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By G. D. Hood

The U.S. Bureau of Mines investigated rapid froth flotation through the use of more efficient bubble-particle collisions that improved the overall flotation kinetics. An in-line static mixer efficiently mixed the air bubbles, water, and ore, rapidly attaching the hydrophobic mineral to the bubble surface. As the mixture entered the flotation chamber, the hydrophobic mineral was already attached to the bubbles and was quickly floated away from the remaining ore pulp. A rougher concentrate containing 4.5 pct copper was obtained from a western porphyry copper ore containing 0.67 pct copper. Over 99 pct copper was recovered from the ore at a rate 7 times faster than conventional flotation.

Jan 1, 1991

By John S. Egbert

Many of you are familiar with the Washington Metro Project. However, inasmuch as some may not be, and others may want to refresh their memories regarding certain facets of the project, I will cite a few summary details about the rail transit system we are constructing. That summary will also give each of us a common point for departure as I develop the principal thoughts that I want to leave with you. The Washington Metropolitan Area Transit Authority (WMATA) is an interstate agency established by Compact between the District of Columbia, Maryland and Virginia and approved by the United States Congress. Currently, the Authority has approval to build and operate a 101-mile (162.5 km) regional rapid rail transit system as shown on Appendix A. Of that total, 38 miles (61.2 km) will be in D. C., 32 miles (51.5 km) in Maryland and 30 miles (48.3 km) in Virginia. (Mileage figures, because of rounding error, do not add to 101 miles.) Referring again to the 101-mile total, 48 miles (77.2 km) will be underground, 42 miles (67.6 km) on the surface and 10 miles (16.1 km) on aerial structures. Of the 48 miles (77.2 km) that are underground, 15 miles (24.1 km) are rock tunnel, 14 miles (22.5 km) are earth tunnel, 0.5 miles (0.8 km) 9.s sunken tube and 19 miles (30.6 inn) are cut-and-cover construction. The system will include 86 stations, with 51 of them underground. Referring again to the full 86 stations, 43 will be in D. C., 23 in Maryland and 20 in Virginia. We currently have 37 miles (59.5 km) of line and 41 stations in operations as shown on Appendix B. Also shown on Appendix B is another 16 mile (25.7 km) increment and 13 stations which have been substantially completed, but are not yet in service. Additionally,

Jan 1, 1981

By G. E. Tooker

This paper (1) presents new data on the use of horizontal curves to optimize the alignment of belt conveyors. The operating concept of a horizontal curve is simple. As an element of the belt passes through the curve, it is motivated to drift toward the center of the curve by unbalanced tension forces in the belt. This lateral drift is resisted by the forces of friction and gravity, which in turn can be stated as a function of drift. Resisting forces, as a function of drift, are stated in terms of variables such as the amount of load on the belt, and parameters such as the trough angle, tilt angle, and super elevation angle, and number of rolls in the idler. The trick is to determine that the belt will drift to a position of equilibrium within predictable limits, without physically constraining the edge of the belt. In practice this prediction is not easy to make without the aid of a computer. The motivating force is easy to calculate, but its value is continually changing throughout the angle subtended by the curve arc. The resisting forces are constant throughout the curve for a given drift and load condition, but the calculations are very complex. Data in this paper will assist the conveyor designer in making preliminary calculations for horizontal curves. Presented is the equation for calculating the motivating force, and a family of curves that show the resisting force as a function of lateral drift for some commonly encountered conveyor design criteria.

Jan 1, 1985

By M. Jaramillo

The Expanded West Pit of the 15 MTY El Cerrejon Mine will reach a maximum depth of270 m. Slope stability evaluations indicated that this depth, combined with structural complexities, would create a serious risk of footwall instability. Stabilization call for anchoring the footwall by installation of cable bolts, 9 to 33 m long, of multi-strand steel cable. For each of the 15 m high mining benches, a row of bolts is installed with variable horizontal spacing, depending on dip magnitude. Cable bolting is now standard operating practice and 11800 bolts will have been installed during the life of the mine.

