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By F. F. Peng, S. S. Peng, V. Z. Sun

A detailed study of ponding phenomenon relative to surface subsidence due to longwall mining was conducted for various surface stream conditions. Based on field monitoring, the stream-ponding process and three forms of stream water are described. Also discussed are the criteria of stream ponding and the influences on the ponding process by two key factors: the angle of stream flow and the change in the stream-bed gradient. A numerical (mathematical) model, which combines trend surface technique that describe the topographical features in the vicinity of a stream and the modified influence function, which represents the profile of the surface subsidence trough (or basin), is constructed to predict the formation characteristics of stream ponding in terms of location, boundary and accumulated water volume of migratory and stationary ponds produced by the stream-ponding process. Three cases are analyzed using the model to predict the interaction of the underground operations and the environmental impact due to stream ponding. The feasibility of various mitigative measures are assessed.

Jan 1, 1997

By M. P. A. Hames, D. R. Beaumont

INTRODUCTION Legend has it that when Isembard Kingdom Brunel was confronted with the cost and consequences of one of his bridges having failed, he responded cheerfully that the event had saved investors and taxpayers a fortune because it prevented him from repeating the same design at several other locations. To meet the demands of rejuvenating and expanding the transportation system in an era of industrial growth Brunel and his contemporaries introduced bold designs and many innovations. Great advances in design and construction techniques resulted, yet these too had their practical limits as the story tells. In the gold mining industry today we face the same challenge of expansion, yet designers must respond mindful of economic restraints. Is innovation the answer? It has two meanings; to bring in novelties and to make changes. Distinguishing between improvement and novelties, or simply recognizing the benefits and limitations of new ideas over proven technology is part of good design. This paper will illustrate the practical aspects of selecting appropriate solutions to a design problem by examining a variety of examples common throughout the industry. Since a variety of engineering disciplines contribute to the design of a gold processing plant, and each has some old wisdom to impart or new development to boast this paper is divided into the different areas of expertise involved.

Jan 1, 1987

By A. S. Azman

In collaboration with Kennametal Inc. and corry rubber corporation, the U.S. National Institute for Occupational Safety and Health (NIOSH) developed a drill bit isolator to address noise overexposures associated with roof bolting machines in underground coal mines. NIOSH laboratory studies confirmed that the drill bit isolator reduces noise during drilling. Field studies were needed to confirm that a noise reduction could be obtained under working conditions and that the device was sufficiently durable. This paper reports results of field tests of the device conducted at five underground coal mines. Noise reduction was assessed by comparing the operator?s noise exposure during drilling with and without the drill bit isolator. Durability was assessed by recording the number of holes and total feet drilled with each bit isolator until either the test period ended or the device failed. The results from these tests showed that the device is an effective noise control in a mine environment. The field-tested drill bit isolators provided a noise reduction of 3-5 db(A). Of nine devices tested for durability, five exceeded 610 m (2,000 ft) drilled and two exceeded 762 m (2,500 ft) drilled before failure. Durability issues found in the field tests led to final production optimizations that have resulted in a commercially available product for drilling with 35-mm- (1.3-in.-) diameter roof bits and hexagonal drill steels.

Jan 1, 2012

By R. Mathur

Cu isotope fractionation in low-temperature aqueous environments has been well documented. Here Cu fractionation was applied to two distinct, yet related, mining problems. First, an exploration geology application, where copper isotope fractionation was measured in oxidized leach caps, chalcocite enrichment blankets and hypogene mineralization in three porphyry copper deposits located in the southwestern U.S. Cu isotope patterns exist in a vertical distribution within the mines, coupled with distinct relationships with the Cu grade. Second, a hydrometallurgical application, where copper isotope fractionation was measured in leach fluids, feed and residues from a quick leach test (QLT) on Cañariaco, Peru ores. The results indicate the relative proportions of the minerals providing Cu during leaching. Interpretations of the QLT test measurements indicate covellite is the dominant ore mineral providing Cu during leaching. Both applications clearly demonstrate the use of the Cu isotope technique to understand low-temperature aqueous processes involving copper.

