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By Floyd C. Bossard

A method of hand-calculating the air-cooling requirements of a conceptualized underground mining operation is presented for the reader's orientation. Separate air heat load calculations were conducted for adiabatic compression, electromechanical equipment, wallrock, broken rock, groundwater, and blasting operations. The air heat sources were calculated for four mining levels under conditions representative of a typical planned mining operation at depth. The total heat gain on each level will approximate the level air-cooling requirements. The results of these hand calculations can be further modified by the use of mine ventilation computer software that refine the heat-source calculations, predict underground ambient air temperatures, and establish the air-cooling requirements of a mine. INTRODUCTION The principal mining method employed is a modified vertical crater retreat (VCR) blasthole operation. Typical scopes range from 20 to 40 feet wide, 75 to 100 feet long, and 100 feet in height. Six and one-half inch blastholes are drilled with an "in the hole" hammer drill. Eight-foot high horizontal rounds are blasted down. Mucking of ore from the undercut to orepass is done with LHD equipment operated from a remote control station. Backfill includes hydraulically placed tailings with cement, and waste rock when available. Stope access is from ramp sub-levels on 50-foot vertical intervals. Crosscuts are ramped down to the first stope cut 25 feet below the sub-level elevation. Then the crosscuts are raised by taking down the back when each stope cut is completed, until an elevation 25 feet above the sub-level is reached. The crosscuts are filled with tailings, and/or waste rock. See Figure 1. Typically, a two-pronged approach to defining the air-cooling requirements is conducted. First, the principal sources that make up the air heat load are individually hand-calculated. Second, projections of mine heat load are calculated by utilizing computer modeling techniques. This paper discusses the first method (hand calculations) of determining the individual components of heat flow into the mine. CALCULATED AIR HEAT LOADS Adiabatic Compression The plans for the mine include delivery of 300,000 cfm of air to ventilate the lower level operation. This is equivalent to approximately 200 cfm/ton of rock produced (300,000 cfm/1500 tpd of ore and waste). As air flows down a shaft, with no heat interchange between the shaft and air and no evaporation of moisture taking place, the air is heated in the same way as if it were compressed in a compressor. Dry air increases in temperature about 5.4°F per 1000 feet. One BTU is added to each pound of air for every 778 feet of decrease in elevation, or is subtracted for the same elevation increase. For dry air, the dry-bulb temperature change is 1/(0.24 x 778) = 0.00535°F/ft., or 1°F/187 ft. elevation. Auto-compression may be masked by the presence of other heating or cooling sources, such as shaft wallrock, groundwater, air and water lines, electrical facilities, etc. The major factors influencing the temperature of the air delivered underground by a shaft are (1) the night time cool air temperature's effect on the rock or lining of the shaft, (2) temperature gradient of ground rock related to depth, and (3) evaporation of moisture within the shaft which increases the latent temperature and decreases the sensible temperature. [Calculation For Adiabatic Heat of Compression] a. Assumptions: 1. Three hundred thousand cfm of fresh air at 3000 level has increased in temperature during the summertime to the point where it has little available cooling power. Air-cooling will be required on 3000 Level and below. 2. Elevation of 3000 Level is approximately +1000 ft. above sea level.

