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By J. B. Knapp, J. W. G. Betts, Scott L. Burreill, J. H. Bassarear, E. A. Lowe, W. J. N. Davis, John L. Bolles, Gaylord P. Dopson, William H. Keener, N. S. L. Steenkamp

This chapter contains moderately detailed descriptions of some typical and well known mills that treat predominantly sulfide copper ores, some containing secondary values in gold, silver, and molybdenum. The descriptions are general in nature, because most of the specific data will be found in tabulations elsewhere in this section, covering a large number of mills in such matters as flotation reagents, crushing and grinding equipment and steel consumption, thickener and filter operating data, power consumption, and many others, in a manner that the reader will find convenient and comprehensive for reference and comparison. On the other hand, the plant descriptions selected for this chapter emphasize the kind of information that cannot be readily tabulated.

Jan 1, 1985

By D. A. Saint Don, M. Prueher, S. Barnwell

With the current changes in technology, multiple methods are now available for aggregate resource identification. From Sonic to Coring, Reverse Circulation to Percussive, there are many choices to make regarding drilling technology. The discussion will outline the methods above with focus on aggregate investigation and highlight the new technologies that will make a difference in your search.

Jan 1, 2004

By Sergey G. Pushkin, Eugeny L. Schastlivtsev, Anatoly A. Bykov

Three years ago the authors and a group of assistants began to develop a GIs-system for integration of comprehensive eco- logical information to be linked to geographical maps of Kuzbass region. And now we have got the working system and some first results that the present paper deals with.

Jan 1, 1996

By John G. Groppo, Bonnie Webb-Groppo

INTRODUCTION The first portion of this chapter is concerned primarily with the contamination aspects of fine coal cleaning that include both the solids and water discharged as waste products after coal preparation. Topics include characteristics of the water and solids, changes that occur to both of these phases during use and after impoundment, as well as the impact that these changes have on unit operations and the environment. Attention is also given to the formation of acid mine drainage and treatment options that may facilitate reuse of the water. The second part of the chapter analyzes the problem of air contaminants from coal preparation plants including analysis of air pollutants and control methods. An overview of air quality regulations is also presented; most provisions have little or no applicability to coal preparation plant operation but are presented to provide an understanding of end-user constraints. The final part of the chapter emphasizes the long-range problem of refuse disposal and control. An overview of characterization, maintenance, and control of the refuse area is presented, as well as design guidelines. A brief review of stability analysis procedures is also included. PREPARATION PLANT WATER Characteristics of Makeup Water for Coal Preparation There are few specific rules governing the quality of makeup water used in coal preparation processes.'" It is generally accepted, however, that the solid matter content should be less than 5%. The solids concentration in plant makeup water typically varies between 0.25 to 0.9 lb per gal (30 to 110 g/L) and may reach concentrations as high as 1.7 lb per gal (200 g/L). The water used in coal preparation is usually obtained from one or more of the following sources: public supplies, rivers and streams, canals, mine water, and wells. In most cases, water is obtained at the lowest cost including any necessary treatment. A major portion of the recirculated water is clarified by thickeners and settling ponds to remove fine coal and clay particles. Sufficient water is generally available to meet the needs of coal preparation in most areas. It has been estimated that approximately 65 to 70% of the makeup water used in a coal preparation plant comes from surface sources. It is important to note that a variety of substances are commonly found in natural water supplies, whatever the available source may be (Table 11-1). Excessive concentrations of some of those substances can have an affect on plant performance; the effect of specific contaminants on unit operations will be discussed