Jan 1, 2012

By John N. Murphy

The past five years have seen communications techniques and hardware developed by the Bureau of Mines and its contractors increasingly brought into use in U.S. mines. The Bureau of Mines philosophy has been to help develop and introduce new hardware that goes as far as possible toward meeting the mining industry's needs. The Bureau of Nines programs are also designed to provide continuous fallout of new communications capabilities which are introduced to the mine operators as they occur. As more progress is made in meeting mine comnications needs, it is hoped that the mining industry will readily adopt the equipment and techniques that it finds most useful in meeting its individual needs. This paper reports on recent advances achieved in the development of improved nine communications systems.

Jan 1, 1976

By Bill Kendall

"In coal mines, handheld air, electric and hydraulic powered drills are used along belt conveyors, in single entry development, in entries where there are obstructions, and in longwall recovering bolting. In hard rock mines, pneumatically powered jackleg drills are typically used in narrow vein mining. Drilling and bolting in narrow confines has traditionally been performed with pneumatically or hydraulically powered hand-held drills or newer electrically powered hand- held hammer drills due to their low cost, simplicity of operation, and extreme flexibility due to their small size. Mechanized equipment has typically been too large, too expensive, inefficient due to the confined space, and too restrictive due to lack of flexibility for use in narrow confines.Although the handheld equipment is effective, it requires operator exposure to certain inherent hazardous conditions in use. These include falling rock from the roof, rib and face; tripping hazards; bending twisting lifting hazards; high noise levels and chemical fumes from the pneumatic exhaust. The operation of the hand held equipment can adversely affect the long-term health of the operator - continual jackleg operation frequently results in longterm damage to hearing, shoulders, arms and wrist joints and back injuries. This results in a reduction in life style for the operator and higher personnel turnover rates for the mining company.Hand-held drills also typically have limited thrust and torque capability that limits their ability to push long resin bolts, inadequate torque to properly install resin bolts or to properly torque mechanical or tensioned rebar bolts. Hand-held drills do not offer any type of protection for the operator from roof or rib falls. They do not offer any means for handling roof or rib mesh. Since the operator stands on the ground the maximum safe drilling height without use of a ladder or "riding the drill" is limited to 2.44 m (8 ft)."

Jan 1, 2014

By Rui Xiang Yang, Cong Li, Jia Zeng, Rong Liang Zhang, Wei Zhang, Chao Fan Tang

This research took the roasted products obtained by leaching copper (Cu) refining slag using waste acids and roasting leaching residues with calcium oxide (CaO) as raw materials and used oxidative leaching to extract Cu from roasted products using waste acids. Influences of the leaching temperature, liquid–solid ratio, and sulfuric acid ( H2SO4) concentration in waste acids on the leaching rate of Cu in the oxidative leaching process of roasted products using waste acids were investigated. The unreacted-core shrinking model of heterogeneous reactions was adopted to reveal the leaching kinetics of Cu in the oxidative leaching process of roasted products using waste acids. Results show that the leaching rate of Cu reaches 91.68% when the leaching temperature, liquid–solid ratio, H2SO4 concentration in waste acids, and injected airflow are 80 ℃, 8:1, 1.5 mol/L, and 200 mL/min, respectively. If the leaching temperature is below 80 ℃, the leaching process of Cu is controlled by the solid-film diffusion, and the apparent activation energy Ea is 11.0835 kJ/mol; at a leaching temperature above 80 ℃, the leaching process of Cu is governed by the interfacial chemical reaction, with an apparent activation energy Ea of 42.3183 kJ/mol. The liquid–solid ratio can positively influence the reaction rate constant. The apparent order of the leaching reaction of Cu is 0.632.

Jan 11, 2023

By C. McKinnon

Climate change is the central challenge facing the mining industry, wherein producers must simultaneously ramp up production to provide the raw materials for clean technologies while transitioning to carbon neutrality. While transitioning to renewables will take time, mines can start improving energy efficiency while improving productivity today. This paper proposes a novel technology ecosystem that can help. By monitoring mine shovels, loaders, haul truck paths, and conveyor belts, the proposed technology enables particle size measurements throughout the operation, truck and belt volume monitoring, ground engaging tools monitoring, and boulder detection.

