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By Ralph J. Anctil, James E. Palmer

According to relatively recent estimates by the US Geological Survey and the US Bureau of Mines, there are more than 0.9 Tt (one trillion tons) of remaining identified coal resources in the seven major western US coal-producing states and about 200 Gt (220 billion tons) in the demonstrated reserve base. Considering coal's current national production level of 0.7 Gt/a (0.8 billion tpy), these US reserve and resource figures are reassuring. Given here is a preliminary analysis of geologic constraints and an estimate of their effects on mining and recoverability of western coal. Mining Constraints USGS and USBM coal figures incorporate only thickness and depth, and are "blind" to many other major geologic constraints on mining and recoverability. The most important of these geologic constraints include: • coal seams too thin or too thick for conventional mining methods; • coal seams that cannot be mined due to insufficient separation from underlying or overlying previously-mined beds; • unstable or otherwise unfavorable roof strata; • soft or heaving floor strata; • coal seams that are too lenticular or discontinuous for conventional mining methods; • coal seams in intensely faulted areas; • coal seams having high angles of dip precluding the use of conventional coal mining methods; • coal seams that cannot be mined due to proximity of water-bearing strata; and • coal quality problems such as excessive ash, sulfur, chlorine, or sodium. Coal Thickness Present USGS-USBM classification for reserve base includes bituminous coal beds more than 0.7 m (28 in.) thick and subbituminous beds more than 5 m (60 in.) thick to depths of 305 m (1,000 ft). Both of these thicknesses are less than the thicknesses now considered mineable in the western US. The lower limit for bituminous steam coal in deep mines is about 1.5 m (5 ft). For subbituminous coal, a thickness of at least 2.4 m (8 ft) is generally required. Accordingly, the Bureau of Mines reserve base includes many thinner coal seams that will not be mined in the foreseeable future. Within the Powder River Basin of Wyoming and Montana many seams-Wall, Upper Pawnee, Lower Pawnee, Cache, Manning-range in thickness from 7.6 m (25 ft) to more than 30 m (100 ft) and occur at depths to 305 m (1,000 ft). These have been included in the Bureau of Mines reserve base. However, the mining technology required to mine these thick seams does not now exist. It will require many years for development. The ultimate recoverability from these thick seams is unknown, but could be as low as 25%.

Jan 8, 1982

By Robert H. Merrill

7.1-STABILITY OF SLOPES IN SOIL AND ROCK ROBERT H. MERRILL 7.1.1-DESIGN ANALYSIS APPROACHES The analyses of slopes in soil or rock can be made through application of soil- or rock-mechanics principles. In most cases, the material in the slope wall can easily be identified as either a soil or a rock. However, there will be instances where, because of unique characteristics, the soil may behave as a rock structure, or vice versa. By necessity, these unique cases must be identified through experience or critical observation by the engineer, because there are no definitive rules clearly establishing the differences. The most important step in a solution to a slope problem is to define it. Normally, once the problem is identified, a solution usually can be found or a reasonable estimate obtained. The design analysis of a slope wall implies an evaluation of the stresses or forces in the wall, and of the ability of the material in the wall to withstand them. Therefore, the factors influencing a design analysis can be divided into two categories: those which influence stresses, and those which influence the ability of the wall to withstand stress. For simplification, the latter is termed the strength of the soil or rock. Perhaps the basic difference between analyses for soils and those for rock is in the identification of the failure surface. Soils generally do not have an identifiable failure plane. However, some soil layers in a slope wall may have different strengths. A soil slope wall may consist of a thin layer of topsoil overlying a thick, uniform clay, or a topsoil over a gravel or other materials. Therefore, the soil generally does not have a potential plane of weakness that can be identified through observation. On the other hand, rocks normally contain planes of weakness in the form of fractures, faults, joints, bedding planes or other discontinuities weaker than the intact rock, which can be found through observation and/or tests upon rock samples. If the failure surface cannot be identified, the principles of soil mechanics should apply because, in soil mechanics, the failure surface usually is assumed. If it can be identified, then the principles of rock mechanics should apply. 7.1.2-ANALYSES OF STRESSES OR FORCES IN SLOPES All existing methods of soil-slope analysis are based upon interpetations of the Coulomb and Coulomb-Navier failure criteria, commonly termed the total-stress 2nd the effective-stress criteria (Sec. 6.1). The total-stress criterion may be illustrated by the diagram in Fig. 7-lA, which relates the total stress, S, to the cohesion, c: the normal stress, u, and the angle of internal friction, 9, by the expression

Jan 1, 1973

By Jan M. Mutmansky

30.1-THE SCOPE OF SYSTEMS ENGINEERING JAN M. MUTMANSKY JAN M. MUTMANSKY 30.1.1-DEFINITIONS The terms "systems engineering," "systems analysis," and "operations research," common in today's technical literature, often are interchanged because of their similar meanings. However, some differences exist which need explanation. The most convenient way to achieve an understanding of the differences is to start with the definition of a system, namely, a collection of components which interact in a specific manner to accomplish a common objective. Accordingly, a haulage truck, a computer, a missile, a mining company or a nation each may be considered a system. Systems engineering generally can be defined as the conception, planning, design and engineering of any system of interacting elements so that the objective of the system is automatically optimized. Thus, systems engineering normally is con¬cerned with systems not yet activated.',' Systems analysis may be described as a scientific method of making decisions based on a quantitative or other objective evaluation of all action alternatives. It is very similar to systems engineering but usually is not considered to be restricted to problems of design prior to use. While the previous two disciplines are somewhat vague and can encompass a great variety of engineering functions, operations research can be more specifically defined as the application of mathematical models to the problem of optimizing the objective of any predefined system. Systems analysis and systems engineering are closely aligned sciences but opera¬tions research differs in the following ways. First, it generally is associated specifically¬with optimization and mathematical models while the other two often are not. In addition, it, is used predominantly on well-defined problems which can be ex¬pressed in mathematical terms. However, systems engineering or systems analysis may deal with highly complex social or governmental systems which may be ill¬defined and unsolvable by the classical methods of operations research. Finally, systems analysis and engineering always strive to look at the entire system to avoid the problem of suboptimization often necessary in operations research because of the necessity for a mathematical model. Because of his more extensive system, however, the systems engineer often is less able to find the optimal system. In summary, operations research, systems analysis and systems engineering are very similar in purpose but one, operations research, is more mathematical in nature and forms an important part of the other two activities. Other definitions which may be helpful are provided by Griffiths 3 Wilson,4 Falkic 5 Hall,2 Quade and Boucher, 6 and Flagle et al. 1 30.1.2-METHOD AND TECHNIQUES The systems or operations research approach makes use of scientific problem solving. The basic procedure comprises the following general steps: 1. Definition of the problem. 2. Construction of a model of the system. 3. Testing of the model under known conditions.