Jan 1, 1991

By John K. (Original by) Berry

The handling and disposal of mine water is a much larger problem than is apparent at first glance. Many more tons of water are removed from underground coal mines in the United States each year than tons of coal. In the old anthracite region of Pennsylvania the average was around 32.7 t (36 st) of water for each ton of coal produced, while bituminous mines average about 4.5 or 5.4 t (5 or 6 st). Water encountered in the mines may come from several different sources. In mines with light cover and without a firm solid roof, rain may seep directly into the workings. This is also true of deeper mines where pillar falls have broken the roof to the surface. The top under creek beds is generally faulty and broken and water comes through the roof at this point. Deeper mines are frequently overlain by water-bearing strata and pillar breaks and faulty top can admit water, while places running to the dip often accumulate water which comes up through the bottom. In all cases, except where passing under a year-round stream, the amount of water made is usually pretty closely connected with the weather-more water during the rainy season, less in the dry, although the deeper a mine is, the longer it is before changes in rainfall are felt. Another source of water, not connected with seasonal changes, is old water, accumulated over a period of time in abandoned workings of the mine, or from flooded adjacent mines. Here the water filters through the coal, the parting, or comes up through cracks in the bottom. Similarly, core holes, drilled to prospect the property, can be a source of unexpected water and modern safety requirements for dust-allaying water at the working faces can turn a pleasant working area into a soggy mess over the weekend if the system is not properly maintained and operated. Why water must be disposed of is obvious. If allowed to accumulate, sooner or later the mine will fill up, or at least it will become "roofed" in low places, blocking ventilation and traffic. It must be controlled long before it reaches this stage. Water standing in the working faces is unpleasant for the men who have to work there and also cuts down on the amount of work they can do, as they must watch where they step all the time. It is especially unpleasant in low coal.

Jan 1, 1981

By Stanley C. Suboleski, C. B. Manula

INTRODUCTION In any mining operation all possible steps should be taken to increase efficiency. One area for improvement is mine planning and design, particularly in the area of equipment selection for room-and-pillar systems. Be- cause of the availability of a wide variety of face machines, a fair degree of selectivity can be exercised in the choice of equipment for a particular job. However, this choice must be made on the basis of quantitative facts and forecasts related to the mining application. The purpose of this section is to develop and analyze the details of the mining process. Some specific areas studied include the relationship of system design to productivity, suboptimization as a result of equipment changes, and measurement of system performance. The plan of work leading to a quantitative description of these study areas is based on the growing interest in total system design using simulation as an analytical method (Manula, 1963). CHARACTERISTICS OF PRODUCTION OPERATIONS FROM ROOM-AND-PILLAR SECTIONS For a given mining method, raw production in a given section of a mine is primarily dependent upon the coal seam thickness, roof and floor conditions, methane emission, the mining methods, and the man-machine element. Average section production varies from 300 to 800 st per shift for conventional and continuous mining in high seams and from less than 100 to 300 st per shift in low seams. Since the reject varies from 0 to 40%, these figures must be decreased by the appropriate percentage to reflect the amount of clean coal mined. Personnel requirements per production section per shift for the various methods are listed in Table 1. Table 1. Production Personnel Method No. Method No. Conventional 12-1 5 Longwall 9-14 Continuous 9-1 2 Shortwall 9-12 MINING VARIABLES To evaluate the constraints and interrelationships for various mining methods, it is necessary to categorize the variables which underlie system production potential. Seven critical independent variables which determine production can be identified and categorized (Suboleski, 1978) : Seam Height The five categories are as follows: less than 36 in.;. 36 to 55 in.; 55 to 100 in.; 100 to 180 in.; and greater than 180 in. Floor Quality Floor quality ranges from : Excellent: Smooth, hard, grades less than 1 to 1 % % , and dry. Good: Smooth, soft but dry, with grades less than 3 % . The floor will deteriorate, but cautious operation can prevent it. There may possibly be heaving at some later time. Fair. Soft and damp. There is occasional interference with equipment operation; requires the use of four-wheel drive shuttle cars; ruts with regular use, and may have adverse grades of 5 to 7%. This may be coupled with slippery bottom and/or occasional steep rolls. Poor: Soft and wet. Requires blocking of the bottom to support equipment. There are frequent steep rolls and grades in excess of 7%. Roof Quality Roof quality ranges from : Excellent: Men are able to work under the unsupported top during the initial production cycle if legally permitted. Good: The roof is bolted on a 4 x 4 or 5 x 5 pattern with short bolts (442 in.) or <seam height if the seam >42 in., or requires posting with no bolts on a 4 x 4 or 5 x 5 pattern. There are no falls. Average: The roof is normally bolted on a 4 x 4 or 5 x 5 pattern, but with long bolts (>seam height or >6 ft.). There are infrequent minor falls or there may be an excellent roof which is difficult to drill. Fair: This type often requires spot bolting in addition to the regular pattern or bolting with planks. The roof conditions require shorter than planned cuts, or narrow cuts. Poor: This type requires bolts plus crossbars and posts, or installation of yielding supports or truss-type support. It is almost certain to fall if this is not done. Methane Liberation This ranges from none detected to low (no buildup at the face, even with minimum ventilation requirements) to moderate (the curtains must be extremely tight and tubing close to the face or methane will build up to 1 % during the loading of the car) to high (methane will build up to 170 if the miner is operated at the normal rate, even with proper ventilation). Hardness of Coal Coal hardness falls into the following categories: Soft: Soft coal is easily cut by a continuous miner. A plow could be used by longwall. Average: Coal of average hardness could be easily cut by a miner, and a shearer would be used in the longwall. Moderate: Moderately hard coal causes difficult cut-