Jan 1, 1993

Hoses provide a variety of critical functions in nearly every mining operation. Care must be taken when selecting, installing, and maintaining hoses to insure maximum hose life. This article provides some tips on selecting, installing, and maintaining hoses, and gives some background on hose construction and safety. In the mining industry, hoses are used for everything from hydraulic systems that power drills and hammers, to suction systems that carry away dust and debris. Hoses are used on such a large scale that a detailed discussion of specific applications would require many volumes. Dayco Corp. has, however, provided these basic and brief guidelines on how to select and maintain hoses for mining applications. Hose Construction Industrial hoses are classified as much by construction as they are by application. So, it is a good idea to start any discussion of hoses with a description of their basic construction. There are four basic components to hoses: the tube, reinforcement, cover, and couplings. The tube is the inner core or lining, and is usually made of smooth rubber or plastic. Rubber or plastic composition is scientifically determined to be compatible with the materials passing through it, such as fluids, chemicals, gases, or substances. Other considerations in compounding rubber or plastic for the tube are temperature ranges and stresses. In practice, the required compound can be precisely chosen from various charts listing them according to application. There are, however, applications where the compounding of tube rubber or plastic must be customized. The reinforcement hose layer immediately surrounds the tube. As its name implies, this layer gives the hose strength and provides protection for the tube component. Reinforcement must be able to contain the inner forces-pressure and shock-transmitted through the tube, and such external forces as twisting, bending, and crushing. So it is usually made of high tensile strength cord, fabric, or wire that is wrapped, wound, or braided around the tube. The cover is the outermost layer, so it may encounter oils, acids, water, dust, abrasion, and other potentially damaging influences depending, of course, on the application. An appropriate cover material must be selected, usually a rubber compound. Couplings, even though they are a hose attachment, are as vital as any other component and should be considered part of the hose. Couplings are usually made of carbon steel or malleable iron; however, bronze or stainless steel might be used if corrosive chemicals are involved and brass or aluminum might be used for flammable fluids. A coupling is usually in two parts: an insert that fits into the hose bore and contains the threaded male, female, or other type of connection, and the ferrule or collar that fits over the hose and secures the coupling in place. There are, however, some one-part couplings that fit securely into the hose without a ferrule. These are usually limited to low pressure applications and are secured to the hose by clamps or metal bands. Some fittings can also be permanently "built-in" as the hose is made. Couplings can be classified as reusable or permanent. The reusables can be removed when the old hose expires and used again on a new hose. The permanent types are crimped or swaged on at the factory or in the field (using special equipment) and must be discarded with the hose. Although couplings are most commonly thought of as having threaded fittings, there is another type called "quick connect." These either have a crow-foot or jaw-type design that permits a fast, "snap" connection. This can

Jan 6, 1982

By Lawrence Pelham

The importance of fluorspar to the United States and the world and the role it plays in industry and society extend far beyond that which is suggested by its measured production and consumption quantities and declared value. Compared to many other industrial minerals, the annual worldwide consumption of 5 million short tons valued at less than $500 million are not impressive figures. Fluorspar is made special primarily because of its status as a critical and strategic commodity; it is essential for the processing of most steels and all aluminum. It also is the raw material for the production of elemental fluorine, which plays a part in uranium refining.

Jan 1, 1988

By Leo Misagi

THE ROLE OF MANAGEMENT IN PLANT SAFETY Safe Production In plant management, safe production is an ideal. It becomes reality when: 1. Plant employees work in healthful and safe conditions. 2. Plant safety is in harmony with productivity. 3. Worker absenteeism and turnover are low-the work force is stable. Plant management must apply its talents, skills, and resources to build and maintain the foundations of safe production on a system of policies, procedures, and attitudes. It must do so by marshaling all managerial functions - staffing, direction, control, innovation, and representation-toward that goal. Ideally, a plant safety program is proactive. Management should plan and implement safety activities that are aimed at prevention. Past accidents-accomplished facts-should not be the sole guide of future actions. In other words, positive preventive measures, not the fewness of accidents are the criteria for safety production. An ideal safety program is one which is cost effective. Management should view plant safety as part of a broader program-loss control. Loss control includes prevention of fires and other accidents, unhealthful conditions, and occurrences that cause bodily injury, illness, property loss, or damage. Management evaluates the employees' productivity and safety. An effective training program aimed at loss control is the hallmark of safe production. Ideally, both labor and management should be well versed in safety. They believe in the premise that work-related injury and illness, as well as property loss and damage from fire and other accidents, are manageable and preventable. Such belief prompts them to learn more about the work process and potential hazards associated with each job. They also learn how to expect and prevent fues and other accidents, and respond to emergencies at the plant. Plant management determines which individuals or groups of employees are accountable for the following functions:

Jan 1, 1991

By E. Milosavljevic, C. H. Bucknam, K. L. Gabby, T. C. Eisele

"Since mercury is chemically similar to silver, it tends to dissolve along with the silver and gold during cyanide leaching. Ultimately, mercury is recovered along with the silver as mercury/silver/gold amalgam, which requires retorting to remove the mercury prior to refining the silver and gold. A possible approach to separating mercury from silver and gold in cyanide leach solutions is to take advantage of the replacement reaction between [Hg(CN)4]2- and Ag2S, which results in precipitation of mercury as HgS while dissolving silver as [Ag(CN)2]-. Results are presented demonstrating that mercury can be removed from cyanide solutions with high effectiveness while, at the same time, recovering additional silver in solution. Initial experiments with static flasks in the laboratory showed a high degree of mercury removal could be achieved. Results using actual metallurgical plant products in bottle roll tests indicate the process works on a larger scale, but also results in significant gold losses if the solution being treated has high gold levels, such as carbon stripping pregnant eluent solution. IntroductionThe dissolution of mercury along with precious metals during cyanide leaching is a long-standing problem, which can result in significant mercury toxicity issues (Washburn, 2003). Mercury is first dissolved by the cyanide solution, and then, since it has a great deal of chemical similarity to silver, it follows silver throughout the concentration and recovery processes, ultimately being recovered as a silver/ mercury amalgam when the metal is precipitated by either the Merrill-Crowe process or by electrowinning.While it would be desirable to either prevent mercury dissolution entirely while still allowing silver dissolution, or to selectively capture mercury before the precious-metals recovery step, there is not yet an effective and economically practical approach to doing this (Grosse, 2003; Ravichandran, 2004; Misra, 1998; Pai, 2000). There are a few processes that have been proposed, such as differential adsorption. While activated carbon, as used in the carbon-in-pulp process, is not highly selective between mercury and silver (Washburn, 2003), it may be possible to achieve some separation by selective elution of the carbon to produce high-mercury and low-mercury eluates (Bunney et al., 2010). Certain synthetic resins were shown to be selective for gold but are more expensive than carbon, requiring special equipment, and have not yet been shown to be selective for mercury (Grosse, 2003). There may be potential for differential dissolution approaches, as thiosulfate leaching does not dissolve mercury sulfides as readily as it dissolves gold (Oraby et al., 2010), and it is possible that it may be similarly more selective for dissolving silver relative to mercury. Some work has also been done with selective precipitants, such as sodium polythiocarbonates (Bucknam, 2007), and 1,3-benzenediamidoethanethiol (Matlock, 2002). Common problems with these approaches include introduction of a solid:liquid separation step, reagent costs, the need to measure the mercury concentration for process control for minimizing silver losses, and regulatory requirements to retort otherwise stable mercury concentrates."

Jan 1, 2015

By Stephen J. Klein, Fernando Orduz, Randall J. Essex

INTRODUCTION The Keystone Tunnel project consists of widening an existing 7.3 m-wide (24 ft) highway tunnel to a width of 13.7 m (45 ft) to provide a three-lane, two-directional highway tunnel. The 13.7 m-wide tunnel was excavated in a staged drill-and-blast sequence through foliated schist, and was supported using pattern rock reinforcement and shotcrete. Notable project features are: the low vertical cover over the tunnel is less than the span of the opening; the project is the second instance where the Federal Highway Administration (FHWA) has accepted rock reinforcement as a permanent support element; and the project is the first instance where the FHWA is accepting shotcrete as the final interior surface of the tunnel. This paper covers the design and construction phases of the project. PROJECT DESCRIPTION The South Dakota Department of Transportation (SDDOT) retained Woodward-Clyde Consultants (WCC) to evaluate the feasibility, prepare construction plans and specifications, and provide construction management services for the widening of the Keystone Tunnel. The tunnel is located in the Black Hills National Forest area of South Dakota near the southwest corner of the state. The Keystone Tunnel is on U.S. Highway 16A, about 0.8 km (1/2 mile) north of Keystone which is approximately 32 km (20 miles) south of Rapid City (Figure 1). Highway 16A serves as an important transportation link between Rapid City and the Mount Rushmore National Monument which is located about 2 miles west of Keystone. The road through the tunnel consisted of two lanes about 7.3 m (24 ft) wide, and is being widened [ ]

Jan 1, 1989

By H. E. Rojas, G. T. Adel

The primary objective of most comminution operations is liberation; however, mill performance is generally quantified in terms of particle size. Only recently has the use of image analysis equipment made it possible to quantify mill performance on the basis of liberation. In this investigation, a series of tests were carried out to study the effect of mill operating parameters on the liberation of sphalerite from a sphalerite/dolomite ore. The parameters studied included mill rotational speed, mill loading and ball size. The results of this work appear to indicate that sphalerite liberation is improved under conditions that promote impact breakage, while dolomite breakage is enhanced under attrition conditions.