Jan 1, 1973

By Richard B. Muter

1. Coal Preparation Terms (General) 1.1 Coal preparation-collectively, physical and mechanical processes applied to coal to make it suitable for a particular use. 1.2 Run of mine (ROM coal)-coal produced by mining operations, before screening, crushing, or preparation. 1.3 Raw coal-coal that has received no preparation other than possibly screening or crushing. 1.4 Raw coal feed-raw coal supplied to a plant or machine, in which it undergoes some form of preparation. 1.5 Coal cleaning-the treatment of coal to lower the mineral matter content (ash). 1.6 Cleaned coal; clean coal-coal produced by a cleaning process (wet or dry). 1.7 Middlings-a product of coal preparation that, by reason of its ash content, is intermediate between coal and discard. It follows therefore that the relative density of middlings is intermediate between those of coal and discard. Middlings may be reprocessed. 1.8 True middlings; bone comparatively high ash particles so nearly homogeneous that their quality cannot readily be improved by crushing and recleaning. 1.9 False middlings; interbanded middlings-middliigs in which the particles consist of bands of coal and shale, and from which the coal may be liberated by crushing 1.10 Refuse-waste material in the raw coal which the cleaning process removes. (See also discard; dirt; stone, No. 1.12). 1.1 1 Reject-the material extracted from the feed during cleaning, for retreatment or discard. 1.12 Discard; dirt; stone--the material extracted from the raw coal and finally discarded. 1.13 Recirculation-the operation in which the whole or part of a product from a process is returned to the feed to the process, e.g., the return of the crushed overflow from a screen to the screen feed for rescreening. 1.14 Foreign coal-coal received at a preparation plant from a colliery other than that to which the plant is attached. 1.15 Imported coal-coal coming from a foreign country, or other state within the country. 1.16 Low-grade coal combustible material that has only limited uses owing to undesirable characteristics (eg., ash content or size). 1.17 Segregation-partial separation of a previously mixed material into its constituents, occurring as a result of differences in either particle size or relative density. 2. Cleaning Characteristics 2.1 Washability-amenability of coal to improvement in quality by cleaning. 2.2 Float-and-sink analysis-the division of a sample into relative density fractions with defined limits, the proportions of the fractions being expressed as percentages of the total sample, commonly with an indication of the ash percentage (and other characteristics, if required) of each fraction.

Jan 1, 1991

By A. Ang-Ug, L. Gurieff

"The Phu Kham Cu-Au mine in Laos is owned and operated by Phu Bia Mining Limited. The deposit is a complex and low grade Cu-Au porphyry ore. The rougher feed has a particle size distribution of 80% passing 110 µm, and the rougher concentrate is reground for cleaning to 80% passing 25 µm. The Cu losses in the cleaning circuit are primarily in the <12 µm fraction. A randomised paired ON-OFF test of magnetic conditioning in one of the two parallel cleaner 1 circuits was undertaken. The results showed that with magnetic conditioning there was approximately a 10% reduction in the <12 µm Cu distribution in the cleaner 1 tail to a 98% level of confidence. The Cu distribution in the <12µm fraction is a less variable measure of changes in the <12µm Cu recovery, therefore, this was used to evaluate the impact of magnetic conditioning. Subsequently, magnetic conditioning was installed in both cleaner 1 circuits and the plant results monitored after this installation. Comparing the plant results immediately before and after this installation showed that there was a 20% reduction in <12µm Cu losses to the cleaner tail. INTRODUCTION The Cu-Au, Phu Kham Mine (PKM) is located within PanAust's Phu Bia Mining Contract Area in Laos, approximately 140 kilometres from the capital city, Vientiane. PKM is PanAust's flagship operation and the significant cash flow generated by PKM has supported PanAust’s growth, while contributing to a strong balance sheet. PKM comprises an open-pit mine and a conventional milling and flotation operation that produces and exports a Cu and precious metals concentrate. The concentrate contains around 23% Cu, 6 ppm Au and up to 35 ppm silver. Process Overview The PKM deposit consists of complex heterogeneous mineralogy horizons of Cu-Au stockwork and skarn mineralisation. It contains areas of soft-leached zones, overlying transition zones and supergene chalcocite dominant secondary Cu mineralisation and clay-rich gangue. Areas of high As and Zn also pose a challenge in treating the highly variable ore. In addition, hardness of the ore increases with deposit depth. The ROM is hauled to a primary gyratory crusher to produce a coarse ore stockpile with size distribution suitable for the 13 MW semi-autogenous mill. The SAG product feeds two parallel ball mills each equipped with a cyclone cluster producing a product particle size distribution of 80% passing 110 µm. The conventional flotation circuit consists of two parallel rougher banks. The rougher concentrates feed the regrind cyclone where the underflow is split and reports to two parallel, 3MW, M10000 IsaMills. The regrind mill discharge combines with the regrind cyclone overflow with a product particle size distribution 80% passing 25 µm and reports to the Jameson cell."

