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By I. Mikov

Established in the 60-ies, the Zlatna Panega quarry is a cement raw materials quarry supplying 95% of the needs of the Zlatna Panega Cement Plant with limestone, marl and other silica-aluminates. The two major types of raw materials are formed by accordingly dipping strata of Upper Jurassic limestones and lower Cretaceous marls. Between them there is a ?transition-zone? or ?contact-zone? of a high-silica content layer, about 5-m thick, defined as ?flint? by investigations including spot sampling throughout the quarry. For many years this ?contact-zone? layer was neglected due its negligible thickness. Further, with the increasing demand on the quality of the raw materials feed to the Plant and the focus on variability of the chemical and mineralogical components, it was decided not to mine any of the parts of the quarry contaminated by this material, for about three years. The latter had an impact on the availability of operating benches that could result in a substantial decrease in the total recoverable re-serves of Zlatna Panega, particularly with respect to marls. There were considerations for resorting to a partial, selective excavation of the quarry pit with a reduced lifetime for the lime-stones and re-opening of an other (currently in-active) marl quarry at a distance of 6 Km from the Plant. The Quarry Team of the Zlatna Panega quarry was assigned to study the problem. The ?contact? layer was duly confirmed with systematic geological (including mineralogy) studies, statistical analysis of available data, surface sampling and shallow exploration drills, resulted into distinguishing a third type of reserves and their credible subsurface delineation by means of geological 3-D modeling and quarry design. A new quarry design was finally elaborated with a view of providing flexibility in mining, utilization of the total volume of all available reserves and maximizing the mine life of Zlatna Panega quarry. An industrial test took place in December 2005, by using the contaminated parts of the quarry (co-mining of the ?contact-zone? materials) at a contribution of 15-20% in the raw materials mix, as dictated by their contribution in the total recoverable resources of the quarry. The purpose was to investigate the viability of this solution ? in line with improved quarry planning ? without compromising the quality of cement production and the efficiency of Plant Operations. Results were overall successful.

Jan 1, 2006

By Jean E. Kiel, Rob Dorey, Dirk J. A. van Zyl

1.1 INTRODUCTION This chapter seeks to provide an overview of important factors involved in the implementation of heap leach technology as a vehicle for gold extraction from its ores. Toward this end, brief discussions of the various elements important to this process are presented here. Topics covered in this chapter include a brief history of the process; basic chemical reactions responsible for the process ; characteristics of ores amenable to heap leaching; location of the greatest concentration of heap leach projects in the U.S.; heap leach components, methods, regulatory/permitting considerations; and economic considerations. Most of these subjects are treated at much greater length in subsequent chapters. Thus, this chapter serves as a summary for those which follow, in addition to providing the reader with a feeling for those topics which may be of most immediate interest. 1.1.1 History of the Process Modern Day Precious metal heap leaching technology has developed over the past decade, although the principles of heap leaching per se, as well as those of gold extraction through cyanidation have a long history (see discussion below). However, it is in the past 10 years that heap leaching has developed into an efficient method of treating oxidized gold and silver ores. It has proven to be both an efficient way to extract precious metals from small, shallow deposits, as well as an attractive way to treat large, low- grade, disseminated deposits. Heap leaching has several advantages compared to conventional milling (i .e., crushing, grinding, and agitation leaching). In general, these advantages include simplicity, lower capital and operating costs, shorter start-up times, and less intensive environmental regulatory concerns. A possible disadvantage of currently existing heap leach technology is a potentially lower percentage of metal extraction being obtained from the ore than would be the case with conventional milling (Potter, 1981; Hiskey, 1985; and Thorstad, 1987) The actual principle of heap leaching has a long history as inferred previously. For example, mines in Hungary recycled copper-bearing solutions through waste heaps in the mid-sixteenth century (Hiskey, 1985), and Spanish miners percolated acid solutions through large heaps of oxide copper ore on the banks of the Rio Tinto around 1752. By 1900, leaching operations were employing such techniques as leach/rest cycles to maximize copper recovery (Taylor and Whelan, 1942; Thorstad, 1987). Copper dump leaching is currently

Jan 1, 1988

By John A. Lombardi

The solutions used in the heap biooxidative treatment of sulfidic refractory Au ores and concentrates generally load free mineral acid and metals and, in time, become bacteria activity inhibiting due to increased acid and/or metals concentrations. The metals­acid loaded, bacteria action inhibiting, solution must then be: (a) processed, prior to re-use, to reduce its metals and/or free acid mineral acid components to bacteria activity compatible levels; or (b) neutralized prior to discharge. The most common solution processing method to adjust acid-metals contents is liming to effect the precipitation of dissolved metals and neutralization of acid sufficient for solution discharge or re-use. The lime treated solution is high in suspended, colloidal precipitates and requires clarification before reuse. Most often the "bleedstream" of biooxidation solution that is lime treated is discharged and req­uisite volumes of fresh water added to the process inventory. Depending on the level of bacterial action in a heap and the acid consumption characteristics of the ore/concentrate being biooxi­dized, the bleedstream taken from the circulating, irrigation on­off, solution, can be large (e.g., 10%-20% and sometimes as much as 50% of the daily solution flow requires treatment-dis­charge). The lime charge for bleedstream treatment is commen­surately large and represents a major operating cost of the bioox­idation process. The patented HW Process Technologies (HWPT) Engineered Membrane Separation (EMS™) method is an alternative biooxi­dation bleedstream treatment method. EMS™ processing removes the multivalent base metals ions in an acid solution to a small volume EMS™ "concentrate" (typically 25% of the feed­stream volume). The EMS™ method coincidentally makes a "clean acid" EMS™ "permeate" (typically 75% of the feed­stream volume). In general, EMSTM processing enables 75% of the bleedstream volume to be recycled; allows up to 2/3 of the bleedstream acid to be recovered as a 10% solution; and delivers dissolved metals in a highly concentrated stream suited to a. metals recovery or b. reduced cost lime neutralization for a given metals precip­itation relative to full bleedstream neutralization (typically a 30% lime use savings). Implicit in the recovery of process water for re-use and reduced reagent addition requirement for metals precipitation is a significant reduction in wastewater-sludge dis­posal facility size.

