temporary support in a 3-m (10-ft) wide, 3700-m (12,000-ft) long storm water drain tunnel by the Minneapolis Sewer Construction Dept. Much of the field development at this construction project was carried out with the cooperation of the personnel of the city of Minneapolis. It also was used for temporary support in a 3.3-m (11-ft) wide, 1800-m (6000-ft) long storm water tunnel built in 1978-1980 for the Minnesota Dept. of Transportation in Minneapolis. In 1975, 2000 m (6500 ft) of 1.5-m (5-ft) wide utility tunnels were spray-grouted for final lining. A 50-m (150-ft) long, 1.2-m (4-ft) wide test tunnel built at the university in 1974 was sprayed for half its length and is being observed for long-term behavior which to date (1981) is good. GROUT MATERIALS The requirements for the liquid grout are more severe than those for injection grouting. Besides being able to penetrate the rock and develop sufficient strength, it must have the following properties: (1) be nontoxic, noncombustible, and have a low odor for spraying under¬ground; (2) have controllable viscosity and setting time for adjusting to the permeability of the sandstone and the required depth of penetration; and (3) must "wet" the sandstone and develop capillary "draw" to penetrate to the required depth. A sodium silicate-based grout with the proper setting agent meets these requirements in the local St. Peter sandstone which has a uniform grain size dis¬tribution curve with D,,, = 0.1 mm, permeability of about 10 to 20 darcys, a porosity of about 25%, and less than 5% of silt size or smaller. It is 97 to 99% pure SiO2. The grout mix consists of Philadelphia Quartz Co. Type N sodium silicate and Celtite 55 Terraset Spray Grade (SG) distributed by Celtite Inc., Cleveland, OH. A volume mix ratio of 100:9:125 for sodium silicate to setting agent to water produces a grout with two to four centipoise viscosity, and initial set at 20°C of 20 min. Setting time is temperature dependent, being longer at lower temperatures, but can be adjusted by varying the setting agent concentration a few percentage points. APPLICATION The application is simple. Spray the grout on the surface until the desired penetration is achieved. Penetration rates of about 5 mm (0.2 in.) per min. are realized in the St. Peter sandstone for up to 152 mm (6 in.) of penetration (30 min. of spraying). The set time must be longer than the spraying time. The spray nozzle should be moved back and forth at a rate that minimizes runoff due to surface buildup. Both a single large hand-propelled nozzle and multiple small machine propelled nozzles have been used successfully. The pumping and mixing equipment used consisted of: (1) a hand-held, hand-pumped, 11-L (3-gal) gar¬den sprayer for test patches (mix by shaking); (2) 76-L (20 gal) batch mixing in barrels with small electric pumps for spraying and mixing; (3) two component pumps consisting of Hypro Model 5300 piston pumps (Hypro Pump Co., St. Paul, MN) on a common shaft driven by an electric motor. The sodium silicate is premixed with a portion of the water and the setting agent is mixed with the rest of the water. The two components are combined and mixed just before the nozzle; and (4) a proportioner metering the setting agent just before the nozzle with the water and sodium silicate premixed and pumped from the surface. The proportioner permits the bulk of the grout to be pumped through a single pipe or hose with only the setting agent stored underground. This is efficient and low in cost. COST The grout mix costs about $0.20/L ($0.75 per gal) in 1978 dollars. Filling all the voids 100 mm (4 in.) deep (25% initial porosity) would have a materials cost of about $5.40/m2 ($0.50 per sq ft). The labor cost of application is usually less than the material cost for normal tunnel jobs. A pump and proportioner cost less than $1000. Sodium silicate storage tanks and pip¬ing costs would depend on the particular site conditions. REFERENCES AND BIBLIOGRAPHY Nelson, C. R., 1977, "Spray Grouting for Tunnel Support and Lining," Underground Space, Vol. 1, No. 3, pp. 241¬ 245. Yardley, D. H., Nelson, C. R., Stocker, T. H., 1974, "So¬dium Silicate Spray Impregnation of Tunnels in the St. Peter Sandstone," Research Report, University of Min¬nesota, 118 pp.

Jan 1, 1982

By G. E. Wasson

Coal in most refuse ponds including those owned by Consolidation Coal Company can be recovered by froth flotation techniques. Laboratory studies show that the flotation conditions required for coal recovery varied greatly from pond to pond. The variables found to be most important for high coal recovery and low product ash were: a) the amount and type of flotation reagents and b) the degree of conditioning before flotation. Coal recovery at a rate of 0.504 kg/s (2 tons/hr) was demonstrated at a pond site employing a mobile froth flotation unit.