Jan 1, 1995

The management of risks associated with slope instability is an essential process in the safe and economic operation of open cut mines. The ?slope stability radar? (SSR) has been developed to better manage those risks. The SSR remotely scans rock slopes to continuously measure any surface movement and can be used to detect and alert users of wall movements with sub-millimeter precision. The high level of movement precision and broad area coverage of the rock face can allow for a better understanding of the geomechanics of slope deformation, including magnitude of potential failures and additional warning time of impending instability. Additionally, radar waves adequately penetrate through rain, dust and smoke to give reliable measurements, 24 hours a day. SSR systems have been deployed in many metalliferous and coal mines throughout the world. Over 80 systems are currently being used with current operations in 17 countries. Greater than 200 rock slope collapses and waste dump failures (from several to millions of tones) have been monitored, with ?warning? precursor movements recorded by the SSR. This technology enables a radical change in the management of risks in open cut mining. This paper will present recent case studies of the SSR operation in coal and metalliferous mines. Slope Stability Hazards and Rock Slope Monitoring in Open Pit Mines Surface mining often creates excavation boundaries which are steeper than those would be created under natural processes in the same rock type. They are also undertake quite rapid stress changes during mining and are influenced by significant external forces such as blast vibrations and changes in groundwater. These conditions are likely to allow some degree of instability to develop in the pit slopes. The hazards associated with this instability of pit slopes can be extremely variable in character; ranging from small cobble sized rocks dislodging and falling from mine benches to complete collapse of entire pit walls. It is usually the position of mine geotechnical or geological staff to assess these hazards for mine operations. The assessment of hazards due to unstable slopes can be done by visual observation of precursor signs such as fracturing or rilling of loose material. A more reliable indicator of instability involves the quantitative measurement of outward movement and acceleration of material as an instability mechanism develops. There is strong evidence that small precursor movements of a rock wall occur for an extended period prior to rock slope collapse (Hoek and Bray, 1981). Development of a monitoring system, adopting acceptable slope deformation criteria coupled with warning systems and design of stabilization or risk reduction measures if appropriate has become a standard method of dealing with slope instability. Framework for Slope Hazard Management using the SSR A framework for the management of slope hazards is shown in Figure 1 (discussed in detail in Harries, 2007). Although this framework is specifically for use with the Slope Stability Radar, a similar framework could be developed for other monitoring technologies for monitoring slope deformation. An important point to consider is the sequential nature of the hazard management framework from selecting the context ? identifying hazards ? analysing hazards ? evaluating hazards ? treating hazards. [ ] Once the context of the system has been developed (for example ? to monitor for rapid bench scale collapse), the identification of hazards is the next stage in hazard management. Identification of slope hazards can be relatively straightforward due to the visual nature of the information that the SSR system produces. The areas of excessive wall deformation are highlighted in color maps of the mine slope with a co-aligned photographic image to review the actual site of interest. The responsible personnel must then analyses this data, usually reviewing the deformation style and rate, as well as the size of area that appears to be failing. This information can then be evaluated against organisation protocols (for example it may be only a small sub-bench failure that will be ?caught? by existing benches and hence can be categorized as a lower risk). Once these hazards are evaluated they can then be treated if required. Two major options are available to reduce the risk during this treatment option. We can reduce the likelihood of failure ? for example by buttressing the failure or de-pressurising of water pressures in a slope. However, the most likely plan of action is to reduce the consequence of the failure. One of the most successful opportunities to reduce the consequence of a slope failure is to ensure that no personnel or mining equipment is present at the location of a slope failure when collapse finally occurs. This has been the major benefit of the SSR system to the different mining operations around the world. SLOPE STABILITY RADAR TECHNOLOGY OUTLINE Slope Stability Radar system specifications and capabilities: ? Measurement precision +/- 0.1mm (std.) ? Perforated dish and other features enable it to operate in high wind conditions ? Measurement range is 1700 metres ? Scans 270 degrees horizontally and 120 degrees vertically ? 1 ? 30 minute repeat rate for area scans (varies according to area being scanned ? RF exposure (<0.05mW/cm2) is a factor of 200 below the acceptable limit for the general population (1 mW/cm2) ? Tolerant to vibration and mining equipment ? The system weighs approximately 1500 Kg ? Remote area power supply enabling unmanned, continuous 24/7 monitoring ? Real time monitoring

Jan 1, 2009

By Paul T. Allsman

The guiding principle in developing this plan resulted from evaluating potential future markets for oil, the quantity of mineable oil shale and the optimum rate of production comensurate with market demand on the available oil from shale. In addition, the rate of production may depend upon a solution to the problems of scarcity of resources other than oil shale. The rate of production finally will tend to establish the choice of technologies with these limits. The principle components discussed will be: 1) The place of oil shale among the various energy sources and its relative adaptability for use in the area of major consumptive demand. 2) A recalculation of the reserves and their classification on the basis of quantity, grade - gallons per ton, - and physical characteristics principally thickness, overburden and depth.