Jan 1, 1998

By P. J. Hyde

Cyanide is a significant contributor to the cost of producing gold. Any reduction in consumption that can be achieved through more reliable measurement and/or more efficient management of dosing rates can have a significant financial impact. An automated analyser, by its mechanical nature, is not prone to the variations of human operators, particularly notice­able from shift to shift. Reducing the standard deviation of the measurement (ie, reducing noise ), the automated measurement will give the metallurgist a truer picture of the underlying patterns of usage. The combination of automated analysis with on-line operation and a control loop provides timely information that is efficiently utilised in the management of the plant. Until the release of the Chemtronics On-Line Cyanide Analyser (OTA3000) about 12 months ago, the Australian experience with automated analysers (especially on-line) had been substantially negative. Since its release, the Chemtronics instrument has outperformed both manual titrations and other automated instruments, gaining an enviable reputation for accuracy and reliability in the process. The instrument is built around the technique ofpotentiomet­ric titrations, in which the potential of an indicator electrode is measured as a function of the volume oftitrant that is added to the sample. As the cyanide in the sample is depleted, the potential on the indicator electrode changes rapidly, producing a typical sigmoidal curve. The OTA3000 takes the first differ­ential of the curve, identifies the apex of the peak as the endpoint and reports the result in parts per million of sodium cyanide. The OTA3000 accepts the peak as valid only if it exceeds a certain minimum height. That minimum height is a programma­ble parameter and is used to optimise the relationship between sensitivity and the rejection of spurious peaks. The instrument will track cyanide trends through both positive and negative slopes of any magnitude, but it will provide a warning if it the slope becomes excessive, not because it affects the instrument but because it means that something unusual is occurring in the tanks. There are a limited number of conditions under which the instrument will produce a result of zero. In these cases the OTA3000 holds any output lines at the last valid result while it repeats the titration. This feature was incorporated for reasons of stability in a feedback loop.

Jan 1, 1995

By Ken Deter

I have some general comments on optimizing profitability based on working in heap leach operations in tropical environments. Many of my planned comments this morning have been stated earlier in the Conference. I get the opportunity to say only a few words, and follow up on some specific comments and concerns. In normal operations, stack the ore higher, where possible. The surface area of the stacked ore controls the total solution volume that can be sprayed on the pad at a specific time period. With higher lifts, solution will contact more ore particles as it percolate through the lift, and therefore the solution grade flowing through the base will be increased, whether it is freshly stacked or almost completely leached ore. This also allows more tonnage to be stacked per lift on the pad, decreasing conveyor movement and downtime. Increasing from 6 to 8 meter (20 to 26 feet) lifts have worked successfully, and some mines are using 10 meter (33 feet) lifts in 10,000-15,000 dmtpd plants. The same concept will work with smaller plants, but initial 6 meter lifts should be used until the start-up period has shown the leaching is working as planned. As the stacked ore lifts get higher, the solution flowing through the highest lift is diluted by solution retained in the lower ones. Eventually no real increase is noted from putting fresh ore in service, and gold production rate suffers. To address the dilution problem, an HDPE liner should be installed between each 2 to 4 lifts, depending on each site specific lift height, recovery, % slump in the lift, recovery rates, and grade treated, among others. The surface of the previous lift should be leveled and sloped to control the effluent draining from the planned lift, then the liner is placed, including berms and discharge pipes. The intermediate HDPE liner does not need to be completely sealed as would be the case with the base liner, and can accept some solution penetrating into the lower lifts. A one-mm (40-mil) thickness liner can be used. This intermediate liner allows the upper lift(s) higher-grade effluent to report to the pad discharge point. This will maintain higher gold solution grades going to the ponds. The slopes of the previous lifts should not be covered with HDPE in this concept, and precipitation falling on the lower lift slopes and benches can be directed to barren or surge ponds.

Jan 1, 1999

By James W. Babcock

Introduction Geologic ore deposit models are key to any exploration program. In the search for ore deposits, the geologist first tries to understand the genesis of known ore bodies. During reconnaissance and evaluation of a prospect, a valid exploration model of a known ore body can help the explorationist focus on critical geologic attributes of the target area. Exploration models can conserve exploration funds that might otherwise be spent collecting data not critical to delineating an ore body. This article examines the benefits of using models, factors to be considered in developing models, and examples of the way recent models have evolved. New exploration concepts and techniques are developed in response to increased competition in the mining industry. New techniques can give an exploration group a slight edge that may translate into a discovery. Mine geology staffs may find previously unlooked for extensions to the ore body by applying new genetic concepts. When all appropriate techniques are applied to exploration and mining efforts, greater efficiency and increased profits may result. Before the 1950s, scientists speculated on the genesis of ore deposits and other natural phenomena. But they presented their results as if universal and final, even though they applied to only a few deposits. Because it did not provide sufficient criteria to distinguish the genetic processes, the classic conflict between hydrothermalists, epitomized by Lindgren, and nonhydrothermalists did not help the minerals explorationist much. Important information about individual ore deposits, such as age of the host rocks and mineralization, composition of the mineralizing fluids, and separation of mineralizing and nonmineralizing events in the paragenetic sequence, could not be gathered with past equipment and methods. But these early concepts of ore genesis contained excellent empirical observations, perhaps the most useful parts of these hypotheses. Through the years, prospectors have found most of the ore deposits in the world using direct observations and their collective empirical wisdom - an empirical model. The geologist's contribution becomes increasingly important after the more obvious surficial ore deposits have been discovered. In Europe and North America that time has arrived. With the introduction of sophisticated equipment and matching techniques in the 1950s in geochronology, geochemistry, geophysics, remote sensing, and computers, many previously equivocal genetic questions can be resolved. This proliferation of highly technical data has allowed the construction of many well-founded genetic models.

