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By Robert R. Dorfler

We ought not look back unless it is to de- rive from past errors, and for the purpose of profiting by dear bought experience. - George Washington During the mid 1960's, the application of automation was introduced to Climax by two process control companies. In 1969 Climax decided to seriously consider improved control of selected areas of their milling process after preliminary research had been completed on two rougher ball mill grinding circuits. In 1971 the Climax Mill Department developed a direct digital control scheme for rougher milling, designed a remote control center for handling process water, planned to automate the central steam heating facility for energy conservation, and convert other plant areas to automation. All twenty rougher grinding mills and the industrial water system were automated by 1977. The idea of automating central heating and other areas was abandoned although some fuel conservation items were installed at the heating plant. During this period from 1971 to 1978, the mill capacity was expanded from 39,000 metric tons per day to 43,000 metric tons per day and cyclone classification was introduced to rougher milling. Ore quality changes degraded the milling recovery from over 90% to less than 85%. This report describes the procedure the mill used in developing the automatic process and the occurances that affected the timing and fina1 control logic. The report deals primarily with the automation of the ball mill grinding, classification and flotation process as this was the bulk of the seven-year project.

Jan 1, 1979

By Hubert Erik Sherer

Determination of the best time to replace an aging unit of equipment is a recurring problem which is common to essentially all surface and underground mining operations. Virtually all mining equipment is subject to wear and the resulting deterioration that increases with equipment life. This mechanical deterioration results in lower availability and, thus, decreasing, production. Due to decreasing production and increasing repair costs, there is an optimum age at which the replacement of a given equipment unit with a new unit will result in the lowest cost per unit of production for the equipment operation in question (Figure 1). Determination of this optimum replacement age is further complicated by factors such as: possible overhaul alternatives, fleet replacement and expansion alternatives, remaining life of the mine, and technological improvements in the potential replacement or differences in the production capacities of these replacements. Determination of the optimum time to replace an equipment unit becomes more critical each successive year due to the rising capital costs of new equipment versus maintaining old equipment. An incorrect choice of the replacement age for an equipment unit can result in excessive costs to the mining operation. If a unit is replaced too early, the capital cost of the unit is not fully utilized; conversely, replacing a unit too late results in excessive maintenance costs and low production. The economic life of a piece of equipment is the equipment age that results in the lowest total discounted cost per unit of production for that piece of equipment, not considering the cost of future replacements. The economic life of a piece of equipment is usually shorter than the physical life, the maximum age at which a machine is capable of doing useful work. The optimum replacement age for a piece of mining equipment may vary from the economic life of that equipment due to the remaining life of the mine, fleet considerations, costs associated with potential replacement machines, and capital rationing constraints. For example, the equipment life of the current machine that may result in the lowest cost per unit of production might be 10 years; however, when the timing of the future replacement costs are entered into the calculation, a different life for the current machine might result in the lowest total cost per unit produced for the remaining life of the mine.

Jan 1, 1982

By R. R. Young

Prior to 1903, the world's supply of sulphur was primarily from surface deposits in Sicily. The discovery of the first salt dome sulphur deposit in 1867 and successful development of the Frasch process for recovering the sulphur resulted in a large part of this essential element being supplied from these mines since 1920. Although the basic process is the same today, engineering and technological developments have evolved to allow its application under many adverse conditions. These and other challenges for the industry in today's environment will be discussed in the paper.

Jan 1, 1979

By John A. Sievert

Sedimentary deposits contain an appreciable portion of the world's supply of uranium. Predictions of nuclear fuel usage in the United States indicate that these reserves will have to be tapped if the domestic industry is to be the major supplier of fuels for the nuclear reactors. Techniques developed within the petroleum industry are a base on which to build economically attractive fluid mining processes for the recovery of these uranium values. The problems central to the success of such mining processes are the containment, control, and recovery of a narrow band of pregnant leach solution as it progresses from the point of injection to the point of production Given sufficient incentive, suitable solutions to these problems can be found

Jan 1, 1970

By Joel Nonnweiler, David Young, John Hawley

The Sacramento Regional County Sanitation District has implemented a billion-dollar interceptor system master plan, including the 31 km (19 mile) long Lower Northwest Interceptor (LNWI). For the two LNWI Sacramento River crossings, pipe-jack/micro tunnel drives in excess of 300 meters (1,000 ft) from 30 meter (100 ft) deep shafts in mixed soil conditions below the water table were considered, as were600 meter (2,000 ft) tunnel drives utilizing pressurized-face tunnel boring machines. Operation and maintenance concerns, predicted flood scour, and levee concerns are discussed. A risk assessment process was used to select both profile and construction method. Construction will begin in mid 2004.

