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By R. Bishop, J. Baggett, J. Monsalve, N. Ripepi, A. Soni

Technology plays an ever-increasing role in improving the safety and efficiency of mining operations. Laser scanning and photogrammetry are two useful methods for capturing 3D digital representations of real world objects. While both technologies have been applied to the mining industry in numerous ways, the practical applications in an operating underground limestone mine has been tested for this paper, including for visualization and site characterization. Each technology is capable of creating highly detailed geospatial point clouds, but are all point clouds created equal? This paper presents a comparison of a laser scan and photogrammetry of a limestone pillar and addresses the strengths and limitations of each method for creating digital models of operating underground limestone mines. INTRODUCTION The United States crushed stone business is a ~$14 billion per year industry representing 1,430 companies which operate 3,782 mining operations across the 50 states. In 2017 alone, 1.33 billion tons of crushed stone was produced with 76% used predominately for road construction and repairs, but also for the production of cement, lime, chemical and agricultural uses. Over 70% of domestic crushed stone producers source their material from limestone and dolomite deposits, with currently only 82 mines (2%) operating underground (USGS, 2018). Underground mining requires a different technical skillset, and poses great challenges for production that must be both safe and economical. Therefore, there are many risks that an operation must mitigate in order to produce underground limestone. Some of the challenges in underground limestone mining involve geological hazards that are unique to the rest of the industry. The geologic features such as faults and weaknesses can be difficult to assess in a production environment (NIOSH, 1998). The consequences of failure due to improperly designed underground mining ground support systems, such as from a roof failure, are not only life threatening for the miners themselves, but can easily shut down an entire mining operation.

Jan 1, 2019

By E Mutaz, M Bahaaddini, D J. Williams, M Serati

The contribution of the mining industry to the global economy has reached (and gone beyond) $1,800 billion in recent years that necessitates (more than before) the need to ensure workability, viability, health, and safety of mine sites and workers to meet the ever-increasing global demand for mineral resources. Amongst the many challenges facing the excavation industries, rockburst and spalling failures in deep mines in hard and brittle rock formations are yet happening that are still deemed as unsolved problems. During unconfined compressive testing of brittle rocks, microcracks initiate at approximately 0.3–0.5 of the peak strength (which is commonly known as the Crack Initiation stress level) mainly in the direction of the major principal stress, then grow unstably and coalesce to form macro fractures at approximately 0.7–0.8 of the peak load at a so-called Crack Damage stress threshold. From a stress-strain perspective, the crack damage stress threshold () is associated with the reversal point of the volumetric strain, which is also proved to be strongly affected by the intermediate principal stress. This paper extends the previous studies on the impact of true triaxial loading conditions on the CD stress level in high porosity rocks, and introduce a new Crack Mode-Changing Stress (CMCS) defined as the minimum principal stress required to change rock fracturing from splitting to a sliding failure mode. A large number of experiments on sandstone specimens were carried out at different true triaxial loading conditions and the effect of principal stresses on the CI and CD as well as CMCS were studied. The results show that not only the intermediate principal stress, but also the minor principal stress can significantly affect the CD and CMCS thresholds.

Apr 26, 2022

By C. L. Lin

A PC image-based on-line coarse particle size analyzer has been developed to measure particle size distributions under plant situations where overlapped particles are being transported on a moving belt conveyor. Although previous testing had shown satisfactory performance of this image-based system for separated particle size measurements (the non-overlapped case), it was necessary to modify original algorithms in order to solve the overlap problem and make the system suitable for plant operations. On-site testing has been conducted at a large copper concentrator in the southwestern U.S. for further investigation of the overlap problem.