Jan 1, 1982

By D. E. Ranta, G. H. Giroux, A. W. Randall, D. A. Barr

Zinc-lead ores of the Pine Point District, Northwest Territories, Canada, are hosted in paleokarst cavities within a Middle Devonian carbonate barrier complex. Of the nearly 100 deposits found, most are irregularly distributed along two main regional stratigraphically controlled trends. The ores, which occur predominantly as open space filling, form deposits most of which average one million tonnes in size with an average grade of 10% combined zinc plus lead. Geological features related to karsting control not only distribution of deposits but also their shape, size and grade. The higher grade ore in these deposits generally occurs in a relatively small core area. Consequently, accuracy of reserve estimation is dependent on drill-hole density and understanding of karst geometry. Several case examples will be reviewed to demonstrate the relationship between shape, size and grade of deposit, geological reserves and drill-hole density, along with a discussion of geological ore reserve estimation methods.

Jan 1, 1992

By J. H. Mackey

"Where can I obtain a reclamation bond today?" "Who is writing your bonds?" These questions are common and heard frequently. The fact that the Society of Mining Engineers has determined that two of the five environmental topics reviewed at the 1989 Annual Meeting relate to reclamation bonding indicates to me that those who are involved in the coal industry, no matter what capacity, remain very cognizant of the reclamation bonding dilemma.

Jan 1, 1989

By D. G. Foot, J. D. McKay, R. A. Ynchausti

The increasing popularity of column flotation as an efficient and profitable mineral beneficiation technique has resulted in the need to perform a study of basic column operating parameters in order to understand mineral behavior in a column. Variation of column rougher flotation parameters on the Nevada Fish Creek fluorite and Utah Collins manganese ores enabled characterization of four beneficiation zones within a column flotation cell. Although specific mineral behavior in each zone was peculiar to the mineral system involved, particle retention time predominantly affected both manganese and fluorite recoveries in the collection zones of the column while enhancing mineral grades in the cleaning zones. Column parameters such as column length, feed injection location, and tailings recirculation velocity were inherently linked to particle retention time and retention time distribution, both of which affected mineral recovery. Froth depth had a discernible effect on fluorite concentrate grades; as froth depth increased, fluorite grades increased. Wash water additions affected both fluorite and manganese recoveries and fluorite grade. Axial mixing drastically reduced fluorite and manganese recoveries due to short circuiting of some of the feed slurry in the column.