Jan 1, 1988

By Frederick W. DeVries

EPA announced, when they proposed to set a potable-water limit on cyanide, in August, 1990, that they had never encountered cyanide in a :sample of potable water, but they felt this was a sufficiently toxic and common material that it should be covered with a specification. The writer worked with both the American Mining Congress and the Aluminum Association in responding to their proposed spec. of 0.2 ppm max. of "cyanide" in· potable water. Although there are many subcategories of "cyanide", they failed to specify which type they meant.

Jan 1, 1993

By C. J. Barbee, P. J. Bethell, C. J. Wood

Heavy medium cyclones (HMCs) are commonly used to clean run-of-mine coal in the 50 x 0.5 mm size range. Unfortunately, field studies suggest that industrial HMC circuits are often poorly designed or improperly operated, resulting in significant losses of recoverable clean coal. To assist plant engineers in overcoming these problems, a spreadsheet-based simulation package has been developed to predict HMC performance based on mathematical expressions originally developed by Wood (1990). A case study is provided to illustrate the predictive capabilities of this simulation package for an industrial HMC.

Jan 1, 2004

By H. Bogart

The Golden Sunlight Underground Mine located near Whitehall, Montana was portalled from the bottom of the existing Golden Sunlight Pit in July 2002. Gold was mined by sublevel stoping in the larger ore zones and longhole back stoping in the narrower ore zones. The stopes varied from 30 feet to 100 feet in width, extended up to 200 feet in height, and were completely unsupported. Underground production averaged 1,500 tons per day with a staff of35 people. The mine operated with a break-even cut-off grade of0.080 ounces per ton.

Jan 1, 2004

"Caterpillar has developed the Cat GH800B longwall plow system, designed for mining very low coal seams at high rates. Building on the attributes of the proven GH1600 plow system, the Caterpillar longwall team in Lünen, Germany, worked with the German Ibbenbüren Mine to design and construct a system for mining an anthracite seam averaging only 0.85 m (33.5 in.) thick. The result is the GH800B plow and associated longwall system, which is now operating in the Ibbenbüren Mine - where the seam does not exceed 1 m (39 in.) in height.The Cat GH800 plow, with an installed power of 2 x 400 kW, has been the plow of choice for seam heights from 0.9 to 2 m (35 to 78 in.), regardless of seam inclination or coal hardness. But the lower the coal seam, the more difficult the extraction process. Flexibility, reliability, efficiency and serviceability of the whole system had to be refined for maximum productivity in such a minimal height, but several key modifications led to success.The armored face conveyor (AFC) modifications for the GH800B plow system include PF4 plow line pans with a replaceable top trough and a new plow guide made of a one-piece casting to provide extra strength and durability while being lower in height than the previous welded version. External dogbone connectors between the line pans and a strong and flexible relay bar between the AFC and the roof supports (shields) further adapt the conveyor for very low seam operation. To increase coal throughput, the engineering team widened the conveyor crosssection with extended side plates."

Jan 1, 2014

From a distance, the 0.75-km (0.5-mile) rail line that sits in the shadows of Vale-Inco?s Stobie Mine outside of Sudbury, Ontario could be mistaken for a standard conveyor belt hauling ore from the nearby mine. But with its large, yellow-painted loop and interconnected green rail cars that invert themselves on the loop before correcting with a smooth rollover further down the track, one might assume that this is some sort of test track for a new roller coaster. The fact of the matter is that it is the proving ground for Rail-Veyor technology, a technology that the company hopes will one day change the face of ore transportation in surface and underground mines around the world. With a line of linked, low-profile rail cars that are capable of carrying up to 1 t (1.1 st) per car of ore by way of simple, above ground, lightweight rail track, the technology is not groundbreaking. In fact, it was developed in the 1960s for the French State Railroad. A demonstration plant for the Florida Institute of Phosphate Research was installed in 2000 and the first commercial under-ground installation was in 2007 in South Africa at Harmony Gold?s Phakisa Mine. There, a 4.7-km- (2.9-mile-) long tram between two shafts operates three Rail-Veyor trains on a shared track. The difference is that the application that is being tested day-and-night in Sudbury is getting closer to wide production, thanks to the use of computer controlled automation that has helped turn an old idea into a cost-effective modern solution.

