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By D. H. Yardley

A major purpose of this contract was to evaluate a variety of dust suppressants to determine their effectiveness in mitigating dust from taconite tailings. During the course of the investigation tailings from pilot-plant tests of Minnesota copper-nickel ores became available and chemical binders were also tested for dust mitigation effectiveness on this mineralogically different material. Field-test plots were established on taconite tailings and wind tunnel laboratory tests conducted. The laboratory tests consisted of comparing treated and untreated samples at wind velocities of 40 mph using a weight-loss method and direct dust emission measurements. It was shown that the emission of fine dust from taconite tailings, in the range of 4.5 microns to 0.85 microns, could be reduced by 99 percent using certain chemical suppressants. It was also demonstrated that a straight water spray on coarse tailings (3/4" to + 10 mesh) reduced dust emission by 95 percent. Although eight different chemicals were tested, most of the tests were made using Coherex, ammonium lignin sulfonate and sodium silicate. These same dust suppressants were used, singly and in combination on almost 75 field-test plots. The Coherex appears to provide the best results in terms of economics and durability. Wind tunnel tests on treated and untreated Cu-Ni tailings demonstrated that applications of dust suppressants reduced fugitive fine dust by better than 99 percent.

Jan 1, 1980

By C. Yi, M. Onifade, B. Genc, S. Dineva, T. Zvarivadza, M. Khandelwal

Deep level hardrock mining faces increasing challenges from elevated in situ stresses and associated rockbursts. This study aims to develop a systematic framework for evaluating destress blasting effectiveness, with a focus on deep Swedish hardrock mines. The methodology integrates a critical review and adaptation of global best practices: advanced numerical modelling (static and dynamic simulations), field-based rock fracture monitoring using ground penetrating radar and borehole imaging, and application of rockburst prediction criteria including the strain energy storage coefficient (F), brittle shear ratio, and burst potential index. Geostatistical simulations, machine learning models, and industrial internet of things-based, real-time monitoring are proposed to enhance predictive accuracy, model calibration, and operational adaptability. Key findings show that effective destress blasting evaluation requires multi-modal integration of numerical outputs, field observations, microseismic trends, and uncertainty quantification, accounting for site-specific geological variability and dynamic stress redistribution. The study advances the field by proposing a predictive, feedback-driven evaluation framework tailored for deep Swedish mining conditions, capable of improving proactive rockburst risk management. It offers both practical tools for mining practitioners to optimise stress management, reduce seismic hazards, and enhance excavation safety, and academic foundations for future refinement of destress blasting models, geotechnical monitoring strategies, and adaptive design protocols. This research contributes to safer, more efficient deep mining operations by bridging gaps between theoretical modelling, empirical monitoring, and real-time, data-driven blast design optimisation strategies.

Jun 27, 2025

By Chris Howell

PotashCorp?s Aurora phosphate mine, in Beaufort County, NC, operates four low pressure boilers and four high pressure boilers to meet its process steam requirements. Maintaining these boilers is critical to meeting production demands. Processing phosphate ore into phosphoric acid requires large amounts of sulfuric acid and steam. Ore is mined from the phosphate deposit with large bucket excavators after 9 to 30 m (30 to 100 ft) of overburden is removed. The phosphate ore is mixed with water to make a slurry that is pumped to the mill. Elemental sulfur is burned in the presence of air to make sulfuric acid. This process is highly exothermic and boilers produce steam while cooling the process. The phosphate ore is reacted with sulfuric acid and the resulting products are refined into different grades of phosphoric acid for fertilizers, feed stock and food additives. Plant operators at the mine were challenged with boiler feed water that was high in silica and other contaminants. As a result, boiler operations were experiencing difficult, labor-intensive operations and equipment deterioration. This resulted in higher operating costs and diminished processing performance. The facility was operating an aged boiler feedwater pretreatment system that included warm lime softening followed by pressure filtration and a cation/anion/mixed bed demineralizer system. The demineralizer required frequent regeneration and chemical usage was high and costs were difficult to predict. Depressurized well water (DPW water), used for plant utility water and boiler makeup, is very high in silica, in the range of 60-70 ppm. Silica can form scale at pressures below 600 psig. Above 600 psig, silica starts to volatize, carrying over with steam to potentially form deposits on downstream equipment and processes. Mine officials decided to upgrade the boiler feedwater pretreatment system to a state-of-the-art reverse osmosis (RO) membrane system. This new system, which started up in March 2009, has improved operating performance and reduced water consumption and overall operating expenditures.

