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By F. F. Aplan

Estimations have been made of the tonnage of raw ores and coal treated by various beneficiation processes. The amount treated by gravity concentration is substantially greater than that treated by flotation, with coal representing the bulk of the material treated by gravity means. A discussion is given of the trends and anticipated improvements in gravity methods. Except for the concentration of iron ores, magnetic and elec- trostatic separation processes find their greatest use in the beneficiation of non-sulfide ores. Exciting new developments in magnetic materials and separation process (and, to a lesser extent, in the electrostatic process) indicate a substantial increase in the application of such processes to ores and a broad variety of non-ore uses.

Jan 1, 1989

By D. R. Nagaraj

The practice of nonsulfide minerals flotation can be best characterized as an evolutionary blend of scientific theory and art. Although extensive fundamental studies have been carried out in the past several decades, and significant theoretical aspects have been developed, plant practice has developed the art form, often ahead of the theory, which is somewhat in contrast to sulfide minerals flotation. Several excellent review articles have been published in the past highlighting both the theoretical aspects and plant practice. An overview of nonsulfide flotation is presented here. Major differences in sulfide and nonsulfide flotation are highlighted. An assessment of the state of nonsulfide flotation, recent advances made in flotation processing, and industry needs are also given.

Jan 1, 1999

By B. Denby, R. Hollands, S. Cooper

Augmented Reality (AR) is a technology that can be used to mix computer-generated images of data with views of the real world. The University of Nottingham is developing AR technology in a major research project which can be used to assist mining companies in visualising complex mine data whilst out in the field. The technology comprises of integrated GPS surveying technology and electronic orientation devices to provide location and viewing information, video capture via digital camera and portable computing hardware to overlay 3D models on views of the real world. The paper describes the concepts of the technology in some detail and concentrates on the possible applications of this in the surface mining and exploration industries.

Jan 1, 2002

By Kostas Fytas, Sebastien Perreault

This paper presents EOLAVAL, an integrated mine ventilation design software. In its new expanded version it is a user friendly and fully interactive computer package. The features of the new version are a graphics interface fully compatible with Autocad files, an integrated database of fan characteristics and the optional use of the thermodynamic approach in carrying out network simulations. This software can be used in underground mine ventilation planning: mine ventilation network analysis and simulation, economic design of airways, auxiliary ventilation design, calculation of a mine air heating plant, and evaluation of the natural ventilation and pressure losses in a mine shaft. This paper presents a case study illustrating the use of the computer package in mine ventilation design.

Jan 1, 2002

By John Jonkman

Many companies in the mining and minerals processing industry are finding themselves in the transition stage from the Industrial Age to the Information Age and are trying to become prepared for the Knowledge Age. To remain competitive in the decades ahead, they are faced with integrating the enterprise through the use of computers and networks. This paper discusses some common human and organizational barriers to enter- prise integration. It proposes how to develop the organizational and human' networks to match the capabilities of the technical networks which are being' implemented. Ten practical considerations are then presented for consideration in the executive suite in addressing the human side of implementing the computer integrated enterprise.

Jan 1, 1989

By Hans Skretting, Abraham A. Dor

This paper briefly reviews the historical development of the Rotary Kiln-Electric Furnace Process, which was first applied commercially to the treatment of qarnieritic ores in New Caledonia in the mid-fifties. It describes various technological improvements introduced, since, in industrial plants or developed in pilot facilities where a wide variety of nickel oxide ores has been treated success- fully over the years. Major categories of ores are characterized. The effect of their physical and chemical properties on their performance in the electric furnace is discussed. The relationships between ore composition, ferronickel grade and slag composition are analyzed and their effect on slag and metal tapping temperatures, nickel recovery, and electrode voltage are discussed.

