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By Z. T. Bieniawski

Results of in situ investigations are discussed providing comparisons of rock mass deformability measurements by three methods: 'petite sismique' technique, the Goodman borehole jack and flat jacks. The tests were performed in exploratory tunnels constructed for site characterization purposes of an underground hydro-electric scheme. Due to the large number of data (41 petite sismique measurements, 45 Goodman jack tests and 19 flat jack measurements), it was possible to evaluate thoroughly the testing techniques employed. The purpose of this paper is to indicate the problems encountered in the interpretation of the results and to offer suggestions for selecting in situ tests. INTRODUCTION In situ tests are expensive and time consuming but they are an important aid in tunneling applications since they can provide deformability data of rock masses in contrast to rock material characteristics obtainable from small laboratory specimens. Some in situ tests are relatively easy to perform but their reliability may be in question unless a comparison of the data obtained by several different test methods is possible. Although many in situ techniques are available for determination of rock mass deformability (1), those more convenient to use are the Goodman jack, flat jacks and 'petite sismique. These methods are considered convenient because the Goodman jack is a borehole device; flat jacks can provide not only deformability data but also information about the stress field around excavations; while 'petite sismique' is based on seismic refraction surveys. Accordingly, these three methods have been selected for comparison.

Jan 1, 1979

By S. S. Peng

High horizontal stress has been well recognized as a significant factor in roof control. It can adversely affect roof stability if the panel is designed improperly. Mining experience has shown that it can cause a cutter roof. The effects of high horizontal stress on roof stability depend on many factors such as magnitude and direction of the maximum principal horizontal stress, ratio of the maximum to minimum principal horizontal stress, structure and mechanical properties of the immediate roof and mining sequence. However, little is known about the intrinsic mechanism of high horizontal stress effect on roof stability except some conceptual explanations. In order to design a stable entry, the redistribution and concentration of high horizontal stress around the openings must be fully understood. Using the 3-D finite element analysis technique, this paper presents the redistribution and concentration of high horizontal stress field for a three-entry system with different orientations of horizontal stress. Based on the horizontal stress redistribution, possible roof failure for different orientations of horizontal stress is analyzed.

Jan 1, 1996

By N. L. Hayward, V. H. Ness

In the past decade Australia has seen the development, construction and commissioning of three complex processing plants for the treatment of lateritic nickel ores. These have all been based on high pressure acid leaching (HPAL) for extraction of nickel and cobalt into solution, followed by solid liquid separation in a CCD circuit. However, flowsheets for separation and purification of the nickel and cobalt diverge markedly. The existing plants have been commissioned with varying degrees of success, with a variety of problems ranging from materials of construction to inadequate design. While the process problems are not insurmountable, there is much room for improvement. Currently there are several new HPAL plants at the feasibility or financing stage. All of these have adopted one of the existing flowsheets. Most of the enhancements have focused on increasing autoclave size. It is now time to assess the success and failures of the three existing plants and incorporate this experience in a design for productivity, efficiency and reliability. Towards this end, this paper discusses aspects of flowsheet optimization, materials of construction and engineering design.