Mar 2, 2022

Brace yourself-what used to be a 1 billion ton deposit is now a 1 Gt deposit. And that 1520-hp ball mill is now a 1.134-MW ball mill. Beginning with this issue, all engineering units and statistics in MINING ENGINEERING will be reported in SI units. The exception to this is the Technical Papers section, where units may still be reported by dual notation. For the past four years, ME has reported all units of measurement first in SI, followed in parentheses by the equivalent English value. This metrication policy, adopted by the SME-AIME Board of Directors in September 1975, was designed to aid magazine readers through a transitional period from English unit reporting to SI unit reporting. In October 1979, the SME-AIME Publications Board completed the English-to-SI transition by requesting the editorial staff, beginning February 1980, to report exclusively in SI. The Publications Board also requested the editors to prepare a chart of SI conversions useful to ME readers; this information follows on pages 156-157. The What and Wherefore of SI In 1960 the General Conference on Weights and Measures met in Sevres, France, and recommended that a new metric system-one that could uniformly cross all international borders-be adopted to replace all other metric systems. The outcome of this recommendation was the SI system of units, an abbreviation for le Systeme International d'Unites (the International System of Units). It replaces the more familiar cgs (centimeter-gram-second) system and the m/ksA (meter-kilogram-second-Ampere) system. SI consists of seven base units, two supplementary units, and a number of derived units. Other units, called allowable units, may be used even though they are not part of SI. The seven base units are: meter (length); second (time); kilogram (mass); mole (amount of substance); kelvin (temperature); ampere (electric current); and candela (luminous intensity). The two supplementary units are radian, a measurement of plane angle, and steradian, a measurement of solid angle. Supplementary units can be used in combination with SI units, e.g., radiant intensity is expressed in W/sr, watts per steradian. Derived units are combinations of base and supplementary units. Velocity is expressed as meters per second (m/s), and acceleration as m/s'. Using the relation established by Newton's second law of motion, force is expressed as kg-m/s2, more commonly known as the newton (N). Energy (work) is force times distance, expressed as kg-m/s2 and known as the joule (J). Power is work per unit time, expressed as kg, m2/s3 and known as the watt (W). Allowable units are the exception to the rule-they are not taken from base or supplementary units but, nevertheless, are used in the SI system. Volume units are normally expressed as the cubic measure of a length, such as cm 3. But a fluid volume of a cubic decimeter (dm3) is called a liter (L); the "L" is capitalized in another special SI rule to keep the lowercase "1" from being confused with a numeric "1". Other allowable units are degrees Celsius (°C) for temperature; and the minute (min), hour (h), day (d), and year (a) for time measurements. The prefixes in the table on the following page are used with SI units to express the magnitude of a measured quantity. A mill that processes 1000 tons of ore per day would be designated a 1 kt/d mill; a mine producing 10 million tons of ore per year would be called a 10 Mt/a mine. The rule here is to use prefixes that keep the value in a convenient range from 0.1 to 1000. In other words, it would be incorrect to write 25 000 kg-the prefix should be changed to indicate 25 Mt. One of the distinct advantages of SI is that it retains the decimal relation between multiples and submultiples of the base units for each physical quantity except for time. One more rule of SI syntax-numbers are separated into groups of three with a space between each group rather than a comma, e.g., 3,839,174 becomes 3 839 174. By the same token, a decimal number is also broken by spaces, e.g. 0.738153 is written as 0.738 153.

Jan 2, 1980

By P. Richardson, B. Sceresini

The carbon-in-pulp and the carbon-in leach process for the recovery of gold and silver from cyanide leaching of ores is now widespread in the precious metals mining industry. However the presence of various base metals can have a detrimental effect on the efficiency and the economics of the process due to high cyani,:••? consumption and a reduction in leaching and adsorption rates resulting in poor recovery of the precious metals. Furthermore the presence of fast leaching and adsorbing metals such as copper can significantly reduce carbon loading capacity for the precious metais thereby necessitating a higher carbon inventory, more frequent stripping and poor bullion quality resulting in higher production costs. This paper describes the development and application of a process for the recovery of copper and and the recovery and re-use of the associated cyanide from the leach circuit at the Mt. Gibson Gold Project in the Murchison Goldfield in Western Australia. The process provides a means to counter-act the detrimental effects of the copper which occurs in the ore as various oxide and sulphide minerals including an extremely fine chalcopyrite powder. It has the added benefit of recovering the base metals in this instance copper, in a saleable form. The process is subject to a patent application by Reynolds Australia Metals, Forsayth N.L. and the author.