Jan 1, 1979

By William Lyman

GENERAL DESCRIPTION Shrinkage or shrinkage stoping refers to any mining method in which broken ore is temporarily retained in the stope to provide a working platform and/or to offer temporary support to the stope walls during active mining. Since ore "swells" when broken, it is necessary to shrink the muck pile a corresponding amount by draw¬ing some of the broken ore out as the stope is advanced-hence the name. Broken ore retained during stoping is drawn out after the stope has reached its limits. The stope may be left empty or may be filled with waste contemporaneous with, or subsequent to, the final draw. Traditionally the method implies conventional overhand stoping methods with miners working between the muck pile and the stope back, in a space which advances updip with mining and is maintained by balanc¬ing "swell" with "shrink." The shrinkage classification is also applicable to so-called "semishrinkage" methods in open pillar-supported stopes where broken ore is temporarily retained as a working platform but offers no wall support; and to various blasthole shrinkage methods which utilize broken ore temporarily retained in the stope for wall support, but which do not require miners to work from muck pile in the stope. The method is generally applied to steeply dipping veins of strong ore between strong walls. APPLICATION Geometry The geometry of a shrinkable vein is described in terms of dip, width, and regularity along dip. Overall strike and dip dimensions and irregularities along the strike generally impose no restrictions on the method. Dip is ideally 1.2 to 1.5 rad (70 to 90°). As dip falls below 1.2 rad (70°), the shrinkage draw begins to strongly favor the hanging wall side, thus leaving a poor working platform for conventional overhand work. This is particularly true in relatively wide stopes. The sup¬port afforded to the hanging wall also diminishes with decreasing dip, reaching nil as the dip approaches the repose angle of broken ore. Dips below 0.78 to 0.87 rad (45 to 50°) are not generally shrinkable except by open stope "seinishrinkage" methods. Minimum mining width is fixed by working space requirements in the stope-generally about 1 m. Shrink¬age in narrower veins requires that waste rock from one or both walls be broken with the ore and the attendant dilution accepted to achieve the minimum width. Nar¬row stopes are less suitable, encouraging hang-ups and bridging of broken ore, with the attendant problems of erratic draw and incomplete recovery of broken ore. Maximum practical width may be 3 m or less to over 30 m, depending upon the competency of the ore and its ability to stand unsupported across the stope back. This is a vital safety consideration in conventional over¬hand stopes, but is much less of a factor in blasthole shrinkage methods. Very wide veins and massive ore bodies have been mined by transverse vertical shrinkage panels separated by transverse vertical pillars which are either abandoned or recovered later by other methods. Regularity along the dip is a prerequisite of shrink¬age as there must be no serious obstruction to the flow of broken ore downward through the stope to the sill level. Gentle rolls along the dip are acceptable if the local footwall dip everywhere exceeds 0.78 to 0.87 rad (45 to 50°). Off-dip hanging wall and/or footwall splits can generally be mined selectively from a conventional shrink stope as they are encountered without ad¬versely affecting subsequent continuation of shrinkage mining updip on the main vein. Vertical offsets or major rolls along the dip which cannot be "smoothed over" generally require that a sublevel be established with new draw control development. Blasthole shrinkage methods are much less flexible (and thus less selective) in their ability to accommodate any of these irregularities. Ground Conditions The wall rock must be strong enough to stand with the minimal support afforded by the dynamic mass of broken ore in the stope. During active mining, local sloughing from the walls is restrained, but the broken ore affords little, if any, useful resistance to closure of the stope walls. Such squeezing, if present, may bind up the stope and cause the loss of much ore. Pillars left between and/or within stopes are effective in preventing closure but reduce overall recovery. Walls may be re¬inforced by bolting after each stope cut in conventional shrinkage but not in blasthole shrinkage. Ore in place must be strong enough to stand with no natural support across the stope width, although tem¬porary artificial support or reinforcement may be used locally in conventional stopes. Some spalling or sloughing is permissible in blasthole shrinkage as men are never present in the stope. Physical and/or mineralogical characteristics of the broken ore may impose restrictions on stope design and/or operational plan¬ning, and may even preclude the use of shrinkage al¬together. Examples include: ores which, when broken, are cohesive or which tend to pack or cement together under the influence of ground water, wall pressure, and/ or chemical reaction. Such conditions precipitate er¬ratic draw during mining and often result in difficult and/or incomplete final draw; pyritic ores which oxidize very rapidly in the stopes and may generate heat, imposing a fire hazard by spontaneous combustion; sulfide ores which oxidize sufficiently in the stopes to adversely affect mill recovery by flotation; and ores (es¬pecially those containing uranium minerals) which ex¬ude radon gas and thereby impose ventilation constraints on stope design. In most cases these problems can be minimized by limiting the size of stopes, by minimizing the duration of mining activity in each stope, and by promptly drawing each stope empty following comple¬tion of mining.