Jan 1, 2010

By T. BoBo

"In the 1990’s metal prices were trending in a long term decline and the usual cost cutting exercises were adopted. Cuts to research and development made innovation difficult and the outlook was grim. But necessity is a strong motivator, and these conditions spurred the Mine-to-Mill movement which optimised across organisational “silos” and utilising new technology tools, innovative software and increased computing power. Many applications of Mine-to-Mill exploited the fact that comminution is usually the site processing bottleneck, and that blasting is more efficient at breaking rock than grinding. This approach sought to: • Understand and characterise rock breakage from mining to the mill • Develop models and simulators for blast design, fragmentation, crusher and mill circuits • Develop tools to measure in real time the particle size distribution of rocks on conveyors and Run of Mine (ROM) muckpiles • Ensure effective communication across the silos between geologists, blast design engineers, mining engineers and metallurgists These methods and the technology tools available in the 1990s were widely adopted at numerous sites around the world with significant benefits. SAG mill throughput increases of 10-20% were common. This was the advance the industry desperately needed. It was low capital cost; it was obvious. It was here to stay. Except, in too many cases, it didn’t. In the minerals price boom of the 2000’s Mine-to-Mill was no longer necessary to “survive”, even though it was still good business. Some operations, including some of the early success stories, slipped back to old habits of optimising within organisation silos instead of across them. Fortunately, other operators still embraced Mine-to-Mill and developed new techniques. Mine to Mill was still successful, but it not as widely adopted as we had expected it to be.Now the boom has ended and operations are again under cost pressure. Since these are the same circumstances that created Mine-to-Mill, it seems time to achieve the wider adoption. Since the 1990s many new or advanced technology tools are available for Mine-to-Mill projects: blast-hole sensors, new explosives formulations, new blasting techniques and modelling, ore tracking devices, improved image analysis to determine size, texture and colour of coarse ore, grade sensors, whole-stream simulation tools, even more powerful computing hardware and data analysis software. If we could achieve so much in the 1990’s, how much more can we achieve today when we have the same imperative, the same potential and a larger number of high technology tools?"

Jan 1, 2017

By William R. Henkle

High-calcium limestone has been intermittently produced from the Monarch quarry since the late 1800s. Initial production was used in sugar processing and for a smelter located near Salida, Colorado. During the 1920s, Colorado Fuel and Iron Company (CF&I) began to exploit the deposit as a source of flux stone for its blast furnace steel operations at Pueblo. The deposit was operated for many years as a captive quarry by CF&I. In 1984, CF&I closed the quarrying operations at Monarch; flux stone was no longer required since the blast furnaces were taken off line in 198 1. In 1994, Colorado Lime Company, a subsidiary of United States Lime and Minerals, Inc. acquired the property. Limestone was mined from the Castle Butte Member of the Leadville Limestone (Lower Missippian). Some high- magnesium dolomite was also mined during WWII from the Redcliff Dolomite Member which directly underlies the Castle Butte. Past production was from both an upper and lower quarry at elevations ranging from 10,000 to 11,000 feet (3,048 m-3,353 m) mean sea level (m.s.1.). The reserves at Monarch quarry are in a dip slope deposit, occupying the northwest slope of Monarch Hill. The structural geology of the deposit is complex on both a region- al and mine scale. The reserves are located on the eastern limb of the Syncline Hill syncline. The beds on this limb of the structure strike to the northeast and dip to the northwest from 30 to 45 degrees. A major reverse fault separates the reserves into an upper and lower deposit. Numerous smaller normal and reverse faults cut the deposit, requiring a flexible mine plan. At the Monarch quarry, the Castle Butte Member con- sists of two stratigraphic units. The Upper unit (Mlu) is a thick-bedded, fine-grained crystalline dolomitic limestone from 0 to 300 feet (0-90 m) thick. The unit generally has an Si02 content of less than 4 percent. A bedded chert horizon and a karst horizon at the base of the unit locally raise the Si02 content as high as 20 percent. The underlying Quarry Stone unit (Mlq) is a light-gray, massive, bedded, calcilutite/micrite limestone from 90 to 130 feet (27-40 m) thick. The Quarry Stone unit is remarkably uniform chemically throughout it's entire thickness. Average analysis is CaC03 - 95 percent, MgC03 - 3.5 percent, and Si02 - 0.5 percent. The Footwall dolomite unit (Mld) of the Redcliff Member of the Leadville Limestone is a dark gray to black, blocky, medium-to-thick-bedded, calcilutite/micrite dolomite from 150 to 180 feet thick (46-55 m). The upper 100 feet (30 m) of the unit is a low-silica facies with an average Si02 content of less than 0.5 percent. MgC03 content is con- sistent at 39.8 percent throughout the upper 50 feet (15 m) of the unit but varies from 15 percent to 39.8 percent in the lower portion (50 to 100 feet [15-30 m]) of the low-silica facies. The lowest portion of the unit, 100 feet (30 m) to base of unit is high in silica and is not considered chemical grade. Current in-place, surface-minable reserves at the quarry are estimated to be 14 million short tons (st) (13 million metric tons [mt]) of limestone and 50 million st (45 million mt) of dolomite. A 1.4 million st (1.3 million mt) stockpile of finely crushed limestone is also present on the property. Colorado Lime is presently processing approximately 2,000 st (1,800 mt) per month from the crushed limestone stockpile. This material is pulverized and sold to underground coal mines for use as rock dust, some of this product is also used for environmental purposes. Limestone mining is scheduled to resume in spring of 2000.