Jan 1, 2018

By Lee Ann Arrington-Webb

Thiele Kaolin Co. has invested much time and money in achieving and maintaining its IS0 9002 certification. Through a regimen of process control and documentation, routine audits and a commitment to quality management, this kaolin producer is a competitive pigment supplier in a worldwide market. Background In 1987, the International Organization for Standards developed the ISO 9000 quality management system. This series of standards is used as a means to ensure customers that a quality management system is in place. Certification of a company does not mean that an individual product is of high quality. Rather, it demonstrates a supplier&apos;s ability to meet its customers&apos; product and service needs. It also shows that a company has documented records to assure product consistency. The ISO 9000 system was begun in England during the early 1980s. During the mid-1970s, the British manufacturing industry was experiencing low quality. In an attempt to boost quality standards, the British government established the BS 5750 standards system. In 1987, BS 5750 was incorporated into ISO 9000. The ISO 9000 series consists of three levels of certification. ISO 9001 is the model for quality assurance in "design/development, production, installation and servicing." This standard applies to companies that rely on research and development to bring products that frequently change to market. ISO 9002 involves "production and installation" only. This standard is relevant to industries that produce a specific product according to specifications - either their own or their customer&apos;s. ISO 9003 is the model in "final inspection and test." This standard is the least detailed of the 9000 series. It is also the most applicable to companies that purchase, inspect or sort and resell. The quality system requirements for certification are outlined in the 18 elements of ISO 9002. These include: 4.1 - Management responsibility. 4.2 - Quality systems. 4.3 - Contract review. 4.4 - Design control. 4.5 - Document and data control. 4.6 - Purchasing. 4.7 - Control of customer-supplied product. 4.8 - Product identification and traceability. 4.9 - Process control. 4.10 - Inspection and testing. 4.11- Control of inspection, measuring and test equipment. 4.12 - Inspection and test status. 4.13 - Control of nonconforming product. 4.14 - Corrective and preventive action. 4.15 - Handling, storage, packaging, preservation and delivery. 4.16 - Control of quality records. 4.17 - Internal quality audits. 4.18 -Training. Continuous improvement, though implied in maintaining these requirements, could potentially be included in the quality system requirements in the next revision. The ISO 9000 requirements and documentation system is shown in the hierarchy pyramid (Fig. 1). A company&apos;s quality policy serves as the top of the pyra-

Jan 1, 1997

By Calvert D. Iles

Open-pit mines in the US and throughout the world are typically characterized by increasing operating costs and decreasing ore grades. Thus, companies must continue to improve their use of human and technological resources to remain competitive in the marketplace. Duval Corp. has historically been very sensitive to mining costs, in part because of relatively low grade copper-molybdenum ore bodies. These concerns have resulted in three haulage system generations at the 218 kt/d (240,000 stpd) Sierrita-Esperanza open-pit mine south of Tucson, AZ. Each generation has exceeded the performance of the one replaced. The latest system of portable crushers and conveyors reduces hauling costs by combining both hardware and technique. Savings are achieved not only bysubstituting conveyors for haul trucks, but also by permitting the crusher-convey or system to closely follow mining as the pit deepens and/or mining areas change location. The Duval system is capable of easy and inexpensive relocation in varying site configurations in relatively short time periods. Its implementation in active mining areas results in minimum disruption to both the mine operation and mine design. This haulage concept allows the mining engineer maximum freedom to design an open pit for profit rather than for equipment. Maximum savings with a portable system are achieved by minimizing truck haulage costs. At Sierrita, for example, the yearly estimated operating cost savings for each thousand feet of +10% truck ramp replaced by conveyors is about $1 million per portable crusher. A savings of $30 million could be achieved fora three-portable system overa 10 year period simply by reducing the average truck Iiftto the crusher by 30 m (100 ft), There is also the added capital expense of buying trucks to cover the additional 30 m (100 ft) of lift. Relocation costs are an important consideration in a portable system. Direct costs can become a significant operating expense if frequent relocations are necessary to take full advantage of potential haulage cost savings. A site preparation cost of $375,000, for example, would add an additional operating cost of 2.8¢/t(2.5¢/st) for a system handling 13.6 Mt/a (15 million stpy), assuming one move a year, In addition to the direct cost of site preparation, such as installation of retaining walls, apron feeder slots, and equipment pads, there is the cost of site renovation after abandonment. Indirect relocation costs, while less tangible and more dependent on specific mine design and scheduling, can be of significance. A slowdown in the mining rate or a less efficient mining plan can result from a site not being immediately available for mining after abandonment. Conversely, mining in narrow areas can be adversely affected if two or three month lead times are required for site preparation.

Jan 4, 1983

By Raymond E. Kary, John W. Mcklveen

The four unit Cholla coal fired power station at Joseph City, Arizona, consumes coal from the McKinley mine located in western New Mexico. Since coals may contain relatively high uranium or thorium concentrations and may release quantities of detrimental progeny of these elements to the environs, there is concern that releases may pose an adverse environmental impact or even a health hazard to the workers. Thus, an effort was made to quantify the radionuclide content of the coal and compare the results with radionuclide concentrations already present in the vicinity of the plant. Additionally an attempt was made to assess any potential for significant enhancement of the levels of radioactivity in the environment or to plant personnel. Background gamma radiation measurements at and near the plant ranged from 8 to 12 [µ]R/hr (70 to 105 mR/yr) with a few localized anomolies measuring to 18 [u]R/hr (160 mR/yr). Levels on top of the coal storage piles decreased to ~ 6.5 [u]R/hr (60 mR/yr). Apparently the direct radiation emitted from the coal is less than the background levels in the vicinity of the plant. Measurements performed on top of several coal seams at the mine confirmed low radiation levels emanating from coal, but readings more than doubled when the overburden was monitored. Laboratory gamma spectrometry confirmed the lower amounts of natural radioactivity in the coal when compared against soil and water samples obtained in the vicinity of the plant and the overburden at the mine. In an effort to see if there is biological enhancement, animal feces was sampled. The radiation variations were found to be influenced primarily by naturally occurring potassium-40 which varied from ~ 1 to 18 pCi/gm rather than fluctuations in the uranium and thorium series. The potassium-40 concentrated in the ash were similar to the concentrations in the environs around the disposal ponds while Radium-226 concentrated in the fly ash was ~ 3 times that found in the immediate surroundings. Radon-222 gas concentrations in buildings at or near the plant ranged from 0.5 to 1.8 pCi/l and were generally lower than indoor concentrations measured in the Phoenix area. The burning of coal at the Cholla station does not appear to increase ambient levels of radioactivity found in the surrounding environs or to the workers. However, the use of byproduct fly ash for construction purposes may need to be examined.