Jan 12, 1984

By Deepak Malhotra, William F. Riggs, Edward C. Dowling

INTRODUCTION Mining companies are In the business of making an acceptable profit on their Investments, It became obvious in the recent recession that economic profitability/viability depends on being the lowest cost producer In order to compete on a global basis. The burden of achieving this objective falls on the mineral manager. He must strive to produce a salable product in an environmentally acceptable process while continuously Improving the profit margln.(1,2) Metal prices fluctuate on a regular basis. They directly Impact revenue and hence operating philosophy. The mineral engineer needs to develop an optimum operating strategy to meet his overall objectives, One would expect that, In the modem era of complex computer models, it would be easy to develop an economic model to maximize profits and/or cash flow. Discussions with mill managers Indicate that one mine/corporation's optimum operating strategy is not necessarily optimum for another. This paper approaches the problems from basics "Economics 101" and Illustrates the complexity of developing a universal model which can be applicable to all mines. Case examples are used to illustrate the need for better communication to avoid the basic economic fallacy “that optimization of segments leads to optimization of the whole system.” TECHNICAL MODELS Technical models are a prerequisite to the development of economic models because they provide the necessary Input data to these models, The flow of material from a mine through the sale of the finished product is shown In Figure 1. The ore Is mined and milled, and the concentrate can be stored for sale or smelted and refined In-house or at a toll facility. The finished product(s) are held In Inventory until sold either through existing contracts or in the spot market. [ ] Fairly reliable technical models have been developed for mining, milling and smelter operations. Mine plans are developed to provide a good Indication of the relationship between production and cut-off grade. It is easier to change surface mine plans than plans for underground mines. However, there are limits to maximum production that can be achieved with existing equipment. It Is possible that mine benches may not be developed to handle additional equipment, even If It was purchased to enhance production. More often than not, the cut-off grade Is lowered to Increase throughput goals.

Jan 1, 1993

By Robert L. Sundeen, Richard V. Wenberg

INTRODUCTION Shotcrete can be defined as spray-applied concrete with a maximum size aggregate in excess of 9.5 mm (0.375 in.). The concept evolved from guniting, which uses smaller aggregate and generally produces concrete with poorer physical properties. The larger aggregate used in shotcrete requires less water in the mix, while offering larger bonding surfaces, thus imparting greater tensile and compressive strengths. TYPES OF MACHINES Two basic designs of shotcrete machines in general use are the wet or prehydrated zero slump and low slump type, and the dry or nozzle hydrated machines. Fig. 1 shows a typical zero-slump machine in a con¬figuration commonly used for small-volume require¬ments. The prehydrated zero-slump machine mixes exactly proportioned batches in a sealed pressurized chamber. "Slugs" of hydrated shotcrete mix and com¬pressed air are sent through a distribution hose to the nozzle where the wet mix emerges as a spray. The spray is directed onto the rock surface by the nozzle operator. The low-slump machine is similar, except that a thick slurry mix of aggregate is delivered throughout the distribution hose to the nozzle. Fig. 2 shows a typical low-slump shotcrete machine used in moderate-volume applications. The nozzle-hydrated shotcrete machine accepts a premixed "dry" (4 to 6% moisture) batch of aggregate and cement. A constant feed of this mix is directed by compressed air through the distribution hose to the nozzle. At the nozzle, water for hydration is introduced into the emerging spray. The hydrated mixture is di¬rected onto the rock surface by the nozzle operator. Fig. 3 shows a typical dry-mix machine. The major difference between the two basic machine designs is the consistency of the shotcrete quality. Pre¬hydrated machines have high degrees of consistency and produce high-strength concrete, resulting from their ability to produce exact mixtures of zero or low-slump concrete within the machine. Nozzle-hydrated machines tend to produce poorer quality concrete because of the difficulty in controlling the water content at the nozzle. Water input essentially is an art. As such it is not precise, and the shotcrete produced is of variable quality. MATERIALS HANDLING Materials handling presents a problem to either type of shotcrete machine. At times considerable quantities

Jan 1, 1982

By B. H. Kaye

Particle-size Distribution Specifications for Particulates. Feed to and finished products from size-reduction operations are defined in terms of the particulate sizes involved. It is also well to know whether the ultimate individual particle is being measured, or, if any aggregation or agglomeration of particles exists, whether this has been created by the size-reduction operation. The most complete description of a powder is given by its particle¬size distribution. This can be plotted in terms of cumulative percent oversize or undersize in relation to the diameters of particles, or it can be plotted as a distribution of the amounts present in each unit of diameter against the several diameters. It is common to employ a weight basis for percentage, but there are some data in the literature where frequency, or number of particles, is used. The basis of percent¬age, whether weight, frequency, or some less commonly used factor, should be specified, as also should be stated the diameter, its units, and preferably whether determined by sieve, settling velocity, or other¬wise. Fig. I presents two sets of distributions-ore cumulative and the other in unit intervals. The slopes of the 5-µ intervals of the cumulative curves are converted to percent per micron and plotted as a block, or histogram, from which smooth curves are derived. Powder A has a narrower, or tighter, size range for the bulk of its weight than powder B. Both materials have the same weights below and above the size marked by the arrow. Simpler treatments of distribution are possible. In some cases the significant value is the top size. Since the 100% point is dubious, some amount, as 95 or 98%, so specified, can be called top size. In other cases the percent on some sieve helps define coarseness. Merely stating that all, or all but a small percentage, passes a given sieve is inadequate for defining the true fineness or a material. Complete parti¬cle-size analysis to show the distribution is essential for most compari¬sons and calculations. Particle-Size Equations. A number of equations have been pro¬posed to correlate the quantity of a particulate material with its parti¬cle size to obtain a distribution relationship. 's 19 In the literature it is often assumed that a powder must follow some distribution, such as the Rosin-Rammler-Bennett:20, 21

Jan 1, 1985

By D. J. Hutchinson, S. Nickson

Cable bolts are used as a means of rock support around underground excavations in many mines across Canada and around the world. A substantial body of research on cable bolting was conducted at a number of Canadian universities and mines from the mid-1980s through the 1990s. Since then, research into selected areas of cable bolting practice has continued in Canada. The research has focussed on cable bolt instrumentation and on the inclusion of cable bolts in empirical stope design techniques. A review of cable bolting practice is now warranted at many mines. It has been five years since the completion of the previous substantial body of Canadian academic research and the associated advancement in understanding of cable bolt design, implementation and verification practice. Attempts were made at that time to transfer the new design procedures and performance assessment technologies to mine-site personnel. In some cases, this was effective, and ground control personnel continue to improve on their cable bolting practice at each specific mine site. However, at other sites, experienced people have moved on, and the new, less experienced personnel do not have access to the training or technical resources required to improve their understanding of the subject. In other cases, the technology was not transferred effectively to practitioners, with the result that substantial improvements can be made to the cable bolting practice at those mines. Guidelines for auditing cable bolting practice are provided in this paper. The guidelines include a discussion of the objectives of the audit, suggestions for audit procedure and a summary of the possible outcomes of a cable bolt audit. The objectives of an audit might include assessing the reasons for unexpected cable bolt failures, evaluating the quality of the installation procedures, quantifying the efficiency of the cable bolt system and assessing training requirements and procedures. A cable bolt audit should include consideration of all aspects of the cable bolting cycle, including the design, implementation and verification phases of the operation. At the end of the audit, a ranking should be assigned to the components of the cycle. The ranking system can then be used to identify items that require immediate attention, others that could be improved and others that are adequate. Following completion of a cable bolt audit, increased confidence in the support system and its application is generally, found, along with the potential for cost savings in many cases.