Jan 1, 1982

By W. M. Ma, Daniel W. H. Su, K. Centofanti, Yi Luo, W. L. Zhong, Syd S. Peng

2.1 INTRODUCTION When total extraction of an opening of sufficient size is reached in a horizontal coal seam, the roof strata in the overburden deform continuously to reach a new equilibrium condition. The severity of deformation decreases upward toward the surface. As the downward saggings of the strata propagate and reach the surface, there will be a depression zone on the surface directly above, but extending beyond the edges of the underground opening. This is the surface subsidence basin or surface subsidence trough. The surface subsidence basin is circular in plan view, if the coal seam is horizontal and the mined-out opening is square in shape. But it is rectangular with rounded corners or elliptical if the coal seam is horizontal and the mined-out opening is a long- and thin rectangle or a short-rectangular, respectively (Fig. 2.1). Most underground openings (e.g., longwall panel) assume rectangular shape when total extraction has been completed. Theoretically the edges of the subsidence basin are the points of zero subsidence. But it is difficult to exactly locate the points of zero subsidence. Therefore in practice the points with vertical subsidence of 0.4 in. (10 mm) are used. The final subsidence basin is that which exists long after the mining has been completed, because its magnitude and shape are quite different from the dynamic subsidence basin formed while the face is moving. 2.2 CHARACTERISTICS AND TYPES OF DEFORMATION IN THE FINAL SUBSIDENCE BASIN For a horizontal coal seam, every point in the subsidence basin moves toward the center of the basin. Subsidence is maximum at the center of the basin. Any cross-section that passes through the point of maximum subsidence and either parallel to AB or CD line (Fig. 2.1) is a major cross-section along which principal directions of surface movements occur. However among those infinite numbers of major cross-sections, two specific ones are of special significance, not only because the magnitudes of surface movements are the largest, but also because they are the most easily identifiable directions, i.e., one that is parallel to the faceline at the center of the basin (CD in Fig. 2.1) and' the other is that perpendicular to the faceline but parallel to the diction of face advance (AB in Fig. 2.1). Nearly all the subsidence data obtained in the US have been derived from these two cross-sections, although some cross- sections parallel to CD but near the edges of the panel have also been included. In addition to moving horizontally toward the center of the basin, every point in the basin also subsides vertically. The magnitude of subsidence increases toward the center of the basin. Therefore surface subsidence is a three-dimensional problem and should be treated so in all cases. On all the major cross-sections, only principal subsidence and principal displacement occur. Since subsidence and displacement vary continuously in every major cross-section, three additional deformation components are de- rived, i.e., slope, curvature, and strain. On all other non-major cross-sections on the other hand the five components are accompanied by two additional components, i.e., twisting and shear strain. The seven components of the surface movement are defined as follows (Fig. 2.2): 1. Subsidence, S. On any cross-section, the vertical component of the surface movement vector is called surface subsidence. It generally points downward. But sometimes it points upward in areas ahead of the faceline or beyond the edges of the opening. In such cases it is a surface heave which is usually less than 6 in. 2. Displacement, U. On any cross-section, the horizontal component of the surface movement vector is called surface horizontal displacement. It generally points toward the center of the subsidence basin. But in steep terrain, it moves along the downdip direction 3. Slope, i. On any cross-section, the difference in surface subsidence between the two end points of a line section divided by the horizontal distance between the two points is called the surface slope of the section. 4. Curvature, K. On any cross-section, the difference in surface slope between two adjacent line sections divided by the average length of the two line sections is called the surface curvature of those two line sections. There are two types of curvature: con- vex or positive curvature and concave or negative curvature. 5. Horizontal strain, e. On any cross-section, the difference in horizontal displacement between any two points divided by the distance between the two points is called horizontal strain. If the distance between the two points is lengthening, it is tensile strain with positive sign. Conversely, if it is shortening, it is compressive strain with negative sign 6. Twisting, T. On the surface of the subsidence basin, the difference in slope between two parallel line sections divided by the distance between the two line sections is called twisting. 7. Shear strain, y. Shear strain is the changes in internal angles of a square on the surface of the subsidence basin or on any major cross-section. It is the summation of the differences in incremental (or decremental) lengths between the two opposite sides divided by the original distance between the two opposite sides. More precisely, the surface deformation indices (i.e., slope, strain, curvature, twisting and shear) are defined by derivatives of surface movement components. For simplicity, the x- and y-axes of the cartesian coordinate system are set to be parallel and perpendicular to the cross-section of interest, respectively. In such a coordinate system, slope and curvature along x direction are the first and the second derivatives of the vertical components (S) of surface movement with respect to x, respectively, or i, = ds/dx and kx = d2s/dx2. Horizontal strain along x direction is the first derivative of the component along x direction of the horizontal displacement,

Jan 1, 1992

By William Hustrulid

INTRODUCTION The Idarado mine lies high in the San Juan Moun¬tains on the divide between the Uncompahgre and San Miguel Rivers and consists of a consolidation of a number of old and prominent mining properties through which course some famous and very produc¬tive veins. Among the better known are the Smuggler, Tomboy, Montana-Argentine, Black Bear, Liberty Bell, Virginius, Flat, Barstow, and Japan, as well as many others. Most of these veins have been extensively mined over the past century and in the last three decades the Montana-Argentine and Black Bear veins have been the backbone of the Idarado mine; therefore, the descrip¬tion of the veins is limited to them. Access to the mine is through either the Treasury tunnel, whose portal is below Red Mountain Pass on US Highway 550 at an elevation of 3244 m (10,643 ft), or the Mill Level tunnel entrance 3.2 km (2 miles) east of Telluride, CO, at an elevation of 2761 m (9060 ft). The Treasury tunnel intersects the Black Bear vein 2643 m (8670 ft) from the portal and the Mill Level tunnel intersects the Argentine vein 2179 m (7150 ft) from the portal. Mining is by shrinkage stoping from slusher sublevels. The size of the scope blocks varies somewhat, but the standard size is 67 to 76 m (220 to 250 ft) long and 61 to 76 m (200 to 250 ft) high. The mine ranks either first or second in Colorado in yearly production of gold, silver, copper, lead, and zinc. It is 9.6 km (6 miles) from the Red Mountain plant to the Pandora plant, via interconnecting drifts and raises. There are engineering offices at both plants as a matter of convenience. The Red Mountain plant in¬cludes the company general offices, warehouse, carpen¬ter and machine shops, and mine change room. The Pandora plant consists of the mill and assay office, ma¬chine shops, and mine change room. The flotation mill has a capacity of 1632 t/d (1800 stpd), making a bul¬lion product and separate concentrates of lead, copper, and zinc. HISTORY OF THE MINE The Montana-Argentine vein was first extensively worked by the Tomboy Gold Mines Co., Ltd., a British concern. This company mined the stoped areas above the Ophir level between 1910 and the late 1920s and most of the stoped areas above the 2100 level between 1900 and the late 1920s. Gold was the principal ore metal mined. The area between the Revenue and Ophir levels was mined chiefly by the Revenue Mines Co. between 1900 and 1910. The ore was worked from the Revenue tun¬nel, which portals in Canyon Creek. Gold and silver were the chief metals recovered. The stopes between the 1700 and Revenue levels, as well as some higher stopes, were mined by Telluride Mines Inc. during the 1940s. The Mill Level tunnel was driven by that company from 1945 to 1948. Lead and zinc then became economically more important than the precious metals. In 1953, Idarado purchased Telluride Mines, which merged with the parent company in 1956. The Black Bear vein was first extensively worked by the Black Bear Mining Co. in the 1900s and by the Colorado Superior Mining Co. from about 1914 until snowslides at the mine camp [altitude 3750 m (12,300 ft)] terminated the company's operations in 1924. Leasers operated at intervals until 1934. The Treasury tunnel, formerly the Hammond tunnel, had been started before 1900 and reached the 1646-m (5400-ft) mark early in the 1900s, at which time activity lagged until the late 1930s. In the early 1940s, Idarado extended the Treasury tunnel from its heading at 1646 m (5400 ft) to the Black Bear vein and established a raise connec¬tion with the 600 level, the lowest level in the old mine. Since completion of initial work in the mid-1940s, systematic development of the mine, both in the driv¬ing of new headings and the utilization of older open¬ings, has resulted in the present extensive network of workings. GEOLOGY Introduction The oldest rocks of the area are the Precambrian metamorphic rocks, the Uncompahgre formation which are massive quartzites, some phyllites and slates, and dolomitic or limestone beds. West-dipping Paleozoic and Mesozoic strata lie on the Uncompahgre formation. These units are separated by a major angular uncomformity from overlying, essentially horizontal, Tertiary formations that include the basal Telluride conglomerate and several thousand feet of overlying volcanic rock. Intrusive rocks, mostly Tertiary in age, are common in the area, and are in the form of dikes, stocks, and sills. Dikes are very closely associated with some of the ore veins (Mayer). Description of Veins The Black Bear and Argentine veins range from 0.6 to 7.6 m (2 to 25 ft) in width, but in most places are 1.5 to 2 m (5 to 7 ft) wide. They vary in character from a well-defined tabular structure between sharp "frozen" walls, or gouge seams, to an irregular zone of quartz and quartz-sulfide stringers. Many gouge seams within the veins make the veins blocky and loose. Common gangue minerals in the veins include quartz, pyrite, rhodonite, chlorite, sericite, clay minerals, epidote, calcite, adularia, rhodocrosite, fluorite, and specularite. Quartz constitutes 60 to 70% of the veins and varies widely in character, ranging in color from clear through white, gray, and green, to amethyst. Chlorite, sericite, the clay minerals, and fine-grained quartz are common alteration products of the vein walls and of wall rock fragments in the veins. Epidote is an altera¬tion of the dike or, less commonly, of tuff-breccia. Calcite, adularia, fluorite, and rhodocrosite, which are wide¬

