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By K. Huttu, A. van Wageningen

"The Kittila underground mine is extracting one of the largest known gold deposits in Europe and has an estimated mine life through 2034. The Kittila mine is located in the Lapland region of northern Finland, approximately 900 kilometres north of Helsinki and 150 kilometres north of the Arctic Circle. The ore at the Kittila mine is refractory, making gold extraction relatively difficult because the gold is generally locked inside the two main sulphide minerals – arsenopyrite and arsenic-rich pyrite. The orebody consists of a complex pinch-and-swell structure including up to 100 different gold mineralization lenses. Fluctuating gold prices and other economic considerations increase the need to maintain and update the mining plans more frequently. Over 3,000 stopes need to be designed and optimized once a year for the LOM-plan and more frequent updates are needed for mid- and short term plans as well as production studies. As a large part of ore body is very sensitive to cost and gold price changes, stope optimization is crucial to find the best possible production scenario for every moment. Kittilä has gone through several methods for stope design and optimization. Most recently the Deswik Stope Optimizer (Deswik.SO) has been successfully used. Dewsik.SO has helped Kittilä to design and optimize stopes in a timely and cost efficient manner. Together with the Deswik suite of planning tools, reliable and auditable plans are created. This paper describes the methods used for stope design and optimization at Kittilä. THE KITTILÄ MINE The Kittila underground mine in northern Finland is extracting one of the largest known gold deposits in Europe. Kittila achieved commercial production on May 1, 2009, becoming Agnico Eagle’s first mine to open outside of Canada. The Kittilä orebodies were initially mined from two open pits – Suuri and Roura – and underground operations were added in October 2010. The pits were mined out in November 2012, so mining is now entirely underground in the Suuri and Roura deposits. Additional, small open pits will mine any remaining near-surface reserves in the future. The Kittila mine is located in the Lapland region of northern Finland, approximately 900 kilometres north of Helsinki and 150 kilometres north of the Arctic Circle. With a mine life estimated through 2034, Kittila is Agnico’s longest-life mine. The mine’s proven and probable reserves contain 4.7 million ounces (32 million tonnes at 4.6 g/t gold). Ore has been mined from underground since 2010. The original 3,000-tonne-per-day operation underwent an expansion project to increase the throughput capacity of the mill by 1,000 tonnes/day, starting by the end of 2014. The region around the Kittila mine is underlain by mafic volcanic and sedimentary rocks of the Kittila Greenstone Belt oriented nearly vertical. This greenstone belt is similar to those hosting Agnico Eagle's Canadian deposits in Quebec’s Abitibi region and Nunavut. At the contact point between iron-rich and magnesium-rich rocks lies a 100- to 200-metre-thick structural zone known as the “Suurikuusikko Trend”. This trend hosts the Kittila deposit, where multiple mineralized zones have been traced over a strike length of more than 25 kilometres."

Jan 1, 2015

By D. Mannino, M. Townsend, E. Daniloff, E. Petsonk

Nonspecific airway hyperresponsiveness is a central feature of chronic obstructive pulmonary disease and has been associated with atopy in the presence of environmental exposures such as smoking. Coal miners may develop emphysema and lung function decrements related to their dust exposure. To investigate the roles of airway responsiveness and atopy in these effects, we evaluated the relationships among smoking, a history of hay fever or seasonal rhinitis, and airway responsiveness (AR) in a group of underground bituminous coal miners and nonmining controls. All subjects had a normal baseline percent predicted FEV1. Increased airway responsiveness was defined as a 15% or greater decrease in FEV1 following methacholine inhalation. 85 miners (45%) and 99 controls (46%) were current smokers, 59 miners (31%) and 55 controls (26%) were lifetime nonsmokers; former smokers were excluded from further analysis. Overall, 25% of 144 miners included in this analysis had increased AR, as opposed to 29% of 155 controls. Thirty-seven percent of both miners and controls had a history of seasonal rhinitis. Significantly fewer miners (15%) had a history of hay fever compared to controls (25%, p< 0.05). Figure 1 shows the prevalence of increased airway responsiveness in miners and controls in association with hay fever. In both smoking and non smoking miners the presence of a hay fever history was significantly (p < 0.05) associated with a higher prevalence of increased AR. Little effect was seen in smoking (p = 0.827) and nonsmoking (p = 0.940) controls. Figure 2 shows a similar higher prevalence of increased AR associated with symptoms of seasonal rhinitis in the mining group. This was significant (p < 0.01) only in the smoking miners. No effect was noted in the controls. These data suggest that the association of a history of hay fever or seasonal rhinitis symptoms with increased airway responsiveness may be more common in coal mining, compared to non-dusty occupations. Although the implication of these data are unclear, further study of miners with a history of hay fever or seasonal rhinitis is warranted. (Supported in part by Grants G1135142 and G1105142, US Bureau of Mines, Generic Mineral Technology Center for Respirable Dust)