Jan 1, 2001

By Glenn A. Shirey, Robert D. Stoessner, S. Gopalakrishnan

Under a contract awarded by the U.S. DOE-PETC, the Pennsylvania Electric Company and New York. State Electric and Gas Corporation are conducting a test program to evaluate and optimize the performance of a two-inch air-sparged. hydrocyclone (ASH) for cleaning fine coal. The test program is being conducted at the Homer City EPRI Coal Quality Development Center, utilizing Upper Freeport coal classified to minus 0.15 mm as the process feed material. A matrix of experiments has been conducted to establish the influence of significant design and operating variables (including overflow and underflow orifices, inlet slurry pressure, airflow rate, and reagent dosage) on the production of deep cleaned coal. Performance of the ASH is being compared to conventional froth flotation, utilizing advanced techniques and selected flotation reagents. This paper focuses on work completed to-date to optimize the performance of the air-sparged hydrocyclone for cleaning fine coal. INTRODUCTION The Pennsylvania Electric Company (Penelec) and New York State Electric and Gas Corporation (NYSEG) have been evaluating advanced fine coal cleaning processes for the production of compliance grade coal (1.2 pounds S02/million Btu). Laboratory bench-scale froth flotation results were compared with results of advanced flotation processes as performed by various organizations, utilizing the same coal feed. Among the processes evaluated, the air-sparged hydrocyclone (ASH) - developed by Dr. Jan Miller at the University of Utah - was considered to offer the best potential for retrofitting into an existing coal preparation facility because of the increased capacity per unit volume. The ASH has been shown to process over 100 times the throughput of conventional froth flotation cells. In May of 1988, Penelec/NYSEG was awarded a contract from the Department of Energy's Pittsburgh Energy and Technology Center (DOE-PETC) to optimize the performance of a five cm (two inch) ASH for producing compliance grade coal from raw classified minus 0.150 mm (1.00 mesh) Upper Freeport coal, and to compare its performance with conventional froth flotation utilizing advanced techniques and selected flotation reagents. The test work is being conducted at EPRI's Coal Quality Development Center (CQDC) in Homer City, Pennsylvania, utilizing a circuit processing 0.10 to 0.25 tons of coal per hour. Other members of the Penelec/NYSEG contract team include the University of Utah, Advanced Processing Technologies, Inc., Davy Dravo, Kaiser Engineers, Ebasco Services, and Management and Technical Systems. ASH DESCRIPTION AND OPERATION The ASH-2 System is a five cm ID (actually 4.76 cm) air-sparged hydrocyclone manufactured by Advanced Processing Technologies, Inc., under license from the University of Utah, for the purpose of research, evaluation and development. A schematic drawing of the air-sparged hydrocyclone is presented in Figure 1. Basically the air-sparged hydrocyclone consists of two concentric right-vertical tubes, a conventional cyclone header at the top, and a froth pedestal at the bottom. The inner tube is a porous tube through which air is sparged radially. The outer nonporous tube simply serves as an air jacket to provide for even distribution of air through the porous inner tube. The slurry is fed tangentially through the conventional cyclone header to develop a swirl flow in the radial direction. Air is sparged through the jacketed, inner porous tube wall and is sheared into small bubbles by the swirl flow. Hydrophobic particles in the slurry collide with these bubbles and, after attachment, are transported radially into a froth phase which forms on the cyclone axis. The froth phase is stabilized and constrained by the froth pedestal, thus moving towards the vortex finder of the cyclone header and discharged as an overflow product. Hydrophilic particles generally remain centrifuged in the slurry phase and are discharged as an underflow product through the annulus created by the froth pedestal.

