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By R. S. Mateer

Linear programming on desktop computers is commonly used to optimize coal blending proportions. The objective function is usually the cost of the blend. The constraints consist of materials and heat balances as expressed as linear functions of the blend. This is illustrated in Example 1. Inclusion of a sulfur compliance constraint in more complex because the maximum permitted sulfur percentage is not a pre-specified constant. Rather, it is a function of the calorific power of the blend itself. The current limit of 1.2 pounds of sulfur dioxide per million Btu may be expressed as: [Sulfur Max%= 0.6 . Hl10,OOO with H in Btu/lb.] A proper constraint equation may be formed by defining for each coal a function that is the difference between the actual sulfur percentage and the maximum sulfur percentage. This may be regarded us an excess sulfur percentage. For Coal A of Example 1. this would be: [ Sulfur Excess%=0.5-0.6 13,000/10,000=-0.28] This value would be used as the coefficient of the weight proportions of this coal in the blend. The complete constraint equation would then be the sum of such terms for each coal set less than or equal to zero. This is illustrated in Example II. [ ]

Jan 1, 1995

By B. H. Kaye

Particle-size Distribution Specifications for Particulates. Feed to and finished products from size-reduction operations are defined in terms of the particulate sizes involved. It is also well to know whether the ultimate individual particle is being measured, or, if any aggregation or agglomeration of particles exists, whether this has been created by the size-reduction operation. The most complete description of a powder is given by its particle¬size distribution. This can be plotted in terms of cumulative percent oversize or undersize in relation to the diameters of particles, or it can be plotted as a distribution of the amounts present in each unit of diameter against the several diameters. It is common to employ a weight basis for percentage, but there are some data in the literature where frequency, or number of particles, is used. The basis of percent¬age, whether weight, frequency, or some less commonly used factor, should be specified, as also should be stated the diameter, its units, and preferably whether determined by sieve, settling velocity, or other¬wise. Fig. I presents two sets of distributions-ore cumulative and the other in unit intervals. The slopes of the 5-µ intervals of the cumulative curves are converted to percent per micron and plotted as a block, or histogram, from which smooth curves are derived. Powder A has a narrower, or tighter, size range for the bulk of its weight than powder B. Both materials have the same weights below and above the size marked by the arrow. Simpler treatments of distribution are possible. In some cases the significant value is the top size. Since the 100% point is dubious, some amount, as 95 or 98%, so specified, can be called top size. In other cases the percent on some sieve helps define coarseness. Merely stating that all, or all but a small percentage, passes a given sieve is inadequate for defining the true fineness or a material. Complete parti¬cle-size analysis to show the distribution is essential for most compari¬sons and calculations. Particle-Size Equations. A number of equations have been pro¬posed to correlate the quantity of a particulate material with its parti¬cle size to obtain a distribution relationship. 's 19 In the literature it is often assumed that a powder must follow some distribution, such as the Rosin-Rammler-Bennett:20, 21