Jan 1, 1991

By Yasutaka Izumi, Tuneyoshi Funasaki, Keizou Miki, Makoto Kanai

INTRODUCTION The Tokyo metropolitan region, inclusive of the three neighboring prefectures (Kanagawa, Chiba and Saitama), has the highest concentration of population and industry in Japan; with an area of 13,500 km2 or only 3.6 percent of the national land, but a population of 31.8 million or 26 percent of the national total, and a Gross Regional Product (GRP) of146 trillion yen (US$1.33 trillion) or 36 percent of the overall GDP of Japan as of 1990. Tokyo Bay, embraced by the metropolitan region as shown in Figure 1 (a), (b), is open to the Pacific Ocean as a marine gateway and is ringed by large cities and coastal industrial zones together with port facilities, all constituting a concentration of prime importance for socio-economic activities. The Trans-Tokyo Bay Highway (TTB highway), which crosses Tokyo Bay and provides a direct transportation link between the west and east sides of the bay, has long been a sort of “dream-come-true” project in Japan. As shown in Figure 2, linked to the Tokyo Bay Shore Highway, Metropolitan Inter-city Expressway, Tokyo Outer Ring Road, Tateyama Express-way, and others, it will play an important role as an integral part of the regional highway network to help promote spatial redistribution and multi polarization of urban functions, thus strengthening cooperation between cities and restructuring the metropolitan region. The TTB Highway is a project of 15.1 km in length designed as a toll highway to cross the bay by tunnel and bridge, and will require ten years to complete at a total cost of more than 1,438 billion yen (US$13.077 billion). Following more than 20 years of feasibility and environmental assessment studies, construction of this giant offshore highway project commenced in May 1989 as the first large-scale highway project with private-sector participation under the prevailing Road Law in Japan, and is now steadily proceeding toward the goal of completion in fiscal1997. This report outlines the entire TTB Highway project, including the waterproofing measures against high water pressure and the rapid construction method applied in the construction of shield tunnel sections.

Jan 1, 1997

By V. F. Hollister

Although it is apparent that porphyry copper deposits may have formed in the Andean orogen from Permian time onward, commercial deposits thus far developed appear to have Cenozoic ages only. Following Chilean useage, the porphyry deposits are divided into either: (a) a stockwork, or (b) a tourmaline breccia model. Stockwork type deposits tend to occur on major strike-slip structures which may be either transcurrent or transverse faults related to plate movements, while the tourmaline breccia type may more frequently form independently of regional faults. Stockwork types may be conjugate fractures that intersect and ramify as a network of veins and veinlets, cutting altered rock, and genetically related to a major regional fault, or in other cases they may reflect the intersection of two large fault structures. Circular fracture patterns may more rarely occur in stockwork deposits in response to a specific intrusive. Tourmaline breccia types are characterized by circular and radial fractures genetically connected in some way with tourmaline breccia pipe development or an intrusive piercement that preceded it. Each structural model contains mineralized intrusive8 which have a close time-space relationship to ore and in each model, mineral paragenesis is similar, regardless of age or location in the plate. Also, each model contains the same type of alterntion halos (phyllic, argillic, prophllitic) as are developed in intru- sive centers in porphyry copper deposits found in North America. The potassic zone is exposed best in the older deposits. Examples of each type of deposit are illustrated, demonstrating the feature evolved.