Jan 1, 1989

By M. L. Harris, I. Zlochower, M. J. Sapko, I. E. Perera

Experiments conducted with limestone dusts and dolomitic marble dusts have indicated that when rock dust is wetted and subsequently dried, it becomes a solid, non-dispersible cake. However, in order to be effective, rock dust must be able to disperse as individual particles to inert. To counteract this, rock dust manufacturers created treated rock dusts that will resist caking after moisture exposure. National Institute for Occupational Safety and Health (NIOSH) researchers conducted a series of laboratory-scale experiments on four base rock dusts and their treated counterparts to assess the effectiveness of various anti-caking additives after being exposed to moisture and then dried. The dusts were exposed to moisture using humidity cabinets having a high relative humidity (99% RH) and by also exposing the rock dust bed to water through bottom wicking. The dusts were then evaluated for dispersibility after drying using the NIOSH-designed dust dispersion chamber. The anticaking additives were different concentrations of stearic acid, oleic acid, and xylene-based surfactants. All results were compared to a reference rock dust used to conduct large-scale experiments in the Lake Lynn Laboratory (LLL), Fairchance, PA. When the untreated dusts were dried after exposure to moisture for one day, no dispersion was measured. However, rock dusts treated with anti-caking agents were readily dispersible even after exposure to moisture for six months. This report details the analysis and characterization of anti-caking additives using the NIOSH-designed dispersion chamber and the 20-L explosion test chamber. INTRODUCTION In order for a coal dust explosion to propagate, five elements are required. In addition to the fire triangle in which three elements are necessary to sustain a fire—fuel (coal dust), heat (ignition source), and an oxidizer (oxygen in air)—a combustible dust explosion requires dispersion of a dust cloud (pressure wave) and confinement of a dust cloud (underground mine entry) (1). Coal dust explosions can originate anywhere in a mine. For example, when a small volume of a flammable methane and air mixture is ignited in a closed end, the high-temperature gases rapidly expand to create a pressure wave, sometimes referred to as a “pioneer wave,” that may steepen into a shock wave as it propagates away from the ignition source. The shock wave produces a wind that disperses the dust preferentially from any elevated and exposed surfaces (roof, ribs, belt structure, cribbing, etc.). Factors that are known to affect the intensity of an explosion are the dust particle size, the location of the dust within the entry, the dust dispersibility, and the volatility of the coal dust (1). Particles less than 75 μm (< 200 mesh) in size are most reactive (2). Dusts that are located on the roof, ribs, and elevated surfaces are more readily dispersed than dusts on the floor. Therefore, application of quality rock dust in sufficient quantities is essential to inert coal dust and to prevent continued flame propagation. 30 CFR 75.402 and 75.403 require 80% total incombustible content be maintained within 40 ft of the face, and 30 CFR 75.2 defines rock dust acceptable for use in underground coal mines. It also defined that “when rock dust is wetted and dried, will not form a cake”.

Jan 1, 2019

By R. H. Smith

The total energy demand in the United States increased about 4 per cent annually during the 1960&apos;s 1 and is expected to increase at the same rate during this decade. Z Fuel oil growth in the 1960&apos;s was only Z per cent per year until 1968, but it increased to 8 per cent in 1969 and 11 per cent in 1970. Fuel oil demand is expected to grow at a 6 per cent rate through the 1970&apos;s. Part of this rapid increase in demand has been caused by restrictions placed on the sulfur content of fuels. High sulfur coal can no longer be burned in many parts of the country and has been replaced by low sulfur fuel oil. An insufficient supply of natural gas has also contributed to fuel oil demand. Air pollution regulations have not only caused an increase in fuel oil consumption, but they have also caused a reduction of sulfur content in residual fuel oils. Current regulations call for 1 per cent maximum sulfur content fuels over much of the country. Sulfur contents of fuels are limited to 0. 3 per cent in certain areas on the east and west coasts. Sulfur in residual fuel oils from high sulfur imported crude oil is reduced to 1 per cent currently. However, sulfur reduction to 0. 3 per cent is difficult and expensive.4 As a result, refiners may be hard pressed to supply the market for low sulfur fuel oil. The oil shale deposits in Colorado and Wyoming potentially are a huge source of low sulfur fuel oil. It will be shown that the low sulfur content and more importantly the sulfur

Jan 1, 1972

By Cary Hirner, Tony Cicchetti, Daniel G. Cressman, Jeff F. Wallace, Prapan Dave, Malcolm Sheehan

The City of Toronto is currently undertaking the design of the Black Creek Sanitary Trunk Sewer Relief Project. The Project involves the design and construction of approximately 20 kilometers of tunnel through shale rock of the Georgian Bay Formation and soft ground soil consisting of glacial till, glaciolacustrine and glaciofluvial sand, silt and clay deposits. The 20 kilometer alignment has been split into four sections of tunnel (1) the 12.9 km, 3.0 meter diameter Keele Trunk Relief Sewer, (2) the 2.9 km, 1.5 meter diameter Keele Trunk Relief Sewer Microtunnel, (3) the 3.9 km, 1.8 meter diameter Jane Relief Sewer, and (4) the 0.5 km,1.5 meter diameter Combined Sewer Overflow Storage Tunnel. Additionally, one 40,000 m3 wet weather flow (WWF) storage tank and numerous WWF connection shafts are to be constructed and connected to the Keele Trunk Relief Sewer to pick up WWF from existing outfalls currently discharging to Black Creek. The anticipated cost of the project is approximately $600 million with construction of the Project’s Phase One Works, the Keele Trunk Relief Sewer and Jane Relief Sewer, scheduled to commence in 2025. This paper provides an overview of the project in terms of the anticipated geotechnical conditions, the scheme developed for design and construction of the tunnel system and the associated schedule and procurement strategy. The consideration for planning of the tunnel system, the risk-based analysis and decision-making process are presented and form the basis for selection of a precast tunnel lining support method in the rock and soft ground.