Jan 1, 1992

By Scott G. Britton

Until now, only the range of typical construction projects which are repeated over and over during the life of a coal mine have been discussed. This chapter looks at underground construction projects which occur only once or twice during the life of any given mine and are major construction projects in both time and cost. These projects usually require the services of an outside engineering and construction firm with hands- on experience in design installation. Because of this involvement with engineering and construction firms (EIC), the role of the construction engineer will change from one of actively de- signing and directing tasks to one of planning, inspection, and quality control. To reflect this role change in the discussions, the focus of this chapter will be changed to a broad overview of project participation. Of course, the projects under consideration in this chapter could easily fill more than one book on their design and installation. In addition, major construction projects are extremely site specific, requiring months of data collection about site geology, physical conditions, and potential excavation and construction requirements. Therefore, no attempt shall be made to discuss any detailed construction aspects as was done in earlier chapters. The purpose of this chapter is to present the fundamental concerns of a staff engineer or supervisor. This engineer does not make the decisions on choosing the EIC firm, but assists and interacts with their project team to accomplish the goals set forth by higher management. The construction topics to be reviewed fall into three main categories. The first category is highlighted by the installation of a main surge bin; the second is typified by the construction of an underground shop; and the third category encompasses special projects such as rotary dumps, crusher stations, and foremen's offices. These categories highlight the differences between degrees of mine responsibilities for project completion. The first topic is usually a turnkey project by an E/C firm for a given mine. The second topic represents the leading role of the mine in the project with a contractor as a support service or equipment supplier only. The third topic, special projects, is an example of a joint venture-between a contractor (leading partner) and mine management. These roles can change with any given project, but the changes in management responsibilities for the construction engineer can be nicely illustrated for the purposes of this text. The first step toward constructing major mine projects is to analyze the mine which influence their design. From the mine management standpoint, these parameters should reflect the role the mine will play in the project construction. In other words, what is the nature of the proposed mine environment and what are the material, labor, and time requirements needed to assist or direct the chosen engineering and construction firm? Answering these questions requires some basic project planning and cost estimation work. Here, only a general treatment is necessary, since detailed discussions of these topics will follow in Chapters 8 and 10. For those persons interested in the role of the EIC contractor, I suggest beginning by reviewing the article by Beall, "The Role of an Engineering and Construction Contractor," presented in the October 1980 issue of Mining Engineering magazine.

Jan 1, 1983

By Wang Yingmin, Xu Zhuyun

The optimum of general mine ventilation networks is very difficult and has a considerable significance. Based on the content model, this paper presents a more general nonlinear programming model whose number of variables equals that of independent meshes. The constrained conditions are linear inequalities only. The model is changed into an unconstrained problem with multiper method. The optimum results are computed with DFP algorithm and the Improved Golden section method is used in its linear searching. The computer program written in Quick BASIC is suitable to the optimum of any type of large mine ventilation networks with different regulating devices. The paper presents two examples for optimum of ventilation networks.

Jan 1, 1991

By Abhiram Kumar Verma, Soumya Ranjan Mallick

Generally, waste such as soil and sandstone is used in mines for road construction or haul roads. In many such mines, it is challenging to design haul road pavement to match the dumper capacity. This paper investigates overburden soil to find its utility as a pavement material. It is found that overburden soil is unsuitable pavement material without stabilization; hence, nine composite materials are developed using overburden soil and RBI. Physical, mechanical and microstructural properties of unstabilized and stabilized soil are determined. Statistical analysis is carried out to assess the effect of soil type, RBI and curing period of strength characteristics of unstabilized and stabilized soil. ANOVA tests assess the influence of RBI percentage on soil geotechnical properties, and regression relations are developed. Two-way ANOVA shows the effect of RBI and soil type on soil geotechnical properties, whereas one-way ANOVA shows that the influence of soil type is more than the RBI. The regression relation is developed for CBR, UCS, Poisson’s ratio and modulus of elasticity with a good correlation coefficient. The performance of RBI is more with the finer material and reduces with an increase in the size of soil grains. The performance of the developed composite is better than the earlier developed composite based on overburden with fly ash, lime kiln dust, lime or clinker.