Jan 1, 1981

By Malcolm J. A. Swallow

In June 1997, Mount Polley Mining Corp., a division of Imperial Metals Corp., completed an 18-kt/d (20,000 stpd) gold-copper mine in central British Columbia. Fast-track construction and extensive off-site construction were combined with an aggressive project schedule to build the mine in one year. This was three months ahead of schedule and 7% below the original budgeted cost of C$123.5 million (US$86 million). The Mount Polley Mine was the first of a crop of new open-pit mines in British Columbia. The Mount Polley deposit has been explored since the mid-1960s when it was known as the Cariboo Bell property. From 1988, Imperial Metals completed drilling and metallurgical testing on the deposit. This led to a positive feasibility study and the issuing of a Mine Development Certificate in 1991. The project was later shelved due to falling metals prices and the strength of the Canadian dollar. In late 1994, Imperial merged with Bethlehem Resources Corp., bringing a strong project team to the company. The deposit was then re-evaluated, ore reserves revised and additional metallurgical testing was performed. The result was a revised project plan. It indi¬cated that a financially feasible gold-copper mining op¬eration could be constructed at Mount Polley operat¬ing at 18 kt/d (20,000 stpd). In May 1996, an agreement was reached with Sumitomo to jointly fund and develop the project. By early 1996, final plans and permits were being prepared for construction, scheduled to begin in June 1.996. On-site work commenced on May 28th, 1996 following lifting of the road restrictions. Mount Polley Mining acted as its own project manager. Construction was completed in June 1997. Plant runup took place in July and full production began in August. Through the use of off-site construction and the careful selection of contractors who cooperated to complete some 12 major contracts, it was possible to reduce the mine and mill construction time from 16 to 13 months. Off-site construction for all possible aspects of the work and the use of used milling equipment allowed significant savings in overall capital costs of construction. Total costs for the mine, mill and tailings dam equated to C$6,100/t/d of throughput. The mill design incorporates several innovative features that will aid maintenance and allow flexibility of operation. The operation is designed as a closed system and meets the latest environmental standards. Concentrate is shipped by truck to the port of Vancouver. Revenues for the mine come from the mine&apos;s production of about 3.1 t/a (100,000 oz/year) of gold and 10.8 kt/a (24 million lbs/year) of copper during the first four years of the 12-year mine life.

Jan 1, 1998

By Tim Neil, O&apos

Finnish mining traditions date back 400 years. The country was then one of Europe&apos;s major iron producers. Most of the mines in current operation, however, were started during the last 20 to 30 years. The Finnish mining industry must make do with limited resources. Deposits are usually low grade and small in size. Open-pit methods are used where possible. But thin ore bodies and harsh winters can impose severe restrictions. Underground mining is often the only option. Despite these handicaps, Finnish metal miners produce copper, nickel, zinc, vanadium, iron ore, cobalt, and ferrochrome. Ironically, the success of Finnish mining is rooted in a crisis the industry faced several decades ago. Mining costs were rising. High grade mineral deposits were nearly exhausted. It was vital that Finland develop new mining machinery and methods that afforded higher productivity. Cooperation between the Finnish mining industry and its equipment manufacturers was good. Partly as a result, the Finnish mining industry has turned itself into one of the most modern and efficient in the world. The Finns did this while continuing to meet the stringent safety and environmental requirements in force in the Nordic countries. Today, the share of domestic raw materials used by the Finnish metallurgical industry is decreasing. So the country increasingly depends on imports of ores and concentrates. To better provide these imports, Finnish mining companies have expanded their foreign operations. And, while importing raw materials, the mining industry has been exporting its technological and equipment know-how and expertise. Mining methods Conventional mining methods and machinery are used in surface mines. Drilling is the exception. It is done with Tamrock hydraulic track drills and 100 to 150 mm-diam (4 to 6-in.-diam) drill holes. Pumpable slurry explosives are used in larger open pits. Underground, sublevel stoping with open stopes is the most common mining method. It accounts for about 80% of all underground stoping. The main benefits of this method are low production costs, high mechanization, and easy control. The best productivity figures exceed 10 t (11 st) per man-hour, including all mining personnel. The development of cable bolting has made it possible to use sublevel open stoping even in relatively weak rock. Cut-and-fill methods, such as concrete-pillar stoping, inclined-wall stoping, and traditional cut-and-fill have also been used, combined with sublevel caving and room-and-pillar methods, according to local demands. Various fill materials and stabilization components have been investigated and tested for cut-and-fill methods. Sublevel open stopes are often later filled, to stabilize the surface or to allow pillar recovery.

Jan 10, 1988

By W. L. Wilson

This article could be entitled "The Rediscovery of the Windfall Mine," since it was originally discovered early in this century. The "Rediscovery" was made by personnel of Idaho Mining Corp. during the course of an investigation of the Windfall mine beginning in 1967. There had been historic production from the Windfall mine sporadically from 1908 until 1951 and there was present during much of this time a mill that vat leached crushed ore. Mining was underground to a depth of 91 m (300 ft). Prior to Idaho&apos;s activities, the mine had been inactive since the early 1950s. Idaho Mining Corp. was an "exploration company," privately owned and based in Reno, NV. The writer was at that time manager of exploration for Idaho Mining. The company held interests in the Cortez Gold mine, then being readied for production, and had conducted exploration for various mineral commodities in Nevada for several years. Idaho&apos;s attention was first directed to the Windfall mine in the Eureka District, Eureka County, Nevada (see Fig. 1, Nevada map) by a search of published geologic literature of the state of Nevada. An attempt was made to acquire a library consisting of all of the published geo-data for the state of Nevada as well as to obtain as much of the unpublished material as could be acquired. A study of these data, and in particular Tom Nolan&apos;s Professional Paper 406 (1962), indicated that the Windfall mine was anomalous for the Eureka District in that it contained deposits of disseminated gold. This source and other sources indicated that the Windfall had produced from underground operations approximately 59 kt (65,000 tons) grading 12.6 g/t (0.368 oz per ton) gold ore that was processed in a vat leaching operation and that the gold was "invisible" to the naked eye and could not be panned. It was ascertained that there were no active operations near the mine and a geologic reconnaissance was made and a few samples were taken to attempt to determine if the mineralization was indeed disseminated. The initial group of samples around the mine area averaged 1.25 ppm [11.5 g/t gold (0.036 oz per ton)] and the mineralization did appear to disseminate into the wallrock. The next step involved researching the ownership of property in the immediate area of the mine, and [ ] a property map was constructed that showed six or seven different ownerships of patented and unpatented claims, as well as some open ground, in the target area. Negotiations were commenced with the claim owners, ultimately resulting in the acquisition of all of the desired property, although some of