Jan 1, 2002

By William A. Turcotte

Introduction The need for water is fundamental to every minerals processing operation. Though this may sound rather elementary, it is one area that cannot be over- looked. Frank C. Quinn in his paper entitled "Water Is Water Is Water" (1967) said, "but it isn't as simple as that. Very often research is slowed down and production is hampered because of a lack of appreciation for the behavior of water." The quantity of water needed for any operation is determined from bench and pilot plant testing. The quality required by the process for both fresh and reuse water is also established during these two testing phases. Quinn's statements are true whether the process involves the paper industry, the minerals industry, or any other industry. It is not the author's intent to cover water quality, quantity, or re- cycle in specific terms, but to provide the reader with a general outline of the topic with reference to several valuable sources of information. The first part of the chapter will cover several of the areas that have been de- tailed by others and also provide a list of reference sources. The references at the end of the chapter are not all inclusive but will provide a good starting point. In the second part of the chapter, a detailed description, with flowsheet illustrations, is pro- vided on the evolution of the reclaim (reuse) water system at the Tilden concentrator. Mineral Industry Water Needs Beresford and Edler (1978) surmnarized the minerals industry's need for water this way: "minerals are necessary to a modern complex technological society. The minerals industry, while its water needs are relatively minor compared to large users such as agriculture, nevertheless requires a large quantity of water to carry out daily operations from mining to final product refining." In 1973 total water use in the United States was estimated at 15.14 billion cubic meters (4 trillion gallons) for the minerals industry. Of this, about 69% was recirculated within the system. In 1982 Watt and Horton stated that the percent of recycled water in the minerals industry had improved to 73%, showing the minerals industry is an efficient user of water and that steady improvement in water recycling is something the minerals industry can look to with pride. The percent of water needed by the United States' minerals industry was estimated in 1962 to be about 2% of the total water required by industry

Jan 1, 1986

By David P. Keller, George Podolsky

Iron oxides are unique in that they are the only significant colored mineral found in a natural state suitable for use as a pigment after being pulverized to pigmentary size. The world production of iron oxide pigments is estimated by the authors to be between 350 and 400 kt. The fact that highly colored natural deposits of iron ore are found throughout the world accounts for its use in prehistoric man's artistic cave paintings. These early artists did not realize that they had established man's first paint test which now has some 20,000 years of exposure. In addition to abundance, iron oxides (about 7% of the earth's crust) have the advantage of low cost, permanency, and are non- toxic. Through the centuries, succeeding civilizations have utilized iron oxides as a major source for decoration and protection when these properties were desired. In the last century the chemical industry has improved on nature by developing a complete range of synthetic iron oxide pigments which surpass the pigments produced from natural iron ores in uniformity, color quality, and chemical purity. In the United States, the combined production of natural and synthetic iron oxides produced sales of over $97 million in 1980 (Spimad, Jr., 1980), increasing to $138 million by 1990 (Templeton, 1990) NOMENCLATURE With the increasing importance of the synthetic iron oxides it is necessary to make a distinction between the natural or mineral pigments and the synthetic pigments. Natural pigments are those products that are derived from selected ores and should not be confused with iron ores mined for steel production. Iron ores that are mined for steel must be capable of being mined and reduced to iron on a competitive commercial basis. These ores are selected on the basis of iron content and processing economics. It is therefore unusual when iron ores for steel production are suitable for use as mineral pigments. Natural pigment ore sources are selected for their special physical-chemical properties and are able to command a premium price over iron ores used for steel manufacturing. Synthetic pigments, and in this instance iron oxides, are those pigments produced from basic chemicals. Through chemical synthesis pigmentary particles are produced as opposed to the use of comminution, the major procedure common to all natural iron oxide pigments.