Jun 13, 2023

By T. P. Weldum, W. Conger

"The objective of the study conducted by the authors is to determine what media size(s) addition will maximize any given plant ball mill’s grinding efficiency (Fig. 1). The functional performance parameters “mill grinding rate through the size of interest,” and “cumulative mill grinding rates” from both plant and smallscale tests are applied to this task. A plant media sizing methodology, and industrial case studies, are provided.BackgroundA previous article (“Ball mill classification system optimization through functional performance modeling,” Nov. 17, McIvor et al., 2017, Mining Engineering) described circuit classification system efficiency (CSE), equal to the percentage of coarse (plus circuit P80 target size) material in the ball mill. It can be measured and then increased through pump and cyclone adjustments. By so transferring the application of ball mill power from fine (minus P80 product size) to coarse material, circuit efficiency is increased and overgrinding is decreased.The other factor that determines overall circuit efficiency is how well the power that is applied to the coarse material grinds it. This is affected by design and operating variables like the media sizing and mill percent solids. From circuit functional performance (McIvor et al., 1990):Circuit production rate of new fine material (t/h) = Power applied to coarse material (kW) x Mill grinding rate of coarse material (t/kWh) (1)The power applied to coarse material is the total mill power multiplied by the CSE. Therefore:Circuit production rate of new fine material (t/h) = Total mill power (kW) x CSE (%) x Mill grinding rate of coarse material (t/kWh) (2) Production rate, mill power and CSE can be measured during a plant circuit survey. The mill grinding rate of coarse material is then calculated. At the circuit target P80 size, this is the mill grinding rate (of coarse material) through the size of interest. The objective is to maximize this rate. The mill grinding efficiency can also be calculated by taking the ratio of the mill grinding rate over the material’s grindability as measured in the lab. Mill percent solids optimization has thusly been achieved (McIvor et al., 2000). For media sizing, maximizing the mill grinding rate on a given ore is equivalent to maximizing the mill grinding efficiency. Ores of different grindability will be part of the discussion."

Jan 11, 2018

By V. Kecojevic

Environmental and economic costs related to equipment fuel consumption and carbon dioxide (CO2) emission present a substantial challenge to the mining industry. Haul trucks are an integral part of the overall surface mining system and they consume a significant quantity of fuel. Consequently, they produce a significant amount of CO2. The objective of this research is to (i) establish a mathematical relationship among a truck?s fuel consumption, power and engine load factors and (ii) determine the amount of a truck?s CO2 emission and the associated cost that may arise from potential CO2 legislation. In order to achieve these objectives, the authors have considered original equipment manufacturer (OEM) haul trucks, which are commonly used in surface mining operations. The research presented here may be used by mining professionals to help determine the cost and environmental burden of the trucks? application and efficiently manage energy consumption.

Jan 1, 2010

By J. D. Noll, J. A. Organiscak, A. B. Cecala

Significant strides have been made in optimizing the design of filtration and pressurization systems used on the enclosed cabs of mobile mining equipment to reduce respirable dust and provide the best air quality to the equipment operators. Considering all of the advances made in this area, one aspect that still needed to be evaluated was a comparison of the efficiencies of the different filters used in these systems. As high-efficiency particulate arrestance (HEPA) filters provide the highest filtering efficiency, the general assumption would be that they would also provide the greatest level of protection to workers. Researchers for the U.S. National Institute for Occupational Safety and Health (NIOSH) speculated, based upon a previous laboratory study, that filters with minimum efficiency reporting value, or MERV rating, of 16 may be a more appropriate choice than HEPA filters in most cases for the mining industry. A study was therefore performed comparing HEPA and MERV 16 filters on two kinds of underground limestone mining equipment, a roof bolter and a face drill, to evaluate this theory. Testing showed that, at the 95-percent confidence level, there was no statistical difference between the efficiencies of the two types of filters on the two kinds of mining equipment. As the MERV 16 filters were less restrictive, provided greater airflow and cab pressurization, cost less and required less-frequent replacement than the HEPA filters, the MERV 16 filters were concluded to be the optimal choice for both the roof bolter and the face drill in this comparative analysis case study. Another key finding of this study is the substantial improvement in the effectiveness of filtration and pressurization systems when using a final filter design.