Jan 1, 1990

By Hugh E. Cronin, Joseph J. Coluccio

The Columbia Slough Consolidation Conduit Project is a 12 ft. 0 in. inside diameter—15 ft. 4 in. excavated diameter tunnel, 8,300 lf. in length. The tunnel is divided in two parts a 4,460 lf. reach to the West and a 3,770 lf. reach to the East separated by a flow regulation structure in the middle. The project was built for the City of Portland, Oregon Bureau of Environmental Services (BES) by the Frank Coluccio Construction Company (FCCC) Seattle, Washington under a $22,600,000.00 contract starting in 1998 and completing the contract in 2000 at a total cost to the City of $25,448,000. An increase of 12.6% from the original contract. The tunnel was constructed in an alluvial sediment that ranges up to 1,500 feet thick. The sediment varies from ?ne silty sand and sandy silt (SM-ML) to coarse sand and gravel (GP-GM) with varying quantities of cobbles and boulders. Most of the tunnel was constructed above the water table although one reach of almost 1,000 lf. had water to above the tunnel crown. The contractor, FCCC, used a Lovat 15 ft. 3 in.900 hp shielded TBM with multiple doors (4) of variable opening to excavate and support the tunnel. Tunnel supports were 5 in. WF 16 lbs. per foot at 4 ft. spacing and4 in. Douglas Fir lagging. Tunnel excavation was worked in 10 hour shifts with additional equipment maintenance and repair on the off time. The original Geotechnical Design Summary Report (GDSR) established a baseline for contractual purposes of slightly over 5 boulders (between 12 in. and48 in.) per foot of tunnel. (Table 1) At the pre-bid conference, three test pits were excavated to approximately tunnel invert, one at each end of the project and one in the middle near the flow regulation structure. The test pits encountered a few boulders but did not seem to support the five boulders per foot (varied from one boulder per foot of tunnel to eight b/lf.). Subsequent to the pre-bid meeting and the test pits, BES issued any addenda which changed the baseline boulder quantity in the GDSR to 1.7 boulders per foot of tunnel (varied from0.2 to 3.5 boulders per foot) a reduction of 67% from the original baseline. It is interesting to note the two separate tunnel reaches and how they were evaluated in the GDSR. Originally the Eastern portion of the work was projected at5.08 boulders per foot of tunnel and the Western portion at 5.30 boulders per foot of tunnel for a variance of less than 5%. In the revised quantity the Eastern portion was projected at 2.01 boulders per foot of tunnel and the Western portion at 1.42 boulders

Jan 1, 2003

By Thomas P. Kwiatkowski, Len Cobb, Christopher D. Orlandi, Judith H. Jones

The Inland System supplies raw water to the industrial and domestic users in the City of Birmingham, Alabama. The Inland System, which is nearly 60 years old, consists of an approximately 57.9m (190 feet) high earthen dam across the Blackburn Fork of the Black Warrior River, a 39.6m (130 feet) high reinforced concrete intake structure, a 1524mm (60inch) diameter Dresser coupled steel transmission pipeline, and various distribution lines extending into Birmingham. The Inland System was owned, operated and maintained by the Industrial Water Board (IWB) of the City of Birmingham, an agency of the city government, from 1958 until 1994. In 1994, the City of Birmingham consolidated all water and distribution functions under the Birmingham Water Works Board (BWWB). In 1988, the IWB determined that the Inland System had reached a point where its reliability could not be assured without an in-depth evaluation and renovation. The IWB engaged Christy/Cobb, Inc. (CCI) of Birmingham as its consultant for the renovation work. CCI evaluated the Inland System and submitted a report to the IWB in April, 1992. The report contained several recommendations for capital improvements to the system including increasing the emergency spillway capacity, reconstructing the fuse plug, relocating public access to the lake away from the dam area, constructing a new intake structure, and constructing a tunnel through the dam abutment to deliver the water from the proposed intake structure to the existing 1524mm (60 inch) diameter supply line. The new intake structure and tunnel would provide for a future maximum demand flow rate as well as provide bypass and level control capabilities. CCI was authorized in 1992 to proceed with the engineering design for the renovation of the Inland System with emphasis on the spillway and intake structure. PROJECT DESCRIPTION The major component of the renovation of the Inland System was the new intake structure and tunnel through the dam abutment. CCI engaged Jenny Engineering Corporation (JEC) of Springfield, New Jersey as a consultant for the tunnel design. JEC was responsible for the design of the tunnel, the connecting shaft between the intake structure and tunnel, and the steel pipe final lining, while CCI was responsible for the intake structure and the connection to the existing supply pipeline as well as overall project management. CCI and JEC evaluated the project site in order to select the location for the new intake structure and set the alignment for the tunnel. A factor in the evaluation process was that the level of the lake could not be lowered to construct the new intake structure since the supply of water from the Inland System had to maintained. Therefore, a location was required which was

Jan 1, 1997

By Richard H. Follis

Activated carbons used for precious metal recovery from carbon-in­pulp (CIP), carbon-In-leach (CIL) and carbon-in-column (CIC) circuits often become fouled with Inorganic precipitates (i.e., CaCOa, CaS04, SI02), and adsorbed organics (I.e., petroleum distillates from lubricants, dissolution products from wood) (1,2). These contaminants occlude carbon pore structure, such that diffusion controlled adsorption and desorption (elution) of Au(CN)2-and Ag(CN)2· are lmpeded(3,4l. Thus, after only short exposure to the hydrometallurgical environment, the characteristic carbon activity is reduced(5l. By frequently regenerating circuit carbons, activity losses associated with fouling can usually be minimized. Regeneration typically Involves (not necessarily Is this order): 1) elution to remove adsorbed ionic species (e.g., until they are removed from the carbon pore structure, Au(CN)2" and Ag(CN)2-are contaminants )(6l; 2) acid washing with 3-15% (wt/wt) hydro­chloric acid, to dissolve soluble Inorganic precipitates such as CaC03; 3) neutralization and/or water rinsing to remove residual acid; 4) thermal regeneration at 350-650 °C to remove adsorbed and/or pyrolyzed organ­ics; and 5) screening to remove fine carbon particles produced by Impaction and abrasion in the various stages of the hydrometallurgical circuit (i.e., adsorption, elution, and regeneration). The overall process Is purposed to reopen the circuit carbon pore structure, and to maintain the activity and qualities of virgin (new) carbon(7,8l. While most CIP, CIL and CIC operators regenerate circuit carbons, few evaluate the effect of the process on carbon quality in a systematic way. Process control usually Involves Iterating critical regeneration parameters such as: acid concentration; time frames for acid washing and water rinsing; and temperature and residence time for thermal regeneration. Carbon quality control frequently comprises only sporadic activity tests performed on non-representative grab samples dipped from conve­niently accessed carbon storage bins. So, regenerated carbon quality may vary and (at times) be unacceptable for effective hydrometallurglcal circuit operation. To verify the efficacy of the regeneration process and assure accept­able adsorption and elution rates for Au(CN)2· and Ag(CN)2·, operators should systematically assess regenerated circuit carbon quality. This paper, then, suggests methods for effectively assessing circuit carbon quality, Including: carbon sampling procedures; standardized physical and chemical test methods; and criteria and general procedures for comparative adsorption rate (activity) testing. Also, a two tier quality control testing scheme is proposed, and techniques for presenting and interpreting test data are demonstrated.