Jan 1, 1982

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 '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's back usually is arched, not square. For example, Fig. 1 shows that if sample intervals are measured from A' to D', 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'-B', B'-C', and C'-D', 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

By S. K. Kawatra, S. K. Valluri, S. J. Root

Common carbon dioxide capturing reagents, such as amines and sodium hydroxide, are expensive, toxic, and energy intensive to regenerate. We have added frothers to a sodium carbonate solution in order to increase the surface area available for CO2 transport within the packed bed of a counter current scrubbing column. Addition of frothers has improved the performance of sodium carbonate scrubbing solutions to be on par with sodium hydroxide scrubbing solutions. This novel enhancement of a relatively inexpensive CO2 capture-regeneration cycle shows potential to allow plants to turn a profit by sequestering a waste product. These results are discussed in detail in the following manuscript. INTRODUCTION The combustion of fossil fuels converts chemical energy to thermal energy. CO2 is produced as a byproduct of the reaction. 80% of the world’s electricity is produced by using this thermal energy to power a generator, accounting for 40% of global carbon dioxide emissions. In an effort to stabilize the rising atmospheric CO2 levels, researchers are investigating the application of carbon capture and storage (CCS) systems in power plants. [1] Efforts to capture CO2 at some power plants have been successful, but the cost of installing and operating the required equipment is high. As such, very few power plants have CCS systems. If CO2 can be captured and processed into a product that can be sold at a profit, installing CCS systems in power plants will be a very attractive business decision. In order for the sale of captured CO2 to become a profitable venture, the cost of capturing the CO2 from a flue gas must be reduced. BACKGROUND The most common method for separating CO2 from a flue gas is by passing it through a counter-current packed bed scrubbing column. The flue gas enters the bottom of a column while a solution of a chemical with properties advantageous to the absorption of CO2 is pumped to the top. The fluids interact in a packed bed, where the CO2 is transported from the gas to the liquid. CO2 is an acid gas, so it is common for the scrubbing solution to be alkaline in nature in order to neutralize the carbonic acid that is formed when it is dissolves into water. The exhaust exits the top of the column free of carbon dioxide, which exits the column in the scrubbing solution, reacting or complexing with the absorbent. Figure 1 is a schematic drawing of a CO2 scrubbing column. It shows the molar flowrates and compositions of the gas and liquid at the bottom [1], the top [2], and an arbitrary height in the column [Z]. [2,15] L is the molar flowrate of liquid at any point in in the column. G is the molar flowrate of gas at any point in the column. Since the CO2 is transported from the gas to the liquid, L and G are variable with respect to height. To simplify calculations, the CO2 free molar flowrates Ls and Gs are often used. X is the fraction moles of aqueous CO2 per mole of CO2 free liquid. Y is the fraction moles of gaseous CO2 per mole of CO2 free gas. Figure 2 is a composition diagram of the scrubbing column depicted in figure 1.