Jan 1, 2001

By Gerald C. Smith

Introduction Newmont Mining Corp.'s Carlin Gold Mining Co. mines gold ore and waste material from four open-pit mines in Northeastern Nevada. Combined production rate is 24.5 Mt/a (27 million stpy). The company is producing gold bullion from the 2.4 kt/d (2600 stpd) Carlin Mill No. 1. and the 2.3 kt/d (2600 stpd) Maggie Creek heap leaching plant. In August 1985, the 6.3 kt/d (7000 stpd) Carlin Mill No. 2 at Gold Quarry came on-line. The total combined throughput of 11.2 kt/d (12,000 stpd) will make Carlin Gold Mining one of the largest US gold producers. Feasibility studies are underway to increase the total gold out-put from Gold Quarry. This will be done by adding a continuous dump-leaching process that will recover gold from lower grade ore. History In 1961, Newmont exploration geologists located claims in the Lynn Creek mining district near Carlin, NV. An exploration drilling program revealed significant gold mineralization. A stepped-up drilling program revealed an ore body with an estimated 10 Mt (11 million st) of ore containing 109 t (3.5 million oz) of gold. Test results in 1963 allowed metallurgists to develop the flowsheet and design for a 1.8-kt/d (2000-stpd) countercurrent decantation (CCD) cyanide plant. Carlin Gold Mining Co. was formed to operate the property. Carlin purchased its first mining equipment in 1965. This consisted of two Bucyrus-Erie 2.7 m3 (3.5 cu yd) shovels and nine 27.2 t (30 st) Wabco Haulpak trucks, two Inger¬soll-Rand track-mounted drills, one Chicago Pneumatic T-650 rotary drill, one rubber tired Michigan 280 dozer, one Caterpillar D-8 36A track dozer, one Caterpillar 16 motor grader, and the necessary ancillary equipment. The fleet was capable of mining 3.2 Mt/a (3.5 mil¬lion stpy) working two eight-hour shifts, five days a week. In 1967, one P&H 1500 3.8 m3(5 cu yd) diesel electrical shovel and three 59 t (65 st) Dart trucks were added to the fleet, along with a Chicago Pneumatic rubber tire mounted rotary drill. This equipment addition increased the mining capabilities to 5.7 Mt/a (6.3 million stpy). Three separate pits, the Main, East, and West, within a 1.6-km (1-mile) radius of the Carlin mill were being mined. The crew consisted of 46 hourly men and three supervisors. This equals 497 t/ total hourly manshift (548 st) or 407 t/total manshift (515 st). Rising gold prices in the early 1970s allowed lower grade deposits to be mined at greater distances from the mill. The Boot-strap pit, 19.3 km (12 miles) north of Carlin, was started in 1973. Two Hough 45.4 t (50 st) front-end loaders, four Wabco Haulpak 45.4 t (50 st) trucks, and a second track type D-8H dozer were added. The higher grade ore was trucked to the Carlin mill using three

Jan 11, 1985

By V. A. Temeng, M. Owusu-Tweneboah, K. Awuah-Offei

The mining industry has long recognized the value of dispatch systems in open pit mines as they reduce load and haul costs. Over the years, researchers have proposed many dispatch systems with various limitations and advantages. The simplest dispatch algorithms are the so called 1-truck-for-N-shovels dispatch strategy. These algorithms are limited by the fact that their objective functions do not consider all the objectives of a mine and cannot be applied to all possible truck-shovel configurations. They are also myopic in nature. However, they are simple and computationally efficient and do not require occasional updates of the upper stage problem as required in multi-stage dispatch algorithms. In this work, an agent-based truck dispatch algorithm that conceptualizes trucks as intelligent agents that make autonomous dispatching decisions to maximize their utility is proposed. The advantages of this algorithm includes utility functions that encapsulate all of management’s objectives and agent’s with broad situational awareness. They are also more suitable for autonomous trucks. We evaluate the new algorithm against a simple 1-truck-for-N-shovels dispatch strategies using discrete event simulation. The simulation results show that the new utility function has significant advantages over 1-truck-for-N-shovels inspired utility functions. Future work will incorporate adaptive behavior into the model via reinforcement learning algorithm. INTRODUCTION In an attempt to optimize truck shovel operations in open pit mines in order to reduce the cost of haulage operation, which according to (Alarie and Gamache, 2002) is estimated to be about 50-60% of the total cost of mining, several effective and efficient strategies have been developed. Optimizing truck haulage systems in an open pit mine is one of the greatest challenges of mining engineering and has been the subject of many research projects. It is important that haulage systems are designed to be efficient, in order to minimise haulage cost, improve profitability and increase the total mine value. Although optimizing the number of trucks required in the operation of open pit mines can help reduce the cost, an effective and efficient truck dispatching system add value by ensuring that the available trucks are used efficiently. With truck dispatching systems, the trucks are normally dispatched between the crusher and dumping station (e.g. waste dumps and stockpiles), and the shovels in order to maximize the total output, obtain grades targets, reduce fuel consumptions and other objectives. Truck dispatching, which can have one or more objectives, is typically subjected to certain operational constraints such as truck and shovel capacities, empty and loaded trucks speed, among others. The problems associated with truck dispatching was first mentioned by Dantzig and Ramser (1959) after they made an attempt to determine the optimum routes of a fleet of delivery trucks between bulk terminals and a service stations. The problem occurs in several industrial systems besides mining (Alarie and Gamache, 2002; Gendreau et al., 2006; Rego and Roucariol, 1995). The dispatching problem, also referred in literature as the vehicle routing problem, has seen a number of modifications and extensions with Bertsimas and Ryzin (1991) presenting a dynamic and stochastic vehicle routing problem, and Ralphs et al. (2003) also adding to the original concept, the concept of vehicle capacity routing.