Jan 1, 2011

By C. D. Ashley, X. P. Garzon-Villalba, R. M. Lopez, T. E. Bernard

Federal standards for refuge alternatives (RAs) mandate that they not exceed a Steadman apparent temperature (AT) of 95 °F (35 °C) at an assumedmetabolic rate of 325W, a limit that appears to be arbitrary. Occupants in an RA spend most of their time at rest (a metabolic rate less than 325 W), and thermal equilibrium can likely be maintained at an AT > 35 °C. The purpose of this study was to examine the upper limit of sustainable heat stress during 4- and 8-h exposures at rest. Five men underwent five 4-h trials (phase 1: AT range 39 to 49 °C, 90% rh) and five 8-h trials (phase 2: AT range 46 to 56 °C, 90% rh) in a semi-recumbent position. Descriptive statistics for gastrointestinal temperature (Tgi), heart rate (HR), systolic and diastolic blood pressure (SBP and DBP), sweat rate, fluid intake, urine color, urine specific gravity (USG), and changes in body mass are reported. Mean Tgi for all trials ranged from 36.9 to 37.3 °C, and mean HR for all trials ranged from 61 to 75 bpm. To determine the effects of increased heat exposure, differences in heat strain between 4- and 8-h trials were examined via repeated measures analysis of variance (ANOVA). As AT increased, there were no significant changes in ?Tgi and ?HR from 1 to 4 h in the 4-h and 8-h trials. During the 8-h trials, as AT increased, there was a difference in ?Tgi between 4 and 8 h. In conclusion, there was support for sustainable exposures resulting in no significant increase in physiological strain at an AT greater than 35 °C AT, with a sustainable limit (no increases in Tgi due to increases in AT) below 46 °C AT. Although the participants were not representative of coal miners, the results provide support for sustainable exposures greater than 35 °C AT at an average metabolic rate of 155 W.

Nov 8, 2019

By B. Genc, C. Gous

Spontaneous combustion during coal mining operations is a major problem that affects the health and safety of workers and causes environmental problems. The phenomenon is associated with the presence of coal, coal shale, and pyrite. In 2020, a premature detonation incident occurred at an iron ore mine where the waste material contains black carbonaceous shale units known to be associated with pyrite. The spontaneous combustion propensity and properties of samples of the black carbonaceous shales from the mine were examined and compared with samples from the Witbank Coalfield. The spontaneous combustion liability indexes of these samples were correlated with X-ray fluorescence (XRF) and proximate and ultimate analyses using linear regression. The Wits-Ehac Index classification results show that the samples were between medium and high risk. The linear regression analysis showed very poor correlations between the Wits-Ehac Index results and the XRF and proximate and ultimate results. The most valuable relationship found is between the presence of relatively high sulphur (greater than 3%) and ground reactivity with nitrate-bearing explosive emulsion.

Jun 5, 2023

By S. Viswanathan, A. S. Sabau

"Dimensional changes between a pattern die and its corresponding investment casting part occur as a result of complex phenomena such as thermal expansion/contraction and hot deformation (elastic, plastic, and creep) during the processing of the pattern material (wax), mold material (shell) and solidifying alloy. Determining the pattern tooling dimensions is crucial to the dimensional control of the investment casting process. To date, computational methodologies have not been applied to the prediction of dimensional changes in investment castings.This paper deals with the evaluation of wax pattern dimensions, which is one important factor in determining pattern tooling dimensions in investment casting. Cerita 29-51, an industrial wax, is considered in this study. The wax pattern dimensions are determined by considering its thermo physical and thermomechanical properties, restraint of geometrical features by the metal die, and process parameters such as dwell time, platen temperature, injection pressure, and injection temperature.IntroductionThe investment casting process involves making a disposable wax pattern by injecting wax into a metal mold, building a ceramic shell mold around the wax pattern by the application of a series of ceramic coatings to the wax pattern, dewaxing the ceramic shell mold, and casting the alloy into the shell mold. Dimensional changes between the pattern tooling and its corresponding cast part occur as a result of thermal expansion, shrinkage, hot deformation, and creep of the pattern .material (wax), mold material (shell), and solidifying alloy during the processing [1]."