Jan 1, 1973

By Ali Ahmadi, Andreas Rausch, Angela Binder, Stefan von der Mark, Dhananjeyan Jeyaraj, Jana Thomas, Nils Nöther, Stefan Wittek, Thomas Ulrich, Eleni Gerolymatou, Mario Kolling, Helmut Mischo, Michael D. Martin, Tugba Sen, Jens-André Paffenholz, Erik Farys, Oliver Langefeld, Wilfried Ließmann

This paper outlines the MOVIE project, which is dedicated to the development of a digital twin for an underground laboratory at the research and education mine “Forschungs- und Lehrbergwerk Reiche Zeche” in Freiberg (Germany). Data from induced rock deformations and temperature fluctuations is collected using fiber-optic sensors. The objective of the project is to establish a virtual laboratory for real-time analysis of the interactions between geology, geomechanics, geometry and ventilation. This will be achieved by integrating models from these disciplines with artificial intelligence, creating a sophisticated hybrid model. The insights gained will enhance safety and efficiency in mining through improved digital planning options, ultimately contributing to more effective resource management and operational decision-making in different mining environments.

Feb 1, 2025

By K. D. Dill, S. D. Rosenthal

"This project analyzed the efficiency of incorporating the use of directional drilling technology into coal overburden blasting. Directional drilling is currently in use in the petroleum industry and it is believed that it will be a valuable asset in the mining industry. This project has shown that directional drilling can be a viable technology for use in the coal overburden removal process resulting in increased drill utilization and potential for cost savings. Future work regarding blasting and geotechnical evaluation should be performed to solidify the concept. INTRODUCTION Determine the Potential Drill Utilization Improvements and Rock Fragmentation Requirements Using Directional Drilling in a Coal Mining Overburden Highwall Application was a project funded by Atlas Copco Drilling Solutions (Atlas Copco) based in Garland, Texas. The project was developed and researched by Montana Tech of the University of Montana. This project was intended to show how the use of directional drilling can increase drill utilization, while decreasing costs, without sacrificing overburden blasting results. This paper discusses a review of the literature, the methods performed, conclusions and future work Background Atlas Copco theorized that an alternative approach to drilling a cast blast pattern would be beneficial to coal mining operations by increasing drill utilization and therefore, decreasing overall costs. Increasing drill utilization results in an increase in the amount of time the drill is being productive, more drill bit time in ground. Atlas Copco’s theory is that by incorporating directional drilling in the design of a cast blast pattern, a drill rig could drill multiple directional holes, which are the length of the cast blast, from a single drill set up. A drill goes through a series of steps to complete a blast pattern, these steps include: tramming to the hole location, leveling the drill, drilling the hole, cleaning the hole, and retracting the drill steel. This process is repeated hundreds of times until the whole pattern is completed. The most time intensive part of this process is drilling out the blast pattern. Integrating directional drilling into this process could decrease the overall time it takes to drill out a blast pattern by minimizing the amount of time spent tramming and setting up for each subsequent hole. Tramming and drill set up are two aspects of the drilling process where the drill tool is not engaged with the ground and, therefore, not being utilized. A cast blast pattern designed to utilize directional drilling will allow the drill tool to be “in the hole” longer and will improve drill utilization. This change in drill utilization was analyzed. Drilling and blasting is an integral part of the mining process, without it mining would be a slower process and less material could be mined. Drilling provides holes for explosives to be loaded and then blasted. Blasting has been found to be the most cost effective method of fracturing rock in mining for material removal. Blasting is also used as a low-cost method to move the rock in open cast mines, commonly referred to as cast blasting. A cast blast is a large blast designed to allow for the maximum amount of overburden to be removed using explosives in a single blast thereby uncovering more coal at a lower cost. Not all the material is blasted to cast-to-final, which requires that material to be moved with large mining machinery. Material that is considered cast-to-final is material that has been thrown far enough to never be moved again in the coal extraction process, therefore it is in its final resting place. A typical cast profile and cast-to-final area is shown in Figure 1."