Jan 1, 1982

By K. N. Han

The kinetics of gold dissolution in iodide solutions with oxygen in an autoclave was investigated. The leaching process was described by a gas-liquid-solid heterogeneous reaction. The effects of stirring speed, temperature, oxygen partial pressure and lixiviant, concentrations on the overall leaching rate was investigated. It was found that the dissolution of gold could be represented by a heterogeneous chemical reaction. Under the conditions studied, the apparent activation energy ranged from 34.7 to 54.3 kJ/mol. Experimental findings were in good agreement with the kinetic model proposed.

Jan 1, 1998

By R. R. Godard

The practice of dust abatement or control is not new to any operation of United States Steel Corporation. This report will review those activities, past and present, within the five operating districts of United States Steel-Coal Operations with specific reference to mechanical and continuous mining, and including both the health and safety aspects of the problem. Obviously, the first principle of dust control is to inhibit its formation If this is not practical, steps may be taken to prevent its becoming airborne. Airborne dust control can be realized by either collecting it at the point of generation, diluting it with large volumes of air to reduce concentration, settling it by various means, or with reference to ignition hazard only, render it incombustible by neutralization with an inert material. These basic principles have been verified by the work of many, including the Bureau of Mines, U. S. Department of Interior, Bituminous Coal Research, Inc., and the Mining Research Establishment of the National Coal Board of Great Britain. Their application to every day mining problems is best illustrated by the following examples.

Jan 1, 1967

By A. Juganda, F. WILSON, H. PINHEIRO, N. SANDOVAL, J. F. Brune, G. E. Bogin Jr.

To evaluate methane explosion hazards in longwall coal mines, a CFD model was developed along with a 1/40th-scale, optically accessible model of an underground longwall coal mining section. In this project, CFD models assisted in the design of the physical model to ensure specifications were met for accurately representing the scaling physics as well as to assist in narrowing the experimental matrix and identifying key locations for sensor placement to measure velocity, pressure, and gas concentrations. This research will help develop strategies for methane monitoring that prevent methane ignitions and explosions in longwall coal operations.

Aug 15, 2022

By Keith G. Wallace

In 1982 the U.S. Congress passed the Nuclear Waste Policy Act (NWPA) to provide for the development of an underground repository for spent nuclear fuel. This development will be managed by the United States Department of Energy. In 1986, the President selected three areas for site characterization to determine their suitability for the development of an underground repository; those sites were: 1) A site in volcanic tuff located at Yucca Mountain in Nevada, 2) a site in bedded salt located in Deaf Smith County in Texas, and 3) a site in basalt located in Hanford, Washington. At present conceptual repository designs are being developed for each site. A key element of a repository design is the underground ventilation system required to support construction, nuclear waste emplacement, and potential waste retrieval. This paper describes the preliminary ventilation system designed for the repository in tuff. The concept provides separate ventilation systems for the construction and waste emplacement activities. The paper further describes the means by which acceptable environmental conditions will be re-established to allow re-entry into previously closed rooms for the purpose of inspection, maintenance or retrieval.

Jan 1, 1987

By J. R. Evans

Mining claims must have a discovery of a valuable mineral deposit to be valid. When required, a Federal government mineral examiner makes a determination of validity and recommendations thereof to authorized management officials. The examiner should show whether or not an existing mining operation, or a proposed operation have economic viability. This is accomplished through an analysis of legal entities, land status and use, geologic and mineralogic relationships, and mining, milling and processing techniques considered in an appropriate economic evaluation framework. A framework involves market and market entry studies, reserve analyses, all costs analysis, royalty buyouts, tax structure, discount rate of return, and escalation and risk factors. An income approach to value is then used by means of a simple or complex discounted cash flow (DCF) analysis to certify validity. Whether one claim is owned by a single miner, or hundreds of claims are owned by a large corporation, thorough, fair, consistent, and appropriate economic concepts should be applied by the mineral examiner.

