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By T. J. Morris

"From November 2014 through September 2016, Westmoreland’s Coal Valley Mine utilized a Marion 7450 electric walking dragline to extract 1.2 million tonnes of a vertically-dipping, 52-meter wide, fault-thickened seam, beneath a conventionally-excavated pit bottom. Mine planning and operational hurdles encountered throughout a kilometer of mined strike length included mitigating highwall instability in both hanging and footwall slopes, water management, and end-of-life machine reliability and maintenance. The most substantial challenge has been the depressed global coal market. However, the successful application of this mining method significantly aided the overall cost of production at the site. The valuable experience gained from this project will be extended to future mining areas where this technique can be similarly applied. INTRODUCTION Coal Valley Mine is located approximately 220 kilometers southwest of Edmonton, Alberta, Canada, in the foothills of the Rocky Mountains. Since commercial production began in 1978, more than 84 million clean tonnes of coal has been produced and sold domestically as well as exported overseas for power generation. While the existing processing plant has a nameplate capacity of 4 million tonnes per year, the site has operated at a 2 million tonne rate since 2015 due to depressed export markets. Westmoreland acquired the mine as part of its purchase of Sherritt Coal Operations in April 2014. The mine operates 24 hours a day, 7 days a week, with an IUOE workforce of 261. GEOLOGICAL SETTING The geology of the Yellowhead Tower mining area differs significantly from previous mining areas at Coal Valley. The Mynheer coal seam discussed herein consists of steeply dipping, highly-sheared, thrust-faulted, and heterogeneous coal pods. Deep set, southwest striking thrust faults were the likely mechanism for the podding and have resulted in overturned, rolled and highly sheared seams with zone thicknesses of up to 65 meters.The hanging wall is characterized by soft bentonitic mudstones and siltstones, overlain by massive competent sandstone. The footwall is defined by a hard volcanic tuff, interbedded with thin, sporadic coal seams. The sandstone above and volcanic tuff below constrained the faulting system within this zone, enabling the mass podding that has afforded the opportunity for large-scale mining."

Jan 1, 2017

By Shular R. Scudamore

There was a time when heavy-duty apron feeders were built to last almost forever. The machines were heavy and built to maintain certain expected longevity and reliability. Increasing emphasis on return on capital leads to cost-effective designs. Costs can be increased or decreased, depending on customer specifications and the design philosophy of the feeder manufacturer. Specifiers can make important contributions to cost effectiveness. To do so, this article describes what should be known about planning with a supplier for heavy-duty apron feeders. There are several factors that should be considered, including capacity, characteristics of material carried, length, method of feeder loading, degree of lift, hopper opening, pan construction, and type of drive. Today's apron feeders are built in various sizes and with a variety of components to handle heavy, lumpy, or abrasive materials that can ruin lighter feeders. Careful feeder selection is important if it is to provide a controlled feed rate, absorb huge impacts, be used with long hopper openings, and have a relatively low power-perton capacity. Variations in systems and materials demand that each feeder be designed for its application, drawing on a wide range of components. Therefore, certain factors must be considered before final specifications are drawn. Capacity The design and application of an apron feeder is based on how material passes through a given system over the feeder into processing machinery. The processing machinery's capacity governs other equipment and productivity of the flow. From this basic information, related sizes of upstream equipment can be determined. Of these, the apron feeder acts as the "mouth and neck of the bottle." No more will pass than the mouth and neck will accommodate. Everything in the apron feeder design, from dimensions to power, is based on processing machinery. Material Considerations Are Vital Just as the nature of materials to be crushed or screened affects capacity and productivity of a crusher or a grinding mill, so does it affect capacity and performance of a feeder. Lump size must be identified closely so the pan width can be selected to accommodate the maximum lump with room to spare. For example, a lump size of 760 mm (30 in.) indicates a pan width of 1.5 m (60 in.) While two times the maximum lump is considered minimum width, if the maximum lump

Jan 12, 1983

By Carl G. Bock

INTRODUCTION The Washington Metropolitan Area Transit Authority (WMATA) has let a contract for the excavation of three rock stations, spaced about 3550' (1082m) apart, beneath Connecticut Avenue in Washington, D.C. Two previous rock stations have been completed, one of which had a crown pilot tunnel excavated to explore geologic conditions that would be encountered during the station excavation. This pilot tunnel was primarily for inspection by bidders and design was based on core hole information. The other rock station had no pilot tunnel and was also designed on the basis of core hole information. The three rock station areas, which are the subject of this paper, have been explored by crown pilot tunnels and an as-built report, with geologic maps and sections, has been prepared, for use by the design engineer as the basis of design for support of the structures. The geotechnical report describes the geologic structure in detail and relates it to excavation and support of the three crown pilot tunnels and the running tunnels which will be enlarged to create the 58'W x 44'H x 700'(+)L (17.6m x 13.3m x 212m +) stations. The information was complete enough to permit the design engineer to visualize the geologic structure and compute loads for the various types of rock conditions. It was therefore possible to determine the size and spacing of the steel sets for the several different conditions of rock structure encountered in the pilot tunnels.