Jan 1, 2019

By J. W. Merks

At mineral processing plants the accuracy and precision characteristics of assays in mill feed, tailings and concentrate largely determine the reliability of the metallurgical balance. The accuracy of measurements is ensured by selecting effective sampling methods and efficient sample preparation procedures, and by effecting each phase without introducing a systematic error while the precision of measurements is determined by the sampling regime. Sampling systems and components, sampling methods and regimes, and sample preparation procedures will be reviewed. Different techniques to determine variance components for sampling, preparation and analysis, and to detect systematic errors in assays at mineral processing plants, will be discussed.

Jan 1, 1986

By G. Deschênes

Following a successful investigation on the Central Zone sample, a cyanidation study was conducted on an ore sample from the AD-MW Zone of Clarence Stream Property, owned by Freewest Resources Canada, to develop a leaching strategy to extract gold. At a grade of 6.48 g/t, gold is present in the forms of aurostibite, free gold and electrum. The ore contains 4.3% berthierite/gudmundite, 1.3% pyrrhotite, 0.7% pyrite, 0.07% stibnite and 0.9% arsenopyrite. Similar to the Central Zone, the extraction of gold from the AD-MW Zone can be performed at atmospheric pressure, pH 10.5 and with a low concentration of cyanide. The grind size had no effect on gold extraction between a P80 of 28 to 71 µm. The oxygen and cyanide requirement significantly increased with a ?ner grind. Consequently, a very close control of cyanide addition is required to minimize passivation of gold. Lead nitrate, added in the pretreatment, increased the extraction of gold by 2.3%. Leaching at pH < 10 improved gold extraction but increased the cyanide consumption by 5%. Passivation of gold occurred when the pretreatment was longer than 2 hours. A longer pretreatment reduced the cyanide consumption by 5%. Mineralogical characterization of the leach residue with no gravity pretreatment prior to leaching indicated the presence of unleached gold grains and passivated gold grains. Removal of coarse gold by gravity increased the extraction by 2.8% (leach residue at 0.74 g/t Au). Increasing the DO to 16 ppm or the lead nitrate from 500 to 1,000 g/t increased the overall gold extraction to 90.4% (leach residue at 0.62 g/t), which is the maximum gold extraction obtained. The leaching parameters were 800 ppm NaCN, DO 16 ppm and pH 10.5 for a duration of 120 hours. The associated cyanide consumption was 4.0 kg/t. Gold leaching was not sensitive to cyanide concentration. Reducing the cyanide concentration to 250 ppm resulted in a gold extraction of 89.5% with a cyanide consumption of 2.6 kg/t.

Jan 1, 2009

By Ray E. Harris

Introduction A resource of more than 122 Gt (134 billion st) of minable trona and mixed trona and halite underlies an area of about 3400 km2 (1300 sq miles) west of Green River, WY (Culbertson, in preparation) (Fig. 1). In 1982, the total domestic consumption of soda ash (refined trona) was 6 Mt (6.7 million st) (Kostick, 1983). About 90% of this came from Wyoming trona. This enormous deposit occurs in the Wilkins Peak Member of the Green River Formation that is of Eocene age. This formation is derived from sediments deposited in a large lake named Gosiute. Lake Gosiute occupied the Green River Basin in southwestern Wyoming some 38 to 54 million years ago. The Wilkins Peak Member represents a restricted, evaporite-rich cycle of deposition when the trona deposits were formed. A total of 42 trona beds are known. Of these, 25 are of sufficient thickness and extent to be considered minable (Culbertson, in preparation). The production of sodium compounds from the trona deposits in the Wilkins Peak Member began in 1896 when wells were drilled within the present boundaries of the town of Green River. The purpose of the drilling was for extraction of sodium carbonate brines. Sodium carbonate (sal soda) was removed from these brines at a plant in Green River. Bedded trona was discovered in 1938 in a core from an oil well drilled south of Westvaco. In 1947, the underground mining of trona began. Production has continued from 1948 to date. Currently, five companies are mining trona and producing soda ash from the Wilkins Peak Member in Wyoming. Additionally, Allied Chemical Co. mines and processes trona for Church and Dwight&apos;s sodium bicarbonate (baking soda) operation in Green River. Competition in the soda ash industry depends primarily on the efficiency of mining and refining, allowing for lower-cost production. Newer plants and mine designs should increase efficiency. Solution mining holds promise as an even lower-cost method of mining. Employment in the soda ash industry is declining since the newer mining and refining methods are less labor intensive. Declines in employment, however, may be offset as production increases. Due to the large resource, the availability of minable trona should not be a problem for perhaps hundreds of years. Geology of Wyoming trona deposits Forty-two beds of trona or mixed trona and halite occur in the