Jan 1, 1985

By C. Widham

Introduction It is generally accepted as dogma that sampling contributes most of the error in gold fire assays. Differences in assay results on pulps from the same sample interval are frequently regarded as evidence of the presence of the so called "nugget effect" of relatively coarse gold particles. It is true that coarse gold particles can contribute to substantial sampling fluctuations. But, while the process of sampling is probably the major source of error, the analytical process cannot be completely ignored as a possible contributor to erratic assay results. To maintain a stable assay process, the analytic part of the system must also be kept in control. One method of monitoring the performance of the analytic system is to systematically assay standard materials, whose sampling characteristics are carefully controlled. Gold assay standards are not prepared, nor can they be prepared, to account for both sampling and analytical errors. It is not possible to send coarse material to a lab for both preparation (i.e., comminution and splitting) and fire assaying and then come to conclusions only about the fire-assay process. Because most gold ores are very heterogeneous, sampling errors would, in most cases, completely mask the contribution of the analytical errors. Assay-standard material is prepared only to assess the accuracy and variability in the fire assay process. Because the objective of the assay standard is to provide information about the fire assaying, it is necessary to control the sampling error of the standard material, so that it is only a minor constituent of the discrepancies observed in any assay results. To do this requires that the particle size of the standard material be reduced to a point where the relative standard deviation of the sampling error (i.e... the standard deviation of the errors divided by the average gold content of the material) is 2% or less. For all but very homogeneous mineralization, this means that the material must be reduced to 100% -150 mesh before the sampling errors are adequately controlled. However, even reducing the particle size can contribute to sampling problems. The liberation of gold may cause segregation that can cause large sampling fluctuations that are not easily controlled while maintaining the desired grade. Because, in most cases, the standard material would already be in the "pulp" state when it is submitted to a lab for assay, it is not possible to entirely conceal the nature of the sample from the lab. This is a problem inherent in using assay standard material. Because of the contribution of sampling to error generation in the assay process, the use of "coarse" material does not solve the problem of submitting a totally "blind" standard to the lab. In the sections that follow, the selection, preparation, testing and use of gold fire assay standard material is discussed. While some may dismiss the production of standard material as folly, it is possible to produce and utilize standard material to stabilize and improve the fire-assay process to produce more reliable assay results. Material selection It is desirable to use material that has as nearly the same metallurgical characteristics as the samples with which the standards will be included. However, this is usually difficult. For many reasons, including the particle size at which a significant amount of the gold mineral is liberated, the sampling characteristics of even -150-mesh material may preclude the use of geologically and metallurgically similar ore as a standard. It is usually easier to get material having desirable grade characteristics with the necessary sampling properties than it is to find geologically and metallurgically similar material with the required sampling characteristics. High-grade standards are especially difficult to find and prepare. This is because, as grade increases, the size of the gold particles usually increases. Larger gold particles are liberated and tend to segregate during comminution, and the homogeneity of the material cannot be maintained. For grades much above 3 g/t (0.088 oz/ton), it is very difficult to find material that has the proper sampling properties. Old mill tailings are likely candidates for assay standards. Some of these have sufficiently homogeneous mineral contents, so that the sampling errors can be effectively controlled. Where mill tailings are either not available or are not acceptable, mineralization that has exhibited homogeneous results in reassays of the pulp material is also a good candidate for the standard. Finally, the mineralized rock being sampled may (and should) be used if adequate homogeneity in the -150mesh material exists. "Adequate" ("acceptable") homogeneity is defined below.) It is important to use standards having a wide range of grades. This alone may preclude the material being

Jan 1, 1997

By James Douglas

For the past several years, inflation has placed increasing pressure on our economic system. The total effect of inflation is complex when combined with other factors which affect the economics of equipment. A mathematical model is described herein which can be programmed on a computer to determine the effect of inflation on economic life and profits after taxes. The computer output results in graphic plots which clearly show the relation of various percentages of inflation to life and profits. Inflation has become a common topic of conversation in the recent past. What is it and how can it be measured? What effect does increased inflation have on equipment policy? Research now in progress at the Construction Institute at Stanford University is beginning to unravel some of the mystery of inflation. As a start, let us answer the first question by defining the word inflation. The dictionary says that inflation is "an increase in the volume of money and credit relative to available goods resulting in a substantial and continuing rise in the general price level. It appears, in fact, that we are in an inflationary spiral which is "a continuous rise in prices that is sustained by the tendency of wage increases and cost increases to act on each other." 3 Inflation behaves just like compound interest. It takes more dollars of future years to buy an equal amount of a given commodity, be it a material or a service. Let us assume an annual inflation rate of 5 percent. Assume that 1Associate Professor of Civil Engineering, Stanford University, Stanford, California. 2Webster's Seventh New Collegiate Dictionary, Springfield, Mass., G. and C. Merriam Co., 1969. 3Ibid.

Jan 1, 1972

By C. Patrick, T. Combs

The purpose of this paper is to provide a method for analyzing the sequencing of mining production processes and predicting the effects of any planned changes. Included is a description of the development and utilization of a project-scheduling simulation that combines analytical scheduling with stochastic-activity durations. The simulation is demonstrated in two realistic examples: a longwall face-to-face transfer sequence and the effect of varying advance rates on longwall development sections. Statistical data that can provide a basis for more informed production decisions are generated in the form of project activity/path criticality and project-duration distributions. In both examples, a quantitative risk assessment, regarding subjective decisions in underground production sequencing, is presented.

Jan 1, 1996

By C. W. Schultz

The Mineral Resources Institute at The University of Alabama is engaged in a program whose objective is to develop a beneficiation system for Eastern oil shales. Prior experience (1,2) at The University had shown that the grind required for liberation of the oil rich kerogen was substantially below 20~m. Based on this fact and the rapidly emerging consensus that column flotation is the preferred method for ultrafine particle separation, efforts were focused on the column flotation of oil shale. This paper, which must be regarded as a progress report on a continuing MRI project, describes the observations made and the results achieved to date. The investigation is being conducted in two phases. The first phase has been completed and consisted of a factorially designed experiment conducted for the purpose of screening a set of primary flotation variables. The second phase, which is still in progress, is focused on optimization of the operating conditions and development of a flowsheet for specific shales.