Jan 1, 1994

By Gordon M. Pugh, David G. Rasmussen

INTRODUCTION The two unit operations in cut-and-fill stoping that would benefit most from increased mechanization are drilling blastholes and moving broken ore to the ore¬passes. Drilling in current practice is done with hand¬held jackleg and stoper drills. Broken ore is commonly moved from the face to the orepass by a small slusher. This type of equipment can be easily moved through the raises. Hand-held drills have a slow drilling rate and require one operator for each machine. Drilling can be done much faster by small jumbo-mounted machines as two or more drills can be operated by one miner. Load-haul-dump (LHD) units move broken ore from the face to one or more orepasses at a much faster rate with less loss and dilution than slushers. These units are mobile and may be used at more than one place during a shift. The problem which must be overcome in applying this higher degree of mechanization is the restricted means of access inherent in current practices. Cut-and¬fill methods are usually applied to mining veins that are narrow, irregular in dip and strike, or where ground con¬ditions require the support of sand fill. The mechanized equipment which is needed to accomplish high produc¬tion cannot be raised or lowered through raises without being disassembled. A jumbo which will give the desired performance would be mounted on a rubber-tired carriage with a diesel engine for propulsion. Twin hydraulic booms run by a small air motor would carry pneumatic drills. Over¬all dimensions of the unit would be about 8 m (26 ft) in length, 1.6 m (5'/z ft) in width, and 1.6 m (5'/z ft) in height. An LHD of the desired capacity would be rubber¬tired and either diesel or electric-powered. The overall dimensions of a 2.3-m (3-cu yd) unit will be about 8 m (26 ft) in length, 1.8 m (6 ft) in width, and 1.5 m (5 ft) in height. It will weigh about 12 700 kg (28,000 1b). It can readily be seen that conventional raises can¬not handle this size equipment. The most desirable means of access enables equipment to be driven into the stoping areas. Ramps can enter the stope by being driven upward in the vein walls or by being developed in the vein on top of the fill as the stope advances. This ramp system not only permits the use of rubber-tired equipment in multiple working places but also enables its removal for repairs without component breakdown. In the proposed method, as in conventional cut-and¬fill stoping, sand fill will replace the mined-out ore. By constructing an inclined timber floor at one end of the stope, a ramp can be maintained on top of the fill and advanced as the stope is advanced cut by cut (Fig. 1). The ramp is, therefore, developed during the stoping process and is only driven as a tunnel when rising over and around the sill and crown pillars. A stoping section consists of three blocks of different geometric shapes. The first and last blocks to be mined will be triangular in the plane of the vein, and the blocks between will have the shape of parallelograms (Fig. 1). The first block will develop the initial ramp by breast stoping and filling to the ramp floor (Fig. 2). The ramp will need to be completed through to the next level be¬fore stoping of the subsequent block can begin. The subsequent blocks are mined by back stoping with the next ramp being developed at the other end of that block (Fig. 3). The length of each subsequent block can be varied without changing the slope of the ramp. As the new ramp is developed, the previous ramp is filled along with the stope cut. Access up and down is maintained through the stope. The final block is again breast stoped as no new ramp will be developed at the end line for jumbo egress. Boreholes provide ventilation between levels as well as utility access. They can also provide a secondary escapeway for personnel during periods of development. These boreholes would be located about 3 to 4.5 m (10 to 15 ft) in the footwall and spaced about 183 m (600 ft) apart along the vein. The orepasses are spaced about 61 m (200 ft) apart in the vein. This spacing allows not only a short tram for the LHD unit but also provides surge capacity for the ore between haulage trains. These passes are raised from the haulage levels. As development progresses, the raises are driven through the sill pillar, and then car¬ried in the fill as stoping progresses. Ventilation and utilities can be maintained through an orepass when it is not needed for ore. The end line orepasses are intended for this purpose as well as for personnel emergency egress. Overall dimensions and criteria assumed to illustrate the ramp method of access in a highly mechanized cut¬and-fill stope are: haulage level spacing of 61 m (200 ft); ramp gradient of 17%; average dip of approximately 1.57 rad (90°); initial stope length of 358 m (1177 ft); subsequent stope length of 179 m (588 ft); average stoping width of 3 m (10 ft) ; sill pillar thickness of 3.6 m (12 ft); crown pillar thickness of 1.8 m (6 ft); bored raises of 1.5-m (5-ft) diam; and cribbed orepasses of 1.2 x 1.2 m (4 x 4 ft) through the sill pillar and 1.2 m (4 ft) square inside the cribbed portion. DEVELOPMENT The ramp access system lends itself well to conduct¬ing the several phases of development concurrently. Haulage Drifts and Raises The first development step will be to drive the 3 x 3-m (10 x 10-ft) haulage drifts above and below

Jan 1, 1982

By Pengfei Wang, Guorui Feng

"With the depletion of easily minable coal seams, less favorable reserves under adverse conditions have to be mined out to meet the market demand. Due to several historical reasons, a large amount of remnant coal was left unrecovered. One such case history occurs with the remnant rectangular stripe coal pillars using partial extraction method at Guandi Coal Mine, Shanxi Province, China. The challenge that the coal mine is facing is that there is an ultra-close coal seam right under it, with an only 0.8 to 1.5 m sandstone dirt band in between the two. The simulation study was carried out to investigate the simultaneous recovery of upper remnant coal pillars while mining the ultra-close lower panel using longwall top coal caving (LTCC). The remnant coal pillar was induced to cave in as top coal in LTCC system. Physical modelling shows that the coal pillars are the abutments of the stress arch structure formed within the overburden strata. The stability of overhanging roof strata highly depends on the stability of the remnant coal pillars. The gob development (roof strata cave-in) is intermittent with the cave-in of these coal pillars and the sandstone dirt band. FLAC3D numerical modelling shows that the multiple seam interaction has a significant influence on mining-induced stress environment for mining of lower panel. The pattern of the stress evolution on the coal pillars with the advance of the lower working face was found. This demonstrates that the stress relief of a remnant coal pillar enhances the caveability of the pillars and sandstone dirt band below them.INTRODUCTIONChina is one of the world’s coal-rich countries. Coal has played a significant role in the growth of China’s economy. Coal reserves have been mined commercially in China for over one hundred years. The demand for coal in China is still going to grow at a relatively high rate (BP Energy Outlook, 2019). Long-term large-scale and intensive coal mining activities have created tension between increased demand and limited resources (Bai et al., 2018). Easily minable coal seams such as sub-horizontal, tabular, and laterally extensive coal seams are being depleted, and therefore the remaining reserves are becoming less favorable. However, in many situations, the coal reserves under adverse conditions have to be mined out to meet the market demand for energy.Due to sedimentary environments and geological conditions, many coal seams are interspersed with sedimentary bands, or partings both within an individual seam and between seams. These bands and partings lead to appearance of ultra-close coal seams, as well as giving rise to considerable variation in their structural and mechanical properties (Ghabraie and Ren, 2016). Difficulties are encountered in the mining of such coal seams. If ultra-close coal seams are mined sequentially, that requires the design of a mining system for each individual coal seam; also, a large number of gateroads have to be excavated, which is not a high-efficiency option. If ultra-close coal seams are mined at a great mining height at once, for instance, using longwall top coal caving—a longwall mining method that is widely used in China for the mining of ultra-thick coal seams—then the location and characteristics of the band would have a significant influence on success of their extraction."