Jan 10, 1985

By Marco Aurelio Piangatelli, Massimiliano Bringiotti, Ferdinando De Angelis

"The construction of the three connecting tunnels between the east side and the west side of station C13, adjacent to the Vistula river, has a long and complex history, as they cross beneath one of Warsaw’s main road arteries (Wislostrada). After the attempt to excavate the first tunnel, the face of the tunnel collapsed on the night of August 13, 2012, and a volume of about 10,000 cubic meters of liquid mass (water, silty soil and sand) entered and flooded the west side of the station, which was filled with more than seven meters of material. Fortunately, the workers managed to escape, but machines and equipment were lost. This is the story of how the tunnels were completed, by adopting soil improvement, jet grouting, GRP bars, ground freezing and a careful choice of work steps. INTRODUCTION The Centrum Nauki Kopernik station on Line II of the Warsaw Underground is located alongside to the Vistula river (Figure 1). This station hereinafter is referred to as station C13.The station consist of three parts: the east and west part, and three connecting tunnels 40 meters in length with a large cross section (170 m2).. It includes the rails along the sides and the platform along the centre. The reason for the underground connection of two shafts with a tunnel is due to the presence of a road tunnel (Wislostrada, hereinafter called WS), which must be underpassed by the new metro line (Figure 2).The existing tunnel (Wislostrada) was built many years before, between diaphragm walls, and some foundation structures interfere with the underground tunnel. The challenge was to build the connectors without stopping the traffic of vehicles along the highway above. All works took place in very difficult geological and groundwater conditions due to the presence of Vistula alluvial deposits with poor mechanical characteristics. The soil profile consists of relatively loose sands on top, and silts overlying high plasticity Pliocene clays highly disturbed by the effects of the repeated over-sliding of glacial moraine deposits."

Jan 1, 2016

By I. M. Morsi, S. Z. El-Tawil, O. A. Fouad, K. A. El-Barawy, M. B. Morsi, M. M. Nasr

Upgrading of a low-grade titanium slag (~55% TiO2) with high silica, calcia, magnesia and iron contents has been studied using sulphation roasting technique using SO2/air gas mixture followed by hydrochloric acid leaching. XRD analysis of the sulphated slag reveals the presence of kennedyite (Fe2O3. MgO. 3TiO2), rutile (TiO2), calcium sulphate (CaSO4) and calcium silicate (3CaO.SiO2) phases. The residue after leaching contains mainly kennedyite, rutile phases and free silica with traces of calcium silicate phase. Synthetic rutile that has about 83% TiO2 can be produced at the optimum conditions. Thermodynamic calculations give rational explanations of the obtained results.

Jan 1, 2005

By F. Dehghani, M. Kebe, H. Askari-Sabzkoohi, T. Ghosh

The depletion of bauxite reserves around the world has necessitated the search for viable alternative sources of aluminum. The major non-bauxite source of aluminum is alunite [KAl3 (SO4)2(OH)6]. Lixiviation is the common industrial process to recover aluminum from alunite. Due to the presence of quartz (40%) as the main gangue mineral, this process causes the formation of Sodium aluminosilicate and Potassium aluminosilicate complexes that leads to loss of alunite, and chemical reagents. Thus, pre-concentration of alunite is required to remove the gangue minerals. Because of the low degrees of freedom (DOF) of alunite, froth flotation, as a preconcentration method, is a viable alternative to increase selectivity in downstream processes. In this work, the possibility of Alunite preconcentration with micro flotation and mechanical flotation cells was studied. The influence of type and dosage of collector, pH, solid percent, particles sizes were considered on the flotation. Flotation tests were designed based on Taguchi method, followed by performing based on surface charge differential between alunite and quartz on direct and revers flotation. In the reverse flotation, Ca+ 2 and Ba+ 2 were used to activate the quartz surface. Under optimum conditions of reverse flotation, alunite grade and recovery were 40% and 62.1% respectively. In the direct flotation, the surface of alunite was activated by F- . The grade and recovery were 50 and 89%, respectively.