Jan 1, 1992

By John W. Rushton

Maintenance and inventory are usually the last items considered in mine development. They need to be addressed properly and early. The micro-computer brings sophistication within the reach of even the smallest mines. Maintenance and inventory programs, if properly installed, will reduce costs 15 to 25% and allow management time to address the mining and metallurgical problems at start up. Shops and warehouses are rarely designed properly. Large engineering firms spend 5 to 20 times more than is necessary for first class facilities. There is an optimum design, size, and layout. Take the time necessary to assure you get the right people, programs, and facilities in that order. Maintenance, inventory, and inventory control must be addressed early if a small mine is to have a chance at optimizing costs during the first two years of operation. This is hard work and is usually put on the back burner until the mine is in some degree of trouble. Hiring an experienced maintenance superintendent is important but it is not the complete solution. Maintenance systems, design of facilities, and use of computers are not skills that maintenance people usually have. In addition, the work of ordering spare parts and working with erection people is a full time job and any maintenance people on site during construction are quickly recruited. The end result is that few mines start up with an adequate Maintenance or Inventory System in place. The months, and in some cases years, that the mine attempts to operate without a system are both painful and expensive. Designs of successful integrated maintenance and inventory programs are virtually 100% supported with some outside assistance. Our system is the approach that we are most familiar with and this paper will isolate how Rushton International would go about setting up systems and designing support facilities. Micro-computers are the only way for a small mine to go. ~ini-computers cost more originally and operating costs are many times higher than even the most sophisticated micro system. Micro- computer software is developed further and recent advances in hardware makes medium priced units equal to most mini&apos;s in RAW COMPUTER POWER. Mines in general, and especially small mines, have no need for main frame power. MAINTENANCE A good maintenance program has: 1. Equipment Master List a. Equipment Detail Reports 2. Work Order and Backlog Control Program 3. Component Control Program 4. Tire Program 5. Wear Materials Program 6. PM Control Program 7. Maintenance History Program 8. Performance Program We also track miscellaneous items such as MSHA training and inspections, timekeeping, etc. Rather than going through the program step-by-step, it makes more sense to go through a typical Maintenance Day and show how the programs are used. The first thing maintenance supervision needs to know is "how we are doing." We

Jan 1, 1987

By Rami Abousleiman, Gabriel Walton, Sankhaneel Sinha

"Although conventional coal mine designs are conservative regarding pillar strength, local failures such as roof-falls and pillar bursts still affect mine safety and operations. Previous studies have identified that discontinuous, layered roof materials have some self-supporting capacity. This research is a preliminary step towards understanding these mechanics in coal-measure rocks. Although others have considered broad conceptual models and simplified analogs for mine roof behavior, this study presents a unique numerical model that more completely represents in-situ roof conditions. The discrete element method (DEM) is utilized to conduct a parametric analysis considering a range of in-situ stress ratios, material properties, and joint networks to determine the parameters controlling the stability of single-entries modeled in two-dimensions. Model results are compared to empirical observations of roof-support effectiveness (ARBS) in the context of the coal mine roof rating (CMRR) system. Results such as immediate roof displacement, overall stability, and statistical relationships between model parameters and outcomes are presented herein. Potential practical applications of this line of research include: (1) roof-support optimization for a range of coal-measure rocks, (2) establishment of a relationship between roof stability and pillar stress, and (3) determination of which parameters are most critical to roof stability and therefore require concentrated evaluation.INTRODUCTIONRecent studies on pillar design have suggested that the assumptions utilized in traditional methods such as tributary area theory (TAT) (i.e. overburden has no self-supporting capacity, infinite pillar array, and CHILE material properties) are invalid in many geo-mining conditions (Esterhuizen, Mark, and Murphy, 2010; Frith and Reed, 2017). Although TAT is conservative with respect to pillar loading, and mine fatalities have steadily decreased since 1990, roof-falls and pillar bursts still occur and injure hundreds of miners per year (Mark, Pappas, and Barczak, 2011). Even though large-scale global mine failure has not occurred in a US coal mine since 2007, the mechanics of global and local stability as governed by the relationship between geologic parameters, mine geometry, and engineering design are still not well-constrained.Numerical methods are a highly effective tool that can be used to gain further insight into these relationships. When using such an approach, calibration to existing empirical relationships is imperative to ensure that numerical model results are realistic. In this study, we utilize the discrete element method (DEM) approach in Itasca’s Universal Distinct Element Code (UDEC v. 6.0) (Itasca, 2010).DEM accounts for explicit separation and contact of individual deformable blocks (Jing, 2003) and captures how discontinuities can affect system behavior. Discontinuous and anisotropic rockmass properties are typical features in coal-measure rock which contain various horizontal (bedding) and sub-vertical (joints) fractures depending on their depositional environment, diagenesis, and historic and current stress states."