Jan 1, 2019

By G. Sanders

Discussion by G. Sanders The studies reported on in this paper were initiated to draw attention to the severe contamination problem in the Section 30 drilling program at Chimney Creek. The lithologic-subset sampling study reached a different conclusion from that presented in your paper, and I wish to comment on your subsequent analysis of the data and your conclusions. Request for more complete data In the section on subsampling, you mention that the subordinate lithologies were separated and sampled, yet only the dominantlithology gold value is plotted in Fig. 4. In a contamination study, the reader is interested in the assay values for the individual subsets. Please include a table of the subsample assay data in your reply. Also, please indicate which analytical methods were used to arrive at the gold values in the subsampling study. Turning barren rock into low-grade ore Figure 5 is very revealing and typical of all of the cross sections in Section 30. Note the long strings of low-grade mineralization spread out for hundreds of feet below the ore zones. There were some very high gold values found in certain contaminated fractions during the subset sampling. The conclusion, here, was that the distinctive, strongly-mineralized dolomite layer was probably loose and crumbly and continued to disintegrate during drilling. This caused salting of the unmineralized rock samples below. Missing the high-grade part of the ore body In your statistical analysis, you directly compare the reverse circulation assays to the diamond drill assays in Section 30. Two points argue against a direct comparison and suggest the differences are greater than the 3 % that you report. First, any core loss in a gold zone most likely means that the true gold values are greater. The drillers lost significant amounts of the clay-rich, Section-30 gold mineralization. Also, the initiated salt-mud system, an attempt to improve the core recovery, met with little success. Second, the practice of not sampling core geologically, but instead sampling on even 5-foot intervals, adds a deliberate dilution to the core assay values by including a portion of nonmineralized rock in the first and last samples of each high- grade intercept. The result is often a pair of low-grade assay values on either side of a high-grade gold zone. In reality, a high-grade gold zone has a very sharp assay wall that is often bounded by barren rock. This sampling method may make the diamond-drill core assays more like the reverse circulation values and may help explain the statistical similarities you found. However, it does not represent the true gold values in the high-grade parts of the deposit. You cannot deny that, by careful geological sampling of the drill core, higher and sharper assay values will be obtained. The low core recovery and the diamond-drill-core sampling method used act together to lower the diamond-core assay values. The 3% difference you found between the reverse-circulation and diamond-core assay values could be much larger when you consider what the true diamond-drill core values would be with optimum core recovery and a geologic sampling method for the core. Should statistics have been applied here? The statement "... that reverse circulation holes have overestimated the values of some ore zones and underestimated the values of others" (p. 345) is not correct. The subsampling confirmed what the cross sections hinted at in Section 30. Namely, beneath the high-grade zones, the reverse circulation holes created, by contamination, large intercepts of low-grade ore in regions of barren rock. Because the low-grade material was not there to begin with, this is not a process of overestimating low-grade mineralization. The next statement that "the average result is similar to that of the diamond drill holes" may apply to the data set numerically, but it is not true when viewed spatially on cross sections. Adjacent reverse circulation and diamond drill holes are almost impossible to correlate, high-grade zone values vary widely and many low-grade intercepts make no geologic sense. The subset sampling and cross sections presented in the first part of the paper show that the reverse circulation portion of the data set has some serious problems, as highlighted above, and should not have been dealt with statistically at all. Conclusion Each ore body is different, and each drilling method presents unique sampling problems. In this case, the diamond drill is the

Jan 1, 1994

By Franz Wilhelmstoetter, Christian Karner

"The Confederation Line is phase one of Ottawa’s light rail transit, which consists of 13 stops and stations. The project’s center piece is a 2.5 kilometer (1.6 miles) long tunnel with three underground stations leading underneath downtown Ottawa and is constructed within feet of adjacent buildings. The downtown stations, Lyon and Parliament, are caverns approximately 16.5 m (54 ft) by 18 m (59 ft) in a horseshoe configuration constructed in rock, while the third station, Rideau, is approximately 22 m (72 ft) by 18 m (59 ft) in an oval configuration and constructed in mixed ground. The paper will discuss design decisions made early on in the project such as excavation method, unique construction sequencing, and equipment selection, and go on to include corresponding progress rates. The paper also documents design and construction challenges of large caverns in very close proximity to existing buildings and varying ground conditions. Finally, an overview of achieved construction milestones and a current construction progress will be presented. Station cavern excavation is expected to be completed in early 2016. INTRODUCTION The number of urban metro construction projects has certainly increased over the last decades. Cities are growing rapidly all over the planet and many are striving to keep up with the increased demand for public transportation. One of the most feasible types of inner-city transportation with the capacity for large quantities of passengers are metro (subway) systems. Still, large construction projects in dense urban areas often affect the city’s residents in mostly negative ways. During construction, residents face the obstacles of road closures, noise, dust, economic loss for businesses, and property value losses are common. In other words, individuals are directly affected by the way new infrastructures are planned, designed, built and operated. The tunneling industry offers unique opportunities of keeping these side effects to a minimum and possibly avoiding them all together. Although challenged by a variety of constraints (some of which will be discussed in this paper), urban tunneling applications offer a chance to keep cities environmentally, economically, socially, and culturally sound during construction."