Jan 1, 1988

By L. G. Stolarczyk

In-mine Continuous Wave (CH) Radio Imaging method (RIM) instrumentation has the capability of measuring EM wave propagation constants in layered formations. The RIM method is based upon the fact that seams of coal, trona, potash, quartzite, gilsonite and oil bearing sands form natural waveguides for low loss transmission of electromagnetic (EM) waves. The waveguide forms whenever the seam is bounded by more conducting claystone, mudstone, shale or in the case of oil/gas traps, by a floor of saline water. Disturbances caused by faults, shear zones, paleochannels, sills and dikes significantly alter the propagation constant. The analytical values of the EM wave propagation constants depend on the physical and electrical parameters of the seam. The values are presented in this paper in a graphical form to assist in the interpretation of RIM tomographic images. This paper compares in-mine geologic mapping obtained during mining to tomographic images of the geologic disturbance zones in a Central Utah coal mine.

Jan 1, 1989

By Edmond Donovan, O&apos

With the increase in complexity of conveyor systems and the inexorable demand for more capacity and availability, the ability of designers to accurately predict performance is becoming increasingly important. Dynamic analysis techniques for modelling wave mechanics in conveyors are finding increased applications. Unfortunately, the reliability of these modelling techniques is dependent on a thorough understanding of the performance characteristics of the components that make up the system. These characteristics include such things as response rate, hysteresis and efficiency. Some parameters such as idler rolling resistance can be easily quantified by measurement. More complicated drive parameters are difficult or impossible to measure, particularly in large horse-power equipment. One method of quantifying or at least bracketing the performance envelope of complicated pieces of equipment is to construct a mathematical model. This can be done quickly and economically, and the influence of various parameters can be determined. This paper sets out to demonstrate by example the usefulness of this technique in gaining an understanding of complicated systems. The model is developed and the influence of a number of variables are examined.

Jan 1, 1996

By V. E. Gall, A. Vonthron, J. Ninic, A. Marwan, M. König, G. Meschke, A. Alsahly

"In current tunneling practice, finite element simulations have become an integral element of the planning and design process. These models are most often manually generated using 2D CAD drawings, which is a laborious and time consuming process. In this paper, we propose a BIM based approach, in which the 3D FE-model is automatically generated through a set of compatible geometries of individual components, i.e. the geology, alignment, lining and the TBM. The BIM model also includes all relevant model parameters of the tunneling project that can then be incorporated into subsequent analysis to be performed during the tunnel drive. The setup and the execution of the FE-analysis are performed automatically utilizing all required data from BIM. The applicability and efficiency of the proposed strategies is demonstrated by means of project data from the Wehrhahn-line metro project in Düsseldorf, Germany. Both the design as well as the construction stage is considered. INTRODUCTIONIt is recommend to employ Finite Element (FE) simulations in order to ensure a high degree of safety and quality for any given project (Guglielmetti et al. 2012). As such, FE models form an integral part in the design and planning stages of modern tunneling projects. These models are often time consuming to construct and require data from many different sources. This data is most often not centrally stored and therefore not easily accessible. Additionally, the formats in which project data are stored are typically not compatible with the formats required by an FE program. This is especially true in the case of geometrical data, such as CAD drawings. Although existing design drawings are most often used as the basis for a FE model, a direct import into a FE program is most often not successful as the imported data generates a geometry that does not fulfill the requirements of a FEM model, such as model connectivity. Even if a geometry is successfully imported, other necessary aspects of an FE model, such as boundary conditions and material properties, must be applied manually. These incompatibilities inevitably result in the creation of a new model, which, especially in the case of complex 3D models, is a time consuming process."

Jan 1, 2016

By Yi Luo

Bolting the mine roof is one of the noisiest underground mining operations. The main contribution to the high noise is the drilling of the roof bolt holes, especially in hard rock. A previous research shows that the specific energy consumed in drilling the bolt holes is inversely proportional to the bite depth (i.e., penetration per drill revolution). Less specific energy at a reasonably high bite depth means less energy is wasted in the drilling process to produce heat, bit wear and noise. Therefore, control of drilling operation to achieve a higher bite depth can provide an opportunity to reduce drilling noise from its sources as a proactive noise control approach. In order to explore the feasibility and to develop the proactive technology for reducing drilling noise, drilling tests have been conducted in laboratory settings. The preliminary results show significant reduction in noise dose can be achieved by drilling in medium hard rock at a reasonably high bite depth. This paper presents the findings from the laboratory experimental studies conducted so far.