Jan 1, 2019

By T. P. Lim, J. A. Cherry, A. J. Vivyurka, R. D. Blair

The first phase of an investigation of the chemical composition of subsurface water and the hydrochemical processes in and near the abandoned Nordic tailings at Elliot Lake, Ontario, was conducted in 1979. The Nordic area, which covers 85 hectares with an average thickness of 10 m and overlies deposits of permeable glaciofluvial sand, contains 15 million tonnes of tailings with 4-7% pyrite. Nests of standpipe piezometers installed at three locations on the tailings indicate that seepage through the tailings is predominantly downward. The water table in the tailings slopes from the north side of the tailings dam, where it is within a metre of surface to the south side near the tailings dam, where it is approximately 10 m below surface. When they were deposited during the period from 1957 to 1968, the tailings contained pore water at a pH of about 8 resulting from neutralization treatment in the mill. At the three piezometer nests, the pH increases from 3 near the water table to 7 near the bottom of the tailings. The pore water in the tailings is gradually becoming acidic because of oxidation of pyrite in the zone above the water table. The concentrations of Fe, Mn, Ni, Co, Zn and Pb are high in the shallow acidic zone and much lower in the deeper neutral zone. Concentrations of 226Ra range from 30 to 230 pCi/L. At the three sites concentrations of major ions in the permeable sand a short interval beneath the tailings indicate 226Ra presence of tailings seepage, but Ra concentrations are only 2 to 8 pCi/L. Greater downward penetration of 226Ra probably occurs in the area closer to the tailings dam. Monitoring of a network of multilevel bundle piezometers installed in permeable sand that extends southward from beneath the tailings indicates the occurrence of a plume of tailingsderived water containing high concentrations of Ca, Mg, Fe and S04 that extend 400 m downgradient from the tailings dam. The total dissolved solids decline from 20,000 at the dam to less than 500 mg/L at the identifiable downgradient periphery of the plume. A much smaller but distinct plume of water with high heavy metal concentrations and 226Ra in the range of 10 to 130 pCi/L extends from beneath the dam for a distance of 15 m. The pH of the groundwater in this heavy metal-radium plume is 4 to 5, whereas elsewhere in zone of groundwater with high dissolved solids, the pH is above 6. The results of this investigat, suggest that the heavy metals an 226 Ra are relatively immobile in the perme-

Jan 1, 1980

By John E. Frost

Duval&apos;s (now Battle Mountain Gold Co.) copper and gold deposits in the Battle Mountain District in Nevada are in two subdistricts - Copper Basin and Copper Canyon - about 18 km (11miles) apart. Location of the two subdistricts is shown in Fig. 1. The two centers of mineralization have distinctive geologic and ore characteristics that required different innovative approaches to exploration and development. Nearly all of the copper from Copper Basin has been recovered by heap leaching. Because of relatively high gold and silver values in the Copper Canyon ores, a significant part of the sulfide copper ores has been recovered by a conventional flotation mill at Copper Canyon. Recently, this mill has been converted to the recovery of gold that is mined from zones peripheral to the copper deposits. Prior to Duval&apos;s entry, small amounts of copper, gold, silver, lead, zinc, and antimony had been mined in the district over a nearly 100 year period, beginning in 1866. The bulk of the $13 million total production to 1962 was derived from copper and gold lode ores and gold placers. Duval&apos;s interest in Battle Mountain-arose directly from a submittal by Asarco, which controlled key claims in both the Copper Basin and Copper Canyon subdistricts. Asarco acquired its holdings in the district in 1959 and undertook limited exploration until mid-1961. At that time, the company decided to dispose of its interests. Duval was invited to evaluate Asarco&apos;s properties and accepted. On Dec. 6 and 7, 1961, J.E. Frost, R.C. James, R.W. Sayers, and H.A. Schmitt, representing Duval, and K. Richard and E.W. New- man, representing Asarco, visited Copper Basin and Copper Canyon. Frost concluded that the properties were of considerable interest and prepared, by Jan. 12, 1962, a voluminous report recommending that Duval acquire and explore Asarco&apos;s properties. Furthermore, he recommended a large initial exploration program that included not only geologic mapping, geochemistry, and geophysics, but also the extensive drilling that was to involve, as the first phase of exploration, 49 drill holes in Copper Basin and 21 drill holes in Copper Canyon. Following property agreements between Duval and Asarco, and Duval and several other claim owners, geologic mapping, geochemical surveys, and drilling began in August 1962. Exploration drilling was completed in July 1964 and engineering studies in April 1965. Development of the Battle Mountain properties was authorized in July 1965, and production commenced in the first quarter of 1967. Because of the unusually high oxide copper content of the ore bodies, particularly in Copper Basin, a much larger than typical proportion of the copper in the deposits has been recovered by heap leaching. The generalized geology and ore zones of the Copper Basin are as shown in Fig. 2 and of Copper Canyon in Fig. 3.

Jan 1, 1991

By C. Anderson, B. Schug, C. Cristoffanini, S. Nazari

The Haile Gold Mine in South Carolina is increasing the automation level of their processing plant in order to improve operational stability. One step towards this goal is the application of a Digital Twin concept to provide Virtual Instruments. Haile has engaged ANDRITZ to implement Virtual Instruments at their grinding area hydrocyclones and at their tailing thickener underflow pump. These virtual instruments provide density values in a traditionally difficult location for reliable slurry density measurement. The Digital Twin is a first principles IDEAS1 ™ simulation model, connected in real time to the control system. Results have been promising, with the virtual instrument values matching the physical instrument under normal operating conditions, and showing better prediction under upset conditions. Haile plans to use the virtual instruments as optional process control parameters. INTRODUCTION OceanaGold’s Haile Gold Mine is located near Kershaw, South Carolina, about 50 miles southeast of Charlotte, North Carolina. This location was first mined in 1827, and has produced gold through several operating periods since then, including a heap leaching operation from 1985 through 1992. In October, 2015, OceanaGold acquired Romarco Minerals, the previous owner of the Haile Operation. The current operation began in 2017 as an open pit mine with additional underground resources. The nameplate throughput is 6,350 tpd ore, and 2017 production was 118k oz gold, with 140 to 155k oz gold expected in 2018. Proven and Probable reserves exceed 3 Moz. The process, as shown in Figure 1, includes a primary jaw crusher, comminution with a Semi-Autonomous Grinding (SAG) mill and ball mill circuit, flash flotation, rougher flotation, regrinding, pre-aeration, carbon in leach (CIL) leaching, carbon elution and regeneration, cyanide recovery, electrowinning and refining. The Haile Operation’s inaugural gold pour took place January 20, 2017. Mine life is expected through at least the year 2030. Based partially on prior experience of some key OceanaGold personnel with automation services from ANDRITZ on other projects, OceanaGold decided to include ANDRITZ on the Haile project. The scope of these services for Haile included design and supply of the process control application software, an ANDRITZ dynamic process simulator for operator training, and ANDRITZ Advanced Control Expert (ACE) ®2 systems for the SAG mill, ball mill, and rougher flotation areas. Since startup, ANDRITZ has continued to provide services for the site, particularly related to PLC and HMI equipment, programming and support, ACE system expansion, and the implementation of Virtual Instrumentation. The implementation of this new Virtual Instrumentation technology at Haile is consistent with Haile goals: increased automation, a change in the role of the operator from interactor to monitor, and improved stability by removing human error.