Mar 2, 2022

By R. Schipper

Two bench scale biochemical reactors (BCRs) were operated over a period of 3 months for the removal of uranium, radium, and selenium from mining influenced water (MIW). A BCR is an anaerobic reactor comprised of organic and inorganic substrate including woodchips, straw, sawdust, and limestone. Uranium and selenium are removed in a BCR by biological reduction of oxidized species (i.e., selenite, selenate) to less soluble forms (Pahler et al. 2007, Blumenstein et al. 2008, Rutkowski et al. 2013). It is believed that radium may be removed through absorption to the biomass in the BCR. The bench scale BCRs were constructed of 6 inch PVC columns with a top to bottom flow configuration. The bench influent water consisted of three different water ratios from two adit seeps. BCR hydraulic retention times of 24 to 36 hours were targeted during the study. Influent concentrations ranged from 19 ug/L to 1.86 mg/L for dissolved uranium, 12 ug/L to 270 ug/L for dissolved selenium, and 4.9 pCi/L to 24 pCi/L for radium 226. The average and maximum dissolved uranium removal was 93.6% and 98.5% respectively, the average and maximum dissolved selenium removal was 93.3% and 98.9% respectively, and the average and maximum radium 226 removal was 89.5% and 100% respectively.

Feb 23, 2014

By Donovan F. Symonds

INTRODUCTION The selection of the optimum coal preparation plant flowsheet is a vital part of the development of a new plant or the expansion of an existing facility. Selection of an inappropriate flowsheet or individual equipment items can result in costly production losses, or needless expenditures or redundant or oversized equipment. The accuracy of the flowsheet selection process is, like most decisions, heavily dependent upon the amount, reliability, and quality of the database. A flowsheet which is based on simple borehole information from an area that has experienced little or no previous mining activity must necessarily include significant allowances for uncertainties. However, a plant which is to be built at an existing mine with good information on run-of-mine (ROM) coal quality, as well as bulk washability data and accurate projections of future quality, can be optimized. The following chapter will describe the processes which should be followed to select the most appropriate flowsheet for each unique operation. They include: 1. Examining and adjusting the washability database to reflect the ROM feed. 2. An evaluation of clean coal specifications. 3. Selection of the level of washing required. This includes no washing, if the coal is already at the desired specification, all the way through to complex total washing circuits for metallurgical grade coals. 4. Selection of the appropriate circuit such as dense medium bath vs. Baum jig. The most appropriate circuit is generally determined on the basis of economics. An approach for performing the evaluation will be desired. 5. The steps that need to be followed to develop a balanced plant flowsheet. ADJUSTING WASHABILITY DATA It is not unusual to have as many as 10 to 20 sets of washability information available for a particular mining property. These may be a combination of core and bulk samples or all bulk samples. In the latter case, they may also be face and/or ROM samples. It is necessary to reduce all of this information to a single set of washability data that is representative of the mean feed to the proposed preparation plant. This set is used to compare the clean coal yield and quality characteristics of the various flowsheet options. The term "representative" has some serious negative connotations. It is recognized that ROM coal quality will vary throughout a mine. However, over a period of 250 to 350 days per year, certain trends can be established. If possible, this "representative" sample should be close to the coal quality of the mean daily sample in terms of raw ash and sulfur contents, clean coal yield, and clean coal ash and sulfur content. If possible, this sample should be within the

Jan 1, 1991

By C. G. Anderson

Today, with a stringent economic and environmental climate prevailing in the copper business, there is increased interest in evaluating new processing alternatives for production. Hydrometallurgical pressure oxidation of copper concentrates is one of the more viable approaches and several technological candidates have emerged. Of these, an overlooked but, ironically the first industrially proven methodology, utilized nitrogen species catalyzation in the oxidizing pressure leach system to produce copper via SX/EW. Given its many inherent process advantages, this may prove to be a feasible process alternative for the future. In this paper, the history of the system and its application to copper concentrates and ores will be outlined. In particular, a novel methodology for effective recovery of precious metals from chalcopyrite concentrates will be discussed. Finally, the perceived economics of chalcopyrite concentrate treatment with this unique industrially proven process will be delineated.

Jan 1, 2001

By Leon Maday, Robert Robinson, Judy Cochran, Don Davis

Over the past 20 years, several soft-ground tunnels have been successfully completed in the Seattle area. These tunnels were built using design-bid-build contracting methods, and have benefited from advancing technologies in exploration, design and construction methods. Owners have utilized various contractual means to share some of the risks that are inherent to tunneling projects. This has helped to foster a competitive bidding environment and to resolve issues related to unanticipated underground conditions in a fair manner. This paper presents case histories of tunnels constructed by King County and the Washington Dept. of Transportation that illustrate a common approach to design and construction issues. The common thread is successful project completion with minor cost changes under the design-bid-build contracting method. Lessons learned from these projects are being applied to the upcoming King County Brightwater Project tunnels, Sound Transit’s Light “Link” Rail tunnels and other tunnel projects.