Jan 1, 2013

By E. M. Frizzell

Decisions affecting ground control design require detailed knowledge of roof rock geology. This is especially true when weak roof materials and anomalies create hazardous conditions and. if not properly supported, result in roof falls. The U.S. Bureau of Mines has developed a system to evaluate geological conditions by using a roof drill that will provide an estimate of the compressive strength of the roof rock in real time. Through the use of sensors and a microcomputer on the roof drill critical drilling parameters can be interpreted and analyzed almost instantaneously, making it possible to inform an operator of hazardous roof conditions such as voids and changes in strata during the drilling process. Geotechnical data can be downloaded to a portable semiconductor memory device, which is brought to the surface so that data can be directly accessed with a personal computer for further analysis. Results from field tests are presented.

Jan 1, 1992

By Mark Board

INTRODUCTION As the world&apos;s supply of near-surface metalliferous ore bodies is depleted, more emphasis must be placed upon mechanized bulk systems of underground mining. In particular, block caving is becoming increasingly important as a method of mining low to moderate grade deposits at depths of somewhat less than 1219 m (4000 ft). 00 ft). The block caving method requires that the ore body be completely undercut, thereby removing its means of support and inducing failure which progresses from the undercut to the ground surface. As the span of the under- cut increases and the weight of the overlying crushed ore body becomes greater, large stresses are induced on slusher and haulage drifts. Stability of these drifts is critical in block caving operations and often proves to be an extremely difficult task (Kendorski, 1976; McWilliams, 1975). Slusher drift maintenance can be costly, often exceeding that of the original support costs. This chapter analyzes the ground support problems of an Arizona block caving mine. The mine presently uses monolithic concrete liners as a support method. Failure of the liners occurs as the undercut approaches, thus initiating a maintenance problem which will continue until the ore body is depleted in that block or panel. To examine this problem, a simple analysis of the stresses induced by the advancing undercut was per- formed. Utilizing a design method developed by Dixon ( 1973, 1974), the calculated stresses were used to deter- mine the minimum thickness required for varying section shapes of unreinforced concrete liners. It is found that the thicknesses required for stable linings are not economic. Recently developed empirical methods of support de- sign are utilized (Barton, 1974; Bieniawski, 1973) and compared to the analytical method. A final discussion of alternative methods of support is given with recommendations for the use of passive or "yielding" support methods. LAKESHORE MINE (GENERAL GEOLOGY) The Lakeshore copper property is located in Pinal County, AZ, approximately 45 km (28 miles) south of Casa Grande. The Lakeshore deposit is a porphyry cop- per similar to others in Arizona. The ore body consists of disseminated chalcopyrite and fracture fillings in quartz monzonite porphyry (QMP), Cretaceous volcanic (KVS), Precambrian diabase, and Mescal limestone. The ore body has been divided into three distinct zones of mineralization which have been termed (South, 1972) the oxide zone, the sulfide zone, and the tactite zone. The oxide zone occurs in the upper portions of the ore body and consists primarily of leached and oxidized quartz monzonite porphyry and Cretaceous sedimentary and volcanic rocks. The sulfide zone underlies the oxide zone and includes all quartz monzonite porphyry and Cretaceous rocks which have not been oxidized. The thickness of the sulfide zone is variable and ranges from about 61 to 122 m (200 to 400 ft). The tactite zone occurs in a series of Paleozoic metasediments consisting of tactite, quartzite, diabase, hornfels, and dikes of quartz monzonite porphyry. The tactite zone is highly variable in thickness and underlies the Cretaceous volcanic and sediments throughout the mine. Together, the tactite and sulfide zones are known as the thick or tall sulfides. The greatest support stability problems occur in mining of the tactite zone and are therefore the subject of the remainder of this chapter. MINING METHOD The Lakeshore mine is a block caving mine utilizing a panel caving method. The present mine consists of two