Jan 1, 2003

By David K. Ingram

The objective of this U.S. Bureau of Mines investigation was to characterize overburden response due to longwall mining. Subsurface strata and surface deformations were monitored during the mining of two adjacent, 900 ft wide longwall panels in southeastern Ohio. The panels were separated by a five entry, four pillar system (about 350 ft wide). The overburden thickness in the study area ranged from 214 to 280 ft, and the extraction thickness averaged 70 in. To delineate the geology of the overburden, six coreholes were drilled along a profile line extending across the two longwall panels. Multi-point borehole extensometers (MPBX) with eight-anchors were installed in six boreholes to measure subsurface displacements. The MPBX units were continuously monitored using a data logger. During the mining of the first panel, displacement occurred in a radiating pattern beginning just above the extracted area in the center of the panel and extending upward and outward towards the surface and margins of the panel. Overburden deformation did not project over the gate road entries. During the mining of the second panel, the MPBX's showed almost simultaneous displacement, indicating that the overburden began to move as a whole unit. Again, overburden deformation was confined to the immediate panel area. At both panels, overburden movement began when the longwall face was approximately 55 ft beyond the line of MPBX boreholes. Moreover, all ground movement was complete when the longwall face was approximately 200 ft beyond the line of MPBX boreholes. This paper discusses the results of the study and identifies the characteristics of overburden deformation.

Jan 1, 1986

By M. Zhang, H. Ni, F. Zheng, M. Guo

"In the present work, we investigate the effects of roasting temperature and time on the transformation of the titanium (Ti)-bearing phase and magnesium aluminate spinel (MgAl2O4) phase in Ti-bearing electric arc furnace molten slag during the molten sodium hydroxide (NaOH) treatment process. With applied temperature of 700 °C for 60 min, and a mass ratio of Ti-bearing electric arc furnace slag to NaOH, or Rslag/NaOH, of 1:1.2, the conversion ratios of the Ti-bearing phase and the MgAl2O4 phase reached about 98.3 percent and 88.1 percent, respectively. Moreover, investigation of the reaction kinetics indicate that the conversion of these phases during the molten NaOH treatment process is a typical solid-liquid reaction, and it can be theoretically described by the unreacted shrinking core model. The conversion ratio of the Ti-bearing phase was found to be controlled by the chemical reaction step, with an apparent activation energy of 41.76 kJ/mol. The decomposition of the MgAl2O4 phase was found to be controlled by the surface chemical reaction in the early stage of 0 to 20 min from start of reaction, changing to diffusion control in the later stage of 20 to 60 min. The apparent activation energies of the process were determined as 55.88 and 22.72 kJ/mol, respectively.IntroductionChina has rich mineral resources of vanadium–titanium magnetite ore deposits, and the iron and steel industry in the country produces millions of tons of titanium (Ti)-bearing slag every year. The problem of how to dispose of the Ti-bearing slag has attracted much attention, and various approaches have been developed, including the acid process (Li et al., 2007; Zhang et al., 2011; Zhang and Nicol, 2010; Li et al., 2011), alkali fusion process (Han et al., 2012; Middlemas, Fang and Fan, 2013; Nayl, Awwad and Aly, 2009; Lasheen, 2008; Zhang, Qi and Zhang, 2009), carbonization and chlorination process (Liu et al., 2008; Tao et al., 2012; Fu et al., 2012; Liu, 2009), selective crystallization process (Lei and Xue, 2008; Li et al., 2011; Zhang et al., 2007a, 2007b). However, these processes have been beset by various functional, economical and ecological problems (Yu and Chen, 2010). With environmental protection regulations becoming increasingly rigorous, and demand growing for the comprehensive utilization of resources, the focus is on methods to treat Ti-bearing slag with high efficiency and mild reaction conditions (Nazari and Riahi, 2011; Wu et al., 2013)."