Jan 1, 1991

By Kelly Sands

"Mitigation may be considered a hurdle when it comes to environmental permitting. For complex projects with site specific mitigation plans, the time to devise those plans and get them approved can be a long and costly endeavor. However, mitigation options are expanding and the mitigation banking industry has grown tremendously over the last two decades, helping projects move forward with less debate. Mining companies are recognizing the value that mitigation banks provide in helping to shore up project timelines and costs. Where possible, companies are taking advantage of the opportunities to buy credits from thirdparty providers offering mitigation bank credits. In some instances, mining companies are taking matters into their own hands by establishing mitigation banks to realize their own mitigation freedom. That can have its advantages and disadvantages, as discussed below. All in all, the expansion of mitigation options has streamlined the permitting process, introduced flexibility, improved project permitting timelines and loosened the permitting girdle.Compensatory mitigationCompensatory mitigation involves actions taken to offset unavoidable adverse impacts to wetlands, streams and other aquatic resources and authorized by Clean Water Act section 404 permits and other permits. As such, compensatory mitigation is a critical tool in helping the federal government meet the longstanding national goal of ‘‘no net loss’’ of Waters of the U.S., both in acreage and function.Section 314 of the National Defense Authorization Act for Fiscal Year 2004 required the Secretary of the Army, acting through the Chief of Engineers, to issue regulations “establishing performance standards and criteria for the use, consistent with Section 404 of the Clean Water Act, of on-site, offsite, and in-lieu fee mitigation and mitigation banking.” The provision also required that those regulations ‘‘maximize available credits and opportunities for mitigation, provide flexibility for regional variations in wetland conditions, functions and values, and apply equivalent standards and criteria to each type of compensatory mitigation’’ (73 FR 19595). The result was the Final Rule on Compensatory Mitigation for Losses of Aquatic Resources issued in 2008 (33 CFR Parts 325 and 332).”"

Jan 6, 2017

By Z. Henderson, J. R. Saylor, L. Rojas-Mendoza

"Diesel particulate matter (DPM) is a major occupational health hazard in underground mine environments, even with available abatement technologies. Prior work has demonstrated that micrometer-scale water droplets, or “fog,” may be effective in removing significant number fractions of DPM from a conditioned exhaust stream. Here, a series of laboratory experiments are described which demonstrate that such a treatment can remove significant DPM mass from an unconditioned stream. The removal mechanism involved coagulation of the DPM and drops. The fog treatment removed about 45 percent more DPM mass on average than no treatment. Varying engine load, flow rate and fog droplet number density did not significantly affect the results under the range of conditions tested. IntroductionDiesel engines have seen widespread use for well over a century due to their relatively high thermal efficiency and fuel economy (Heywood, 1988). Recently, however, the adverse health risks of diesel exhaust have become increasingly clear. The term diesel particulate matter (DPM) is used to refer to the solid components of diesel exhaust, which are an ultrafine mixture of elemental and organic carbon (EC and OC) and minor constituents including sulfates and metal ash (Kittelson, 1997). DPM is generally considered to occur almost entirely in the submicrometer range. It is classified as a carcinogen (Occupational Safety and Health Administration, 2013), and epidemiological studies have demonstrated a positive correlation between long-term exposure to DPM and other combustionrelated fine particulates and increased cardiovascular and pulmonary diseases (Pope et al., 2002; McDonald et al., 2011). Many of the risks of diesel exhaust are associated with the physical and chemical properties of exhaust components (Heywood, 1988; El-Shobokshy, 1994; Kittelson, 1997). Exposures are generally measured and regulated on the basis of mass concentration. However, DPM number density and particle size are increasingly recognized as critical factors in terms of health outcomes (Bugarski et al., 2012; Kittelson, 1997; Occupational Safety and Health Administration, 2013; Pope et al., 2002)."

Jan 11, 2017

By J. A. Organiscak

Enclosed cab filtration systems are typically used on mobile mining equipment to reduce miners? exposure to airborne dusts and diesel particulates generated during mining operations. Various filter configurationscan be used within the heating, ventilation and air conditioning (HVAC) systems of cabs to remove the airborne contaminants. The air cleaning performance of cab filtration systems is highly dependent on the efficiency of the air filters, the number of filters used and their placement within the HVAC system. This paper shows how to mathematically model cab filtration systems and illustrates a methodology for developing other cab filtration system design models. Node diagrams of several filtration system circuits are illustrated with their mathematical models at steady state airflow and concentration conditions. Laboratory and field data are presented to demonstrate the validity and utility of modeling enclosed cab air cleaning performance. The paper concludes that incorporating multiple filters throughout the HVAC system improves the air cleaning performance and robustness of the cab filtration system design