Jan 1, 2001

By Steve Dillingham

Whenever misfired holes or portions of misfired holes remain after a blast is fired, a hazardous situation exists until the unfired explosive materials have been disposed of properly. Yet, there are so many factors that can contribute to misfires that it is impossible to give blanket instructions on how to handle them. To complicate matters, misfires today are rare due to reliable blasting products. As such, it isn’t very surprising that many blasters have never had to deal with a misf i re condition. While the likelihood of a misfire occurrence is minimal, you should always be pre p a red to handle a misfire should you be involved in such an event. To begin, every blaster and blasting org a n i z a t i o n should develop a misfire policy which contains the company’s safe operating pro c e d u res, emergency contact list, as well as any regulatory compliance re q u i re m e n t s . Tailgate and/or company-wide safety meetings should communicate the company’s contingent misfire pro c ed u res and any special site-specific precautions BEFORE the start of any project.

Jan 1, 2004

By H. A. Kramer

The Flin Flon mine is at present hoisting 5,000 tons of ore daily. In 1943, a peak production of 6,188 tons daily was maintained. Before discussing secondary blasting, it is felt that a short review of stoping procedure will be of value. A list of papers dealing with various phases of the mining operation is appended. Stoping Practice Stopes vary in width from 30 to 150 feet, in length from 200 to 450 feet, and in height from 200 to 500 feet. A typical stope would be 300 feet long, 300 feet high, and 90 feet wide. The dip of the orebodies varies from 60° to vertical, with an average of 82°. The ore is heavy, having a specific gravity of 4. Both sub-level benching and radial drilling are practised. Blast-hole diamond drills are responsible for 85 per cent of the total production. Two sizes of hole are drilled at present, 11 7 6 -inch and 1%-inch. For bench drilling, the burden on the 1 7/16-inch hole is six feet, with holes on 8-foot staggered centres; in radial drilling, the burden is six feet, with nine feet between the toes. The burden on the l 5/8-inch hole is eight feet with holes on 11-foot centres, or 13 feet on the toe of fanned holes. Holes are blasted with two No. 8 instantaneous electric caps per hole, using 403 Polar Forcite. A powder charge of 0.15 pounds per ton is used, with a tonnage factor of 5 and 6.4 per foot drilled for the two sizes. Two rows of holes are blasted together. Fragmentation is satisfactory with the above drilling pattern and powder charge. However, some overbreak and fall-off locally results in oversize material. It is estimated that, after primary breaking, 75 per cent of the ore will pass through the grizzly openings and 15 per cent will pass the draw points, leaving 10 per cent that must be broken further to pass the draw points.

Jan 1, 1946

By S. Chatterjee, F. Suparno, A. Paithankar

Estimating reserves and quantifying resources stand as pivotal and intricate endeavours within the realm of coal mining operations. Intricate geological formations compound the challenges in resource estimation, thereby complicating reserve calculations. The uncertainties tied to geological attributes of coal, encompassing parameters like tonnage and coal quality, wield significant sway over resource and reserve computations within coal mines. This research delves into the domain of geological uncertainties, with a specific focus on calorific value, aiming to numerically characterise resources and reserves within an open-pit coal mine situated in Indonesia. To quantify resources, the coal seam geometry underwent simulation via a multipoint geostatistical technique known as single normal equation simulation. A geologically established coal seam served as the training image for generating 20 equiprobable coal models. To simulate CV, 50 realisations were generated for each simulated coal seam, utilising sequential Gaussian simulation. Deviations of the simulated coal seams ranged from -0.07% to 5.48% in comparison to the training image. The CV simulation yielded an average value of 5,920.29 kcal/kg, accompanied by a standard deviation of 586.54 kcal/kg. However, the average CV spanned from 5,305.26 kcal/kg to 6,526.55 kcal/kg across diverse simulations. For reserve calculation within the context of geological uncertainties, an algorithm rooted in maximum flow graph theory was employed to construct the ultimate pit for the coal mine. Within this final pit, the average stripping ratio was 1.62, coupled with a CV value of 6,019.66 kcal/kg. When juxtaposed with the deterministic model, the findings underscore that the stochastic ultimate pit delineates a more expansive excavation, accompanied by a heightened undiscounted cash flow.