Jan 1, 1989

By Eugie Kabwe

The velocity of detonation (VOD) is a significant property considered when rating an explosive. It may be quantified as a confined or unconfined velocity. The confined velocity is the rate at which the detonation wave propagates through an explosive within a blast hole or any restrained space. The unconfined velocity emanates when an explosive is detonated in the open. However, the confined velocity is more significant as explosives are usually employed under a definite degree of confinement. This paper evaluates the increment of air deck volume and its effect on the VOD. The air gap (air deck) volume is increased stepwise from 10 to 30% of the total blast hole volume by reducing the stemming height in the collar zone. Thereafter, the VOD estimation in the 172-mm-diameter explosive column is conducted using the Dáutriche technique. It is observed that the VOD reduced with the increasing air deck volume and detonation failure is high when 30% of the blast hole volume is replaced with air deck. It is deduced that the increment of the air deck volume has an effect on the attenuation of the steady-state VOD.

By Zhan-fang Cao, Hong Zhong, Li-feng Li, Shuai Wang, Xiao-yu Cao, Guang-yi Liu, Ming-ming Wang

"In recent years, there has been an increasing focus on the hydrometallurgical processing of molybdenite, with the decline of high-grade and easily handled molybdenum minerals. In this work, a novel method of selective extraction of complex molybdenite concentrate, which is characteristic of high-grade copper ore, has been investigated. The results show that the leaching of chalcopyrite was impacted by the pH of the electrolyte when sodium carbonate and ammonium acid carbonate were added to the electrolyte. The leaching rate of molybdenum reached 99.39%, but chalcopyrite could not be leached at all in 240 min under experimental conditions of liquid-to-solid ratio 25:1, concentration of NaCl 4 mol/L, pH=9, room temperature and agitation speed 400 rpm. The grade of copper in the residue was 21.84%, with a recovery 99.93%. The technology of selective electro-oxidation leaching of molybdenum and recovery of copper from complex molybdenite concentrate is an effective separation method. The diffusion control model can be used to describe the process of leaching of molybdenum. The activation energy of the leaching reaction was found to be approximately 9.21 kJ/mol.IntroductionWorldwide, high-grade and easy-to-handle ore is decreasing, while the proportion of low-grade and complex ore is increasing. Porphyry copper ores with associated low-grade molybdenite are very abundant in China. Molybdenum is a valuable byproduct in the processing of porphyry copper deposits where the molybdenum is present in low quantities in the form of molybdenite (MoS2). In the conventional processing of copper sulfide ores by froth flotation, most of the molybdenite in the ore floats along with the copper sulfides. In this way, a bulk copper-molybdenum concentrate is produced. Afterward, the bulk coppermolybdenum concentrate is treated by differential flotation to obtain a rougher molybdenum concentrate, which requires several successive cleaning steps to produce a final molybdenum concentrate. However, it is widely recognized that flotation alone cannot economically produce a high-grade molybdenite concentrate, i.e., with less than 0.5% Cu. Therefore, in most cases, it is necessary to use a final step of chemical purification of the molybdenite concentrate to reduce the copper content to a level suitable for the market (Antonijevic and Pacovic,1992; Gerhardt et al., 2001; Fu and Zhong, 2003; Reza et al., 2006, Rafael et al., 2013)."