Jan 1, 2016

By Fulian He, Jianhang Chen, Shoubao Zhang

"Fully grouted rock bolts have been used in mining industry for many years. Significant research has been conducted to evaluate the load transfer behavior of fully grouted rock bolts with experimental programs. However, compared with that, less work has been conducted with analytical modelling. Therefore, in this paper, the authors used an analytical model to study the load transfer behavior of fully grouted rock bolts. To confirm the credibility of this analytical model, an in-situ pull-out test was used to validate this model. There was a close match between the experimental result and the analytical result. Following this, a parametric study was conducted with this analytical model. The influence of coefficients, Young’s modulus of the rock bolt and the diameter of the rock bolt on the load transfer performance of rock bolts were studied. Furthermore, the load distribution along the fully grouted rock bolt was analytically studied. The results show that the axial load in the rock bolt decayed from the loaded end to the free end independent of the pull-out load. However, the trend of the load distribution curve was influenced by the pull-out load. This paper was beneficial to better understand the load transfer mechanism of fully grouted rock bolts.INTRODUCTIONRock bolts have been used in the mining industry for a long time. They are used as the reinforcement tendons to keep the stability of openings in underground mining and the slope in open-pit mining. Furthermore, rock bolts are commonly used not only in the mining engineering, but also in civil engineering (Thompson, Villaescusa, and Windsor, 2012).The rock bolt is a tendon which can be a rod or a cable inserted into a borehole drilled in rock masses or soils. Traditionally, the material of rock bolts was steel. However, engineering practices showed that the steel rock bolts can be easily corroded by the moist environment in underground mines. Next, Fiber Reinforced Polymer (FRP) was used to manufacture rock bolts, which are termed as FRP rock bolts. Experimental results proved that the FRP rock bolts can effectively reinforce the surrounding rock masses without being corroded (Wang et al., 2018)."

Jan 1, 2019

By R. D. Williams

INTRODUCTION In a very real sense, Butte, Montana is where the copper came from that won two world wars. The price for that unrestricted mining and smelting of copper came due in 1983 when Butte was declared a Superfund site. One of the most complex remnants of this mining history is the Parrot Tailings site. The Parrot Smelter operated from 1881 to 1910 moving Upper Silver Bow Creek to the south and depositing approximately 370,000 cubic yards of waste material adjacent to Upper Silver Bow Creek. Additional mixed waste from other processing sites deposited below the Parrot Tailings at the Northside Tailings, Diggings East and Blacktail Creek Berm total approximately 224,000 cubic yards of contaminated material. As part of the Superfund remedy, EPA proposed leaving all these wastes in place. This proposal prompted a detailed criticism by several geoscientists familiar with the site. The EPA’s 2006 Record of Decision (ROD) did in fact leave these wastes in place which set the stage for a series of studies which continue to refine how best to protect extensive completed Superfund Remediation downstream of Upper Silver Bow Creek. HISTORY OF THE PARROTT MINE AND TAILINGS Placer gold was discovered in Silver Bow Creek, the headwaters of the Clark Fork River only a few miles from the Continental Divide in 1864. This discovery lead to a stampede of placer and lode claims. One of the first lode claims located was the “Parrot” lode, located in the fall of 1864. Progress for developing a mine at the Parrott was slow and it wasn’t until 1876 that one of the original locators encountered a rich vein of chalcocite. Butte had its first copper mine! The Parrott mine was patented in 1879 and by 1881 the Parrot, having lost a “t”, concentrator and smelter were built. The smelter and concentrator were an immediate success with a considerable market for processing the District’s rich copper ore. In 1885 the employees of the Parrot Smelter bought the Butte Reduction Works, another smelter. In 1885 dams were built on Silver Bow Creek to contain various tailings and related processing byproducts (Figure 1).

Jan 1, 2019

"Outotec’s TankCell® e500 flotation cell was commissioned during 2014 at the Kevitsa Mine in northern Finland. The TankCell® e500 is the first cell in this size class (500m3) to be run in an operational duty and it has been working as the first rougher in the copper flotation circuit. Treating around 850tph with an average grade of 0.3% Cu the TankCell® e500 has run with an average copper recovery of 71% and an average enrichment ratio of 50. In addition the cell has proved to be very energy efficient, operating with a specific energy of around 0.4kW/m3 under production conditions. This presentation will provide a review of the installation and operation of the TankCell® e500 at Kevitsa and will summarize the results found in the published literature from the various testing campaigns conducted around the cell. INTRODUCTION The TankCell e500 was installed and commissioned at the Kevitsa mine in northern Finland in 2014. Post startup the cell was initially run for testing and optimization with various test campaigns undertaken. but was subsequently taken into a production duty (Mattson et al., 2015a). At Kevitsa the cell is run as the first copper rougher replacing the previous two copper roughers (both 160m3) in this duty. The unit is running at between 38-40% solids at a feed rate of approximately 850tph. Since operational the average copper recovery from this single cell has been around 71% copper. This is typically with an enrichment ratio of around 50 being fed with an average head grade of 0.3% copper. More details on the circuit configuration and production performance of the e500 at Kevitsa can be found in Mattson et al., (2015a) and Murphy et al., (2016). In addition through the optimization program with the use of a variable speed drive controlling rotor speed it was found the the e500 was very energy efficient under production conditions. During this work the cell was found to run at a specific energy of around 0.4kW/m3 which exceeded design expectations (Xia et al., 2015)."