Jan 1, 2017

By J. P. Vandenhoeck

Sizing devices used commercially are unfortunately not 100% efficient, the splits that can be obtained are usually inferior to the results obtained in the laboratory. Often the capital and operating costs of equipment capable of high sizing efficiency cannot be justified by the economics or requirements of the process. To reconcile the desired sizing efficiency needed to obtain optimum process requirements with the sizing efficiency of equipment economically practical, it is often necessary or desirable to predict the gradation and weight splits of the products obtained based on the expectant efficiency of the contemplated sizing device or method. The less efficient the sizing device one must employ, the more important it is to know the results that may be expected for the establishment of the optimum flow sheet and the more difficult it becomes to estimate the results accurately. Air classification equipment which sizes materials by opposing the drag force created by an air stream acting on a particle with a combination of gravitational, inertial or centrifugal forces are substantially less efficient than, for example, a laboratory sieve analysis. Air classifiers allow oversize material to contaminate the fine fraction and leave undersize material in the coarse fraction. The techniques employed to predict the gradation and weight splits obtained from air classification outlined below are adaptable to other sizing devices less than 100% efficient such as screens, sifters and wet classifiers.

Jan 1, 1965

By Ben H. Daud

The systems optimization approach to mine planning has led to the adoption of a number of mathematical and computer models for short, medium and long range planning of mining operations. In the realm of open pit mine planning, the most common uses of such an approach have been in the areas of optimal pit design and planning of materials handling operations. Since a truck-shovel combination is the most widely used method of materials handling in open pit mines, a number of models have been presented in the past to analyze and optimize such systems. Broadly, these models can be divided into two categories. The simulation model, which is primarily used as an aid in analyzing the performance of either an operating or a proposed haulage system under the influence of externally alterable parameters, and the linear or integer model which is used as a tool for short or medium term optimization of the system under the operating constraints. In the first part of this paper, a brief discussion is presented of the linear/integer optimization models presented in the past, and in the second part a new model is described which is actually a combination of the simulation and the integer programming models. In its bare essentials, the model consists of an integer optimization algorithm, but the constraints are regularly and dynamically updated by incorporating the conclusions drawn from a stochastic simulation model of the mine haulage system. Since the successive optimization is performed by the mine operator who manipulates the constraints to guide the model towards the optimum, the mystical black box approach is minimized and should help in adoption of the model by the mine personnel. Also, by combining the stochastic simulation model and the deterministic integer model, some of the advantages of each are acquired.

Jan 1, 1976

By L. A. Garfield

Since the start of mining operations at White Pine in 1953, over 22 miles of drifts have been driven for the purpose of developing the ore body. The drifts, normally driven in multiples of three or five headings, serve as conveyor belt routes, ventilation ducts, equipment and haulage roads, and initial access to mining blocks. Development openings in the flat-lying, bedded ore body are driven in rectangular section from 16 to 32 feet wide and from 8. 5 to 16 feet high. The floor level normally coincides with the bottom limit of the chalcocite values and the roof or back is carried at the most competent stratum adjacent to the top of the ore. A development front consisting of five headings and associated crosscuts is advanced at an average rate of about 100 feet per month using conventional drill and blast techniques and equipment. The cyclic nature of the functions involved in driving the openings and the need to include them in production schedules preclude significant improvements in advance rates using current drill, blast, muck, and haul sequencing.

Jan 1, 1969

By Daniel C. McLean

Chemistry plays a more important role in milling operations than is generally recognized, particularly in fluid- solid separation processes. Chemical aspects are also very important insofar as plant water quality is concerned, and in the manner in which this water quality effects plant design and operation. Aside from the inherent chemical characteristics of plant water, the mineral preparation industry is becoming increasingly more dependent on the use of mining chemical reagents--relatively new to the coal industry, but items of I standard usage in complex metal recovery plants. These chemicals have many useful properties and can produce significant improvements in operations and metallurgical efficiencies when used properly and intelligently. They can produce poor results and sometimes severe losses when used improperly. Most reagents are relatively expensive on a per pound basis, hence costs can be excessive if they are used wastefully or inefficiently.

Jan 1, 1968

By W. L. Price

The intelligent selection of wire cloth and perforated plate for screening surfaces seems a rather confusing subject to the newcomer in aggregate processing. The proper application of various screening surfaces is a complicated one and one could consider himself an expert only after many years of working with all applications. However, part of the subject matter could be simplified considerably for, the average screen user. In this paper an attempt will be made to do this, by presenting some comparative specification data along with cost figures Unfortunately, it is impossible in the limited scope of this paper to give actual service comparison records between different cloths or cloth and plate. Anyone who has considered making actual service comparisons for competing screening media is well aware of the almost insurmountable problems Their solution would form the basis for a separate paper on that subject alone. Therefore, this paper will be limited to comparative data up to the point of installation.