By Frank Florez

INTRODUCTION The Magma mine, located at Superior, AZ, some 100 km (60 miles) east of Phoenix, was discovered in 1875, and was originally mined for silver. The property was acquired by W. B. Thompson who formed the Magma Copper Co. in 1910. Prior to 1965, square-set mining and cut-and-fill methods were used to mine steeply dipping veins at Magma. Ore was moved to the chute by gravity in the square-set stopes and by slusher in the cut-and-fill stopes. The stopes were gobbed with waste from the surface and development headings in the mine. The original square-set method of mining was modi¬fied in order to extract replacement ore bodies in limestones that dipped at 0.52 rad (30°) and varied in thick¬ness from 4.6 to 38 m (15 to 125 ft). Mining of these ore bodies was done by starting at the upper lateral ex¬tremities and retreating to the access openings. The timber square sets and cribbing were used as temporary support, the ground being allowed to cave as the mining retreated. It was found, however, that mining in such a manner was very inefficient; safety hazards increased be¬cause of the blocky and fractured characteristics of the ore and the weakness of the hanging wall. GENERAL DESCRIPTION The undercut-and-fill method of mining is now the primary stoping method used at Magma, and no pillars are left for later extraction. Four ore bodies are pres¬ently being mined. These ore bodies (see Fig. 1) dip conformably at 0.52 rad (30°) to the east with the host limestone and shale sediments, their average strike length is about 122 m (400 ft), and their thickness varies from 9 to 38 m (30 to 125 ft). The ore bodies are separated by varying thicknesses of barren rock. Three of them come together at various elevations primarily in the area of the main feeder fault, but there is an apparent horizontal displacement of about 23 m (75 ft) on the main feeder fault which complicates stope layout. The method as originally applied has been modified in order to mine more efficiently ore bodies that are 9 m (30 ft) or more in thickness. The modification was made to mine all the ore at a given elevation before dropping down for the cut below. The advantage of this is that mining is done on a horizontal plane and there¬fore muck removal can be mechanized. Also multiple faces are made available, thereby improving overall min¬ing efficiency. Another major change made was to carry a 6.7-m (22-ft) face when mining the top sill and a 6.4-m (21-ft) face when mining under fill rather than the former standard 3.35-m (11-ft) high face. This change lengthens the mining cycle and cuts mat prepa¬ration costs in half. Stopes are laid out (see Fig. 2) so that maximum mining panels are available and yet one stope crew will not tie up too much of the ore reserve. One stope might consist of as many as 20 panels, and these panels are divided evenly on opposite sides of the extraction cross¬cut. The stope layout is planned so that the maximum tramming distance to the orepass will not exceed 30.5 m (100 ft) on the crosscut and 38 m (125 ft) in the panel. Stope panel distances are sometimes exceeded when min¬ing to the extremity of the ore body. Tramming distance limits are determined by the efficiency of the com

Jan 1, 1982

By YOUNIS S. MUSTAFA, GHORASHI ZAIN, ABDEL-ZAHER M. ABOUZEID, AHMED A. S. SEIFELNASSR

The present paper attempts to characterize the breakage properties of a brittle rock, quartz, at different size fractions when compressed at various loads in a confined piston die.  This is to understand the interaction between the various parameters affecting the breakage behavior of a brittle solid in a confined zone. Application of the results may lead to improving the energy efficiency in the size-reduction units, which use compression as the main stressing force. 