Mar 19, 2024

By Gordon Miner

GENERAL DESCRIPTION At the Star mine, Hecla Mining Co., located in Burke, Shoshone County, ID, the mining method is horizontal slicing (timbered or untimbered) cut-and-fill using hydraulic sand backfill. Back stoping as well as breast stoping is used; see Fig. 1. The annual produc¬tion from all cut-and-fill mining is 259 816 t (286,400 st). The average depth ranges from 366 to 2408 m (1200 to 7900 ft). A majority of the tonnage comes from an average depth of 2240 m (7350 ft). Average travel time to the stope is 1 hr. GENERAL ORE BODY REQUIREMENTS AND LIMITATIONS Size and Ground Conditions The ore body is 610 to 1219 m (2000 to 4000 ft) along the strike and 2408 m (7900 ft) on the dip, with an average vein width of 2.6 m (8.5 ft). The dip is essentially vertical. The ore body is heavily sheared and broken on the main fault structures; it is more competent on the wall veins but everywhere subject to rock bursting at depth. The hanging and footwalls are strongly fractured on the main fault structures, again more competent but still fractured on the wall veins. All host rock is Revett quartzite. Value and Production Requirements The ore value is $56 per ton. Production require¬ments are 1089 t/ d (1200 stpd), five days per week. Why Cut-and-Fill? The slusher extraction method is the most satisfac¬tory and practical for narrow vertical veins with sheared walls. The method must be quite selective to be com¬mercial. The veins are too narrow for rubber-tired ex¬traction and the grade is not high enough to make resuing practical. DETAILED DESCRIPTION Development Access to the main shaft is by the 3.2-km (2-mile) main haulage tunnel. The level intervals in the vertical timbered shafts are 61 m (200 ft). Vein development is by crosscut and laterals located 30.5 m (100 ft) into the vein wall. Vein crosscuts are on ±61 m (200 ft) centers. Access to the vein is by raises carried up with the stope. The raises are equipped with four wood¬cribbed chutes (two orepasses reinforced with angle irons, one manway, and one timber slide, Figs. 2-6). The initial cut is located 6.7 m (22 ft) above the main level (Fig. 1). Stoping Cycle Detail Sampling: Grab samples are taken from the mine cars. Channel sampling of the faces and backs is done when necessary. Drilling: For stope drilling, Ingersoll-Rand 300M percussion drills are used, with Ingersoll-Rand 300M and Gardner-Denver RB83M stopers, used for back stoping. Breasting down is done with Gardner-Denver S83FM jacklegs. The back stoping rounds are 0.9 to 1.8 m (3 to 6 ft) and the breast stoping rounds 1.5 to 3m (5 to 10 ft). Blasting: For stope blasting DuPont Tovex 100 watergel with fuses and igniter cord is used primarily. Approximately 60% of the blasting is done with am¬monium nitrate-fuel oil (ANFO). For development headings and shaft sinking, DuPont Drivex watergel with Ensign Bickford Nonel Primadet primers is is used. Mucking: Broken ore is removed entirely by slusher in the stopes. These are mostly 3.7-7.5-kW (5-10-hp) Ingersoll-Rand air-powered or 7.5-37.3-kW (10-50-hp) Ingersoll-Rand electric-powered. In the horizontal de¬velopment headings Eimco 12B and 21 and Atlas Copco LM-56 overshot mucking machines and a 0.76 (1 cu yd) hydroelectric Wagner HST-1 LHD (load-haul-dump) unit are used. Rockway for Extraction: The cribbed chutes (Fig. 7) are nearly vertical and hexagonal in shape and are made of spaced 152 x 203-mm (6 x 8-in.) timber, reinforced with varying weights and sizes of angle iron, depending on wear requirements. Rectangular cribbed chutes are used where ground pressure is not a problem. Cribbing is added in 2.4 or 3-m (8 or 10-ft) extensions accord¬ing to the floor height (whether timbered or untimbered). See Figs. 8 and 9. Backfilling: Backfilling is done with hydraulic sand consisting of classified mill tailings at 65% solids. The fill contains approximately 20% 325-mesh fines. The tailings are pumped 3353 m (11,000 ft) at 50-55% solids 22.7 dry t/h (25 dry stph) with a duplex piston pump. Cycloning increases the tailings density to 65% solids. Fill storage is in a 998-dry-t (1100 dry st) agi

Jan 1, 1982

By Dean H. Herrick

The production of crushed stone in the United States has exceeded 1.0 Gtpy since 1987. Eighty percent of the crushed stone produced is used in the aggregate industry, accounting for over half of the quantity of natural aggregate consumed in the United States. Approximately 718 of the crushed stone produced in 1989 was limestone and dolomite followed by granite, traprock, sandstone and quartzite, miscellaneous stone, marble, shell, calcareous marl, volcanic cinder and scoria, and slate in order of descending volume (Table 1). Of the remaining 20% crushed stone produced in the United States and not used as aggregate, 12% is used in the manufacture of cement which is an important construction material. Only 8% of crushed stone is used in specialty products not associated with the construction industry. Therefore, this chapter will concentrate on crushed stone as used in aggregates, with occasional reference to other uses or to chapters which discuss specific nonconstruction uses of crushed stone. Construction aggregates are hard, essentially inert materials suitable for being formed into a stable mass by the addition of cementing or binding materials that produce a concrete by compaction, or by natural weight to produce a road or foundation fill. Aggregates are usually from 80 to 100% of the material volume of the construction masses in which they are used. Conditions necessary for a rock deposit to be developed for construction crushed stone include: Quality-should pass rigid specifications for strength and durability; Quantity-adequate volume of rock is needed to have a production life of at least 10 years, preferably 20 or more years; Market-must have adequate market to support a new crushed stone operation; Transportation-transportation cost must be competitive for the intended market; Environmental-the impacts of mining, processing and transportation of crushed stone on the environment at or surrounding the mine site must be within acceptable limits; Permitability-mining, processing and transportation of crushed stone must usually be permitted by governing agency. The following discussion will touch upon all facets of crushed stone operations from the mineral deposit and its geology to technology, uses, economic factors and end with governmental regulations of mining crushed stone. The sections are not designed to answer all questions, but rather to be an introduction to the crushed stone industry. Where more information is required, The Aggregate Handbook by Barksdale (1991) is recommended as the best general source.