Jan 1, 2017

By J. S. Hall

The performance of a full-scale balling drum circuit is well known to fluctuate, and is difficult to control. The fluctuations stem from causes that are not readily quantifiable. Results from plant tests show that surging of the recycle tonnage is dependent, inter alia, on the method of bentonite addition, small variations in drum or cutter bar speed and the method of drum water addition. This paper presents detailed observations from experiments on full-scale drum circuits, including preliminary results of a computer measurement and control scheme. Size distributions measured at points along an industrial drum are presented, together with the results of a mathematical model fitted to one plant data set. Model predictions are also included emphasizing the importance of screen performance on circuit stability.

Jan 1, 1983

By Richard H. Johnson

White Pine Copper Company began operating a plant designed to recover copper matte particles from its discarded reverberatory furnace slags in April, 1972. Seasonal operation, generally the months of April through November, continued through October, 1974, shutting down due to feed exhaustion. White Pine reverberatory furnace slags contained economically recoverable amounts of copper for a variety of reasons. Among them: 1. White Pine concentrates yield extremely high matte grades (65+% Cu) and viscous slags. The viscous slags hinder matte settling causing matte "prill" entrapment in the slag 2. White Pine&apos;s total refined copper output is accomplished through fire refining requiring special techniques for impurity removal, especially small amounts of nickel introduced from purchased pyrite and chalcopyrite fluxes. These special techniques result in high-copper slags, refractory to smelting, being reintroduced into the reverberatory furnace. 3. Furnace bottom build-ups resulting in mechanical matte entrapment through the formation of "dams" hindering matte flow. 4. Occasional concentrate fluxing and furnace atmosphere control problems.

Jan 1, 1975

By M. Pope

By 1950, packaged oil and gas fired boilers were on the market for capacities up to 30,000 lbs./hr. By 1962, capacities were up to 100,000 lbs./hr. and even higher when separate heat traps (economizers and air preheaters) were attached to the boiler packages. During this period, complete packaged coal fired boilers were developed into the 10-15,000 lbs./hr. range. Oil and gas fired plant designs became very compact, "basementless", and far: less costly than the coal fired plants of any capacity. Coal fired plants remained large and expensive, required field erection of boilers and required basements. With the spectacular strides made in automation, oil and gas plants became far less costly to operate, requiring only the part-time attendance of a single operator. Some advances were made in coal fired plants, but not to the degree attained by the fluid fuel competition.

Jan 1, 1965

By Edmund C. Bitzer

The unit costs of dry crushing, within the efficient reduction ranges of equipment now available, are considerably less than those for the reduction of the same material in the same size range by wet mills using grinding media. This is a commonly recognized fact which cannot be disputed; still we find considerable imbalance in costs between the crushing and grinding sections of many operating mills. In other words, there are many instances where more reduction could be accomplished in the lower cost area of dry crushing. In addition, equipment is now available which enlarges this area and makes it possible to consider much finer crushing than has been heretofore economically feasible. There has been a tendency for the past two decades to utilize wet rod-¬milling in the reduction step. This actually amounts to a fourth stage of crushing in a size range formerly served by roll crushers which have fallen from favor on account of high operating costs, Some investigations have indicated that there is not sufficient difference in the operating costs of rolls and rod-mills to give a clear cut advantage to either method, There has for some time been a need and an opportunity for a dry crushing process to improve costs in the range now served by these two systems.

