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By Dan A. Brown

The New Mississippi River Bridge (MRB) employed conventional design-bid-build procurement, but the bidding process included the opportunity for pre-qualified bidders to develop and propose confidential Alternative Technical Concepts (ATC?s) for evaluation and approval in advance of bidding. The ATC process provides a mechanism for value-engineering prior to bid, with incentives and opportunities for both the owner and the bidding contractor. The contract described in this paper is for the river bridge composed of a cable-stayed crossing with two pylons and the anchor piers on each side of the river. The Massman-Traylor-Alberici (MTA) team developed a re-designed foundation system for the main pylon and anchor foundations which utilizes their equipment and experience along with field load test measurements to optimize the foundations. The paper describes the experience with the use of the pre-bid ATC process on the project from the perspective of the bidder, and outlines some of the benefits and limitations of this concept with respect to foundation design and construction. Some keys for success are identified and impediments to the effective implementation of the process are described.

Jan 1, 2011

By W. G. Pariseau

Conventional barrier pillars in underground coal mines are intended to protect mains and sub-main entries for relatively long periods of time. Inter-panel barrier pillars that may be used in especially deep coal mines, say, below 800 m (2,400 ft), are intended to protect longwall panel faces and adjacent panel entries for the life of nearby panels. Barrier pillar failures have serious consequences, especially when failure is sudden and catastrophic. Design of barrier pillars is thus of considerable importance to safety and efficiency of operations. Locally, failure occurs at the limit to elasticity when stress tends to exceed strength. Strength of rock is stress dependent, so an analysis of stress is central to rational design of barrier pillars. There are several approaches to stress analysis for coal mine pillar design that range from the simplified but popular tributary area method including variations that use empirical definitions of the tributary area (volume) to more sophisticated numerical methods such as the boundary element method and the finite element method. Despite ease of use, empirical methods are limited and always suspect when applied to conditions beyond their scope of definition, while boundary element methods do not generally include pillars in the solution domain. The finite element method is not so restricted and thus is a natural choice for conventional and inter-panel barrier pillar safety analyses. Caving of roof strata certainly affects the distribution of stress about longwall faces and nearby barrier pillars. Conventional wisdom that considers roof strata to respond as cantilever beams indicates the longer the span the greater are loads on longwall faces and face supports. Failure of roof beams, caving, shortens spans and relieves face and support loads. The same view is applicable to barrier pillars in that long roof spans outside of barrier pillars leads to higher pillar loads than short spans. Hence, caving should be beneficial to barrier pillar safety. However, the benefit may be small depending on the extent of caving and only in the form of a slight reduction in overburden weight transmitted to the barrier pillar, especially in case of inter-panel barrier pillars. FINITE ELEMENT MODEL Finite element models for stress analysis in rock mechanics are formulated in consideration of physical laws, kinematics, and material laws. These models meet the essential requirements for equilibrium of stress, satisfy strain displacements relations, and in a general sense, relate stress to strain through material laws, often referred to as constitutive relations. The last may be complex and replete with nonlinearities. A de facto engineering material law is the elastic model in the form of Hooke?s law for linear, homogeneous materials that may be anisotropic. A limit to elasticity is imposed by strength in the form of a failure criterion or yield function that accommodates triaxial stress states. Examples used in rock mechanics are the well-known Mohr-Coulomb (MC)criterion, the popular Hoek-Brown (HB) criterion, and the Drucker-Prager (DP) criterion. The first is linear in stress, while the second is nonlinear. Both do not include any effect of the intermediate principal stress. The last is also linear in stress but includes the intermediate principal stress. All show an increase in the major principal stress at failure with confining pressure. In the UT2/3 finite element programs, the failure criterion Y is a nonlinear, anisotropic form and is also the plastic potential [1]: [YJI=+ 11NYJI=+- (1) ] where N is an exponent, usually 2, and J2 and I1 are nonlinear anisotropic forms of the second invariant of deviatoric stress and the first invariant of principal stress. The form (1) reduces to DP when N=1and isotropy prevails. When Y = 1, the limit to elasticity is reached and plastic deformation is possible. The material constants that enter (1) are all computed from unconfined compressive, tensile, and, in case of anisotropy, shear strengths. These latter strengths may be estimated from conventional laboratory test data. Of course, joints in field-scale rock masses induce different values. How different is often a matter of conjecture. For comparative studies, laboratory rock properties seem adequate. Inelastic, that is, plastic behavior occurs with strain beyond the elastic limit. Strain hardening, ideal plasticity, and strain softening are possibilities. Ideally brittle behavior with complete loss of strength is an extreme form of strain softening [2]. Caving may also be considered as strain softening with a less drastic reduction of material properties. As intact roof strata fail, fall, and form a rubble pile below, a new material is formed with greatly reduced elastic moduli and strengths, say, 1/100th of the original intact moduli and strengths. From another viewpoint, caving is simply a natural cut and fill process. Input for stress analysis by the finite element method includes: 1. Preexcavation stress state 2. Geology (stratigraphic column) 3. Structure (major faults, joint sets) 4. Elastic moduli 5. Strengths, and specific gravities of each rock type 6. Excavation sequence Output from an analysis includes distributions of: 1. Stress 2. Strain 3. Displacement 4. Local factor of safety (element fs) The last is defined as the ratio of strength to stress and in view of the failure criterion (1), is the ratio [fsJstrengthJstress=] where the numerator is the maximum possible value and the denominator is the actual value of the square root of the second invariant of deviatoric stress. This invariant is a measure of shear stress intensity but also reduces to uniaxial tensile and compressive strengths as the case may be. An fs greater than 1 indicates an elastic state of element stress; a value of 1 indicates the limit to