Jan 1, 2016

By J. Trimble, B. M. Davis, D. McClure

During initial mining operations at the Colosseum gold mine, discrepancies between the average blasthole grades and mill-head grades were quite large. In December 1987, mill-head grades averaged 39% lower than corre¬sponding estimates from the blastholes. The correlation between average blasthole grade and mill-head grade had an r-value of 0.015. The magnitude of the discrepancy and the lack of correspondence between estimated and actual grades led to a series of investigations of the sampling system, the effect of blast movement of the ore, and the estimation methodology. The discrepancy between blasthole grades and mill-head grades had been reduced by about half on a percentage basis at the conclusion of these studies. Ongoing studies have produced changes that reduced this discrepancy to its present level of less than 10%. Geology The Colosseum mine is within the Clark Mountain mining district about 68 km (42 miles) southwest of Las Vegas, NV in San Bernardino County, CA. Gold mineralization at the deposit is hosted in two 152 m x 213 m (500 ft x 700 ft) interconnected, rhyolitic intrusive breccia pipes that intruded Pre¬cambrian crystalline basement rocks and overlying Paleozoic marine sedimentary rocks. Subsequent erosion (deroofing) of all of the Paleozoic section and a portion of the Precambrian section has exposed the pipes to the present level (probably a significant depth in the original breccia pipe system). Lithologies exhibited within the breccia pipe complexes include rhyolitic felsite, and breccias of felsite and wall/ country rock. Breccia types grade from felsite plus Precambrian granite-gneiss fragments in an aphanitic felsite matrix to crosscutting breccias of felsite, Precambrian granite-gneiss, and Paleozoic sedimentary rocks in a matrix of finely comminuted material of the same composition. All the lithologies in and around the breccia pipes exhibit moderate to intense quartz-sericite-pyrite alteration, with sericite and pyrite the dominant visual alteration minerals. Gold at the Colosseum mine is generally submicroscopic and associated with sulfide mineralization, chiefly pyrite. It occurs as free gold, with minor alloyed silver. It is primarily in contact with pyrite in fractures or along pyrite grain edges. Secondarily, it appears as isolated particles in quartz and other gangue minerals but spatially always close to pyrite, and rarely as particles encased in euhedral pyrite. The pyrite mineralization and minor base metal sulfides occur in three distinct styles: as disseminations, as open space or vein/fracture filling, and in breccia clasts replacing dolomite. The pyrite megascopically ranges from absent to 30% of rock volume within the breccia pipes. Although the gold at Colosseum is apparently spatially and geochemically associated with pyrite, there is only a general proportional relationship between pyrite content and gold content. Sometimes, a high volume percentage of pyrite has only geochemically anomalous gold. This makes visual pyrite only a general grade control tool and not a specific tool. This further makes the ore control process at Colosseum all the more difficult. Initial grade control procedure Initial grade control was based on blasthole samples collected in two stages by an "automatic" sample collection device (rocket sampler). A sample was collected for the top 3 m (10 ft) in a hole. The bag was then changed so that a sample was collected for the lower 3 m (10 ft). Each sample was prepared and fire assayed to determine the grade. Large discrepancies were sometimes noted in the assays of top and bottom samples. The samples were averaged together and a plot was made showing the blasthole locations and their associated grades. Discrimination between ore and waste was based on as-

Jan 1, 1989

By Thomas A. Boyce

Introduction Although innovative coal extraction technology can be approached from many angles, this paper focuses on three: innovation in response to rising costs, to changing mining conditions, and to regulatory issues. Before exploring these three subjects, however, a basic question needs to be raised. Why Innovation? Coal is widely heralded, and quite correctly so, as the United States' "ace in the hole." It has been said that the U.S. is to coal what the Arabs are to oil. The implication is that we have nearly inexhaustible reserves - that coal is cheap, plentiful, and easy to extract. Indeed, because our current productive capacity is estimated to be 15-20% greater than demand, we have been given a false sense of security about the ease of future coal extraction. Therefore, innovative measures are being concentrated on developing technologies for the demand side of coal, including power generation, synthetic fuels, and industrial applications. What is often forgotten is that the unit cost of producing energy from these new technologies depends critically on the cost of the coal feedstock -in some cases the coal consumed accounts for more than one-third the total cost. Thus, while the developers of the end-use technologies struggle to improve the overall thermal efficiency of their process by one or two per¬centage points to make the energy produced competitive in the marketplace, they often are unaware of the risk of a 3 to 6% increase in coal costs wiping out the expensive, hard-earned results of their research. Energy planners are just beginning to realize that coal is not a uniform commodity available at low cost in infinite quantities. Coal is a heterogeneous material whose cost and quality vary widely with local conditions. To produce the massive amounts of projected coal-based energy at reasonable costs, we urgently need devel¬opments in extraction technology, as well as in end-use technology. Thus, coal extraction, coal conversion and utilization, and end -use energy consumption must be viewed as highly interdependent parts of an energy supply network (Fig. 1). And current emphasis on conversion and utilization must be balanced by more emphasis on upstream technologies, including coal extraction. Even more important, the impacts of innovation (or the lack thereof) in any one technology area in the energy supply network need to be discussed and understood by those working in each of the other areas. Why innovation in coal ex¬traction technology? Figuratively, we are all in the "coal boat" together - geologists, miners, transporters, processors, power producers, distributors, and consumers. Innovation is required from each of these sectors to overcome the obstacles that prevent greater use of our most abundant national energy resource. All other links in the coal energy supply network depend on large volumes of coal produced at reasonable costs. Declines in productivity, therefore, must be reversed; the more difficult mining conditions that will be encountered call for new methods and equipment, and regulatory issues must be resolved if the coal ex¬traction link is to be a strength rather than a weakness in the planned U.S. transition to an economy based more on coal and less on imported petroleum. Innovation Tailored to Regional Conditions Once the importance and urgency of innovative coal extraction technology has been grasped, the next step - implementing a research program - becomes quite