By M. A. Trevits, A. Ozment

"According to the Office of Surface Mining Reclamation and Enforcement, Abandoned Mine Land Inventory System, in 2013 there were 98 underground mine fires in nine states. This is considered to be an underestimate for the actual number of fires nationwide. Many mine fires are started by people burning trash where the coal seam or an abandoned coal mine is close to the surface. Other fire ignition sources include lightning and forest fires. Once ignited, a coal mine fire can easily spread into the remaining coal pillars and mine entries. Once established, the fire creates its own ventilation system supporting further combustion by drawing air down into the workings through unsealed mine shafts, fractures and surface subsidence depressions. As the coal left in the workings from the past mining operations burns, the mine void can collapse, causing subsidence and creating dangerous voids, damaging overlying surface structures and roadways. The products of combustion include smoke and noxious fumes such as carbon monoxide gas. These products are released to the atmosphere through fractures that develop within the ground surface, killing vegetation and creating serious health hazards (Richmond, 2017).Abandoned mine fires, if left uncontrolled, can burn for years and, in fact, one of the most wellknown mine fires in the United States is the mine fire in the Borough of Centralia, PA, that began on May 27, 1962. Reportedly, officials decided to clean up the local landfill for the upcoming Memorial Day holiday by burning the garbage on the site. Unfortunately, the landfill was located on the site of an abandoned strip-pit mine on the edge of the Borough (Dekok, 2009). The strip pit had been left open after being excavated around 1935 and was approximately 23 m (75 ft) wide and 15 m (50 ft) deep. It is believed that the lack of a properly constructed noncombustible shale barrier in the strip pit enabled the fire to spread to adjacent carbonaceous refuse material and then to nearby coal mine workings (PA DEP, 2017). As time went on, the fire grew and spread under the Borough and directly affected the residents despite various efforts to extinguish the fire by the state and federal governments (Fig.1). In 1984, a voluntary program was begun to move residents from their homes. Many accepted buyout offers for their properties and moved elsewhere and after leaving, their homes were demolished. In 1992, the state used eminent domain to take control of all the property within the Borough. In 2013, after years of litigation, the eight remaining residents were allowed to stay for as long as they lived. Today, only a few buildings remain within the Borough (Centralia, PA, 2014). This fire has been burning for more than 55 years."

Jan 5, 2018

By Robert E. Mclvor

Classification system efficiency (CSE) is a simple metric that can be used to measure and increase ball milling circuit efficiency. It is defined as the percentage of the ball mill energy being used to grind target (oversize) particles. This paper describes CSE analysis conducted for a variety of operations with related improvements in circuit efficiency ranging from 3 to 30 percent. The achievable CSE for any given ball milling circuit is described, providing the means to determine the extent to which every plant can benefit from increasing grinding circuit CSE.

By Michael J. Fiannaca

Disseminated epithermal gold metallization occurs within a breccia unit interpreted to be a submarine debris flow. Cold is associated with oxidized pyrite, fluorite, and jasperoidal silica. A minimum of 9 million tons of open-pittable mining reserves will be cyanide heap leached as run-of-mine ore at a rate of 1.0 million tons per year. Careful evaluation of the combined results of successive sampling projects has established the pit plan, mineable reserves, and process metallurgy of the operation. The sampling procedure utilized for a given test depended upon geological interpretation as well as the procedures, successes, and failures of previous tests. This paper summarizes the sampling and testing procedures of the development effort in light of the project goals and attendant problems in achieving them.