Jan 1, 2019

By W. M. Ma, Daniel W. H. Su, K. Centofanti, Yi Luo, W. L. Zhong, Syd S. Peng

1.1 INTRODUCTION When underground mining involves total extraction, it induces overburden strata movements. If not properly planned it causes surface subsidence and affects surface environmental conditions. Total extraction usually refers to longwall mining and room and pillar mining with pillar extraction. Surface subsidence has long been a subject of intensive research for scientists all over the world and considerable achievements have been obtained. However due to its difficulties and complicated nature, research into overburden movement has been thus far incomplete as compared to that into surface subsidence. Since surface subsidence is a manifestation of the results of overburden movement, the processes and mechanisms of overburden movement must be fully understood in order to establish the mathematical prediction models of surface subsidence. In this chapter the processes of overburden movement and its zones of movement in the overburden will be discussed. 1.2 PROCESSES OF OVERBURDEN MOVEMENT When total extraction is used, it produces a large void in the coal seam and disturbs the equilibrium conditions of the surrounding rock strata. The roof strata bend downward while the floor heaves. When the excavated area (or gob) expands to a sufficient size, the roof strata will cave. As a result, the overlying strata continue to bend and break until the piles of the fallen rock fragments are sufficiently high to support the overhanging strata. At this tie the overhanging strata no longer cave, but bend and rest on the underlying strata. Strata bending and subsidence develop upward until reaching the surface and forming a subsidence basin. The whole overburden strata and the surface subsidence basin will further go through a period of compaction and gradually become stabilized. Current knowledge regarding the process of overburden movement has been derived from several sources. One is the direct observation of the mined sections and their surfaces, another from field monitoring of strata movements in the overburden, and others from computer modeling and scale modeling in the laboratory. For example, Figs. 1.1 and 1.2 show the results of a borehole monitoring experiment (Borehole B-2) using the full profile borehole inclinometer and full profile borehole extensometer for lateral displacements and vertical subsidence, respectively (Conroy and Gyarmaty, 1983). The panel, 400 ft wide by 5000 ft long and 630 ft deep, was extracted from the Pittsburgh No. 8 seam which had an average thickness of 54 in. The face advanced from east to west. Curve A in Fig. 1.1 shows the inclination of the borehole when it was 45 ft ahead of the face at Position A (i.e., on the solid coal side). The total deviation of the borehole was 1/7000. The first shear movement occurred at 130 ft above the coal seam and appeared to have occurred along the bedding plane with large contrast of rock strata on both sides, i.e., sandstone vs. shale. When the face had passed the borehole 79 ft at Position B, the borehole deformed to assume a bow shape with a maximum deviation of 2.8 in. from the center line. Numerous shearing planes occurred and extended further upward. When the face had passed the borehole by 75 ft at position C, strata subsidence was measured. Fig. 1.2 shows that a strata separation as large as 4 ft occurred at 40 to 100 ft above the coal seam. Above this level, the strata subsided more uniformly with bedding separations within 1 to 2 in. Another example is the longwall mining with complete caving in the Soviet Union's Karaganda Coalfield (Kolebaeva, 1968). The coal seam was 6 ft thick and 154 ft deep. The immediate roof was the fine-grained sandstone interbedded with coarse-grained sandstone. Above this, there was sandy shale and shale. In order to monitor the process of overburden movement, 15 stations were established in a borehole (Fig. 1.3). The elevations of those stations were measured when the face was at various distances from the borehole. Fig. 1.3 shows the movement history of each station. For convenience of comparison, the movement history of each station was plotted on a common reference point as shown in Fig. 1.4. Clearly, the movements of strata from Station # 1 to #6 were nearly simultaneous, i.e., they behaved as a single unit. The strata from Station #7 to # 15 had differential separations, the maximum of which was 40 in. As the face passed by and moved away, strata separations in general reduced gradually. But strata separation between Stations #12 and #13 remained the same until the end of the monitoring period. The total bed separation between Station #7 and # 15 reached 11 in., approximately 18% of seam thickness, when the face was 132 ft beyond. Based on the above observations, the author developed a conceptual model of overburden movement due to underground longwall mining (Fig. 1.5). The above-mentioned two case studies illustrated that subsidence in the overburden strata propagates upward and subsidence velocity decreases from the bottom to the top. When the subsidence velocity at the surface reaches the maximum value, the subsidence velocity at the bottom portion of the overburden strata has decreased considerably and the strata have begun to compact. Bed separation occurs within a certain distance above the coal seam and reduces from the bottom to the top. When the face has moved away, bed separation reduces gradually, so that eventually some beds completely close and others partially close. Bed separation reaches the highest level when the subsidence velocity at the surface is at its maximum value. For instance in Fig. 1.4 surface subsidence velocity was maximum when the face was 38 to 44 ft past the point. At this time the total bed separation was maximum. Overall the strata directly above the opening are subjected to tension in the vertical direction. But above a certain level, all strata move nearly simultaneously. There is also significant shear movement along some bedding planes. 1.3 ZONES OF MOVEMENT INTHEOVERBURDEN After the extraction of a longwall panel or room and pillar section of sufficient width the strata in the overburden are subjected to various degrees of movement from the bottom to the top. According to the movement characteristics, the damaged overburden can be divided into four zones (Fig. 1.6). Caved Zone. After the extraction of coal, the immediate roof caves irregularly and fills up the void. The strata in this zone not only lose their continuity completely, they also lose their stratified

Jan 1, 1992

By M. A. Trevits, A. Ozment

"According to the Office of Surface Mining Reclamation and Enforcement Abandoned Mine Land Inventory System, in 2013 there were 98 underground mine fires burning in 9 states. This is considered by experts to be an underestimate for the actual number of fires nationwide. Abandoned mine fires, if left uncontrolled can burn for years and in fact, one of the most well-known mine fires in the US, the mine fire in Centralia, PA, which began in 1962 has been burning for over 55 years. In Centralia, the mine fire won the battle, despite suppression and control efforts, as most of the residents were bought out by the State of Pennsylvania and moved away. However, the world record for the longest-burning coal fire, which may have been started around 6,000 years ago in New South Wales, Australia is still smoldering. Suppressing a coal mine fire requires cooling the hot zones and removing any source of oxygen. If the workings are shallow, the fire zones can be unearthed and the burning rock mass can be quenched on the surface. If the workings are too deep to excavate, then the fire must be fought remotely through boreholes using a variety of agents including water, gas-enhanced foam and grout. Access to surface areas for drilling can be problematic due to topographic and property constraints. When this occurs, large areas of burning may go unaddressed or simply left to burn. Fire Solutions, Inc. has been investigating the use of gas-enhanced foam in concert with directional drilling technology. Gas-enhanced foam has the advantage of using less water and adds inert nitrogen gas to displace oxygen to infiltrate and suppress a fire. Directional drilling has the capability to steer a borehole to a specific place underground. Directional drilling has many advantages over conventional drilling technology as it can provide the least disruption to the ground surface, minimize surface preparation and reclamation costs, multiple sites can be accessed underground from a single site and thus treated simultaneously and it offers increased efficiency because it is not constrained by difficult terrain. This paper discusses the combined use of these two technologies and it is hoped that with the use of these two technologies the efficiency and capability to address mine fires will be significantly improved. INTRODUCTION According to the Office of Surface Mining Reclamation and Enforcement, Abandoned Mine Land Inventory System, in 2013 there were 98 underground mine fires in 9 states. This is considered to be an underestimate for the actual number of fires nationwide. Many mine fires are started by people burning trash where the coal seam or an abandoned coal mine is close to the surface. Other fire ignition sources include lightning and forest fires. Once ignited, a coal mine fire can easily spread into the remaining coal pillars and mine entries. Once established, the fire creates its own ventilation system supporting further combustion by drawing air down into the workings through unsealed mine shafts, fractures and surface subsidence depressions. As the coal left in the workings from the past mining operations burns, the mine void can collapse, causing subsidence and creating dangerous voids, damaging damage overlying surface structures and roadways. The products of combustion include smoke and noxious fumes such as carbon monoxide gas. These products are released to the atmosphere through fractures that develop within the ground surface, killing vegetation and creating serious health hazards [1]."