Jan 1, 2009

By Z. A. Md-Nor

Risk management is an established loss-control methodology that has been applied successfully in many industries. Recently, interest in this structured approach has grown in the mining industry. The main objective of this research was to develop a risk-assessment process, which is a part of risk management, that can be used by the U.S. mining industry to more thoroughly characterize risks associated with haul truck-related fatalities. The assessment is based on historical data obtained from the U.S. Mine Safety and Health Administration (MSHA) investigation reports, which includes 113 fatal incidents that occurred from 1995 through 2006. The risk-assessment process used in this research involves the following basic steps: identification of the risks, risk analysis and risk evaluation. The preliminary hazard assessment (PHA) method is used in identifying and quantifying risks. Risk levels are then developed using a pre-established risk matrix that ranks them according to probability and severity. The resulting assigned risk value can then be used to prioritize control strategies. This paper is a part of a detailed study on risk assessment for equipment-related fatalities in mining sponsored by the Western U.S. Mining Safety and Health Training and Translation Center.

Jan 1, 2008

While ring-blasting techniques offer substantial benefits, there are intrinsic problems that are exacerbated when large blocks of ore are mined in attempts to improve the production economics. These problems relate to both the convergence of holes around the ring drive and the charging of up-holes. The size and shape of the ore block concerned should be carefully considered so that good drilling, patterns can be achieved. The hole diameters chosen, both affect and are affected by the size of the ore block. The explosive selected depends on the hole diameter and together with the diameter, determines the nominal drilling pattern. Ring-drilling patterns are best derived through the use of computer software together with engineering judgement. The loading of explosive is necessarily high, and charging patterns require careful thought if overcharging is to he minimised and adequate energy is to be available at all parts of the ore block. Initiation systems are critical to the blasting results, but sub-optimal systems must be adopted if collar priming is practised. In view of all these factors, idealised guidelines can be given for ring design, but it should be recognised that developments in technology may in time change some of these parameters.