Jan 1, 2015

By J. E. Moyle

Introduction In October 1983, Rio Algom Ltd. decided to develop the East Kemptville tin deposit. This decision was the result of intensive geological and metallurgical evaluation work, as well as careful studies of environmental and marketing aspects. The property is located about 54 km (34 miles) northeast of Yarmouth, Nova Scotia, near the village of East Kemptville. Geology The geological setting is a granitic intrusive at the contact with metasediment. The tin mineralization is in the form of cassiterite (SnO2), which is about 80% tin. Sulfides are associated with the deposit containing silver, copper, and zinc, which are recoverable as byproducts. The mineralization is associated with quartz veins and greisen alteration. In the center of the deposit, the zones are so close together they have the appearance of a massive or porphyry type deposit. Ore Reserves The mine plan is based on 56 Mt (62 million st) at 0.165% tin. Up to the ninth year, mill feed will average about 0.2% and lower grade material will be stockpiled. In later years, the average grade will be 0.16% tin. A 0.08% cutoff grade is planned for the entire 17 years of production. East Kemptville is a relatively low grade tin deposit, and only a combination of favorable circumstances has made it economic. • The ore body is situated at, or very close to the surface, permitting mining by open-pit methods at a very low stripping ratio. • The tin occurs in relatively large crystals that allow for good recovery by gravity techniques. • It is ideally located in a well-settled area, with all essential infrastructural requirements already in place. • The positive and cooperative response from various governmental agencies involved are encouraging. Mine Plan The mine plan incorporates an initial pit, followed by three expansions to the ultimate pit. Mining will be at 14 kt/d (15,000 stpd) and will include 9 kt (10,000 st) of ore plus 5 kt (5,500 st) of waste and low grade. After five years, the mining rate will increase to 21 kt/d (27,500 stpd). It is planned to drill vertical 165-mm (6.5-in.) blast holes on a 6-x 6-m (20- x 20-ft) pattern. Bulk slurry explosives will be used to break the rock. Loading will be with hydraulic shovels into 70 t (77 st) haulage trucks. The trucks will deliver the ore to a 13 x 20 m (42 x 65 ft) gyratory crusher, where it will be crushed to 150 mm (6 in.). Secondary and tertiary cone crushers will be used to further crush the ore to 13 mm (0.5 in.). Process Design Tin extraction technology is practically unknown in North America, but Rio Tinto Zinc, which has a 53% ownership in Rio Algom, is involved in tin production in the UK. Rio Algom has used this experience and expertise in metallurgical evaluation work and in tin-related aspects of plant design. Mineralogical and metallurgical test work have been performed, both on laboratory and pilot plant scales. Mineralogical work was done largely in the UK, the laboratory test program was carried out at the Lakefield research facility in Ontario, and pilot plant work was conducted in Australia. This program has resulted in significant improvements in recovery and concentrate grades, over best projections that could be made before this work was carried out. Concentrating Process The concentrator is being designed to process about 9 kt/d (10,000 stpd). The first step is to reduce the size of the ore to -700 µm (-24 mesh) in a rod mill/ball mill grinding circuit. The ball mill will operate in closed circuit with high capacity banana-type screens. The screened product will be deslimed to 30 µm (560 mesh) with cyclones, then passed over the pri-

Jan 4, 1984

By Basil S. Fee

INTRODUCTION Probably the most important family of chemicals used in mineral processing today is a category of basic commodity chemicals loosely denoted as pH regulators. Typical chemicals which are referred to as pH regulators include lime, magnesium hydroxide, soda ash, caustic soda, ammonia, sulfuric acid, and hydrochloric acid. These chemicals are often used in very significant amounts in almost all of the major mineral processing operations such as flotation, hydrometallurgy, etc. (in dosages up to ten pounds per ton of feed ore treated). While cheaper in cost per unit weight of chemical than more specific chemicals such as collectors, frothers, extractents, etc., the overall cost to the mill operator is generally higher with pH regulators per ton of ore processed than with any other given processing chemical. For example, a rough rule of thumb in sulfide mineral flotation is that the cost of lime is double that of the collector(s) used. The symbol pH is used to designate hydrogen ion concentration. When acids, salts, and bases are dissolved in water, individual molecules are dissociated into their constituent radicals or ions. The strength of an acid or a base increases with the extent of such dissociation or ionization. An alkaline solution is one in which the number of Qydroxyl ions (OH ) exceeds the hydrogen ions (H ). In an acid solution, the hydrogen ions predominate. In either case, both ions are always present as water itself ionizes to a limited extent, that is: [ ] The pH scale is logarithmic and the pH value is the negative of the logarithm (base 10) of the molar concentration of hydrogen ions per liter of solution. For example, a pH of 5 means that the molar concentration of hydrogen ions per liter is 0.00001 (1 x 10-5). Likewise, a pH of 9 means that the molar concentration of hydrogen ions per liter is 1 x 10-9. Normally, the relationship between hydrogen ion and hydroxyl ion concentration is based on the relationship: Concentration of H ion x concentration of OH- ion = constant. Eq. (1) In dilute and/or moderately concentrated solutions that are normally used in mineral procygsing processes, the constant at 25°C is 10-14. On the pH scale, the value of pH equal to 7 represents the hydrogen ion concentration of a neutral solution. pH values lower than pH 7 indicate increasing acidity and higher values than 7 indicate alkalinity. Table 1 shows the nature of the pH scale at 25°C. Temperature affects the extent of ionization of dissolved acids, salts, bases, and water so that the hydrogen ion concentration (hence pH) of a solution is also affected by temperature. As a means of demonstrating this dependency, Table 2 shows the change of the exponent of the constant (base 10) in Eq. 1 and the pH value corresponding to neutrality as a function of temperature. It is also important, for example, that alkalinity or acidity expressed by pH not be confused with total alkalinity or total acidity. For example, total alkalinity is commonly determined by titration with a standard acid solution (usually HC1). pH is a measure of the hydroxyl ion concentration of an alkaline solution, whereas titration is a measure of an alkaline solution's acid neutralizing capacity. Thus, if one takes a series of various alkaline solutions prepared using different chemicals but all of exactly the same pH (and temperature) and then subsequently carries out a titration on each solution with a standard acid, it would be observed that the various alkaline solutions would neutralize entirely different amounts of acid. As an example, 0.08 grams of caustic (NaOH) , 6.32 grams of soda ash (Na2CO3), and 8.17 grams of ammonium hydroxide (NH4OH) all have a pH of 11.3 at 25°C. One liter of each of the above solutions neutralizes 0.073, 4.348 and 8.496 grams of HC1, respectively. Therefore, depending on the specific use of any given pH regulator, special tests need to be run by the mill operator to determine factors such as: the specific technical goals to be accomplished by