Jan 1, 1961

By R. Jacobson

This paper presents comparison of proposed future solution and in-situ mining methods for in-situ leaching and recovery of uranium, gold, silver, and copper. We propose to show that in the future new extraction methods for each metal will eventually evolve for each hydrometallurgical process step, i.e. in-situ leaching, recovery, and concentration; precipitation and reduction; and drying, when required--without generating waste or bleed streams. In the future, to be highly profitable, these forms of mining will develop into an integrated system to eliminate waste streams using safe, powerful lixiviants. This integration will include, with some exceptions, improvements in the following areas of process development: 1. Lixiviant selection and in-situ leaching technique, 2. Target metal recovery and concentration (at a surface hydrometallurgical facility) from the circulating pregnant lixiviant stream, 3. Target metal precipitation or reduction, where decant water from the precipitate dewatering step is reusable for resin or solvent strip liquor (eluant), 4. Final preparation of target metal for sales, i.e. drying, purification, casting, packaging, 5. Restoration where most of the groundwater restoration process is performed at the same rate as multi-wellfield mining.

Jan 1, 1987

By J. R. Knudson

A major contribution to the art of blasting was made at the Cleveland- Cliffs Iron Company?s Hawkins Mine in Cashwauk, Minnesota during April of 1956. It was here that the first experiments with fertilizer grade ammonium nitrate were conducted-experiments which led to the development of the ammonium nitrate and fuel oil method of blasting which has been eagerly adopted by open pit and quarry operators throughout the United States. While this new blasting agent was a more or less natural outgrowth of the Akremite Process devised by Maumee Colleries of Terre Haute, Indiana; no one, to our knowledge, had used this particular combination of ingredients prior to this time. Certainly this was a "break-through" as far as the Lake Superior District was concerned. Akremite consists of a mechanical mixture of grained ammonium nitrate and a solid carbonaceous material packed in a plastic bag. At the time of its introduction to the Mesaba Range, Akremite sold for 10.5 cents a pound as compared to the 18 cent powder previously used in both wet and dry large hole blasting. The discovery of this agent was indeed a remarkable achievement, one that afforded a substantial reduction in blasting costs. However, Akremite proved to be only an intermediate step in the evolution of our present low cost explosives. Cleveland-Cliffs utilized prilled ammonium nitrate and fuel oil and then eliminated the mixing and packaging plant entirely by combining the two raw materials in the field at the drill hole collar or in the hole itself. Thus we have approached the ultimate in simplicity, for the blaster and his helper now manufacture their own explosives not only as they are needed but where they are needed, and at a cost of under 4 cents a pound. This two-man crew may charge and explode as much as 8 tons of ammonium nitrate in an eight-hour shift and in so doing break some 65,000 tons of ore.

Jan 1, 1959

A Mark IV surface force apparatus (SF A) was used to conduct direct force measurements in dodecylamine hydrogenchloride collector solution between mica surfaces at varying pHs. It has been found that the long-range hydrophobic forces reach a maximum at pH 9.5, which also corresponds to the maxima attained by the contact angle and flotation recovery of quartz. The charge densities are comparable on both mica and quartz surfaces (0.40-0.48 run2), making it possible to draw conclusions on the flotation of quartz from the direct force measurements conducted on mica. Therefore, it may be suggested that the maximum flotation recovery of quartz m the range of pH 8.0-10.3 may be attributed to the appearance of long-range hydrophobic force, which in turn can be related to the formation of a close-packed monolayer (or hemimicelle) in the presence of neutral amines formed as a result of hydrolysis.

Jan 1, 1996

By M. A. Alan

The Bellavista Gold Deposit lies 75 Kms north west of San Jose, the capital of Costa Rica, Central America and 3 Kms north east of Miramar, in the Province of Puntarenas. The total road distance from San Jose to the deposit is 120 Kms all paved road except for a 2.5 Km graded, gravel road on the property (Figure 1). Infrastructure services in the Miramar-Puntarenas area are well developed with electrical power distribution, a modern port facility and community services being available. The deposit lies on the south west flank of the Tilaran Mountain Range between an elevation of 550 and 800 meters above sea level. The local topography is steep, originally covered with pre-montaine, tropical wet forest vegetation which has been largely destroyed due to clearing for cattle grazing. The annual rainfall is approximately 2,950 mms, 90% of which occurs between May and October.