Jan 1, 1976

By Enrique Rubio Esquivel, Matías Fuentes Bustamante

The transition from open pit to underground mining represents a key strategic challenge in modern mine planning. This paper presents a multicriteria and multivariable methodology to plan such a transition, integrating technical, economic, environmental, and social considerations. The proposed framework focuses on risk reduction (de-risking) through the identification of alternative development paths, the use of stochastic simulation to model geological and technical uncertainty, and different mining strategies such as methods, production rates, sequence, etc. for ultimate underground implementation. The study summarizes the limitations of traditional planning methods—which usually treat the open pit and underground mine designs as independent processes—, also often it considers a single mining method and production rate for the underground option. Instead it is proposed an integrated approach that leverages synergies between both mining methods and scalable production rates. Also the proposed methodology incorporates the definition of critical variables (geotechnical, economic, and social, among others), the construction of multiple operational and technological transition paths, and the evaluation of each path under criteria of economic attractiveness and practical do-ability (capacity to be executed). Using simulation tools (SIMPLAN for operational performance and SIMVAL for economic evaluation), probabilistic distributions of the different path are obtained, enabling the quantification of the project’s Value at Risk (VaR). A subsequent multicriteria analysis weights the expected profitability against the feasibility of execution of each alternative, producing a decision matrix that provides a framework for evaluating the different project path. As a result, the proposed approach identifies OP→UG transition strategies that maximize business value by capturing remaining resources while minimizing financial and operational risks. Finally, strategic conclusions and recommendations for the industry are provided, emphasizing the importance of early transition planning even challenging the open pit sequencing of the latest push backs by introducing early underground operations. Future work includes strengthening the methodology—such as integrating real-time analysis and adaptive decision-making—and providing recommendations for its application in various mining operations. In summary, this paper delivers a technical and professional framework to guide mining companies on when and how to carry out a successful transition from open pit to underground mining, balancing resource recovery with financial, operational, and social security.

Feb 22, 2026

By William R. Taylor

The information I have concerning the early in situ uranium mining activities in Texas is "sketchy." Apparently, the first test was conducted between 1967 and 1971 in Gonzales County, north of Yorktown, at the Dueterstat Site, by, City Service. Other tests were run in Webb County at the Bruni Site, and in Live Oak County at the Burns and Clay-West Sites in the early 1970's. During this period the Texas Water Quality Board maintained a general surveillance of these tests. In mid-1974, Mr. Glen Davis walked into our Austin offices and requested a permit for Atlantic Richfield's first in situ uranium mine. This permit, issued in January, 1975, was for 3 acres of production, and was the first such permit issued in the United States. It formalized the regulation of in-situ uranium mining in Texas. The 4 years since then have been characterized by growth and change. Today there are 20 permits approved for approximately 2,000 production acres. These permits are issued to 8 companies at 13 sites in 6 counties. There are 7 additional permits being processed, or on the verge of being applied for. These will involve possibly another 2,000 production acres, 3 more companies in one more county.

Jan 1, 1979

By Don Reining

The sand and gravel industry in Southern California has made a sizeable contribution to the construction development of the 15 County area in this part of the State. Facing this mineral extraction industry, in the next decade, is a problem of meeting the market demand. The question is, "how can sand, gravel and rock operations be made compatible with goals for sound and attractive community development?" Three basic principal characteristics of the land needed for this-mining industry are: 1. The land must have workable sand, gravel or stone deposits. 2. The land must be properly located for transportation purposes close to the market in order to sell the aggregate at a minimum cost. 3. The land must be zoned so that it will be permissible to mine.

Jan 1, 1967

By Thomas D. Boyce

Minewater from many western U.S. uranium mines often contains radioisotopes in concentrations higher than permissible for discharge to unrestricted areas. .Radium 226 has the lowest permissible concentrations of all radioisotopes and usually greatly exceeds this limit in untreated effluents. This radionuclide is derived from the natural radioactive decay of uranium: [ ] The treating of minewater effluents has been studied previously (1) and the classical removal method for radium has been adsorption on barytes or freshly precipitated barium sulfate. This paper describes the removal of radium from minewater by utilizing a selective complexer based on a crosslinked organic polymer matrix.