Jan 1, 2018

By S. H. Amini, A. Noble

"While many classic mineral processing circuit design approaches rely on deterministic methods, numerous studies have shown that the uncertainty inherent to the various input parameters can significantly influence design decisions. In a separation circuit, many of the technical uncertainties can be mathematically represented as statistical distributions for the unit recovery and feed grade. These uncertainties are then propagated by the circuit design and reflected in the circuit performance indicators, such as product grade and global recovery. An effective design strategy entails identifying which units are most influential in this uncertainty propagation and thus merit further consideration. Unfortunately, few analytical tools provide this information, particularly those that rely solely on deterministic methods. However, one promising approach is through the application of Taguchi’s method, in a fashion similar to statistical tolerancing of manufactured products. Since the technical challenges are fundamentally similar, Taguchi’s method shows promise in evaluating the level of compounded uncertainty by various circuit designs while determining the role of each separation stage on the overall selectivity and variability of separation circuits. This paper demonstrates Taguchi’s method in several two and three unit circuit designs. INTRODUCTION AND METHOD DESCRIPTION Uncertainty in Mineral Processing Circuit Design Analysis and estimation of uncertainty propagation in mineral processing separation circuits is an essential and significant aspect of an optimal circuit design procedure. Most of the current circuit design methodologies have been constructed, implemented, and validated using a deterministic modeling approach. However, numerous studies have shown that the inherent uncertainty induced by uncertain design parameters can be a significant factor in many mineral processing applications (e.g. Kraslawski 1989; Xiao and Vien 2003; Ghaffari et al. 2012). Thus, an effective circuit design methodology must be able to quantify uncertainty in separation circuits with limited input data. This paper describes Taguchi’s method, as an experimental design technique, to approximate uncertainty in mineral processing applications, and then the similarity of results between Taguchi’s method and Monte Carlo Simulation is investigated. Monte Carlo Simulation The term “Monte Carlo Simulation” has been used to refer to a variety of stochastic simulation methods spanning a broad array of business, technical, and scientific domains. Most generally, Monte Carlo Simulation refers to numerical analysis which is based on (pseudo) random sampling from a domain of input parameters with a known statistical distribution. This methodology allows one to determine the distribution of expected results from a standard deterministic process model. Execution of a Monte Carlo simulation begins by identifying the domain of input parameters, statistical distributions for each parameter, and a deterministic process model which describes the anticipated result as a function of the inputs. Next, one discrete random sample is taken for each input parameter, considering the input distributions. The process model is then solved for this mix of input parameters, producing a single simulation result. This procedure is then repeated a “large” number of times, utilizing new randomly sampled input parameters for each simulation iteration. Once the desired number of iterations is reached, the results of each independent simulation are aggregated to produce descriptive statistics such as minimum, maximum, mean, standard deviation, kurtosis, and other distribution parameters. Given enough iterations, the Monte Carlo method fully explores potential synergies between the input parameters. The methodology does not require the process model to be differentiable, continuous, or even analytically defined. While the number of required iterations varies in relation to th"

Jan 1, 2017

By J. Duncan Wilkins

Introduction In difficult times such as these, there is a strong reaction to the current way of doing things. Typical reactions that we have all heard are "There has to be a better way," "We're pricing ourselves out of business," "We have to improve our productivity," and "We have to have more cooperation," between union and management and employees and management. All of these comments have at their roots one common factor - getting the maximum amount for your dollar. The other factor inherent in these statements is productivity. I have not yet met a person in our industry who has not expressed the opinion that we should, and that we can, improve our productivity. Reducing Labor Costs In light of these factors I suspect we have all spent a fair amount of time examining our labor costs. For some of us, labor costs are a high proportion of our total cost of doing business. To alleviate the impact of these costs, we have generally done four things. We have reduced our number of employees, shut down operations for appropriate periods, sought concessions from union employees, and placed freezes on wages, salaries, and other benefits on nonunion employees. These approaches have been made to improve current shortfalls in our cash positions - to tide us over, as it were - or to provide us with a chance for survival. These are short-term measures that help to bring immediate relief, but can pose significant problems (or challenges), for the longer term. For instance, what do we do when times improve? How much do things have to improve before we do anything? By seeking concessions from unions in bad times, what do we do when unions come to us in good times? It is a rather sad and critical fact that we have grown too fat during the good times and too thin during the bad. In the first case, we have failed to optimize our earnings. In the second, we have cut ourselves too far. Consequently, when good or better times have arrived, we have had to bulk up our requirements to meet production commitments. In mining, for example, when times become tough, we tend to reduce our development plans so that when times improve we have to really "sock it to 'em," so that we can maintain productive capacities. We should plan ahead a little more to reduce the amplitude of our cyclical wavelength, so that in good times profits are optimized and in bad time we are better able to take the strain. Of course, forecasting cycles is not a refined art, but if we properly control our work force levels and costs at all times, and therefore optimize our productivity, we would be more able to withstand the problems we face today. Unfortunately, it appears that it takes bad times to bring us to a realistic appraisal of our way of managing our businesses. Labor Practices Not all of the problems now faced are due to low metal prices. Inflation has played a major role in bringing costs to a frightening level. Trade unions alone can not be blamed for high inflation levels over the past several years, popular though that notion is. Contract negotiations, after all, require two parties. Indeed, our current levels of labor cost are due to two factors: • We have felt obliged to keep our employees whole, relative to the cost of living. • We have felt obliged to maintain our competitive position relative to our peers, in order to maintain our ability to attract and retain a skilled, efficient work force. In the first case, our felt obligation has been applied without due recognition of the factor of performance, either in individuals or groups. Average, even mediocre performers, have been amply rewarded for their average work and mediocrity, while the good performers, no doubt receiving more for their good work and effort, have not perhaps appreciated the slight premium for their effectiveness. The rather predictable result of this practice has been that the average and mediocre stay that way (why change?), while some (certainly not all) of the good performers have said "It's not worth it," and slipped into the warm, cosy pool of the average and mediocre. In the second case, that of maintaining competitiveness with other companies' employees, we have compounded an already serious problem of lack of skilled tradesmen by paying higher and higher prices for the existing pool to maintain comparability without improving the flow of more skilled people through sound and sufficient training programs. We have the rather dubious pleasure of paying more and more for the same problems. Compensation is either mutually agreed, as with unions, or it is unilaterally applied, as with nonunion employees. Generally, there is more flexibility available to the employer when dealing with nonunionized groups than there is with unionized groups. Sometimes that flexibility has not been used well, most often be-