Jan 1, 2000

By Eugene Frey

"INTRODUCTION The Iron Mountain deposits were examined in the summer of 1943 by engineers of the Bureau of Lanes, and a diamond-drilling project was set up under the Strategic Minerals Program. Actual drilling operations, under the supervision of a Bureau engineer, commenced November 9, 1943, and were completed on November 1, 1944. When the Bureau began detailed exploration of the Iron Mountain deposits, the Federal Geological Survey participated. Dr. I. W. Newhouse, of the Survey, and Dr. A. F. Hagner, of the Wyoming Geological Survey, were assigned to the investigation.ACKNOWLEDGMENTSIn its program of exploration of mineral deposits, the Bureau of Mines has as its primary objective the more effective utilization of our mineral resources to the end that they make the greatest possible contribution to national security and economy. It is the policy of the Bureau to publish the facts developed by each exploratory project as soon as practicable after its conclusion. The Mining Branch, Lowell B. Moon chief, conducts preliminary examinations, performs the actual exploratory work, and prepares the final report. The Metallurgical Branch, R. G. Knickerbocker, chief, analyzes samples and performs beneficiation tests. Both these branches are under the supervision of Dr. R. S. Dean, assistant director.With respect to this report, acknowledgment is made to S. R. Zirrmerley and F. T. Allsman, Bureau of Mines, Salt Lake City, Utah, who supervised the work of this project."

Nov 1, 1946

By R. K. Matheson

THE Nivloc mine is 9 miles west of Silver Peak, Esmeralda County, Nevada. It has been operated by Desert Silver, Inc., since the summer of 1937. The cyanide mill treats 19o tons of silver-gold ore per day. In March of this year the main operations were about 2000 ft. west of the main shaft, which served as the downcast air passage for the mine. The mechanical blowers in use were sufficient to ventilate the west end at that time, but it was foreseen that as the mining area was extended ventilation would be an increasingly difficult problem, especially during the summer months. It was decided, therefore, to continue a raise from a point 100 ft. above the highest mining level (440-ft. level) to the surface; a distance of 443 ft. at approximately 65° From this point 40 ft. of drifting would be necessary to connect with a shallow shaft on the surface. (Fig. I.) All the raise was serviced from the 440-ft. level; costs and progress obtained in the initial 100 ft. are included in the following pages. CHOICE OF RAISE A raise in this locality would have four objects: (I) to provide an ample supply of fresh air to the working places; (2) to explore the vein from the highest mining level to the surface; (3) to provide waste fill for stopes in the area, and possibly provide a transfer for waste from the surface; (4) to provide an emergency exit in the west end of the mine. Before the mining of an ore shoot in the eastern part of the mine, a similar raise had been driven for purposes of ventilation and exploration. This raise and subsequent mining showed the vein to be noncommercial for a distance of about 250 ft. below the surface. However, in the area under operation, old surface workings showed occasional values. Thus from a development standpoint the raise was an excellent prospect. The raise chosen was a three-compartment, stulled raise having two chutes and a manway. The manway inside timber measured 4 ft. 8 in. by 4 ft. 6 in.; it carried air and water pipes, a ladderway and a slide for a timber skip. The two chutes would be used for drawing off the muck and the cleaning of the bulkhead, as well as for ore and waste passes, which would be especially useful for lateral exploration from the raise. The fairly large cross section of the raise (6 by 15 ft.) would provide a large quantity of waste for backfilling, and more could be obtained from a glory hole on the surface. TIME SCHEDULE As speed in completing the raise was the main requisite, three shifts were considered, but two shifts were decided upon when it was found that three would necessitate considerably larger fans than those available for clearing the raise. The main diffi-