Jan 1, 2005

By Charles D. Litton, Ronald S. Conti

This report describes the results of research conducted in the Bureau of Mines Experimental Mine at Lake Lynn Laboratory to determine the effects of air velocity, sensor spacing, and stratification of CO levels on the detection of slowly-developing coal/conveyor belt fires. In general, the time to detect a developing fire is calculated assuming an average CO level that has mixed completely with the ventilated airflow. The data indicate that CO sensors placed near the roof of an entry take advantage of buoyant- induced stratification of combustion products, resulting in earlier detection of developing fires. Even during the low-temperature, smoldering stage of a fire, the combustion products will tend to stratify near the roof, and once flaming occurs, the degree of stratification increases markedly. As one would expect, at higher air velocities, the degree of stratification is less than at lower air velocities, although still significant, especially during the flaming stages of the fire

Jan 1, 1993

By B. Jensen, B. K. Abraham, G. J. Tuck, A. D. S. Gillies

A dust concentration meter has been developed with Australian coal industry support for the measurement of total airborne dust concentrations capable of causing explosions in underground coal mines. The instrument, The University of Queensland Dust Meter (UQDM), measures the effective dielectric constant of a mixture of coal dust and air that is drawn between the plates of a capacitor, the capacitance of which is measured by a ratio-transformer bridge and synchronous detector system. This technique measures small changes in capacitance to approximately 1 part in 106. The unique design of the capacitor sensor permits a precise volume of coal dust and air mixture to be sampled. This arrangement overcomes difficulties inherent in optical dust measurement. The meter is portable, battery powered, incorporates data logging and has intrinsic safety approval. The operation of the meter is discussed. The system is designed to monitor dust concentrations along a coal mining face and provide concentration profiles leading to the development of suppression strategies. If data from underground tests are available, the meter&apos;s potential to contribute to establishing safe working conditions underground will be discussed. Other areas of potential application of the measurement technique are also examined.

Jan 1, 1993

By B. K. McMahon

Block caving is an efficient mass-production method for the underground mining of large orebodies. However, the block-caving method requires a substantial investment in developmental works, based on the assumption that the rock will cave. Efficient block caving is prevented if the rock forms stable arches or breaks into fragments that are too large to pass easily through the finger raises. Under these conditions, the cost of the additional labor, drilling, and blasting required to assist caving may be sufficient to make marginal orebodies uneconomic or alternative mining methods more attractive. Reliable methods for the prediction of the block-caving behavior of an orebody are clearly needed at the initial planning stage. These methods should be capable of considering both geological and engineering aspects of a block cave, as experience has shown that the manner in which a block cave is initiated and developed may be almost as important as the geological conditions such as the nature of the rock material, degree of alteration, characteristics of the joints and faults, etc. The initial state of stress in the rock mass should also be considered.

Jan 1, 1969

By D. M. Steelman

Digital systems have been employed in industry as an aid in process evaluation and control for about 20 years. The earlier systems were quite simple and did little more than periodically scan a multitude of process variables, perhaps comparing them to alarm limits, and print out the results. As time passed, the needs of industry and the imagination of process engineers became more demanding. Rather than perform only alarm checking and logging, operating guides were generated to help the operator gain optimum performance from the process. Eventually, systems were entrusted with the job of manipulating the set points of the local analog control loops in a (DDAC) digitally directed analog control mode. Today digital systems are used to gather all-the process information, including both analog and digital variables; to make the necessary data manipulations; and to control the process end elements in a (DDC) direct digital control mode. During these years of transition from the very simplest systems to the complex systems of today, the demands upon the system engineer have been every increasing. The process engineer has required, among other things, more computational ability at high speeds, faster input and output rates, greater flexibility, better communication facilities (man-machine interface), (process oriented software), and finally, greater reliability at lower cost.

Jan 1, 1969

By Lance Kolbush, Alex Vaughn, James Fisher, Michael Barish, Tony Toney

Somerset International has worked to enhance the recovery of sub-20 micron particles in our patented Sub325 Systems by introducing chemicals in the systems prior to the centrifuge. This paper details four cases of such enhancement. Two on the fine coal recovery side at a 325 tph plant in Central PA and Warrior Met Coal 7 plant. Two on the tailings dewatering side at a 1200 tph Southern WV plant and at Warrior Met Coal 7 plant.

Feb 22, 2026