Jan 1, 1986

By P .. F. Ziemkiewicz, T. E. Rymer, J. J. Renton

Prediction and quantification of acid mine drainage (AMD) from a given rock unit is in its infancy. Acid Base Accounting is the current procedure. II is a useful screening procedure and is designed to indicate whether a rock unit will or will not produce acid. In this sense, Acid Base Accounting is qualiwtivc. Researchers at West Virginia University arc developing a prediction method to estimate how much acid will be produced over the life of a rock unit's AMD production. TI1c AMD prediction procedure can also evaluate the required rates and effectiveness of various AMD control amendments. This presentation discusses the WVU AMD prediction procedure in the context of historical approaches and indicates the mechanics and applications of this new technology.

Jan 1, 1991

This section is addressed to the computing needs of corporate (top) management which are less specialized than the computing needs of operating or divisional management described in the previous sections. It is essential that the computer be used as a tool to serve corporate management and that information and results are presented in a succinct form. A key responsibility of the manager of computer applications is to communicate well with corporate management. A common trap for computer management is to allow the development and implementation of sophisticated and academically interesting computer applications which do not enlighten corporate management in its decision-making task. Thus it is essential to devote sufficient time and effort to designing and specifying a computer system before developing and implementing the system.

Jan 1, 1979

By V. P. Petrov, V. S. Znamensky

There are a great number of asbestos deposits in the USSR, and some are quite large. The main deposits of chrysotile asbestos are shown in Fig. 1. The Bazhenovo deposit, which is located in the Sverdlovsk region, is the largest and has been mined since 1889. Systematic studies of asbestos deposits were begun from the very be- ginning of Soviet government and are most often connected with the names K.E. Tarasov, P.M. Tatarinov, B.J. Merenkov, N.D. Sobolev, F.V. Syromjatnikov, V.N. Lodochnikov, V.P. Petrov, V.P. Eremeev, L.A. Sokolova, V.R. Artemov, and V.F. Dybkov. Investigations by these people and others helped the USSR be one of the most important producers of asbestos in the world. In this chapter, the main asbestos deposits of the USSR are explained in a general outline from the point of view of their genesis, common features, and differences that must be taken into account for forecasting, exploration, and the mining of asbestos. All chrysotile asbestos deposits of the USSR were formed from magnesium rock as a result of comparatively low grade temperature, hydrothermal alteration, and free crystallization of fibrous chrysotile in open cavities that served as catalysts. Chrysotile, a widely distributed rock-forming mineral, is not in general an important mineral for industry as it is a long fibrous and main vein variety. However, it must serve for geologists as an important prospecting criteria. And, the geologist should search for the mineral not only in places of possible chrysotile generation, but in places where asbestos may be crystallized in its long fibers. The Soviet geologists distinguish chrysotile asbestos deposits in two groups: (1) ultrabasic rocks and (2) sedimentary dolomite terranes, taking into account how the maternal rock was affected by hydrothermal solutions. The largest deposits are related to ultrabasite. In dolomite, the largest deposits are small and their asbestos contains almost no iron, which makes it a better filler for insulating plastics. There are a few types of asbestos deposits within the limits of the ultrabasic group, the largest being a Bazhenovo type to which the deposits of the Urals are attributed. This Bazhenovo asbestos is characterized by the most complete zonality of disposition for various kinds of asbestization. These types of deposits are, for the most part, the only ones presently working in the USSR. The Bazhenovo deposit is connected to the gabbro-peridotite intrusive belt that is about 180 km long and is broken into a number of fault blocks as shown in Fig. 2. Granitoid veins of intruded rocks are found along the deepest faults, and were sources of hydrothermal solutions that affected the maternal rock. The nearest granitoid vein, ultrabasite, is substituted by listwanite that consists of quartz, talc, carbonate, and some other essentially talcous rock. The next zone of hydrothermal alteration is composed of serpentine. The boundry between serpentinite and unaltered ultrabasite is as irregular, diffuse, and rough as the boundaries between all other zones. Nearer the vein of granitic rock, ultrabasite is entirely serpentinized into massif serpentinization that penetrates through cracks and fissures. There is an unaltered ultrabasite space between fissures, and the biggest block is preserved unaltered in the central part of the massif. The Soviet geologists call this ultrabasic sopka. Preservation of the unchanged blocks is of great importance for deposit formation as volume changes are impossible during the process of asbestization and serpentinization. It may be confidently said that asbestization in Bazhenovo type deposits is possible only when alteration of ultrabasite has not completely occurred. Under conditions of massif volume, a very big mass of substance must be carried out of the ultrabasite for the progressive stage of serpentinization. However, in the regressive stage,