Jan 1, 1942

By R.K. Matheson

The Nivloc mine is 9 miles west of Silver Peak, Esmeralda County, Nevada. It has been operated by Desert Silver, Inc., since the summer of 1937. The cyanide mill treats 190 tons of silver-gold ore per day. In March of this year the main operations were about 2000 ft. west of the main shaft, which served as the downcast air passage for the mine. The mechanical blowers in use were sufficient to ventilate the west end at that time, but it was foreseen that as the mining area was extended ventilation would be an increasingly difficult problem, especially during the summer months. It was decided, therefore, to continue a raise from a point 100 ft. above the highest mining level (440-ft. level) to the surface; a distance of 443 ft. at approximately 65°. From this point 40 ft. of drifting would be necessary to connect with a shallow shaft on the surface. (Fig. I.) All the raise was serviced from the 440-ft. level; costs and progress obtained in the initial 100 ft. are included in the following pages. Choice of Raise A raise in this locality would have four objects: (I) to provide an ample supply of fresh air to the working places; (2) to explore the vein from the highest mining level to the surface; (3) to provide waste fill for stopes in the area, and possibly provide a transfer for waste from the surface; (4) to provide an emergency exit in the west end of the mine. Before the mining of an ore shoot in the eastern part of the mine, a similar raise had been driven for purposes of ventilation and exploration. This raise and subsequent mining showed the vein to be noncommercial for a distance of about 250 ft. below the surface. However, in the area under operation, old surface workings showed occasional values. Thus from a development standpoint the raise was an excellent prospect. The raise chosen was a three-compartment, stulled raise having two chutes and a manway. The manway inside timber measured 4 ft. 8 in. by 4 ft. 6 in.; it carried air and water pipes, a ladderway and a slide for a timber skip. The two chutes would be used for drawing off the muck and the cleaning of the bulkhead, as well as for ore and waste passes, which would be especially useful for lateral exploration from the raise. The fairly large cross section of the raise (6 by 15 ft.) would provide a large quantity of waste for backfilling, and more could be obtained from a glory hole on the surface. Time Schedule As speed in completing the raise was the main requisite, three shifts were considered, but two shifts were decided upon when it was found that three would necessitate considerably larger fans than those available for clearing the raise. The main diffi-

Jan 1, 1943

By R. K. Matheson

The Nivloc mine is 9 miles west of Silver Peak, Esmeralda County, Nevada. It has been operated by Desert Silver, Inc., since the summer of 1937. The cyanide mill treats 190 tons of silver-gold ore per day. In March of this year the main operations were about 2000 ft. west of the main shaft, which served as the downcast air passage for the mine. The mechanical blowers in use were sufficient to ventilate the west end at that time, but it was foreseen that as the mining area was extended ventilation would be an increasingly difficult problem, especially during the summer months. It was decided, therefore, to continue a raise from a point 100 ft. above the highest mining level (440-ft. level) to the surface; a distance of 443 ft. at approximately 65°. From this point 40 ft. of drifting would be necessary to connect with a shallow shaft on the surface. (Fig. I.) All the raise was serviced from the 440-ft. level; costs and progress obtained in the initial 100 ft. are included in the following pages. Choice of Raise A raise in this locality would have four objects: (I) to provide an ample supply of fresh air to the working places; (2) to explore the vein from the highest mining level to the surface; (3) to provide waste fill for stopes in the area, and possibly provide a transfer for waste from the surface; (4) to provide an emergency exit in the west end of the mine. Before the mining of an ore shoot in the eastern part of the mine, a similar raise had been driven for purposes of ventilation and exploration. This raise and subsequent mining showed the vein to be noncommercial for a distance of about 250 ft. below the surface. However, in the area under operation, old surface workings showed occasional values. Thus from a development standpoint the raise was an excellent prospect. The raise chosen was a three-compartment, stulled raise having two chutes and a manway. The manway inside timber measured 4 ft. 8 in. by 4 ft. 6 in.; it carried air and water pipes, a ladderway and a slide for a timber skip. The two chutes would be used for drawing off the muck and the cleaning of the bulkhead, as well as for ore and waste passes, which would be especially useful for lateral exploration from the raise. The fairly large cross section of the raise (6 by 15 ft.) would provide a large quantity of waste for backfilling, and more could be obtained from a glory hole on the surface. Time Schedule As speed in completing the raise was the main requisite, three shifts were considered, but two shifts were decided upon when it was found that three would necessitate considerably larger fans than those available for clearing the raise. The main diffi-

Jan 1, 1943

By Hans G. Kok

At this time in the history of engineering and construction, cost indexes give weekly evidence of the sharply rising costs of labor and materials. Paralleling the rise in construction costs are the costs of plant maintenance. Automation of the modern industrial plant has permitted a substantial reduction in the total number of employees required for plant operation. Of this reduced total, however, up to one-half of the employees may be engaged in plant maintenance. Management's decision to risk investment capital for a pro- posed plant or improvement is increasingly dependent upon the engineer's ingenuity in designing plant systems and structures which will require a minimum of materials and labor for manufacture, installation, and operation and which will be maintenance-free to a maximum degree. This paper concerns itself with the structural elements of bulk material handling systems in which the foregoing factors influenced the design to a considerable extent.