Jan 1, 1984

By Lok W. Home

HISTORICAL DEVELOPMENT At the beginning of the 1970 more than 90% of pro¬duction drilling in open-pit mining was obtained with the use of rotary drills. At the same time, the use of rotary drills underground was virtually nonexistent. Several progressive mines began questioning whether efficient rotary drilling could not be adapted to under¬ground applications. In the late 60s and early 70s at¬tempts to use rotary drilling were made by Inco, Ltd. and Meramac Mining, utilizing drill units designed as a combination of small diameter raise drills and rotary drills. These initial attempts were unsuccessful primarily due to the immobility of the equipment and low degree of dependability. At the same time, experiments con¬ducted by Inco and Meramac using large diameter in¬the-hole percussion drills produced immediate encourag¬ing results. The first serious attempt at utilizing rotary drilling underground was made by Texasgulf in 1971 at Kidd Creek mine in Canada. An experimental stope 15 m wide, 36 m long, and 30 m deep was drilled using 200-mm diam rotary drilled holes. The machine used was a Robbins' Model 11 D. The Model 11D is a rotary drill designed to drill drain and utility holes under¬ground. The results of this test revealed that large diameter blasthole stoping was feasible. The indications were that the major areas of concern-fragmentation and pillar damage-could be controlled and a new low-cost stoping method resulted. At this point Texasgulf and Robbins entered into a cost-sharing agreement to develop a rotary drilling unit specifically for this stoping method. Drilling of the trial stope required pouring a concrete pad at each drill hole plus anchoring the derrick to the pad by means of rock¬bolts. Utilization expressed in terms of drilling time to available time was less than 20%. It was obvious that the required unit had to be mobile, setup time had to be minimized, and rod handling had to be improved. The first attempt at an underground production-type rotary drill consisted of mounting the basic 11D on an air-powered Eimco 600 Series crawler. The crawler was equipped with four screw-type leveling jacks operated by an impact wrench; rod rack facilities for 60 m of 1.2-m long, 170-mm diam rods; a rod handling facility consisting of an air-lift swing boom and a pipe loader which automatically positioned the rod under the float¬ing drive box; a turnbuckle anchor system requiring in¬stallation of four rockbolts in the drill drift floor and placing turnbuckles from the rockbolts to the corners of the drill; and trailing power units containing the electrics and hydraulics required to operate the drilling unit. Operation of this unit revealed a vast improvement in drilling time but some problem areas remained. The anchoring system required secondary drilling and proved difficult to adapt to the localized irregular floor surface and the trailing power unit was difficult to maneuver and was causing unnecessary maintenance costs. Addi¬tional improvements resulted in the standard 11MD rotary blasthole drill. DESCRIPTION OF UNIT This standard model consists of the drill and the power unit integrally mounted on an air-powered crawler chassis (Fig. 1). The rod-handling and leveling¬jack systems are similar to those used on the prototype Texasgulf unit and remained relatively the same. The anchoring system has been replaced by the roof jack, initially screw type but now hydraulically actuated. The anchoring system is necessary as the drill exerts more thrust than the weight of the unit. During operation it is found that the two front jacks adequately anchor the machine. The complete unit's operational dimensions are: height, 3.2 m; width, 2.1 m; and length, 4.2 m requiring a 3.5- x 3-m drill drift. The derrick is pivot mounted and raised and lowered by a hydraulic cylinder reducing the transport height to 2.5 m. The derrick is also turnbuckle mounted to allow dip angle drilling from vertical to 0.52 rad (30°) off vertical. The drill derrick can be mounted to allow dip axis to be parallel or perpendicu¬lar to the long axis of the unit. The standard rods are 1.2 m in length and are contained in racks of 10 which can be removed for transportation. Drilling torque is produced by a 56-kW (75-hp) electric motor driving a double vane pump which in turn drives a Robbins high torque gear-type hydraulic mo¬tor. In addition to the hydraulic motor, the machine drive system has a 3.9:1 planetary gear reducer. This results in six speeds with the following minimum and maximum range: Maximum rpm 112 Torque 3725.5 J (380 kg m) Minimum rpm 10 Torque 14 706.0 J (1500 kg m) Drilling thrust is provided by two bootstrap cylin¬ders pressurized by a variable volume pump. This sys¬tem can produce from 0 to 267 231.2 N thrust in either direction. Bailing air (approximately 17 m3/min at 382.2 to 490.3 kPa) is introduced through a rotary swivel pass¬ing directly through the center of the drive system to keep efficiency and losses to a minimum. Hole annulus is also kept to a minimum by using 170-mm rod when drilling a 200-mm hole. A small amount of water is added to the air for dust suppression. The machine has to be supplied with air, water, and electrical power. It does not appear feasible to provide an integral air power supply with the rotary drill such as is done with open-pit rotary drills. Air is required and installed during the primary development of the drill drift. The machine can be adapted to drill hole diameters from 170 to 228 mm for blasthole operations. The drill string consists of the rotary bit, crossover stabilizer, roller stabilizers, and standard rods. The operator controls speed and thrust pressure from an operating console which can be mounted in a conve¬nient position. Mount Isa has made an adaption to the operating console to permit remote collaring of drill holes in an uncleared floor.

Jan 1, 1982

By John A. Shuster

INTRODUCTION Controlled ground freezing for mining and construction applications has been in use for over a century. Despite the great technological evolution which has occurred during this period, ground freezing is still used on projects today with much the same basic technology as that used originally over 100 years ago. Major developments have occurred in the separate fields of frozen soil mechanics, refrigeration, and heat transfer analysis. However, except for a few projects in Europe and Asia, these developments have not been combined to upgrade ground freezing techniques to their potential level. In view of this situation, it is believed desirable to examine the art of ground freezing in light of recent technical developments, together with highlights on some of the apparent advantages, disadvantages and economics of the various alternative approaches. The purpose of this paper is to attempt such an examination, with particular emphasis on the practical application of presently available technology. GENERAL ADVANTAGES AND DISADVANTAGES Ground freezing may be used in any soil or rock formation, regardless of structure, grain size or permeability; however, it is best suited to soft ground rather than rock conditions. Freezing may be used for any size, shape or depth of excavation and the same physical plant can be used from job to job despite wide variation in these factors.

Jan 1, 1997