Jan 1, 1986

By P. J. Hyde

Cyanide is a significant contributor to the cost of producing gold. Any reduction in consumption that can be achieved through more reliable measurement and/or more efficient management of dosing rates can have a significant financial impact. An automated analyser, by its mechanical nature, is not prone to the variations of human operators, particularly notice­able from shift to shift. Reducing the standard deviation of the measurement (ie, reducing noise ), the automated measurement will give the metallurgist a truer picture of the underlying patterns of usage. The combination of automated analysis with on-line operation and a control loop provides timely information that is efficiently utilised in the management of the plant. Until the release of the Chemtronics On-Line Cyanide Analyser (OTA3000) about 12 months ago, the Australian experience with automated analysers (especially on-line) had been substantially negative. Since its release, the Chemtronics instrument has outperformed both manual titrations and other automated instruments, gaining an enviable reputation for accuracy and reliability in the process. The instrument is built around the technique ofpotentiomet­ric titrations, in which the potential of an indicator electrode is measured as a function of the volume oftitrant that is added to the sample. As the cyanide in the sample is depleted, the potential on the indicator electrode changes rapidly, producing a typical sigmoidal curve. The OTA3000 takes the first differ­ential of the curve, identifies the apex of the peak as the endpoint and reports the result in parts per million of sodium cyanide. The OTA3000 accepts the peak as valid only if it exceeds a certain minimum height. That minimum height is a programma­ble parameter and is used to optimise the relationship between sensitivity and the rejection of spurious peaks. The instrument will track cyanide trends through both positive and negative slopes of any magnitude, but it will provide a warning if it the slope becomes excessive, not because it affects the instrument but because it means that something unusual is occurring in the tanks. There are a limited number of conditions under which the instrument will produce a result of zero. In these cases the OTA3000 holds any output lines at the last valid result while it repeats the titration. This feature was incorporated for reasons of stability in a feedback loop.

Jan 1, 1995

By Robert L. Sundeen, Richard V. Wenberg

INTRODUCTION Shotcrete can be defined as spray-applied concrete with a maximum size aggregate in excess of 9.5 mm (0.375 in.). The concept evolved from guniting, which uses smaller aggregate and generally produces concrete with poorer physical properties. The larger aggregate used in shotcrete requires less water in the mix, while offering larger bonding surfaces, thus imparting greater tensile and compressive strengths. TYPES OF MACHINES Two basic designs of shotcrete machines in general use are the wet or prehydrated zero slump and low slump type, and the dry or nozzle hydrated machines. Fig. 1 shows a typical zero-slump machine in a con¬figuration commonly used for small-volume require¬ments. The prehydrated zero-slump machine mixes exactly proportioned batches in a sealed pressurized chamber. "Slugs" of hydrated shotcrete mix and com¬pressed air are sent through a distribution hose to the nozzle where the wet mix emerges as a spray. The spray is directed onto the rock surface by the nozzle operator. The low-slump machine is similar, except that a thick slurry mix of aggregate is delivered throughout the distribution hose to the nozzle. Fig. 2 shows a typical low-slump shotcrete machine used in moderate-volume applications. The nozzle-hydrated shotcrete machine accepts a premixed "dry" (4 to 6% moisture) batch of aggregate and cement. A constant feed of this mix is directed by compressed air through the distribution hose to the nozzle. At the nozzle, water for hydration is introduced into the emerging spray. The hydrated mixture is di¬rected onto the rock surface by the nozzle operator. Fig. 3 shows a typical dry-mix machine. The major difference between the two basic machine designs is the consistency of the shotcrete quality. Pre¬hydrated machines have high degrees of consistency and produce high-strength concrete, resulting from their ability to produce exact mixtures of zero or low-slump concrete within the machine. Nozzle-hydrated machines tend to produce poorer quality concrete because of the difficulty in controlling the water content at the nozzle. Water input essentially is an art. As such it is not precise, and the shotcrete produced is of variable quality. MATERIALS HANDLING Materials handling presents a problem to either type of shotcrete machine. At times considerable quantities