Jan 1, 1982

By Anshumali Gangwar

This paper outlines the chance constrained binary integer program developed by the author in his doctoral dissertation at Columbia University, New York, 1973. The optimum planning of a mining program involves the determination of a production schedule which maximizes the net present value of the mining property. The actual mining process has an inherent stochastic and discrete nature. In particular, the demand for the mined product and the grade of the mineralized material in the ore are the two most important stochastic mining variables. The decision to mine or not to mine any block of a mining property is a binary integer variable. This paper outlines the formulation of pit production scheduling problem as a chance constrained binary integer program. The deterministic equivalents of these chance constraints are nonlinear in binary variables which have subsequently been linearized. The planned objective is to meet with a stipulated high degree of reliability, all the random demands as they materialize with a random supply of the commodity being marketed as determined by the eventuating random grade values of the exposed ore blocks and the mining system&apos;s production constraints such that the discounted total production costs over the entire planning horizon are minimized or the profits maximized. The emerging surface of the mine in each scheduling period is determined by the values taken by the binary decision variable associated with each block of material in each scheduling period.

Jan 1, 1982

By Malcolm Reeves

The heavy oil reservoirs of Alberta and Saskatchewan include sequences of unconeolidated silts and fine sands which respond in an unusual manner when subjected to the high temperature and pore-fluid-pressure associated with steam injection and other thermal enhanced recovery techniques. Apparent transient connections are suddenly established between producers and injectore and there is an associated enhancement in oil (and sand) production. These phenomena have been attributed to local fluidization and piping processes which are known to occur in situations where high pore-fluid-pressures prevent the mobilization of intergranular friction. These breakdown processes may be associated with the penetration of tongues of the heated reservoir fluids with suspended sand into the creep sensitive intact reservoir since the viscosity ratios are sufficiently high for viscous fingering instability of the fluid-fluid interface. A program of laboratory tests has been carried out to measure the strength characteristic of saturated oil sands at a variety of temperatures and fluid pressures. These tests suggest, that in common with the coarser Athabasca sands, the Waseca sands have very little cohesive strength and show the characteristic dilational failure response of dense sands. Some special tests, devised to determine the relative ease with which samples can be fluidized, confirm the expected susceptibility of the fine- sand silts. Both piping and fluidization can be readily induced at in-situ reservoir temperatures and pressures. Analysis of stresses around wellbores suggests that fluid-pressure gradients may be sufficient to initiate piping and that backward erosion along streamlines may link producers and injectors. Such pipes may act as well “adits” gathering oil from the heated reservoir and contributing to enhanced production rates. Mechanical breakdown of the reservoir by drilling, injection and production processes provides a consistent possible explanation for the sand-cut response of production wells.

Jan 1, 1986

By David Morrison

On the Corramandel Peninsula of New Zealand, the Waihi Gold Mine has undergone a major redevelopment. The upgrade included the construction of two very large overland belt conveyors, both rated for a capacity of 6000tph. The first of these conveyors (CVR2) traverses a total of 1600m over undulating terrain, including a significant section running through tunnels. A large hill is climbed, just after the mid-section of the conveyor run, before dropping back down prior the discharge point. The second of these conveyors (CVR8) traverses 1100m along an alignment with a horizontal curve, a number of vertical curves and a significant lift before the discharge point. The CVR8 conveyor also included a shuttling head pulley arrangement. The mine they service is inside the town of Waihi. The conveyor alignment runs through the town and out to the metallurgical process site on the outskirts of the municipal area. This paper documents the engineering process that took place in the design of these conveyors. The Martha Hill mine in Waihi, New Zealand is unique. The mine is placed in the township. Consequently there are quite stringent environmental limitations on the mine&apos;s operation. In [Figure 1] the proximity of the mine to the town is shown. In the top-centre of the photograph the alignment of the existing smaller overland conveyor can be seen leading from the mine to a tunnel through the hill in the background.