Aug 1, 2016

By E. H. Christiansen

Gem-quality red beryl in the Wah Wah Mountains of southwestern Utah formed as a post-magmatic mineral in a topaz rhyolite flow. Unlike topaz, this beryl occurs along fractures in devitrified rhyolite, rather than in lithophysal vugs. The flow occupies a graben and perhaps a significant paleo-drainage. Beryl occurs at medial depths within the eroded flow, about 170 m below overlying mafic lava flows. Both beryl-bearing fractures and the host rhyolite show an unusual amount of argillic alteration compared with other topaz rhyolite flows. Alteration may be related to the incursion of surface water along shrinkage fractures within the flow. Whole-rock concentrations of CaO within the host rhyolite are very low (<0.01 % to 0.18%) compared with other topaz rhyolites and Be concentrations (about 20 ppm) are about average. Low Ca activities prevented fluorite precipitation and allowed Be transport as fluoride complexes. Beryl probably grew at subsolidus temperatures (-300-650°C), but above the temperature of kaolin development (200-300°C) as fluoride-rich vapors, released by devitrification, encountered fractures and mixed with vapors derived from surface water flowing through the fractures. Beryllium fluoride complexes reacted with water vapor, silica minerals, alkali feldspar, water, and Fe-Mn oxides along fractures to produce red beryl. Continued reaction of the cooling flow with surface water at lower temperatures probably produced a boiling, more acidic fluid that produced the argillic alteration. Consequently, eruption of low-Ca topaz rhyolite followed by entrance of surface water into medial portions of a devitrifying and still hot lava flow appear to be critical factors that led to red beryl formation.

Jan 1, 1996

By B. Simser, T. Butler

Glencore’s Nickel Rim South Mine, located in Sudbury, Ontario, Canada, is a primary/secondary blasthole operation with ore zones lying between 1,100 and 1,700 m below surface. One of the main benefits of the primary/secondary mining method in high stress ground is that up to 50% of the mining will take place in secondary stopes that should be in a yielded rock mass state. Every future secondary mining stope begins as a primary stope abutment pillar. As with many mines, the stress loading history and rock mass conditions can vary considerably in space and time. Safe mining practices cannot allow for an assumption that the planned secondary stopes are indeed yielded. An incorrect assumption can result in mining development and stoping taking place in high stress ground rather than the expected yielded rock mass conditions. This can result in higher levels of seismicity than expected and incorrect ground support recommendations in areas where workers could be exposed. Seismic monitoring is the mines’ premier tool to track the actual rock mass response to mining. Yielded zones have a reasonably clear seismic signature/pattern over time that is visible in the collected data. The ability to accurately locate small magnitude seismic events can help define and delineate different stages of rock mass failure. The vast majority of the seismic data recorded by the microseismic monitoring system is below moment magnitude (Mw) 0.0. Examples are provided showing how small seismic events (Mw < -1.5) can help provide feedback from the rock mass deformation that allows the ground control engineer to gain a better understanding of how the rock mass is responding to mining processes.