Jan 1, 1995

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 A. T. Iannacchione

Unrecognized roof beam failures resulted in 69% of the falls of ground injuries occurring in underground U.S. stone mines from 1990 to 1996. Field investigations at 45 underground stone mines suggest that excessive beam sag or deflection is the principal unstable roof behavior characteristics associated with miner injuries. Laminated roof structures, common in most underground stone mines, allow the immediate roof to separate into distinct beams. The timely determination of roof beam deflection could provide stone miners with advanced knowledge of potentially hazardous roof. Existing methods to assess roof conditions consist of observational and monitoring techniques. Observational techniques include open drillholes, sounding, wedging, and borescopes. Monitoring techniques include scratch tools, extensometers, telltales, Miners Helpers, and Guardian Angels. 1 Miners have difficulty utilizing these techniques because roof beams are located an average of 8 m (26.5 ft) above the miner's head in environments that are dark, dusty and noisy and these deflections are measured in millimeters of movement. To help mitigate the potential for falls of ground injuries to underground stone miners, the National Institute for Occupational Safety and Health (NIOSH) has developed the Roof Monitoring Safety System (RMSS) to aid in monitoring dangerous levels of roof beam deflection. If enough site-specific information is collected and placed on highly visible maps, individual mining operations could establish general guidelines for responding to specific roof beam deflection occurrences. This would help in implementing a pro-active, comprehensive roof control plan for reducing falls of ground injuries.

Jan 1, 1999

By William Gleason

"Mining roots run deep in Glenrock, WY, a town that was founded in part, because of the coal mining that began in the late 1800s. But that rich history also now threatens large parts of the town that were built over the top of the mines. Like many towns in the Western United States, mining is a proud part of the historical fabric and heritage of the town, but the legacy issues from past mining activities continue to present challenges to modern-day engineers.In the case of Glenrock, a small town about 40 km (25 miles) east of Casper, it is flooding from a century of water that has accumulated in the down geologic dip of the furthest northeastern extent of 87 ha (217 acres) of mine workings that has made the town susceptible to sinkholes and other issues (Fig. 1).Full commercial coal mining production began in Glenrock in 1887 during the development of the Fremont, Elkhorn and Missouri Valley railway. Due to poor overlying roof rock and continuous flooding within Glenrock, Mines No. 1 and 2 were fully closed in 1909. During its history, an estimated 87 ha (217 acres) were mined with 50 to 60 percent extraction ratio within the eastern town limits through room-and-pillar mining. It is estimated that the mines produced 22 kt/a (25,000 stpy) during their lifetime. A century later, the water pressure built up in the northeastern most portion of the mine has created serious concerns for residents.In 2016, the Wyoming Department of Environmental Quality’s Abandoned Mine Lands Divisions (AML) investigated a number of subsidence events in the town including sinkholes following heavy precipitation. The AML investigations confirmed that most events were related to the previous mining in the area.Mitigation efforts on the mines were done in the 1980s and 1990s, but the problems in the town and in a nearby subdivision persisted, leading the AML to search for alternative solutions."

Nov 1, 2019

The fundamental objective of mine plans are usually to maximize the mine profit by extracting ore at the lowest possible cost over the mine life while some technical constraints are satisfied. Since the costs associated with the operation of trucks and shovels as resources are significant, determining the optimum number and allocation of these resources is an essential issue. This paper presents a Mixed Integer Linear Programming (MILP) model to determine the optimum number of shovels and trucks required to meet the short-term plan?s goals. Also, the model takes into account the allocation of trucks and shovels to mining faces. This model minimizes operational costs, while attempting to meet the production demand and to satisfy technological constraints. This paper also analyzes the operations of trucks and shovels in an open pit mine using ARENA simulation software.

Jan 1, 2012

By Desiree Willis

At 6 percent downgrade to a low point beneath California?s Sacramento River, the Lower Northwest Interceptor Sewer was a difficult project from the outset. Project owner Sacramento County Regional Sanitation District (SRC-SD) planned to excavate the V-shaped tunnel in two sections using an earth pressure balance (EPB) machine in soft ground. Muck removal, due to the steep downgrade, was a tricky proposition. ?We did look at muck cars, but determined they were not the best option due to the relatively steep slope of the tunnel alignment. The safety benefits of conveyors on this job were obvious, as there were not the potential hazards associated with muck cars traveling on an incline or derailing,? said Steve Norris, senior civil engineer for SRCSD The successful project represented one of the first instances of a conveyor system being used for a soft ground tunnel boring machine (TBM) application. Several projects in California, including a recent tunnel completed in 2009, have highlighted a growing acceptance of conveyor muck removal in a variety of ground conditions. Conveyors offer higher system availability than more common methods, such as muck cars, as they allow for continual operation of the TBM. The design of conveyor systems must, however, take into account the differences in material between soft ground and hard rock, including belt cleaning features, sealed transfer points and other elements.