Jan 1, 1969

By D I. Cliff, K Johnstone, N LaBranche

The resurgence of silicosis and coal worker’s pneumoconiosis (CWP) in mining and the recent ban of engineered stone products underscore the need to comprehensively address mineral dust exposures and the associated health hazards. Worker exposure to particulate matter, remains a pressing concern demanding nuanced exploration as all dusts are not created equal. The established link between silica and respiratory diseases necessitates a more detailed understanding of the characterisation of particulates and their health impacts. While current monitoring relies on total mass, it is increasingly evident that the size, shape, mineralogy and agglomeration of particles play a crucial role in determining health impacts. This has been evident in the engineered stone industry with workers not only facing a greater risk of silicosis, but also seeing a faster progression of the disease and a higher likelihood of dying from it. Simultaneously, the surging demand for critical minerals, integral to low-carbon economies, has seen government and industry fast tracking the exploration, mining and processing of these commodities. Critical minerals, crucial in technologies like lithium-ion batteries and high-performance alloys, are central to achieving global environmental goals. Significant funding is currently being allocated to advance the development and processing of these materials, however there remains a considerable gap in understanding potential health risks, urging comprehensive health risk assessments. To tackle these challenges, the University of Queensland has developed innovative dust characterisation methodologies, including the use of the Mineral Liberation Analyser (MLA), a scanning electron microscope, to characterise respirable and inhalable dust samples. This advanced technique, initially applied in coal and metals mining, is expanding to critical minerals mining, unravelling diverse mineralogical components and particle size distributions. This conference presentation will showcase the latest research, emphasising the use of advanced analytical techniques to deepen our understanding of the size, shape and mineralogical composition of particulate matter. By synergising insights from coalmines, metals mines, smelters, engineered stone and critical minerals assessments, we aim to proactively manage health risks, contributing to safer workplaces and responsible production practices in the mining sector.

Apr 16, 2024

By A. J. Kraemer

For a number of years the Bureau of Mines has conducted semi-annual survey to determine the changes in the characteristics of the rotor gasoline being sold in the United States. Certain cities have been decided upon as points of origin of samples. Tare choice of these cities has been governed by their situation as the representative of a more or less definitely bounded marketing territory. Samples have been taken on or about the fifteenth day of January and July of each year. In the collection of the samples an attempt has been made to get a representation of the important types of motor fuel being sold, and also to tame into consideration the proportion of tie total consumption that is supplied by the different marketers. It is therefore felt that the survey affords a reliable indication of the quality of the gasoline being sold throughout the United States at the time of sampling. The figures in Table I, below, were obtained by averaging the end points of the least volatile and most volatile samples of gasoline. For example, of 850 samples analyzed in 1919, the 10 per cent (85 samples) having the highest end points, and the 1C per cent having the lowest end points were averaged. These averages are represented in the table as the "Maximum 10 per cent" and the "Minimum 10 per cent", respectively. The same scheme was used in the succeeding surveys.

Jan 1, 1925

By S. Dwarapudi

Pelletizing of Indian chromite ores is more challenging due to their high refractory nature. High Cr/Fe ratio and high MgO content in these ores demand high firing temperatures and longer firing cycles but often result in low strength fired pellets. Aim of this study was to develop cold bonded chromite pellets for smelting in submerged arc furnace (SAF) from chromite fines using suitable binder that induce less gangue into the pellets but cures quickly. Different binders were studied through laboratory pelletizing experiments for their suitability for cold bonding the pellets. As result, a composite binder comprising organic and inorganic binding agents was found to be suitable and pellets made from the same were tested for their low and high temperature behavior. High temperature reduction studies revealed that pellets were resistant to disintegration up to 1200 °C.