Jan 1, 1994

By C. J. Johnson, C. J. Bise

Abstract - As the US coal industry moves into the next century, it is becoming more apparent that the effective application of new technology is the only way it can remain competitive in the energy marketplace. Examples of this emphasis are with us today in such diverse areas as face- ventilation and excavation-equipment applications. The focus of this new technology is directed toward improvements in health and safety, cost control, and productivity. Room-and-pillar mining accounts for approximately 68% of the nation's underground coal production; this percentage is not expected to decrease markedly in the near future. As such, this paper analyzes the impacts of the changing techno- logical climate on future room-and-pillar operations by comparing a base-case section (current) to three scenarios incorporating new or emerging technology. A microcomputer-based comparative analysis will reflect the changes in section production and costs.

Jan 1, 1990

By Mark E. Smith, Gerald J. Gierszewski

Introduction The purpose of this paper is to present a procedure used by the authors for evaluating and designing soil liner systems. This method is particularly valuable in evaluating various treatment schemes for chemically amended soil liners. A tabulation of laboratory test results on various soil types are presented to quantify the effectiveness of certain treatments. A typical liner design program includes developing and proving soil borrow sources, designing the cross-section of the liner system, developing construction specifications, and providing construction services to ensure the intended product is achieved. Material Source Development The first step in designing a soil liner is to identify and evaluate suitable borrow sources within an economical haulage range. This is best done in a two step approach: a reconnaissance level investigation to identify target areas and a detailed evaluation of those targets. Reconnaissance: The goal of the preliminary investigation is to locate potential borrow sources for liner quality soils. This includes all natural materials which can be compacted, chemically treated, or otherwise amended to yield an installed permeability at or below some target value. This requires utilization of all available data sources: Soil Conservation Service, BLM, aerial photos, USGS geologic maps, and project geologist records. The goal at this stage is to locate shallow deposits of favorable soil types. The Unified Soil Classification System provides an excellent first pass grouping. Clays, clayey sands and silts are the most favorable soil types, although silty sands and occasionally clayey gravels can make excellent liners, and are often amenable to chemical modification. The lowest permeabilities are generally achieved with CH, CL and MH soils. Once preliminary targets have been identified using visual examination, laboratory classification tests should be performed to further refine the selection. Testing at this stage should include gradation, plasticity and hydrometer analyses. Additionally, "preg-rob" testing should be done as early as practical. Preg-rob is a phenomenon where gold or silver ions in solution associate with the clay, or other, minerals. When this occurs, a portion of the gold or silver leached from the ore is actually tied-up by the clay and thus a reduces recovery. Testing for this consists of agitating a small sample of the soil in a solution containing dissolved gold or silver, preferably of similar chemical make-up as the solution which will contact the actual liner. The solution and soil are assayed before and after agitation to determine loss to the clay. A reliable estimate of the hydraulic conductivity, commonly referred to as permeability, can be developed from the D10 value by the use of Hezen's formula: K = 100 (D1012 This relationship is limited to soils where the finer particles do not move due to the force of flowing water (i.e.: "hydrodynamic stabilitym)(1). Additionally, the effect of platty particles on permeability is not as predictable as the effect of equidimensional particles. D10 is the sieve opening size at which 10% of the material is finer. Plasticity is also important from several standpoints. Constructability is directly related to plasticity. Very plastic clays and non-plastic silts both tend to be difficult soils, while medium plastic clays and clayey sands are generally very desirable. Post construction performance is also related to plasticity (e.g. swelling, shrinkage cracking, frost heave, etc.). Additionally, low plasticity silts and silty sands generally do not respond well to chemical amendment. Source Development: The result of the reconnaissance evaluation should be an estimate of the relative probability of developing a suitable borrow source within an economical haul distance. Of course, "economical distance" depends on the degree of handling and treatment the borrow material requires, as well as the cost of synthetic alternatives. The purpose of the detailed investigation is to prove out quantity and quality of material sources, and determine design parameters such as degree of compaction, mixing, treatment and thickness of liner. The emphasis of the testing program should be permeability and strength. Strength becomes increasingly important as the slope of the liner and the height of the heap increase. Permeability testing should evaluate the effects of compaction, water content, mixing and chemical treatment where appropriate. The effects compaction and water content during compaction have on