Jun 1, 2010

By Thomas P. Meloy

Comminution is an old art. Unfortunately the analysis of the unit operations has been handicapped by the lack of both theoretical and experimental information. Two types of infor¬mation are needed, one an understanding of what happens when a particle breaks in a brittle manner; and two, an understanding of what happens when the daughter fragments pass through the mill. This paper is concerned with the former, the catastrophic fracture of a brittle solid. Recently, successful theoretical attacks on the frac¬ture of brittle solids have yielded significant-results. New experimental techniques are yielding insight into the physical phenomenon occurring during brittle fracture. Many of the re¬doubtable shiboliths of the last century are either being laid to rest or verified.

Jan 1, 1965

By R. Rodgers

The Pinto Valley concentrator of BHP Copper Inc. restarted its sulfide mining operations in September 2007 to produce copper and molybdenum concentrates. During the plant production ramp up, the copper/molybdenum circuit quickly reached the copper grade and recovery targets, but the production and quality of the molybdenum concentrate continued to be a challenge. Even though several improvements in the molybdenum circuit were completed, the final molybdenum concentrate did not meet contractual specifications. After reviewing several options to improve grade and reduce contaminants, the installation of three flotation columns were chosen. The column design and construction were done using in-house craftsmen and engineers. Ten days after commissioning, the columns successfully reached a steady operation and began producing molybdenum concentrate to the required specifications, that is, high molybdenum tenors and low copper and iron content. The Pinto Valley operations, are comprised of the Pinto Valley concentrator, the Pinto Valley dump leaching and a solvent extraction/electrowinning (SX/EW) plant. The concentrator operated from 1972 to 1998 when mining and milling operations were curtailed for economic reasons. The SX/EW plant continued operations for copper cathode production. After a cessation of mining operations of almost 10 years, Pinto Valley operations began a restart project in January 2007, to bring the mine and the concentrator back into production. Copper concentrate production started in October 2007 after refurbishing and upgrading the facilities with a new DCS system and filter plant. The daily targets of the concentrator were to treat 56.7 kt (62,500 st) of ore, produce 762 t (840 st) of copper concentrate containing 27.5 percent copper at recoveries of 86 percent and produce 2.2 t (2.5 st) of molybdenum concentrate containing more than 80 percent MoS2.

Jan 1, 2010

By D. Mohapatra

Extraction of Al(III) from aqueous sulfuric acid solutions into organic Cyanex 272®-carbon tetrachloride systems was investigated from a multi-component mixture containing aluminum, cobalt and nickel. Al(III) extraction efficiency was quantitative with 1.0 M Cyanex 272® in two stages at an aqueous: organic (A:O) phase ratio of 1:1 and an equilibrium pH of 3.2. The stripping of Al(III) from the loaded organic (LO) with different acids indicated sulfuric acid to be the best stripping agent. More than 99% stripping of Al(III) from the LO was achieved using 2 M H2SO4 at an A:O phase ratio of 1:1 in a single stage. The stoichiometry of the extracted species of Al(III) was determined by the slope analysis method. The extraction reaction proceeded via the cation exchange mechanism and the extracted species was assumed to be AlA3?3HA, where HA refers to the dimeric and A the anionic form of the extractant, Cyanex 272®. The extraction of Al(III) was studied in the presence of various ions to ascertain the tolerance limit of individual ions.