Jan 1, 1980

BP Minerals America (BPMA) has completed its $400 million modernization of Kennecott Utah Copper's Bingham Canyon mine. Modernized facilities include the in-pit crusher, conveyor, ore storage area, concentrator, ore pipe-line, and tailings pipeline. With modernization, Kennecott will achieve substantial cost reductions. Bingham Canyon is now one of the lowest cost copper producers in the world. And the mine will continue to provide jobs and income in Utah for another 30 years or more. The Modernization Project began in February 1986. The work took 33 months. In late 1987, the construction force peaked at 1100 workers. Kennecott Utah Copper now employs a permanent work force of 2100 employees. Less equipment and more efficiency The modernization work has increased the scale of the facilities that transport and process copper ore. A large, continuous transport system has replaced multiple haulage units. Fewer but larger pieces of equipment have replaced many crushing, grinding, and flotation units. The result is less maintenance, fewer employees, and improved metal recoveries. About one-third of the $400 million modernization project is directed at improving ore movement by eliminating rail transport. The new concentrator accounts for the remaining two-thirds of the modernization investment. The modernized plant was started up in stages. In February, ore was fed to the crusher. By the end of March, the three grinding lines were operating and the flotation circuit was producing copper concentrate. In July, the design through-put rate of 70 kt/d (77,000 stpd) was reached. October saw the achievement of targeted metal recoveries that corresponded to the fully modernized operation. It is estimated that Bingham Canyon will annually produce 181 kt (200,000 st) of refined copper, 9.3 t (300,000 oz) of gold, 71 t (2.3 million oz) of silver, and 5.4 kt (12 million lb) of molybdenite. Current Bingham Canyon production is about 97 kt/d (107,000 stpd) and a similar amount of overburden. Some 70 kt/d (77,000 stpd) of ore are processed through the in-pit crusher and 27 kt/d (30,000 stpd) of ore are transported by rail. The Bingham Canyon mine has yielded more than 11 Mt (12 million st) of refined copper. In 1863, precious metals - silver-bearing lead - were first discovered in Bingham Canyon. What was once a mountain is now a huge open pit. In fact, the mine is the world's largest man made excavation. The pit is some 4 km (2.5 miles) wide and 0.8 km (0.5 miles) deep. Since the pit began operating in 1906, more than 4.5 Gt (5 billion st) of material have been removed. Modernized facilities Electric 21 m3 (27 cu yd) or 23 m3 (30 cu yd) shovels are used to load the haulage trucks. Both 154 t (170 st) and 181 t (200 st) trucks are used to haul the ore to a mobile, in-pit crusher. The crusher replaces three crushers located at the old concentrators. There are two dumping locations at the new crusher, which can handle two trucks a minute. The 1.5 x 2.7 m (60 x 109 in.), 745 kW (1000 hp) Allis-Chalmers gyratory crusher reduces the ore to 254 mm (10 in.) size or less and delivers it to the conveyor. The crusher and conveyor systems are capable of handling 9 kt/h (10,000 stph) of ore.

Jan 11, 1988

By Melvin E. Poad

INTRODUCTION The placement of concrete by means of pumping is one of the fastest growing practices in the construction industry and is widely used in mining applications. In¬troduced in the United States in 1933, this method has been adapted to many aspects of the industry, especially where space for equipment is either limited or inaccessi¬ble. Concrete pumping offers many advantages over other methods of placement. Using pumping equipment, the concrete mixture can be transported in a single operation from a central mixing site to numerous forms in many remote locations, eliminating the need for rehandling the mixture and reducing the labor costs to a minimum. Pressurized pumping provides an extra mix¬ing and blending step, resulting in concrete of improved uniformity and greater density. In turn, this produces high compressive strength and increased durability. Various placement defects can be reduced or pre¬vented by pumping. The steady and even flow of the concrete minimizes the form distortion and delays caused by vibration problems that are common to fast dumping of the concrete mixes. Segregation, planes of weakness, bleeding, and surface faults are diminished considerably. Perhaps the greatest advantages of concrete pump¬ing are the distances and volumes that can be handled. Depending upon the equipment used, pumping rates of from 8 to 115 m'j/h (10 to 150 cu yd per hr) can be achieved, with a horizontal range of up to 460 m (1500 ft) and a vertical range of up to 140 m (450 ft). Theoretically, almost any concrete mixture could be pumped. However, consistency is an important consid¬eration for efficient operation. Mixes that are too wet, too dry, too hard, or poorly graded can cause difficulties that increase the costs. Similarly, mixes containing aggregate material that is over the maximum size or that is porous and highly absorptive can also result in diffi¬culties and increased costs. For the best efficiency, the mix formulation must have a well-graded sand and aggregate combination, with enough fines to provide flowability. If additional water is required for pumping, additional cement must be used to meet the design criteria of the structure. Each pump manufacturer can provide recommended mix formulations for the specific equipment furnished. Since the 1950s, many improvements have been made in pumping technology. Originally, only rigid steel pipe of 168 to 245 mm (6.625 to 9.625 in.) diam was used; today, both rigid steel pipe and flexible conduit are available in diameters of from 76 to 203 mm (3.0 to 8.0 in.). Improved pumps have been developed, and the introduction of power-coordinated booms has in¬creased the versatility of concrete pumping. The older and more conventional placement methods can no longer compete with the flexibility, efficiency, and econ¬omy of the pumping method, and continued research and development can only improve the system and expand its use throughout the mining industry. PUMPING EQUIPMENT The three major types of concrete pumping units currently available are the piston pumps, the pneumatic pumps, and the squeeze-type pumps. Table I provides a comparison of capacities, horizontal ranges, and vertical ranges for pumps of all three types. Piston Pumps The simplest of all concrete pumps are the piston¬type pumps; they are available from many manufac¬turers. As shown in Fig. 1, each unit consists of a

Jan 1, 1982