Sep 1, 2012

By L. B. Pankratz

The Bureau of Mines measured heat capacities of alpha and beta spodumene (LiAlSi2O5) and of beta eucryptite (LiAlSiO4) over the temperature range 51° to 298° K. The entropies at 298° K were found to be 30.9 ± 0.2 cal/deg mole for alpha spodumene, 36.9 ± 0.3 cal/deg mole for beta spodumene, and 24.8±0.2 cal/deg mole for beta eucryptite. High-temperature heat content measurements above 298° K were conducted for the same compounds. Measurements were made to 1,150° K for alpha spodumene, 1,600° K for beta spodumene, and 1,470° K for beta eucryptite. Tables of heats and free energies of formation are given from the elements and from the constituent oxides.

Jan 1, 1967

By P. Archer

The Blondeau Formation has been examined in a stratigraphic drill section at Barlow Lake. The formation contains pyroclastic flows and volcanogenic sediments throughout, but epiclastic sedimentary rocks dominate the upper parts of the section. Five distinct types of pyroclas tic flows are known in the section, as well as minor mafic lava flows. Volcanogenic sediments are commonly intercalated with the true volcanic rocks. Sedimentary rocks include fine-grained intraclastic sandstones, heterolithic sandstones and green and black shales. Interpretation of the origin and relative disposition of the volcanic and sedimentary rocks favours a model of near-emergent vulcanism in an upward shoaling environment.

Jan 1, 1984

By R. F. Favreau, Patrice Favreau

"A blast with explosives creates vibration waves in the zone around the blast.The prediction of the intensity of vibrations is important because people in the cinity demand that the vibrations do not disturb them, and society imposes limits on the vibrations. The U. S. Bureau of Mines (USBM) showed that the easurement of the particle velocity U predicts the risk of damage from vibrations. A currentapproach of blasters to monitor vibrations is to measure them with a seismograph at several distances S, and create a graph of U vs. S and the explosive charge “W” according to the equation U = K / (S / W)N where K and N are constants that depend on the blast method and the nature of the ground over the distance S. Though useful, this equation is empirical, and the results predicted from it are very variable, as shown by figure 1 which indicates that a prediction of U can deviate significantly from the actual value. Moreover, a minimum of two seismograph measurements at two values of S are required to establish the curve. However, here now exists another very convenient method to predict U; it is with the simulator Vibmas, which is based on the fundamental principles of how a blast reates vibrations. This simulator takes into account not only W but rather all the blast parameters, the type of explosive and the properties of the rock. Moreover, to establish the calculation of U as a function of S, only one seismograph measurement is required rather than two as for the USBM method. This last difference is important because it makes it easy to predict U not only as a function of S, but also as a function of the orientation of S. On the site this allows much safer control of the vibrations, and the use of larger explosive charges which reduce excavation costs. The article presents results from a field case at a hydroelectric project in Canada, where the deviation was very moderate. Use of this vibration simulator on the Web is very easy."

Jan 1, 2015

By T. Bednarski, T. McCallum, L. Moya, L. Nsenga, B. Mbao

The accumulation of solids in solvent extraction (SX) circuits can cause problems with the physical and economic performance of the operation. These solids are responsible for the formation of crud, a solid stabilized emulsion containing organic, aqueous, air, and fine solid particles. The build-up of crud in SX settlers can limit SX operation throughput, require plant downtime, lead to higher organic losses, or result in higher aqueous in organic entrainment (all of which can have a significant, negative economic impact). Current practices for crud management are typically reactive. Crud is physically removed after it has been allowed to build to excessive levels in the circuit and processed mechanically using centrifuges or filtration. A new proactive solution has been developed to prevent crud formation and build-up. Low ppm dosage of an additive, ACORGA® CR60, into the pregnant leach solution (PLS) has resulted in reduced crud formation in pilot and commercial scale for a variety of PLS streams. This paper will review results from operations in North America and Africa, covering heap leach, concentrate leach, and agitated leach feeds. INTRODUCTION Crud formation and build-up in solvent extraction (SX) circuits has been a long time topic discussed and claimed to be crucial to the whole hydrometallurgical process (Liu, et al, 2006). Crud formation is present in all SX systems be it copper, uranium, zinc, etc., and Ritcey (1980) defined crud as a stabilized emulsion containing organic, aqueous, and fine solid particles. The ore type and feed composition is a major determining factor of crud formation and typically contains Si, Al, Fe, P, and sulfates (Ritcey, 1980). It typically builds at the interface between the organic and aqueous, but can also be observed as floating crud on the surface of the settler, or bottom crud, which is common to find in circuits treating agitated leach feeds such as in Africa. As crud formation has been unavoidable, many of the current crud management techniques are reactive, allowing crud to form and build in SX circuits. Once it does, it is removed and treated via different mechanical techniques, typically through some type of filtration or three-phase centrifuge, to remove the solids and recover the organic phase trapped within the emulsion (Ritcey, 1980). Automated pumping systems have also been designed and are frequently used in North American operations. The time and capital required to react to crud formation can be an overwhelming aspect to SX operation.