Jan 1, 2008

By Donald L. Freed

Design, development, factory test, and in-mine evaluation plans are presented for an Automated Extraction System (AES) that has been designed to fill needs which are present in the Mining Industry. The philosophy of design and proposed operation of the machine, as well as the relationships between these factors, and final machine configuration are reviewed. The AES is designed to overcome production problems resulting from the regulation of roof control, clean-up, ventilation, dust suppression, and operator protection required by the Federal Coal Mine Health and Safety Act of 1969. The design permits more continuous operation and greatly improves operator safety. The machine is a full-face drum-type continuous miner which includes self-advancing, continuous temporary roof support. Onboard self-advancing face ventilation, dust suppression, and air scrubbing equipment are also included in the design. A further improvement over existing equipment is the incorporation of an automatic programmed cutting cycle control system. All mining functions including movements of the roof support system are handled by this automatic system.

Jan 1, 1977

By Kent von Maubeuge, Charles R. Landis

The global bentonite industry is a dynamic market balancing a range of end uses with an ever growing number of bentonite producers. In addition, markets in developed countries are elevating the level of detail and transparency for materials reporting in programs, such as the Registration, Evaluation, Authorization and restrictions of Chemicals (REACH) program in the European Union. Bentonite is impacted by programs of this nature. The majority of the bentonite products sold around the world are sodium bentonites. These sodium bentonites consist either of natural sodium bentonites, or naturally occurring bentonite dominated by divalent interlayer cations (Ca+2 andMg+2) subsequently exchanged with sodium cation. The latter variety is often called “sodium-activated“, or “soda-ash treated” bentonite. Once exchanged with sodium, these bentonites are often difficult to distinguish from natural sodium bentonites using common commercial testing protocols. Due to their similar colloidal and bulk chemical properties; the lack of test procedures designed specifically to identify this product feature; and the failure to completely or effectively explain their differences, sodium-activated bentonite is often considered to be identical to natural sodium bentonite in the industry. This article defines both general types of sodium bentonite and presents a technique that could serve the industry as a tool to confirm the variety of sodium bentonite in question. Specifically, geochemical properties such as carbon (C13/C12) and oxygen (O18/O16) isotopic ratios can distinguish activated, even slightly activated bentonites, from bentonites not introduced to additional sodium during manufacture.

Jan 1, 2004

By Patrick Sullivan

Over the past few years, successful operations for processing phosphates employing the Grate-Kiln System have been achieved. Control and instrumentation philosophy followed in these installations are described and analyzed. Concepts utilizing digital techniques and computer systems appropriate for similar installations are also discussed.

Jan 1, 1967

By Brendan Demich, Linda L. Chasko, Brendan Macdonald, Timothy J. Orr, Jessie J. Mechling, Cassandra Hoebbel, Jason Navoyski, William Helfrich, Jessica Cohen, Paul E. Schmidt, Jennica L. Bellanca

The National Institute for Occupational Safety and Health (NIOSH) in partnership with the Mine Safety and Health Administration (MSHA) developed VR Mine Rescue Training (VR-MRT). It is a training platform that allows mine rescue team members to explore virtual emergency scenarios collaboratively. Team members use a head-mounted display to visualize and interact together with equipment, hazards, and dynamic ventilation. All player actions are recorded and displayed for debriefing. Researchers at NIOSH demonstrated the training platform at various mine rescue competitions, including the International Mine Rescue Competition, and collected feedback. Participants’ responses were overwhelmingly positive. This paper describes the development of the software platform and the results of the usability surveys and interviews.

Jun 25, 2023

By George Saratsis, George Exadaktylos

In the marbel industry, ore reserves refer to the produced tonnage of orthogonal prismatic, defect-free blocks of commercial sizes and of marketable aesthetical quality. On the contrary, in the mining sector, ore reserves refer to the tonnage of ore with grade above some cut-off value. While in the mining sector ore reserves are reported according to existing standard guidelines, the marble industry has no standardmethod. The purpose of this paper is to present a method to estimate white marble reserves using drill coring or borehole wall logging and limited joint mapping on exposed marble surfaces. Data pertaining to fracture orientation, fracture frequency, size, and whiteness were collected. The primary phase of marble characterization is followed by the discretization of the orebody into a three-dimensional array of orthogonal prismatic blocks, where the marble whiteness and fracture intensity are attribute values. Both these variables are considered to obey continuous and not discrete probability distributions. The discretization is performed using computer aided design and the Block Kriging techniques. An outline to improve the description of a given marble ore deposit at the exploration stage is presented for the prediction of marble block size distribution from joint count measurements along diamond drill cores. An example case of a white dolomitic marble quarry is used to demonstrate these methods.