Oct 2, 2025

The Australian Gas Light. Company was established in 1837, the first gas being sent out from the Kent Street Works, then the headquarters, in 1841. Mortlake Works were opened in 1886, and has since grown to the vast undertaking inspected by members of the Institute.The Kent Street Works were closed in 1921. The area of the, company's property at Mortlake is 117 acres.Some interesting figures from the statistics of the company's activities for 1937 are:-

Jan 1, 1938

By Donald Bruce, Alex Reeb, George Filz, Wesley Schmutzler, Mitsuo Nozu

"One million seven-hundred thousand cubic yards of deep mixing were installed to support an elevation raise of an earthen levee known as LPV 111, which is located in east New Orleans. The deep mixing was done by two contractors using somewhat different technologies, but producing elements of the same size. As part of the quality control and quality assurance program, 507 deep-mixed elements were cored and 5,082 specimens were strength tested in unconfined compression. This paper presents data from the coring and testing program. Very high percentages of the core run lengths were recovered, except during one period for one mixing technology. The coring methods are described, and reasons for the isolated episode of low recovery are discussed. The strength data show that the specification acceptance criteria were satisfied. Statistical parameters of the strength data are provided, and trends between strength and the following factors are discussed: depth, soil type, binder factor, and project-specific experience. Photographs of longitudinal cuts of core specimens are presented. Important lessons learned from this unusually large project are described.INTRODUCTIONThe Lake Pontchartrain and Vicinity Hurricane Protection system included a project known as LPV 111, which is an approximately 5-mile long segment of levee and floodwall located in east New Orleans along the Gulf Intracoastal Waterway and adjacent to Bayou Sauvage National Wildlife Refuge (Cali et al. 2012). The majority of the project consisted of a raise of an existing earthen levee from an existing crest elevation of about 18 ft to a new crest elevation of about 28 ft. The underlying soils were weak and compressible, providing inadequate support for the raised levee. Ground improvement in the form of deep mixing was selected to provide the necessary stability and settlement control for the raised levee.The US Army Corps of Engineers (USACE), New Orleans District, was the governmental agency in responsible charge of the project, including approval of design, construction, and payment. The lead designer was URS Corporation, and the general contractor was Archer-Western-Alberici (AWA), working in alliance with TREVIICOS South (Trevi), who, together with Fudo Construction (Fudo), performed the deep mixing.As part of the quality control and quality assurance program, over 500 production deep-mixed elements were cored, the cores were carefully logged, and over 5,000 unconfined compression tests were performed. In addition, selected specimens were carefully photographed before and after longitudinal saw cuts to expose the interior of the specimens. This paper provides a summary of the coring and testing performed on the production deep-mixed elements."

Jan 1, 2015

There is an increasing demand for refrigeration as mines progress deeper. Mine refrigeration plants have special requirements in terms of their duty and application. There are no national or international standards applicable for the testing of these plants. There have been numerous examples of plants failing to meet their required nominal capacity, but with the plant owner being unable to enforce the design duty on the manufacturer, due to a failure to unambiguously specify testing or accepting criteria. This paper discusses some of the reasons for performance shortcomings and describes a methodology for developing and specifying performance criteria, when inviting design and/or construction tenders, and a method of performance and acceptance testing once the plant has been built and commissioned. These guidelines encompass the full range of performance duties expected from a mine plant, including stable operation at maximum and minimum plant capacity, and under the various combinations of extremes of high and low condensing and evaporating pressures. A description of a methodology for developing a computer-based mine refrigeration plant model is also presented as part of the performance testing criteria. The methodology is illustrated with reference to the actual performance and acceptance testing process used on the worldÆs largest bulk air cooler (R67) at the Enterprise mine at Mount Isa, Australia.

Jan 1, 2005

By Donald A. Voigt

This publication is intended as a guide for persons responsible for obtaining classifications and approvals Tor explosive compounds, mixtures, or devices from the U.S. Department of Transportation (DOT). It deals with the procedures for obtaining a recommendation for classification of an explosive compound, mixture, or device through the Bureau of Mines. Completion of the procedures detailed in this guide is necessary to obtain a final classification of an explosive compound, mixture, or device and assignment of an "EX" number by DOT, which will allow the producer to offer the explosive for shipment within the United States. This guide also gives procedures for obtaining United Nations (U.N.) classification for shipping outside the United States.