Jan 1, 1982

By Edward D. Elkins

CalMat Co. is one of the largest producers of construction aggregates in the United States. The CalMat Reliance plant is located in Irwindale, California, and is situated on an alluvial fan formed by the San Gabriel River. The fan deposits are of Quaternary age and were derived from granitic and metamorphic bedrock exposed in the San Gabriel Mountains to the north. The sediments are poorly sorted and range in size from large boulders to clay size particles. The coarse material is hard and durable and consists of quartz diorite, granodiorite, granitic gneiss, schist and minor amounts of volcanic rock. The matrix is comprised of silty sand with varying amounts of clay. Thin silt and clay lenses occur sporadically throughout the deposit. The clay content of the matrix and the degree of weathering of the coarse material increases with depth. Based on seismic data, the alluvial fan varies from 200 to 400 feet thick and overlies crystalline bedrock. The Reliance plant was constructed in 1971 and 1972. It has a nominal capacity of 1200 to 1400 tons per hour, and typically produces more than three million tons of aggregate products per year. Alluvial fan sediments are mined from a gravel pit situated adjacent to the plant. The material is fed through a mobile, 48- by 60-inch, in-pit jaw crusher, then transported by conveyor to the main plant to be processed. The plant is modern and highly efficient. It has a typical wet and dry circuit and is capable of producing a wide range of aggregate products. To extend the life of the plant, a 9,000 foot conveyor system is being constructed to connect it to a granite quarry operated by Azusa Rock Inc. The quarry is located north of Irwindale, and is 50 percent owned by CalMat Co. The conveyor will be used to export crushed rock from the quarry to Reliance. The existing plant will be modified to process the crushed rock.

Jan 1, 1995

By Jurgen H. L. Stuve, Rick E. M. Widelski

Any one who has been involved with the mining or the treatment of gold-bearing ores will agree that at the end of the total process, the quantity of gold recovered is never as much as was expected. These short falls are usually explained away by claiming that:- the original evaluation of the ore being mined was inaccurate the plant treatment process is not at its optimum theft is taking place excessive dilution occurs during mining The truth of the matter is that the gold is being "lost". During all the stages of its production, starting from the mining, right through to the smelting process, a certain amount of gold is not recovered. When a newly found gold deposit is being evaluated and the drilling data being analysed, cautionary factors are applied so that the grade is not over-stated. Once the decision has been taken to start mining and development is in progress, the exposed ore is sampled. When Again, when this data is being analysed more safety factors are used to ensure that when a piece of ground is given a block-value, the gold content is not over estimated. mining actually takes place and the ore transported to the plant for treatment, the metallurgical department also introduces certain safety factors when predicting how much gold will be recovered. This shows that the quantity of gold that one expects to finally recover is already a conservative figure. The actual quantity recovered invariably falls short of this figure, in fact, South African gold mines are satisfied when achieving a Mine Call Factor (MCF) of 85%. In the last few years a great deal of work has been carried out to investigate how this unaccounted for gold is lost and what steps can be taken to remedy the situation. Previously mined areas were re-swept using the traditional hand brush method; experiments were carried out with water jets sometimes in conjunction with compressed air and even small portable vacuum cleaners were tried. None of these methods were totally successful in recovering the left over gold but the use of vacuum systems showed the most potential. the In South Africa this has resulted in introduction of high-powered vacuum systems on the gold mines to suck-up this lost gold. These were initially introduced in the mining areas, particularly old areas, to recover any left over ore. The results proved extremely profitable, So much SO that presently, well over 120 suction systems are doing this type of work. This success encouraged plant operators to look at areas of potential loss in the plant itself. This has resulted in an increasing number of machines being supplied to gold plants mainly for clean-up applications with the bonus that the material sucked up invariably contains gold.