Jan 1, 1965

By J. M. Mutmansky

Procedures for the determination of the operating point that results when a fan is connected to a mine and placed in operation have been a source of confusion for decades. A variety of techniques has been proposed, often to the detriment of the practicing ventilation engineer trying to apply the techniques to the multitudinous mine/fan matching problems encountered in the engineering of mine ventilation systems. This paper proposes some different assumptions and definitions that simplify the solution of problems in determining the operating point. As a result, a set of procedures is provided for determining the operating point of a mine/fan system when the outlet areas of the mine and fan are known. To establish a major simplification in the development of procedures for the determination of the operating point for a mine/fan combination, the authors suggest that the mine pressures be redefined and then used in such a manner that the conservation of energy is ensured. This provides the basis for the recommendation that the fan operating point be determined using the fan total pressure and the system total pressure characteristics. This procedure can be applied to fans in the blower, booster, and exhaust positions equally well and to both single-fan and multiple-fan networks. Another advantage is that the method is applicable to fans with or without an evase.

Jan 1, 1984

By Aditya Rana, A. K. Mishra, M. M. Singh, N. K. Bhagat

Many cases of derailments, train accidents and traffic interruptions due to the slope failure and rockfall have been reported in the past on Konkan Railway trackline because of unstable rock slopes. The authorities installed stability enhancement measures such as wire netting, rock bolting and shotcreting which were found insufficient. The failed geotechnical measures prompted authorities to excavate the unstable slopes. Rock blasting is generally not preferred for the excavation of slopes along the transportation route. Stabilisation and widening of slopes by blasting need closure of the route for a considerable time. Discontinuing traffic on Konkan Railway route was undesirable as it connects two important port cities Mumbai and Mangalore. The paper presents an innovative directional controlled blasting (DCB) technique for excavating three unstable slopes (i.e. cuttings) in hilly terrain without disturbing the traffic. New empirical formulae for small geometry blasting to estimate burden and throw have been developed. A unique sequence of firing and excavation with powder factor varying between 0.04 and 0.39 kg/m3 irrespective of the actual free face was used for rock blasting. The process controlled the throw of blasted material over the track and resulted in a smooth and stable wall. The slope angle has been changed from 80 to 82° to 45–47° and probability of wedge, planar and toppling failures was eliminated. About 100 such accident-prone slopes (52 km cumulative length, 7% of total route) situated at discrete locations along Konkan Railway route were stabilised using the technique. No incident of rockfall and slope failures along these slopes has been observed after the stabilisation.

By Rick Wright, Gregory L. Raines

San Diego’s need to access underground space has dictated a long history of tunnel construction. The subsequent need for water, coupled with San Diego’s natural attractiveness, has led to significant population increases and the need for the construction of even more tunnels. This paper will provide a historical summary of the progression of mines, tunnels and trenchless technology construction projects that contributed to the advance of current technology, feasibility, and constructability. A bibliography of relevant reports and data is included.