Jan 1, 1990

By Larry A. Haas

The Bureau of Mines is investigating the use of complex gas mixtures, such as would be produced by coal gasification, for reducing iron oxide pellets. A semiempirical model was developed to predict a priori the reduction rate of an iron oxide pellet in a laboratory tube furnace with simulated coal gases at a total pressure of 1 atm. The model uses partial pressures of H2, CO, CO2, CH4 (methane), water vapor, and a temperature-dependent term to predict the rate of reaction. The reduction rate (weight loss per unit of time) was determined with a thermo-gravimetric apparatus (TGA) in the temperature range of 800° to 1,100° C. The ranking of the reductants, in order of decreasing reducing potential, is H2, CO, and CH4 at temperatures between 800° and 1,000° C; however, CH4 became a better reductant than CO at 1,100° C. Water vapor did not decrease the reduction rate as much as did CO2, Carbon deposition was observed at temperatures below 1,000° C with gas mixtures with high CO and low H2 concentrations. With low-Btu gas at 900° C, over 90 pct of the iron oxide was reduced to iron in less than 3 h. This com-pares with a reduction time of 1.5 h with medium-Btu gas and 1 h with reformed natural gases.

Jan 1, 1985

By D M R Stone

Failings di,posal presents many challenges at metal mines in Peru where the extreme relief yields very little useable land area. These challenges are amplified by other hazards such as the high seismic environment, high rainfall, and frequent debris torrents down the steep hillsides. As a result many mining operations have resorted to in-river disposal of mill tailings. This paper presents a case history of one underground metal mine located in the high Andes mountains of Peru where in-river disposal of tailings has been used for some 30 years. The paper outlines a comprehensive engineering study that was aimed at increasing the utilisation of tailings disposal underground. The scope of this study included modifications to the grinding and flotation circuits to produce a high sulphide and low sulphide tailings for underground and surface disposal respectively: mapping and cataloguing of abandoned stopes for hackfilling: modifications to the existing backfill system to improve icliahility and throughput, and a preliminary evaluation of alternatives Such as silicated and paste fills.

Jan 1, 1998

By M. Board, C. Alexander, M. Ramström, W. Johnson

"Hecla - Lucky Friday is an underground mine that has been in operation for 75 years and produces from depths over 2.3 kilometers (7,500 ft) below the surface. Technical issues at this depth include high rock temperature and seismicity. The use of new innovative mechanization is being examined for ore production. Principal advantages of mechanization are the reduced exposure of personnel, reduced seismic triggers, and increased efficiency. Lucky Friday has partnered with Atlas Copco to design, build, test, and deploy a mechanical miner for primary ore production. We will discuss the background, concept, development and deployment strategy for a mechanical miner. BACKGROUND The Coeur d’Alene mining district in northern Idaho was first discovered over 130 years ago and has been the most prolific silver producer in the US. The Lucky Friday mine of Hecla Mining Company (Figure 1) is the oldest (76 years) and deepest mine in the district, producing approximately 725 tonne per day (800 tons per day) from narrow, vertical lead-zinc-silver veins. Cut and fill stoping is currently employed at depths to about 2.3 km (7,500 ft) below ground surface, making it one of the three deepest mines in the western hemisphere. The #4 Shaft, a winze collared at approximately 1.8 km (5,900 ft) below surface and extending to a depth of nearly 2.9 km (9,600 ft), was recently completed to exploit the high-grade 30 Vein, whose resource has been identified to at least 3 km (9,700 ft) in depth.As with other deep mines worldwide, a number of challenging engineering issues occur, including: high in situ rock stress, squeezing ground, seismicity, high in situ rock temperature that requires refrigeration, diesel equipment that generates diesel particulate matter (DPM) requiring high ventilation airflow rates, and long travel times for worker transport to mining areas."

Jan 1, 2018

By Floyd C. Bossard

Major air-cooling facilities have been installed and are operating underground in U.S. copper mines (at Butte, Montana and Superior, Arizona), South African gold mines, European coal mines, Zambian copper mines, Kolar gold mints in India, Coeur d'Alene mines of Idaho, plus other locations. Mining districts that utilize spot cooling and local air-cooling facilities would increase the above listing many times. The principal objectives of underground mine air cooling systems are to; 1. Extract heat and water vapor from the mine air and provide for their removal from the mine via (a) water pumped to surface or (b) discharged mine exhaust air. 2. Provide acceptable artificial ambient conditions for efficient working mine development and production areas. 3. Recondition used air, thus supplementing the limited mine ventilation balance. In deep South African gold mines, when rock temperatures exceed about 105°F, cooling of underground air becomes necessary for the poorly ventilated sections of the mine. By U.S. standards, air conditioning in poorly ventilated mines usually commences when the virgin rock temperature exceeds 100°F. On the other hand, well ventilated deep vein-type mines in the U.S. have been able to forego general cooling of mine air until the mines have been extended to such depths that the virgin wallrock temperatures have exceeded 115° - 120°F. It must be noted that spot-cooling of selected development headings may be required in many mines where rock temperatures exceed 100°F. Such local air-cooling practices permit the driving of long headings for exploration or extended development prior to holings which result in establishment of primary ventilation circuitry. The following summary reports on current practices of cooling and dehumidifying mine air. General system descriptions with important details will hopefully permit the reader to gain an understanding of the subject. A list of references is provided for this serious investigator who desires a more thorough knowledge-base and who is responsible for the design of a mine air-cooling system.