Jan 1, 1993

By H. Nematollahi

The comminution laws The grindability of the ores or minerals is an important factor that could not be deduced from their mineralogical characteristics (Testut, 1958). Many theories have been proposed to determine the amount of work required in a comminution operation as a function of the nature of material, size and shape of the minerals before and after of comminution. The most important laws are proposed by Rittinger, Kick and Bond. According to Rittinger (1867), the work is directly proportional to the new surfaces produced during the comminution. Kick (1885) believes that the work needed to produce similar modification is proportional to the volume or mass of the material. Bond, during his 25 years of experience and research, reached to the conclusion that on the average the energy required for a known comminution is between the Rittinger's and Kick's laws (Bond, 1952). According to Bond's Third Theory, the work input is proportional to the new crack tip length produced during the particle breakage, and equals the work represented by the product minus that represented by the feed. The work input is given by the following equation: [111W = 1 Owi( FP FF(1)] where W is the The energy input to the mill, (kWh/st); F is the Feed size based on 80% passed, (µm); P is the Product size based on 80% passed, (µm); and W; is the Work Index, (kWh/st) Methods to determine the Bond Work Index Various methods have been proposed by Bond to determine the Work Index. There are some problems with the Standard Bond Grindability Test: •Requiring a special mill. •Requiring a relatively large amount of sample. •Being relatively time consuming and tedious. [ ] During the recent years, numerous attempts have been made to install faster and easier methods to determine the Work Index. A complete description of these methods is presented by Yap et al. (1982). Determining the Bond Work Index by a new size ball mill During the recent years, researchers at the University of Tehran determined the Bond Work Index, using the Anaconda simplified methods by its Denver 305- x 127-mm (12 x 5-in.) laboratory wet ball mill. The equation used to calculate the Work Index for the above mill containing 16.12 kg (35.5 lbs) of steel balls is as follow: [1 _ 1-1W;=0.54x [P [)(2)] In addition, an attempt has been made to determine the Bond Work Index similar to standard Bond ball mill procedure but by a different ball mill, smaller and in consequent portable and less laborious than the Bond ball mill. Method description Just as the standard Bond grindability test, this method is a locked-cycle dry grinding and screening process that is carried out until steady state condition is reached. Test equipment The test is conducted in a special ball mill, called the New Size Ball Mill (NSBM). The mill is scaled down with a coefficient of two-thirds from the standard Bond ball mill. It measures 200 x 200 mm (8- x 8-in.) inside (instead of 300 x 300 mm or 12 x 12 in. in Bond ball mill). Just like the Bond

Jan 1, 1995

By O. H. Tuovinen, T. M. Bhatti

Microbiological leaching has been used as an alternative approach to conventional hydrometallurgical methods of uranium extraction. In the microbiological leaching process, iron-oxidizing bacteria oxidize pyritic phases to ferric iron and sulfuric acid, and uranium is solubilized from the ore due to sulfuric acid attack. If uranium in the ore material is in the reduced, tetravalent form (UIV), a red ox reaction is involved whereby uranium is oxidized to the hexavalent form (UvI) upon dissolution. In acid-leaching systems, the primary oxidant is ferric iron, which is reduced to ferrous iron by its chemical reaction with UIV. The ferrous iron thus formed is reoxidized to ferric iron by iron-oxidizing bacteria such as Thiobacillus ferrooxidans and Leptospirillum ferrooxidans. Nutritional requirements and responses to environmental extremes of acidophilic iron-oxidizing bacteria are appraised. The S-entity in Fe-sulfides is oxidized to sulfate by bacteria similar to Thiobacillus ferrooxidans and Thiobacillus thiooxidans. Pyrite and marcasite oxidation is a sulfuric acid forming reaction. Heap, dump and in situ leach techniques are feasible as bacterial leaching systems.

Jan 1, 2000

By Gunnar Nord, Per-Anders Persson, Duri Prader

The paper gives a short resumé of the findings of the joint US - Swedish "Mechanical Boring (MB) or Drill and Blast Tunneling (DB) ? Workshop" in Stockholm December 1976. New developments and trends in both MB and DB techniques in Europe since 1976 are described, and information is given on some interesting new projects. Differences in contracting principles between the US and Europe lead to different practices for preinvestigations, design, excavation and arbitration and legal affairs. It is shown that this often leads to a higher cost in the US than in Europe for similar excavation jobs. A typical US tunnel job also has more men on the site than a similar job in Europe. This adds to the cost difference and is only partly balanced by a slightly higher rate of excavation in the US. A comparison is made between two quantitative methods of predicting MB penetration rates. These are in reasonable agreement. A probable future development in machine design is an increase in cutter load towards perhaps 25 tons per disc where the tunnel diameter is large enough to house the machine size dictated by the size of the very large electric motors required. Recent European developments in drill and blast technique include improved methods for cautious blasting, precision hydraulic drilling of small diameter holes, and non-electric interval blasting techniques. These developments still keep the drill and blast method ahead of mechanical boring costwise for tunnels in igneous rock whereas for most sedimentary materials mechanical boring has a lead, particularly for long tunnels.

Jan 1, 1979