Jan 1, 2001

By J. Peterson, P. G. McKneight, R. W. Riggle

"In today’s tough mining environment, the key difference between surviving and thriving is how effectively miners optimize their mining operations. This paper explores the compounding effects material fragmentation has on mining costs and productivity. Starting with Drilling & Blasting and moving through Loading, Hauling and Crushing, we will illustrate the opportunities to improve operating costs and thus reduce the Total Cost of Ownership (TCO) of the total mining system. Material size and volume requirements are defined by the needs of the mill or plant, and drive the rest of the mining operation from the bench through the primary crusher. At the start of the cycle, where the rock mass is drilled and blasted, the degree of fragmentation and the muck pile configuration is determined, as are the effects on cost and productivity of all subsequent operations. The cost of managing oversize material, including secondary handling, overloading, reduced diggability, higher load factors, extra pass loading, increased wear factors and reduced throughput rates are all opportunities for improvement. INTRODUCTION Having visited many mine sites, customers are asking about the end to end mining value stream from loading to the crusher and how to measure the cumulative effects on each part of the mining value chain. All too often improvements with the drilling and blasting may impact fragmentation but do not improve the diggability so the net change on cycle times or fill ratios is perceived to be negative at the point of loading but they may have a measured improvement going through the primary crusher. This makes it imperative to measure performance at each segment of the overall cycle and not just to the crusher but thru it. The mine material process originates with solid rock, in-situ, and progresses to the output of the primary crusher. Once the material has passed the primary crusher the fragmentation size will be fairly consistent and promote consistent throughput of the plant or mill. In late 2015 a team was conducting a reassessment visit to a mining site where we had conducted the original assessment the previous year. Though many recommendations had been implemented some had been so recent they had not resulted in the expected improvements. Several of our discussions kept returning to the impact fragmentation has to both the individual components of the mining cycle but also the effects compounded throughout the end to end process. Let’s move on to the big picture before we zoom in on any specific operation. To get from native solid rock in the ground to a concentrate or a saleable rock product we have the same steps from one end of the chain to the other. We start with a good geological understanding of the rock type, characteristics and structure to design a program to drill and blast the rock. The height of the bench is relevant to the annual production requirements and the specific digging equipment models best sized to make that production possible. The equipment selection influences the bench height which has a minimum and maximum hole diameter to maintain a safe blasting operation. The diameter of hole burden and spacing all contribute to a good distribution of explosives in the rock mass that maximize the fragmentation and preconditioning of the rock when blasted. The timing and sequencing of hole initiation will yield a fragmentation ratio and diggability factor. Fragmentation size adjustments will change the meters drilled per cube and the explosives consumption will change accordingly. Drill cost is measured by meter drilled so a direct relationship exists between cost and fragmentation sized achieved."

Jan 1, 2017

By R. R. Rollins

Field data combined with theoretical derivations were used to design a surface mining operation using explosive casting. Calculated casting results varied from 44 to 65% depending on bench height. This model was then applied to field tests at an eastern, contour, coal stripping operation. Results from five experimental casting shots are summarized and compared to conventional mining methods. The overburden material removed by casting ranged from 35 to 65% with an average of 50%. Fragmentation was improved and it was demonstrated that this procedure is economically attractive with a resultant cost savings of $.59/m3 or $5.93/t of clean coal.

Jan 1, 1987

By Donald B. Buddecke

The large walking dragline has been the most commonly used overburden-stripping equipment in Texas lignite mines. These draglines, operating on the highwall side of the pit, can excavate to a depth of 45 m. Considerable overburden rehandling is required beyond 30 m and may reach 100 percent by 45m. As pit depth increases, dragline operation also becomes more complex, with two or three passes being necessary to make a cut 30 to 40 m wide. Multiseam operations also require extra passes. At pit depths greater than 45 m, the dragline must be operated also from a spoil bench. This can be difficult in the unstable alluvial deposits in which lignite occurs.

Jan 1, 1985

By William J. Cavanaugh

In 1967, Lone Star Industries, Incorporated, established a program to carry out the necessary Research and Development work to bring the Snyder Process to commercial application. The purpose of this paper, is to present the detailed metallurgical results gathered in comparing the Snyder Process of Comminution to conventional grinding techniques for the milling and beneficiation of certain ores and other materials. The ore primarily dealt with in this paper is a fine-grained magnetite ore, containing approximately 23% acid-soluble iron, from the Lake St. Joseph ore body of Steep Rock Iron Mines, Ltd., (Canada). This particular example was selected for presentation in this paper because extensive comparative results with conventional methods of grinding were available. Also briefly described in the paper are the comminution of other ores and materials such as porphyry copper ore and alumina.

Jan 1, 1973

By Kanhaiya Mishra, Phanil. K. Mandal, S. K. Behera, P. S. Paul, Prashant Singh, C. N. Ghosh