Jan 1, 2019

By L. R. Duncan, W. H. McCracken

Magnesium is the eighth most abundant element in the earth&apos;s crust and the third most plentiful element in seawater. It is found in more than sixty minerals and in brines and seawater as a magnesium salt. One of the more important magnesium minerals is magnesium carbonate, MgCO3, with a theoretical maximum magnesia, MgO, content of 47.6%; this carbonate form represents the world&apos;s largest source of magnesia. The next most used sources for magnesia are magnesia-rich brines and seawater, where it occurs as magnesium sulfate and magnesium chloride. Other commercially important magnesium-bearing minerals are dolomite, CaMg(CO3)2, which serves the aggregate industry as well as the chemical industry; brucite, Mg(OH)2, which is used in the production of both caustic and sintered magnesia; olivine, ([MgFe]2-SiO2), which serves the refractory and heat storage industries; talc, (H2Mg3(SiO3)4), which serves several industries as a filler and as an ingredient in cosmetics; and serpentine, (H4Mg3SiO9). All these minerals are of commercial value because of the desirable characteristics given to them by magnesia and by its placement in the crystal structure. These minerals are the starting raw materials for a wide range of products. These include the production of magnesium metal, and several grades of magnesia used for the production of both sintered magnesia for refractory manufacture and lighter fired caustic magnesia, the latter is used in fluxes, fillers, insulation, cements, decolorants, fertilizers, chemicals, in the treatment of waste water including pH control, and in the removal of sulfur compounds from exhaust stacks. Magnesite is one of several major minerals possessing a dual character, i.e., non-metallic industrial mineral because of its principal end use as a calcined/dead burned raw material in non-metallic end products and magnesium metal. A large percentage of the magnesia used to produce magnesium metal comes from seawater1 brines. The total primary metal production worldwide is about 350 ktpy. This report reviews magnesite solely as an industrial non- metallic raw material. Magnesite for most refractory purposes must have a content of sintered magnesia above 85%, with a preference for those sources that have the potential for 90 to 98% magnesia values. Dolomite, used in large tonnages because of its physical properties in the construction industry, also has a use in the chemical and the refractory industry. In its calcined form, dolomite is used as a slag addition in the making of steel, and it is used in a high-purity calcine to precipitate magnesium hydroxide from brines or seawater. The precipitated magnesium hydroxide is intermediate for the production of deadburned clinker for refractories and the production of magnesium metal. In many cases the high-purity magnesium hydroxide obtained by precipitation with calcined dolomite, or in some cases limestone, has supplanted much of the world&apos;s production of magnesia from the natural mineral magnesite. This is especially true of the United States and in major industrialized areas of Europe and Japan. The best known of the minerals directly and widely exploited for its magnesia content is magnesite (MgCO3), one of the calcite group of rhombohedra1 carbonates, which includes calcite (CaCO3), siderite (Fe2CO3), and rhodocrosite (MnCO3), among others. The members of this group enter into a wide range of substitutional solid solutions when the positive ions have similar radii. The radii of magnesium and iron ions are within 6% of each other; hence magnesite and siderite form a complete series, of which the mineral breunnerite (ferroan magnesite) is a well known end member. The radii of calcium ion is 36% larger than that of magnesium ion and only limited substitution exists at each end of the MgC03-CaCO3 series. Dolomite is not a member of the calcite group, but it occurs when calcium and magnesium ions alternate in equal number in an ordered structure among carbonate ions. The result of this relationship is that calcite and dolomite are commonly found intermixed with magnesite. They occur, commonly, as identifiable crystal entities, which can be separated to a varying degree from the magnesite by benefication techniques. Magnesite, when pure, contains 47.8% MgO and 52.2% CO,. The pure mineral is found occasionally as transparent crystals resembling calcite. Magnesite principally contains variable amounts of the carbonates, oxides, and the silicates of iron, calcium, manganese, and aluminum. Although the genesis of natural magnesite deposits can be complex, it is distinguished in nature in two distinct physical forms, namely crystalline, (with a wide range of visible crystal sizes) and cryptocrystalline, sometimes referred to as amorphous, where the crystal size is not detectable to the eye and will range from 1 to 10 pm. The two types not only differ in crystal structure but in the sizes of the deposits and in modes of formations. The crystalline form has a Mohs hardness of 3.5 to 4.0. The color range is from white to black with shades of yellow, blue, red, or gray. The color is not a significant indicator of purity but in a given deposit an experienced person can often roughly grade the magnesite by observing color and crystallinity. Macrocrystalline deposits occur in relatively few, but generally large, deposits on the order of several million tons. The ore shows a marble-like crystal structure and belongs to the sedimentary or metasomatic groups of origin. The cryptocrystalline variety of magnesite frequently occurs in many small deposits, although there are exceptions. Cryptocrystalline magnesite is typically massive with no cleavage and is sometimes descriptively called bone magnesite. The fracture is usually conchoidal, with a hardness of 3.5 to 5.0. The color is mainly white, but it can have tints of yellow, orange, or buff. Accessory silaceous minerals such as serpentine, quartz, or chalcedony are usually present. Calcium minerals are normally absent or in low concentration, this contrasting with the macrocrystalline form where calcium minerals are almost invariably present and some- times in high concentration. The specific gravity of cryptocrystalline magnesite ranges between 2.90 and 3.00, whereas the value of pure crystalline magnesite is 3.02. In actuality, the normal specific

Jan 1, 1994