Jan 1, 1973

By H. A. Doerner

"Statistics show3 that in 1927 the United States consumed 60 per cent of the world's production of chromite and produced none. The uses of chromium are rapidly expanding, and a dependable supply is essential to our industries. Since the political and commercial control of our supply of chromite is now in foreign hands it is prudent to determine how this nation can best prepare to supply its need for chromium if the imports were cut off or restricted and whether it is now possible and desirable to stimulate production of chromium from domestic deposits. We have only small amounts of chrome ore of high enough grade to meet present trade requirements. Therefore the Bureau of Mines is making a study of methods for beneficiating our low-grade ores.Chromite (Cr203.Fe0), is the only commercial source of chromium. For refractories, a minimum of 40 per cent Cr203 and a low silica content (5 to 8 per cent) is usually specified. In most instances bricks or lumps having a physical structure that will not crack when heated are required. The only way an inferior ore could be adapted to this use world be to concentrate it and develop a satisfactory binder to make bricks out of the concentrates.For making ferrochrome the requirements are a minimum of 45 per cent Cr203 and an iron content not exceeding one-third of the chromic oxide. On some low-grade ores gravity concentration will give a product meeting these requirements; other ores yield concentrates containing too much iron or other impurities to be saleable."

Jun 1, 1930

By E Amini, N Beaton

Application of ore pre-treatment technologies such as Grade Engineering and in-pit ore treatment require comprehensive analysis to show proven value is being generated by switching away from the current practices. Considering most mining operations last several decades; the pre-treatment technologies must be investigated over life of mine (LOM) to truly show value was generated or lost. Therefore, a value chain flowsheet model has been usually developed to run those analysis/simulations over LOM using a set of defined ore characteristics and the operating equipment capacities. The Integrated Extraction Simulator (IES) has been used for this type of LOM simulations as it has the necessary flexibility to accommodate the whole mining value chain flowsheet in one platform and provide essential computation power to run mass simulations for the given scenarios using the resource block model. This allows to forecast the value chain response over the LOM for all given scenarios in a feasible amount of time. The traditionally developed flowsheets for the analysis/simulations are usually limited to the mineral processing plant and were developed based on a single audit conducted on the plant in a very short time, e.g. several hours. The process plant is being operated with a certain ore at a certain condition during the short audits which would not reveal the true capacity of the operation equipment when ore characteristics changes. Therefore, it would be impossible to define the true capacity of the equipment and identify the constraints along the value chain. In this study, a methodology was developed to identify the constraints along the value chain and implement those in the IES flowsheet using process information. Also, the equipment models were fine-tuned to represent the true capacity of the equipment in the flowsheet using the same data set. The methodology was tested over five process plant in two operations and showed the performance forecast was significantly improved over the traditional methods being used so far. Then those high-fidelity flowsheets were used to evaluate the ore pre-treatment technologies proposed by the relevant resource company experts.

Nov 10, 2020

By Muhammad Zia-ur Rehman, Sajjad Ahmad Baig, Gao Lei Fu, Muhammad Mohsin, Sobia Naseem

The environmental sustainability is linked back to human sustainability; the sustainability development comprises the environment, society, and economics. This study mainly focused on mining sustainability and its impact on the environment of Pakistan. For measuring the level and environmental mining sustainability nature to Environmental Impact Assessment (EIA) of the world’s second largest salt mine “Khewra,” Folchi and Phillips environmental sustainability mathematics (PESM) models are applied. The results indicated that Khewra was deemed to be potentially sustainable for the environment and quality of salt. The results suggest the delicate balance of maintaining sustainability successfully, regarding the mining process in Pakistan which has been affecting the local environment and community so badly. The paper lays emphasis on the potential significance of the model’s application for the attainment of sustainable mining goal. This study is essential for researchers,

May 11, 2020