Jan 1, 1987

By William M. Greenslade

As the search for minerals leads to deeper mines the need to control water inflow into mine shafts and mine workings will grow. As shal¬lower more readily available minerals are exploited, ore bodies which were previously uneconomic be cause of their depth will now become attractive targets. In known mining districts new ore bodies are being discovered down dip of existing ones. In the west many metal deposits are being developed in rocks of the mountain front pediments, frequently covered by several hundred feet of saturated alluvium. Greater depth often means that workings are developed well below existing water tables, creating high hydrostatic heads adjacent to mine workings. The purpose of this paper is to review some of the methods available for determining if water control will be necessary and to present some of the techniques for locating water bearing zones and the amount of water to be expected. The paper will concentrate on problems associated with water control in mine shaft development, although many of the exploratory techniques are similar to those needed for assessing overall mine water inflow problems. A brief case history on water control problems in mine shafts in the Grants mineral belt of northwestern New Mexico is presented. Water control during shaft sinking presents some unique features. First, the problem is a relatively short term one, since water control is often necessary only while the shaft is being sunk. Control measures may be necessary only while the shaft is open into the water bearing zone and the lining installed and cemented. Secondly, limited working space within the shaft area presents difficulties in handling large amounts of water inflow. The presence of water in the shaft working area increases sinking time and the potential hazard to miners. Finally, there is the economics of shaft sinking. While the shaft is being sunk no ore is being produced and there is every incentive to minimize costs.

Jan 1, 1979

By R. J. Smith

Tar sand deposits in Utah represent more than 25 billion barrels of in-place bitumen, 96% of the known U.S. reserves. The technological development of a hot water processing strategy for the Canadian tar sands and the commercial Production of synthetic crude oil from this natural resource suggest that similar success may be possible for the utilization of Utah tar sands to provide our country with a new domestic source of synthetic crude oil. In this regard an active research program to develop a hot water processing strategy for Utah tar sands has been sustained at the University of Utah for the past five years. The hot water process for Utah tar sands differs significantly from that used for Canadian, tar sands due to the inherent differences in the respective bitumen viscosities and the nature of the bitumen-sand association. These differences have led to the identification of certain design criteria necessary to achieve satisfactory phase disengagement during digestion. (8,9) Subsequent flotation of the digested tar sand is a subject of current attention and is considered in this contribution for the Asphalt Ridge tar sand deposit. Although contact angle measurements of pure Asphalt Ridge bitumen indicate moderate hydrophobicity, air bubble attachment to the bitumen concentrate taken from a 37 liter flotation cell is not possible. This surprising result suggests that the flotation separation is dependent on air bubble entrapment rather than on attachment due to surface hydrophobicity. The occlusion of air bubbles in the bitumen is apparent from visual examination of the concentrate, especially at lower temperatures. It seems that a bitumen bubble agglomerate forms in the impeller region of the flotation cell, the effective density of which is such to allow the agglomerate to float to the surface. A factorial design of the major operating variables in the flotation separation indicated that the quality of separation was significantly dependent on the flotation temperature and to a lesser extent on the degree of agitation. For a flotation temperature of 77C recovery of 96.7 percent was realized at a grade of 61.0 percent bitumen. The improved separation at higher flotation temperatures was found to be due to the decrease in bitumen viscosity resulting in more effective rejection of coarse sand from the concentrate.

Jan 1, 1980

By Stuart K. Nicol, Malcolm D. Engel, Neville T. Moxon

Ion flotation of gold is a process continuing to be developed to selectively beneficiate dilute aurocyanide liquors using a novel surface­active complexing agent. Aeration of these liquors causes accumulation of the precious metal species in the froth concentrate produced in a flotation column. Successful development of ion flotation for concentrating solids-free liquors would offer a low--cost method for recovering specific metals from heap leach runoff or waste tailings solutions. The cost of constructing and operating a 100 tph gold ion flotation facility has been compared with the relevant carbon adsorption plant costs. Reductions in operating costs of 25% and installed capital of 40% are possible in relevant expenditure areas. For an 'average' processing plant, this equates to annual operating savings of A$ 170 000 and a capital saving of A$ 515 000, with a three-point increase in total (mine and plant) site DCF ROI. High gold inventories are eliminated and the number of processing steps is reduced. The application of gold ion flotation will initially be confined to heap leach operations in the western U.S.A., although use of the process in mine abandonment programmes elsewhere is also an area of future development.