Jan 1, 1992

By Thys B. Johnson

This paper presents a new algorithm for solving the Controlled Flow Mine Ventilation Problem in an optimal manner - minimum power consumption. The use of booster fans and restrictions on the locations of ventilation regulators are dealt with in a straightforward manner. This new approach is based on a common Operations Research algorithm for which there are many fast computer programs available.

Jan 1, 1982

By Dongxu Jia, Chen Cao, Shuangwen Ma, Jun Han, Huibin Ma

Coal bump, or coal burst, is one of the catastrophic disasters in underground coalmine. Among them, approximate 80% of recorded destructive burst events were occurred in mining roadway. The bump panel is often formed by hard coal seam and hard main roof; however, the cause and course of the bump events are not understood yet. This study provides a fundamental theory for the causality and progress of a coal bump event occurred in longwall roadway. Considering hard roof condition commonly existed in bump panels, we think that the abutment pressure induced by suspended roof behind the working face is the trigger of a coal bump event occurred in the roadway. Once high abutment pressure is concentrated on the hard coal seam around the working face, the roof and the coal around the working face form a seesaw-like structure, which leads the roof in front of the working face moving upwards. The movement of roof uplifting in front of the working face is a well evidenced phenomenon in practice, namely, roof rebounding, which causes unloading of the normal stress on the roof-coal interface around the roadway. A rib coal block bump mechanical model has been developed to establish the critical equilibrium equation of the coal block. Result shows that, the force source of coal bump is the horizontal ground stress perpendicular to the rib surface; and the slippage of the coal block is caused by the unloading of the normal stress acting on the roof-coal interface. Based on the causality and consequence theory of coal bump developed in the paper, several new prediction and prevention measures have been developed, including roadway design and in-tunnel reinforcing methods. The commonly used prevention measures in present have also been discussed. This paper presents a comprehensive explanation of coal bumps occurred in longwall roadways, from abutment pressure at working face forming seesaw structure to roof rebounding leading slippage of the rib coal mass into roadway, which provides a fundamental theory for coal bump causality mechanism and prediction and prevention technology.

Apr 26, 2022

By Harry M. Parker

Exploration for sandstone-type uranium ore bodies which will be exploited using underground mining techniques is usually executed in sequential fashion. Geostatistical methods can be used at the earliest exploration stages to design drilling programs which have a high probability of success in the location and delineation of mineable ore-grade mineralization, should such mineralization exist. In succeeding development stages, geostatistics may be used to outline and define the potential grade of reserves and to apply confidence limits on the grade estimation. Delineation of mineable reserves for mine planning requires extensive development drilling; using such drilling data geostatistical methods may be used to minimize the drilling needed to achieve a desired level of confidence in the calculation of mineable uranium ore grade and tonnage. Because sandstone-type uranium deposits tend to be pod-shaped and are generally aligned along a mineralized trend, this paper proposes a geostatistical method of calculating uranium ore reserves which accounts for this fact and predicts the area within a trend which is underlain by ore. This method uses an indicator value; that is, drill holes which contain mineable ore-grade intercepts are coded as 1, those that do not are coded 0. The indicator may be kriged over the mineralized trend to obtain an estimate of the area underlain by ore. The thickness and grade-thickness may be kriged, and the average ore grade may then be obtained by dividing grade-thickness by thickness, as is conventional for stratabound deposits. The mean grade is then multiplied by the kriged mean thickness to obtain an estimate of the volume, and thence tonnage, of mineable ore within the trend.

Jan 1, 1979

By C. D. Hebert, T. Gregor, M. Ghazi, G. J. E. Kramer

This paper describes the recent application of gasketed, precast concrete segmental lining design that has improved tunneling productivity and quality. The speci?c segment arrangement includes: six-piece ring with equal 60 degree gasketed segments, juxtaposed key and counter-key segments, trapezoidal and parallelogram shaped segments, inclined and skewed radial joints and plastic circumferential joint connectors. This allows rapid, but accurate ring erection, reducing mine cycle times. The design avoids cruciform joints, minimizes potential gasket damage and eliminates the need for separate left and right ring molds. The implementation on the Los Angeles ECIS and NEIS sewer projects is described.

Jan 1, 2003