Jan 1, 1993

By Lyle W. Talbotr

The DeLamar Silver Mine is a Joint Venture between Earth Resources Company, the Canadian Superior Mining Company and the Superior Oil Company. The property is a consolidation of several old operations in the district. The more widely known are the DeLamar and Sommercamp Mines. The location is 20 miles northeast of Jordan Valley, Oregon and 55 miles southwest of Boise, Idaho. The mine is on the south slope of DeLamar Mountain in the Owyhee Mountains of southwest Idaho Initial interest started in 1969 with serious exploration beginning in 1970. The feasibility study was completed in 1974, and in 1975 Mountain States Engineers were retained to design and build the plant. Processing of ore began in March of 1977 with the first bullion poured in April of that year. The plant is currently processing some 2200 TPD of ore with a grade of four ounces of silver and 0.03 oz. of gold per ton of ore. Cyanide consumption is 2.5 lb/ton, zinc consumption of 0.3 lb/ton and flocculent consumption of 0.5 lb/ton. A flowsheet is shown as Figure I. The grinding circuit is of two stage design. The primary mill is an 18 ft. dia. by 9 ft. long semi-autogenous unit. The secondary mill is a 9 1/2 ft. dia. by 15 ft. long overflow ball mill. The mill discharges are combined and sent to 15 in. cyclones for classification. Cyclone overflow reports to leaching with the underflows reporting to the ball mill. The cyclone feed pump is an 8X6 metal case pump. The primary mill has 1000 installed horsepower and rotates at 68% of critical speed. The ball mill has 700 installed horsepower and rotates at 72% of critical speed. Because the primary mill is horsepower limited the ball charge is varied to maintain the desired throughput. The mill power is held at 980-990 HP by modifying the feed rate. Four inch forged steel balls are added any time the ore feed fails to reach 100 tons per hour for four hours. During periods of extremely hard ore, the ball charge has been as high as 14 v/o and at other times the mill has operated fully autogenous for weeks.

Jan 1, 1980

By J. Winterton

Cripple Creek and Victor Gold Mining Co. is developing a full-scale pad injection program as part of the overall operational plan for the Valley Leach Facility. As part of the development process the mechanics of the injection methodology are evaluated in the context of physical injection measurements, outcome from geophysical monitoring, and gold leaching theory. The results indicate that hydraulic fracturing analogies advanced heretofore do not accurately describe injection phenomena in unconsolidated and noncohesive material typical of leach pads. A new interpretation is presented and extended to define operational strategies for maximizing gold production.

Feb 27, 2013

By Thomas V. Falkie

In bygone days, the mineral engineer spent most of his/her time on the technical aspects of designing, building, and managing mineral operations. In the last few decades, mostly since the 1960s, increasing amounts of engineering man power and time are spent dealing with federal, state, and local government agencies, laws, and regulations. It is not possible to plan, design, build, or operate mines, mills, smelters, or even office buildings without detailed consideration of government regulations and permitting requirements dealing with environment, worker health and safety, etc. The four chapters of this section provide an overview of the social, legal, political, environmental, and health and safety aspects of mineral enterprise. The section sets the stage for the working engineer/geologist in industry, government, consulting, or academe and points the direction for more detailed study in each of these areas.

Jan 1, 2008

By R. H. Weiland

The dewatering of flocculated clay suspensions by compression in a centrifugal field has been investigated. The results are evaluated with regard to both direct applicability to centrifugal dewatering, in solid-bowl centrifuges for example, and indirect applications to gravity sedimentation systems. It is well known that sediment consolidation under gravity depends on structural characteristics such as permeability and on the "self-weight" of overlying sediment layers. An important objective of the study has been to evaluate the self-weight effect using centrifugal forces to simulate consolidation in deep sediments. In this way, small-scale laboratory tests can be used to predict the rates and limits of sediment compression in industrial systems such as thickeners and settling ponds. Experiments have been conducted in which the sediment mass and initial structure were maintained constant while the effective sediment weight was varied by centrifuging at different speeds. The effects of initial structure have also been studied using controlled variations in flocculation conditions.