Jan 1, 2000

By N. L. Wesis

General History The first jaw crusher in the US was patented about 1830, but the Blake crusher that has maintained a substantial advantage over other types was invented in 1858 by Eli Whitney Blake. For going on 150 years the jaw crusher has been an invaluable machine; even today, when gyratory crushers have assumed most of the burden of crushing large tonnages of large rock and ore, the jaw crusher has a firm place in the mining industry. It is interesting that soon after the invention of the Philetius Gates gyratory crusher in 1881, a contest between the first No. 2 Gates and Blake jaw crushers of equal gape showed the gyratory to be 3.2 times as fast. This foretold a day not far in the future when the jaw crusher would not be adequate on the largest jobs. In his Textbook of Ore Dressing (1909) R. H. Richards recom¬mended that the word breaker be used for all machines breaking to relatively large sizes (say to 5 in. or greater), and crusher be used for finer work. Even in the 1968 US Bureau of Mines dictionary of mining terms the work breaker is defined by Richards in this context. However, in modern US usage it is nearly limited to coal breaking. Over the years the jaw crusher has been developed in a variety of forms but it appears today in these three general forms: (I) Blake type, (2) Dodge type, and (3) single toggle (or overhead eccentric) type. All of these crushers have a fixed jaw and a moving jaw, between which the coarse rock fragments are intermittently caught and crushed. The three types are differentiated by the manner in which the movable jaw is moved in relation to the fixed. Types The Blake-type crusher is shown in cross section in Fig. 1. The simplicity and strength that have made it first among jaw crushers are evident. The movable jaw is suspended from a cross-shaft at its upper end, this shaft being carried in bearings on the sides of the frame. The actuating mechanism consists of an eccentric shaft, also supported in frame-mounted bearings, which imparts through a pit¬ man and a pair of toggles a reciprocating motion to the bottom of the swing jaw, the return movement being effected by several spring¬loaded rods. The whole is built into a box frame with the crushing chamber at one end. As in all jaw crushers, the eccentric shaft is equipped with a heavy flywheel that maintains an even speed through¬out each stroke. The Dodge-type crusher illustrated in Fig. 2 is simple, mechani¬cally. Its movable jaw, being pivoted below the discharge opening, has minimum movement at crusher discharge and maximum at crusher feed. Because the choke point coincides with the point of least motion, these crushers are of relatively low capacity, and the rapid action in the feed area gives the machine the advantages of large reduction ratio and closely sized product. Thus, it fits well into low-capacity operations like sample reduction. Besides low capacity, its principal disadvantages are tendency to pack and make fines. The single-toggle or overhead-eccentric type of jaw crusher (see Fig. 3) has gained in usage, starting as a modification of the Dodge idea in that the greatest movement of the movable jaw occurs at the top, and being gradually improved and strengthened to its position today where it covers just as wide a field of application as the Blake type from the standpoint of feed opening. The motion of the movable jaw is the result of the circular motion of the eccentric shaft at the top of the swing jaw combined with the rocking action imparted to the bottom by the inclined toggle plate. Because of light weight and mechanical simplicity, it is economical in small mills and portable plants where the rock is not as tough and abrasive as to punish the machine with excessive impact from shocks, or excessive wear result¬ing from the pronounced vertical component of motion.