The techno-economic and social benefits of cemented paste backfill (CPB) resulted in its wide acceptance by the mining industry. The Ordinary Portland Cement (OPC) remains the key binder but to diminish its economical constraints, suitability of alternate binders has been examined worldwide. The present study aimed to investigate the effect of partial replacement of OPC with fly ash on the CPB’s strength and to determine the most optimal mix to achieve the required strength (1 MPa at 28 days of curing) at the most cost-effective way using the Bayesian network (BN). The CPB mixes were prepared at 72 wt.% solid concentration with mill tailings (87–91%), OPC (6–13%), and fly ash (0–4%), and instantly after mixing, fresh (slump, bleeding, density) CPB properties were measured. The strength was tested at 7, 14, 28, and 56 days of curing and initially analysed through traditional model. The traditional models follow the aleatory principle and are considered not appropriate for geotechnical engineering. Hence, the BN model was developed and tested. The reliability of two classifiers in learning model structure was compared which gives Naïve Bayes as the highest reliable tool. The CPB’s strength is most sensitive to the OPC content. The most consistent mix(s) is mill tailings: 87–88%, OPC: 9–11%, fly ash: 1–4%. Adding fly ash at 89–91wt% mill tailings possesses high failure probability of the CPB. The collinearity test indicates that the fines percentage and chemical composition of CPB’s ingredients are highly correlated with its slump, bleeding, and strength development.

Jul 14, 2022

By T. M. “Mike” Gurr

"U. S. marketable phosphate rock production for the 2018 crop year (July 1, 2017 to June 30, 2018) showed an increase to 27.7 Mt (30.5 million st) from the 2017 crop year production of 27.1 Mt (29.8 million st), as reported to the U. S. Department of the Interior by the mining companies operating in the United States during crop year 2018 (Table 1). Domestic consumption decreased in the 2018 crop year to 23.9 Mt (26.3 million st), as compared to the 27.2 Mt (30 million st) consumed in crop year 2017. U.S. imports of phosphate rock increased to 3 Mt (3.3 million st) for the 2018 crop year from the 1.81 Mt (1.9 million st) for crop year 2017. Imports of phosphate rock were from Peru and Morocco. There was an increase of 2.85 Mt (3.1 million st) in producers’ stocks from 7.45 Mt (8.2 million st) in crop year 2017 to 10.3 Mt (11.3 million st) in crop year 2018 (Table I).For calendar year 2018, Florida and North Carolina phosphate rock production accounted for 75 percent of the total U.S. production, with the balance being produced in Idaho and Utah. The Florida and North Carolina percent of U.S. production is declining. U.S. production of 25.7 Mt (28.3 million st) in calendar year 2018 was produced by five companies at 10 mines in four states and accounted for approximately 9.6 percent of the world production of 267.7 Mt (295 million st) for calendar year 2018 (Table 2).In calendar year 2018, U.S. phosphate production decreased by 2.2 Mt (2.4 million st), while companies increased stocks of phosphate rock by 1.7 Mt (1.9 million st). The domestic production for calendar year 2018 of 25.7 Mt (28.3 million st) represents 78.4 percent of the domestic production capacity of 32.8 Mt (36.1 million st).At present, China reportedly produces more than 50 percent of the world’s phosphate rock. As a result of planned expansions, Morocco production has increased from 27 Mt (30 million st) in 2017 to 33 Mt/a (36 million stpy), which is approximately 12.3 percent of the world’s production, moving ahead of the U.S. production. In calendar year 2018, the U.S. phosphate rock production was 25.7 Mt (28.3 million st), which is less than 10 percent of the world’s production of phosphate rock (Fig. 1)."

Aug 1, 2019

By Molood Saeedi, Nima Sadeghi, Eskandar Keshavarz Alamdari

This study showed the kinetics of gold chlorination leaching from an anode slime complex containing copper, silver, barium, selenium, and tellurium. AgCl and BaSO4 control the rate of gold leaching, and the interface transfer and diffusion across the product was selected as the desirable kinetic model. The concentration of hydrochloric acid, chlorine gas flow rate, pulp density, and temperature have been precisely studied. The results showed that temperature and hydrochloric acid have a favorable influence on the leaching rate. However, the chlorine gas flow rate is inactive in the reaction rate. Furthermore, the leaching rate decreased with an increasing pulp density. The leaching process was controlled by the interface transfer and diffusion across the product, and the activation energy was determined to be 13.66 kJ mol–1.

Oct 19, 2021