May 27, 2020

By Yuting Xue, Mark Van Dyke, Zoheir Khademian

Borehole scoping has been widely used in the U.S. underground coal mines as a method for geologic mapping and stability analysis because it is very useful for identifying rock types, voids, fractures, and formation boundaries through the direct observation of the borehole wall. An image stitching technique was proposed to generate panoramic borehole images from side-view borescopes in a previous study. However, the generation of panoramic borehole images can be affected by the borehole condition because the image stitching process highly depends on the key point detection and matching. The presence of dust and water could have a significant influence on the detection of key points. In this study, the influence of borehole conditions on panoramic borehole images was studied from the success of image stitching and the visibility of geologic features and lithology changes. Borehole videos recorded at different times, when the same borehole was freshly drilled, freshly washed with water and 12 days after washing, were processed with the proposed technique. The comparison of the stitched images shows that the presence of dust has minimal influence on the image stitching process but shows significant negative effects on the visualization of geologic feature and lithology change. In addition, the presence of water can significantly hinder the image stitching process, which in turn helps in identifying lithology change. Borehole scoping aids in identifying and mitigating hazardous conditions in underground mines, and this study provides preliminary inputs into identifying optimum borehole conditions for geologic mapping with panoramic borehole images.

Feb 1, 2025

By J. L. Ellenberger

The U.S. Bureau of Mines has cooperated with Cyprus Casa Grande Corporation Cyprus Casa Grande Mine, Cas a Grande, Arizona, since 1988 to assess the hydrology of an in situ leaching test facility. This work has resulted in an in situ design for leaching copper oxide ore that takes advantage of the increased mineral access associated with poor quality rock surrounding mine drifts. Blasting and bulkheads are used to produce readily-leachable rubble and contain and control solution movement, respectively. The design reduces the need for handling solution under high pressures in an injection/recovery system for in situ leaching of underground ore bodies. Early work at the site recognized a stress redistribution zone and increased transmissivity near drifts. Fracture data from a corehole oriented parallel to the dip of the ore body indicated an increase in fracture frequency around drifts a distance of 21 m (70 ft) away from the 5 m (16 ft) diameter drifts.

Jan 1, 1994

By J. Mark Richardson

The use of an integrated software package to simulate and optimize the performance of mineral processing circuits is illustrated by three examples. Each represents application of the software and associated techniques to optimization of an existing plant. In the first example, a South-western U. S. copper company employs the method to simulate and improve performance in crushing, screening and SAG milling. A second copper producer evaluates equipment changes and flowsheet alternatives in a closed-circuit ball mill-cyclone operation. A Southeastern U. S. aluminum refinery optimizes performance of a dry milling and classification circuit for production of carbon particles used in anode casting. Graphical and tabular data describe the circuits simulated and the results achieved in each example.

Jan 1, 1990

By Ross Glanville

The Optimum Production Rate (OPR) is one of the most important parameters in the evaluation of a mineral deposit. The OPR cam also be expressed as the Optimum Mime Life (OML) in years since the expected mime life is determined by dividing the OPR per year into the estimated ore reserves. Unfortunately, very little time and effort has been directed towards the determination of the OPR. Instead, "rules of thumb" are often applied to select a mime life without due consideration of the economic implications of such a selection. A "justification" for a particular production rate is often based on a pre-conceived arbitrary requirement for a mime life of 5, 10, or 15 years, for example. In this paper it is demonstrated that such arbitrary selections for high-grade/low¬-tonnage mines (such as many underground gold/silver deposits) often lead to sub¬optimal results. In fact, am apparently uneconomic deposit at am arbitrary production rate may be economic at the OPR. Consequently, investment opportunities may be overlooked if one analyzes mime' properties based on arbitrary production rates.

Jan 1, 1984