Jan 1, 1985

By C. Alan Tapp

Introduction Mine sampling determines the practicality of any mining operation. Improper sampling can result in an incorrect appraisal of present production and future potential. Therefore, the mine department in charge of ore-reserve calculations and mine sampling should be overseen by competent and experienced professionals with technical backgrounds qualifying them to produce accurate results. Sampling is a process by which portions of an ore body are collected and analyzed to estimate the average mineral content of the entire ore body. It is incorrect to assume that a large number of samples eliminates any errors in the sampling method. To obtain unbiased samples, proper sample location with respect to rock type and mineralization is just as important. The sampling procedure must yield correct results for the type of mineral deposit, and careful consideration should be given to whether or not the sampling technique has been developed to an extent sufficient to eliminate as much human error and bias as possible. Only after the ore has been mined and milled is the sampling accuracy known. Sampling also provides information about the bulk composition of the ore for mineralogical and metallurgical tests that determine the economic ore-waste boundaries and the geologic trends for exploration. Actual mining plans can be developed from this information to maximize profits. Accurate sampling is critical, and thus must be approached in a scientific manner. Sampling Techniques Sampling practices and techniques are as varied as the mines in which they are used. The method(s) chosen must be tailored to suit the company and mining needs. For instance, tabular uranium deposits, vein gold deposits, and porphyry copper deposits pose special problems in conducting unbiased sampling. The mine geologist or engineer in charge must develop a sampling method, test it in a sample area, and then critically evaluate the results. If the results from the test area are accurate within the economic limits established by the company, they then may be adopted for general use in the mine. Four routine sampling methods are suitable for specific sampling objectives in the daily mine routine: (1) channel sampling, (2) chip sampling, (3) grab sampling, and (4) bulk sampling. The final sampling results depend upon how the four methods are combined to accurately determine the grade of the ore body. When used in conjunction with each other during different stages of mine development, the channel, chip, grab, and bulk sampling methods provide an in-house check or a comparison by which mining methods and sampling procedures can be evaluated. However, the most valid check is based upon the daily mill production. Channel Sampling Channel samples consist of cuttings collected from a groove cut into the rock about 102 mm (4 in.) wide and 19 mm (0.75 in.) deep. Various tools ranging from a 1.8-kg (4-lb) hammer and moil to a pneumatic chisel can be used to cut the sample. Accessibility and rock hardness determine the applicable sampling tools. Before attempting to take a sample, the rock surface must be cleaned thoroughly; the method of cleaning depends upon the amount of mine dust accumulated on the surface or the degree of alteration of the rock surface. Typical cleaning methods &apos;employ a wire brush, water, or chipping a fresh surface. Next, the sample outline is marked on the prepared surface, taking care to choose appropriate sample loca- tions. After determining the proper sample outline and length, the sample can be chiseled out, catching the rock fragments on a canvas tarpaulin on the floor, in a powder box, in a canvas bag to avoid contamination, or by some other suitable means to capture all of the sample. Vertical veins present a special problem because the drift&apos;s back usually is arched, not square. For example, Fig. 1 shows that if sample intervals are measured from A&apos; to D&apos;, the length of the sample interval is greater than the true vein width, resulting in an incorrect calculation of the ore reserves. The correct method would collect three samples at A&apos;-B&apos;, B&apos;-C&apos;, and C&apos;-D&apos;, using either actual measurements in the mine represented by A-B, B-C, and C-D or trigonometric calculations to determine the true ore thickness. Then, each sample can be weighted by grade and true thickness for the vein. Channel lengths usually are a maximum of 1.5 m (5 ft), and it is good practice to divide longer samples into smaller intervals according to structures, changes in rock types, or differences in rock hardnesses. The influence of those features on the mineralization then can be determined. Chip Sampling Chip sampling is a variation of channel sampling used when the rock is too hard to channel sample economically or when little variation in the mineral content indicates that this sampling method will yield results similar to those of channel sampling. Rather than cutting a channel in the rock, small chips are flaked off at regular intervals over the entire face or area being the mine. sampled. Care must be taken to assure that the sample is representative of variations in the rock hardness and type. hi^ method is fast and useful in preliminary evaluations, but it should not be used for quantitative ore-reserve calculations. Grab Sampling Grab sampling is a fast method for double checking either channel or chip-sampling procedures, and, in some instances, mine production can be estimated from carefully taken grab samples. Grab sampling takes equal amounts of material at selected intervals over a mine dump, a muck pile, or from an ore car to estimate its Channel Sampling mineral content. Generally, this method is not considered reliable. Many independent variables can affect this type of sampling process. Thus, if the ore occurs in the softer fraction and a proportional amount of the result-

Jan 1, 1982