Jan 1, 1992

By John C. Draper

The principal object of an underground coal mine transportation sys- tem is to move coal from the face where it is produced to the outside of the mine where it is prepared for market. In addition to the movement of coal, a transportation system must also move supply materials to the inside; it must move men in and out of the mine; and it must move rock and other debris out of the mine for disposal. Usually, mine haulage is divided into three categories named for the various areas of the mine served : 1) Pace Haulage. This is used at face from the mining machine to the intermediate transport system. 2) Intermediate or Secondary Haulage. Generally this is used between face haulage and main haulage. Sometimes this is merely an extension of the main haulage. 3) Main Haulage. This extends from the secondary haulage to the outside and is usually either a track or conveyor system, sometimes a combination of both. Transportation modes of coal in the various areas pose different problems, but the whole system must be synchronized. For instance, at the face the coal should be removed quickly from the various places to facilitate the movement of machines, whereas the intermediate system just passes the same coal on to the main transportation system. Selection of the transportation system must be based on the seam height, mining methods used to extract the coal, capacity required, ease of coal handling, and haulage efficiency. When driving rooms, mine cars can be used. They often simplify supply haulage and rock handling because they can also handle these items. Conveyors could be used and would have a greater haulage capacity but they cannot handle supplies easily or segregate the rock. When using a different mining method, such as longwall, mine cars cannot be used because there is not sufficient space to move the cars at the face nor can the cars haul the coal without delays which would cause excessive losses in production. As for ease of handling at a longwall face, conveyors are convenient because of their low height, ruggedness, and ability to operate while being advanced. Similarly, in low seam heights, conveyor systems are easier and more economic to operate than the regular track haulage because roof has to be taken down for handling mine cars of adequate size.

Jan 1, 1981

By R. S. Rickard, B. Ball

INTRODUCTION Pyrite responds to the flotation reagents normally used in sulfide flotation systems. For example, it is well known that xanthates are good collectors for pyrite, and cyanide, sulfide and hydroxides are depressants. The particular response of pyrite to these and other reagents is used to advantage in several areas. By careful control of depressant and collector additions, pyrite can be separated from other sulfides by flotation. This is of obvious importance in base metal sulfide flotation where pyrite depression is often required to produce a marketable grade concentrate. Flotation of pyrite has long been employed in gold ore processing, and it appears that pyrite flotation from coal may become important in the future (1). It is probably a direct consequence of this relative ease of separation, plus the low economic value of pyrite, that detailed investigations on the chemistry of its flotation and depression are relatively limited. Gaudin (2,3,4) and others, in the mid-1950's published data on the flotation of pyrite, but detailed or consistent mechanisms were lacking. In 1955, Sutherland and Wark's book (5) was published, in which a large amount of empirical information, especially earlier work of Wark and Cox, was presented. More recently, in 1968, significant contributions have been made by Fuerstenau et al. (6,7) and Majima and Takeda (El. These papers introduced electrochemical explanations of the effects observed with xanthate and cyanide ions.

Jan 1, 1976

By William C. Paschen

Damen Avenue to Roosevelt Road, the Tunnel and Reservoir Plan (TARP), a Deep Tunnel Project for the Metropolitan Sanitary District of Greater Chicago consists of 25,189 lin. ft. of 32 ft. tunnel with 1 ft. 0 in. concrete liner throughout. The 17,100 ft. main tunnel connects at both ends with two other Deep Tunnel contracts. A branch tunnel connecting to the mainstream tunnel, commences approximately 4,140 ft. from the construction shaft and runs 8,089 ft. south and terminates at the Racine Avenue Pumping Station. Shaft work consists of one 28 ft. 6 in. diameter construction shaft, one 3 ft. 6 in. diameter access shaft and 25 drop shafts-four of which have vent shafts. The drop shafts vary in diameter from 4 ft. 0 in. to 17 ft. 0 in. All shafts vary in depth to invert from 225 ft. to 235 ft. Work has been underway for 16 months and as of this writing the construction shaft and access shaft are complete. Seven shafts have been excavated and concreting operations have started in four of them. The starter tunnel is complete, the mining machine has mined 250 lin. ft. and the muck removal system is in full operation. The contractor for the project is a joint venture of Paschen/Morrison-Knudsen/Kenny. Paschen Contractors, Inc., Chicago, Illinois is the sponsor. The total contract amount is $107,837,300.00.

Jan 1, 1979