Jan 11, 1983

By C. W. Fitzgerald

INTRODUCTION Continuous mining machines are used in regular flat ore beds having uniform thickness. Since coal is the most common type of material occurring in such beds, most continuous miners are used for the extraction of coal from underground mines. The first continuous mining machine for under¬ground coal mining was introduced in 1948. In one machine, the continuous miner combines the functions performed by a conventional face drill, a cutting ma¬chine, and a loading machine. It breaks or digs the material out of the surrounding solid and transfers the material back through the machine chassis by means of a chain conveyor. The conveyor carries the material to a haulage system, usually a rubber-tired vehicle known as a shuttle car. Generally, continuous miners are used in seam heights from 890 to 3660 mm (35 to 144 in.). For seam heights of 1070 mm (42 in.) or less, the haulage system may be a multiple-bridge conveyor or battery-powered tractors that haul low-bed trailers. Proj¬ects are being undertaken by private industry and through government contracts to develop a continuous haulage system for all seam heights. As of 1976, the majority of continuous miners were of two types. The first type is the drum miner, which mounts the cutting elements on a series of drums across the cutting head of the machine, sometimes using a cutting chain to cut clearance for the gear drive. In the drum-type machines, the cutting head rotates parallel to the material face. The second type is the auger miner, which utilizes cutting elements mounted on two auger-type cutting heads. One head is located on each side of the machine, with the heads mounted per¬pendicular to the material face. Chain-head ripper-type miners/miners having rotary cutting elements with oscillating motion and twin-head boring-type miners have all been popular within the last 15 years. Although these are considered obsolete, some machines are still being supplied for specialized applications. GENERAL DESCRIPTION This chapter describes only continuous mining ma¬chines available as manufacturers' production models as of September 1976. Figs. 1 through 11 illustrate some typical machines currently in use. Cutting Elements The cutting elements are the interface between the machine and the material being mined. In describing the cutting elements, consideration must be given to the differences between the drum- and auger-type machines, the bit-tip speeds, and the bit types. Drum-Type Continuous Mining Machines: Three of the four principal manufacturers of drum-type miners use multi-motor designs that provide separate a-c motors for driving the cutting elements. Gear trains are used to reduce the speeds of the motors to the rotational speeds of the cutting elements. One of the three suppliers drives the cutting elements of its low-seam models with a heavy-duty drive chain that is equipped with cutting bits; all other functions of this machine have separate motor drives that are either electrical or hydraulic. The fourth principal manufacturer generally provides two a-c electric motors, utilizing gear reducers, drive shafts, and hydraulically operated engaging clutches to power all of the operating functions, including the cut¬ting elements. Some drum-type miners use trim chains, equipped with cutting bits, to cut clearance for the gear drives. These trim chains vary from two 305-mm (12-in.) chains to one 760-mm (30-in.) chain per machine, de¬pending upon the manufacturer's design concept. In all machines other than the low-seam models that use chains to drive the heads, the trim chains are driven by sprockets on the drive shaft of the final cutterhead drum; therefore, the trim chains are not used for power trans¬mission to the cutting elements. In addition to cutting clearance for the gear drives, the trim chains provide uniform bit spacing all the way across the head and assist in moving the cut material to an area where it is either loaded directly into the miner's conveyor or is easier to load into the conveyor. On a drum-type mining machine, the cutting ele¬ments all rotate from the top to the bottom. Therefore, the recommended procedure is to raise the cutterhead to the maximum seam height, sump the entire machine forward to maximum depth by means of a crawler drive, and then shear downward until the cutter elements reach the floor level. The maximum sump depth varies ac¬cording to the drum diameter, the type of material, the floor conditions, and the operator's abilities. It normally varies an average of 460 to 760 mm (18 to 30 in.). Cutter drums are supplied with two types of cutting bit holders. Weld-on holders are attached permanently. Pin-on holders are replaceable at the face. Cutter-drum diameters over the bit tips vary from 660 to 965 mm (26 to 38 in.), depending upon the chassis height of the machine and the range of cutting heights. Auger-Type Continuous Mining Machines: As cut-

Jan 1, 1982

By M. J. DeJong, R. J. Mair, S. Ritter, G. Giardina

"For urban tunneling projects it is essential to predict and prevent building damage. Although various case studies and experiments have shown that buildings considerably modify greenfield soil movements, widely accepted damage assessment methods neglect this soil-structure interaction and simplify structures as linear elastic beams. This paper summarizes an experimental investigation of the response of more realistic structures to tunneling-induced deformations. Small scale structural models with façade openings and brittle material properties were 3D printed and tested in a geotechnical centrifuge. Soil and structure displacement data were obtained by image-based measurement. Results demonstrate that structures notably mitigate differential greenfield ground displacements. It is also shown that maximum soil settlements, horizontal soil displacements beneath the structure and structural damage in the form of cracking significantly depend on the position of the structure in the settlement profile. The results provide a basis from which to predict building settlement response with greater certainty. INTRODUCTION With continuing population growth and increasing urbanization expected in the next decades, there is an urgent need for a next generation of infrastructure. It is likely that a large part of this infrastructure will be underground. Such creation of urban underground space frequently involves the excavation of shallow tunnels in close proximity to existing structures and utilities. For example, along the Crossrail project in London more than 1,200 structures are located within a possible influence area of the tunnel construction (Torp-Petersen and Black 2001). The assessment of building damage plays a vital role throughout the planning, design and construction stage of urban tunneling projects. However, while predictions of tunneling-induced deformations in the greenfield can generally be made with some confidence, much uncertainty exists about how structures interact with ground movements arising from tunnel construction."

Jan 1, 2016