Apr 26, 2022

By D. Nilsson, B. Aaro

To find out if there are any potential savings in "in-the-wall haulage for open pits," the Swedish companies ASEA and Kiruna-Truck in 1984 gave us financial support to study this solution in more detail. In the study, the use of in-the-wall haulage was studied for three hypothetical open pits with different sizes and shapes. Annual ore production rates also varied. A part of this thorough study was published by the Swedish Mining Research Foundation (see Report F 8444, "The Use of Electric Trucks and In-the-Wall Haulage in Open-Pit Mining," in Swedish with an English Summary). Although in-the-wall haulage is of interest in some open pits, the local terrain is important for profitability. The authors do not think in-the-wall haulage is of any major interest for the mining industry. The following is a summary of some of the author&apos;s findings. To place the haul road in the wall is not of interest in open pits with declining metal contents in the bedrock. In such mines, the volumes from the haul roads will not yield revenues. The cost per m3 for an underground haul road is much higher than for a haul road in the pit. This means that it is only of interest to place the haul road in the wall when mining gets deeper than about 50 m (165 ft). In-the-wall haul roads will reduce flexibility in the pit, and it will make necessary the use of smaller equipment with lower productivity and higher costs per ton. As Hustrulid, et al. show (Table 4), excavation savings of material hauled are very low, $0.10 to $0.25/t (0.09 to 0.27 per st). The extra operating cost, due to lower productive equipment, will normally be much higher and thus destroy the whole idea with in-the-wall haulage. If electric trucks are profitable for in-the-wall haulage, it is normally also profitable to use electric trolley assist for trucks on the haul road in the pit. But the profitability of using electric power is different in different countries, and depends on the relation between the cost for electric power and diesel fuel. In the US, diesel fuel is inexpensive compared with electricity, but in a country like Sweden, diesel fuel is much more expensive. In most open pits, the trucks have to move a considerable distange from the loader until they reach a final haul road in the wall. A trolley line along such temporary haul roads will be exposed to flyrock. It is normally less expensive to perform rock support from the open pit than from underground ramps in the wall. In Fig. 8, Hustrulid et al. gives the impression that the underground haul roads will be very close to each other. This is seldom the case. Figure 1 shows a haul road in the wall. The haul road passes through each cross-section only once. Arranging a reliable dewatering and rock support system from only one underground ramp is probably impossible. Many more drifts are therefore necessary. In our report, we also studied what would happen if the final pit slope is increased by 5°, using underground drifts in the wall. Our conclusion was that, the extra cost for drifts, rock support, pumps, etc. destroys the whole idea, and that it was better to accept a higher stripping ratio with the haul road in the pit and to use conventional low-cost open pit equipment. Finally, we think it is would be difficult to try to reduce the safety factor when determining the slope angle by moving the haul road in the wall. Minimizing the risk for pit wall collapse is also important with in-the-wall haulage, primarily because men and equipment will be working in the pit, but also to guarantee many accesses between the pit and the haul road in the wall.

Jan 1, 1990

By D. Nilsson, B. Aaro

To find out if there are any potential savings in "in-the-wall haulage for open pits," the Swedish companies ASEA and Kiruna-Truck in 1984 gave us financial support to study this solution in more detail. In the study, the use of in-the-wall haulage was studied for three hypothetical open pits with different sizes and shapes. Annual ore production rates also varied. A part of this thor¬ough study was published by the Swedish Mining Research Foundation (see Report F 8444, "The Use of Electric Trucks and In-the-Wall Haulage in Open-Pit Mining," in Swedish with an English Summary). Although in-the-wall haulage is of interest in some open pits, the local terrain is important for profitability. The authors do not think in-the-wall haulage is of any major interest for the mining industry. The following is a summary of some of the author&apos;s findings. To place the haul road in the wall is not of interest in open pits with declining metal contents in the bedrock. In such mines, the volumes from the haul roads will not yield reve¬nues. The cost per m3 for an underground haul road is much higher than for a haul road in the pit. This means that it is only of interest to place the haul road in the wall when mining gets deeper than about 50 m (165 ft). In-the-wall haul roads will reduce flexibility in the pit, and it will make necessary the use of smaller equipment with lower productivity and higher costs per ton. As Hustrulid, et al. show (Table 4), excavation savings of material hauled are very low, $0.10 to $0.25/t (0.09 to 0.27 per st). The extra operating cost, due to lower productive equip¬ment, will normally be much higher and thus destroy the whole idea with in-the-wall haulage. If electric trucks are profitable for in-the-wall haulage, it is normally also profitable to use electric trolley assist for trucks on the haul road in the pit. But the profitability of using electric power is different in different countries, and depends on the relation between the cost for electric power and diesel fuel. In the US, diesel fuel is inexpensive compared with electricity, but in a country like Sweden, diesel fuel is much more expensive. In most open pits, the trucks have to move a considerable distange from the loader until they reach a final haul road in the wall. A trolley line along such temporary haul roads will be exposed to flyrock. It is normally less expensive to perform rock support from the open pit than from underground ramps in the wall. In Fig. 8, Hustrulid et al. gives the impression that the underground haul roads will be very close to each other. This is seldom the case. Figure 1 shows a haul road in the wall. The haul road passes through each cross-section only once. Ar¬ranging a reliable dewatering and rock support system from only one underground ramp is probably impossible. Many more drifts are therefore necessary. In our report, we also studied what would happen if the final pit slope is increased by 5°, using underground drifts in the wall. Our conclusion was that the extra cost for drifts, rock support, pumps, etc. destroys the whole idea, and that it was better to accept a higher stripping ratio with the haul road in the pit and to use conventional low-cost open pit equipment. Finally, we think it is would be difficult to try to reduce the safety factor when determining the slope angle by moving the haul road in the wall. Minimizing the risk for pit wall collapse is also important with in-the-wall haulage, primarily because men and equipment will be working in the pit, but also to guarantee many accesses between the pit and the haul road in the wall.

Jan 1, 1989