Jan 1, 2019

By P. K. Hofmeyr

Base Metal (Cu, Ni, Mo, Zn, Co etc.) analysis traditionally has been very slow in conventional analytical laboratories which can lead to additional costs being incurred in waiting time of drill rigs or mining excessively off-reef. Recent developments in the automation of analytical laboratories now allow for very rapid turn-around times for base metal analytical data which can help reduce such costs as well as reduce the unit costs of the analyses themselves. 2.0 CONVENTIONAL ANALYTICAL LABORATORY PRACTICE Conventional analytical laboratories are necessarily are very manual. Sample preparation is slow and laborious, samples are moved from stage to stage manually and, typically, all samples are processed in batches so a sample which is the first to be, say, milled has to wait for the complete milling of the batch before it advances to the next stage eg. mixed with wax prior to pelletising for XRF analysis. Inevitably this results in long sample data turn-around times for laboratories handling large numbers of samples. Manual sample preparation can only be accelerated by increasing staff which increases costs and can lead to additional errors being introduced. Cleaning of sample preparation equipment, eg. crushers and mills, between samples is essential to avoid cross-contamination between samples and this is slow and labour-intensive. In manual sample preparation the correct sample numbers staying with the samples is totally dependent on the efficiency of laboratory operators. It is extremely difficult to detect when errors have occurred due to the switching of sample numbers during the sample preparation stages. Running duplicates and QC samples at best is only partially successful in detecting when sample number switching errors have occurred. Errors of this nature and cross-contamination problems due to incomplete cleaning of sample preparation equipment tend to occur when laboratories are working at pressure due to high sample numbers and when supervisors are not present during 24/7 laboratory operations which is often required on plant process control laboratories. These problems are common to all conventional analytical laboratories and their magnitude is dependent on the level of control and efficiency of QC systems practised by individual laboratories. The disadvantages of conventional manual analytical laboratories can be summarized as follows:- 1. Labour-intensive which is costly especially when multiple shifts have to be used during 24/7 operations.

Jan 1, 2005

By Roe-Hoan Yoon

Froth flotation is widely used for separating particulate materials from one another in aqueous media. The basis of this important separation process is to control the hydrophobicity of particles. However, it is difficult to predict the kinetics of flotation based on contact angle (?), which is a thermodynamic property. A solution to this difficulty would be to develop a relationship between ? and the gradient of the hydrophobic forces across the solid-liquid and bubble-liquid interfaces. Such information can be obtained using the surface force apparatus (SFA), the atomic force microscope (AFM), and the thin film balance (TFB). The present communication reviews the results of the measurements conducted using these methods, and offers a methodology for deriving a rate equation for bubble-particle interactions

Jan 1, 1999

By W. G. Davenport

Most pre-t 988 Outokumpu flash furnaces were built with blast heaters. Many continue to use them. Blast temperatures in these smelters· are typically 400 to 700 K (Table 2. 1) This chapter: (a) shows how hot blast is included in our calculations (b) evaluates the effects orhot blast onauthothermal65% Cu matte production.

Jan 1, 2001

By John M. Karpinski, David S. Cahn

For many years the green and dry compressive strengths of pellets have been considered as possible criteria of pellet performance during induration. Little attention, however, has been given to standardizing the unfired pellet compressive strength test or to determining the characteristics of this test. The purpose of this study was to compare the constant deformation rate test with constant loading rate tests. In addition, the dependence of these tests on such pellet properties as size and weight was to be investigated.

Jan 7, 1969