Jan 1, 1993

By J. McMurray

Introduction Uranium was recovered from mining of collapse breccia pipe deposits in the Arizona Strip district in northwestern Arizona (Figure 1) intermittently between 1956 and 1990, with cumulative production having totaled 23.3 million pounds of U3O8 (8960 tU) at an average grade of 0.6% U3O8. Production was suspended in 1990 because of low uranium prices, but the recent resurgence in market price once again makes the breccia pipe deposits potentially attractive exploration and development targets. This paper describes: (1) formation of the pipe structures and their metallogenesis; (2) mining and milling of breccia ore; and (3) potential roadblocks to development of new breccia pipe mines. Breccia Pipe Formation Formation of the breccia pipes began in cave systems formed during development of karst topography in the Mississippian Redwall Limestone. The karst Redwall surface was eventually covered by up to 900 meters of shallow marine to continental sediments ranging in age from Pennsylvanian to Triassic. The pipes were formed by progressive upward stoping caused by groundwater ascending under hydraulic head along near-vertical faults, joints and fractures. The stoping process involved both chemical disaggregation and mechanical failure of roof and wall rock. Upward stoping of successively younger units allowed for upward propagation of a pipe and resulted in a chimney-like column of breccia between the steep walls of the pipe. Syndepositional thickening of sedimentary units of varying ages indicates that pipe formation began as early as early Pennsylvanian time and continued intermittently through Triassic time.

Jan 1, 2007

By QUANWEN PU, Yonggang Wei, Bo Li

The spent catalyst is one of solid waste resources, which contains precious metals with considerable recovery value. However, due to the different compositions of catalyst, the recovery process is difficult. Most of the catalysts are supported by alumina which has good alkali-soluble properties, while platinum is insoluble in alkali. Platinum is enriched by pressurized alkaline solution which is a kind of hydrometallurgical process. The novelty of this study is the separation of alumina and platinum from the spent catalyst and the enrichment of platinum in the solid phase. In this study, a leaching process using NaOH as leaching agent under high temperature and pressure was employed. The effects of alkali concentration, leaching temperature, leaching time, and liquid–solid ratio on leaching efficiency were investigated. The optimum process conditions were determined as follows: alkali concentration of 240 g/L, temperature of 200℃, pressure of 1.20 MPa, time of 1 h, liquid–solid ratio of 5:1. It was concluded that the leaching rate of alumina can reach 99.69%. The content of platinum increased from 0.1749 to 21.12%, the loss of platinum was 3.39%, and the enrichment ratio of platinum increased by 120.82 times. The leaching kinetics of alumina was studied based on the shrinkage core model. The activation energy of alumina leaching is 19.53 kJ/mol, and the leaching process is controlled by chemical reaction. Pressurized alkali solution can not only efficiently leachate alumina from the spent catalyst and enrich platinum, but also reduce the reaction time and increase reaction rate.

Feb 11, 2022

By Jonathan G. Price

Silver-copper-lead deposlts in Precambrian, Permian, and Cretaceous red-bed sequences near Van Horn, Texas, have the following features In common: (1) mineralization controlled by steeply dipping fractures in sandstones, (2) lack of ore In overlying and interbedded limestones, (3) absence of igneous intrusions, (4) sulfide ores without typical hydrothermal gangue minerals, and (5) associated arsenic, molybdenum, zinc, and cadmium but not gold. Ore was probably deposited along fractures by reaction between low-temperature, oxidizing, metal-rich ground waters and sulfide-bearing waters or rocks. Detailed mapping and studies of fracture patterns indicate that mineralization occurred during Basin and Range extensional deformation, thereby postdating major Trans-Pecos igneous activity.

Jan 1, 1983

By Jeffrey A. Pontius

The Cresson gold deposit is in the southcentral portion of the Cripple Creek mining district in central Colorado (Fig. 1). The deposit is hosted by a series of 32 Ma (Kelley, 1995) felsic to mafic, alkaline volcanic rocks that were deposited in a diatremal environment. Mineralization in the Cresson deposit is structurally con-trolled. It occurs as disseminated and vein type. Gold deposition is closely associated with potassium-pyrite metasomatism and quartz-telluride vein development(Pontius, 1992). The Cresson deposit is owned by the Cripple Creek and Victor Gold Mining Co. (CC&V), a joint venture between Pikes Peak Mining (67%) and Golden CycleGold (33%).

Jan 1, 1996

After several years of laboratory and field evaluations, it has been established that each frother has its own unique relationship among bubble size reduction, gas hold-up increase, froth stabilization and frother concentration. It has been shown that a frother?s hydrodynamic properties can be characterized using the relationship between the froth zone property froth height and the pulp zone property gas hold-up. The relationship is established in a bubble column run at a given air rate over a range of frother concentration. The use of relationships in frother selection and plant diagnostics is illustrated in a series of plant case studies.

Jan 1, 2009