Jan 1, 1982

By M. M. Poulton

Universities are not immune from the large number of retirements facing the mining industry. Department heads anticipate that we will need to replace all of the senior mining engineering faculty in the U.S. by 2020. Long term studies of Ph.D. completion times and success rates indicate that in Engineering disciplines in the U.S., the ten year completion rate is 65% at best and the four year completion rate is less than 8%. We do not have accurate data on completion rates and times for mining or metallurgical engineering Ph.D. students in the U.S. but students self-report their expected completion times are around 4 years. National data on doctoral students indicate that younger, full time students with sufficient financial support have a higher probability of success in completing their degrees in a timely manner. Students in Eastern U.S. mining engineering programs are on average 32 years old, 85% are full time, and the gap between support and cost of attendance is $5,000 per year. Students in Western U.S. programs are on average 38 years old, 78% are full time, and the financial gap is $8,483. Further, it appears that domestic graduate students are not very geographically mobile. The current demand for new faculty is higher in Western programs but the longer term retirements are higher in Eastern programs. While all Ph.D. students who responded to the 2010 survey indicated an interest in a faculty position at some point in their careers, very few were applying for the open positions. Over the past 35 years for which data could be obtained, the maximum number of Ph.D. students enrolled in mining engineering in the U.S. is 115 with a maximum of 27 degrees granted in one year. On average 89 students are enrolled in doctoral programs in mining engineering in the U.S. with average degree production of 15 Ph.D.s per year.

Jan 1, 2012

By P. I. James, M. R. Baker

This paper examines new copper production from waste-grade ore using Blue Planet Strategies’ DEMET™ (Dynamic Electrode Mining Effluent Treatment) technology, which aims to bridge the practical weak pregnant leach solution (PLS) processing gap in heap leach-solvent extraction/electrowinning (SX/EW) operations by electrochemically removing metals from weak streams and generating product streams strengthened in target metals in a functional analog to SX. Weak PLS treatment was studied and used to estimate the economic and production impacts for a representative hypothetical brownfield copper heap leaching and DEMET-augmented SX/ EW operation to utilize "waste ore." Copper removal at 86% plating efficiency to below 0.1 g/L copper with power consumption of 0.55 kWh/lb copper and product strengths of over 10 g/L copper are presented. Modeling yielded approximately $2/lb total copper production and = $1/lb treatment costs with an immediate operation gross margin increase of 12% for the augmented brownfield. Potential mine life extension with an average 38% gross margin on "waste ore" augmented production at copper values of $3.25/lb was also indicated.

Jan 1, 2014

By R. J. Sheets

"INTRODUCTION What is the primary responsibility for geo-engineers in mining? Early in my career, I was asked this question by an experienced engineer. After explaining some various tasks, I was informed that the main responsibility was to monitor. Although this proved true for many cases, I came to understand that monitoring is merely one means to achieve the primary responsibility of a geo-engineer – ensure the safety of all personnel. The potential for groundwater to impact stability allows this responsibility to be shared with hydrogeologists as well. Inherently, we monitor to identify risks; we analyze designs and develop models to mitigate risks; and we are provided opportunities to explain to supervisors the geo-risks to personnel and mine plans that will result from particular courses of action. A couple years ago, I took an opportunity to share experiences from the first several years of my career as a geotechnical engineer with junior geo-engineering staff within my company. The intent was to share situations that were a part of my development, so they could learn and understand their responsibility to the safety of all personnel. Similarly, the purpose of this paper will be to share these thoughts with new geo-engineers in the industry in order to help prepare them for the challenges they will face. WHAT IS OUR PURPOSE? A senior hydrogeologist shared an experience that occurred in their career. The experienced hydrogeologist had spent some time with the hydrogeology staff, most of them relatively new to the mining industry, to understand how they viewed, and carried out, what they perceived as their responsibility at an active mining operation. After a couple months of observing and mentoring, the senior hydrogeologist met with everyone and posed the question above. I am sure that most junior geo-engineers would start explaining different tasks or projects that they must complete. The list would include monitor groundwater levels, track slope movement, evaluate designs, build models, oversee drilling programs, etc. These individuals did as well their supervisor. After going around the room allowing each individual to answer, the experienced hydrogeologist provided them the real answer to the question: SAFETY. If you examine the root intent of nearly every responsibility geotechnical engineers and hydrogeologists have, it is to apply our professional skills to ensure the safety of our workforce. When the objective of safety from the geotechnical and hydrogeology disciplines is achieved, inherently accomplishing the mine plan should follow suit, or it should significantly minimize the impact by gravity or groundwater. The objective of a hydrogeologist is to ensure the soil or rock around an open pit or underground mine is dewatered; this specific objective will reduce pore pressures which could lead to a possible ground failure. They ensure a working bench/UG level is dry, so operations are not dealing with adverse conditions while mining. A geotechnical engineer is expected to use all available information and engineering judgement to provide design recommendations, so that the mine will not catastrophically fail. Additionally, they will implement the monitoring systems to identify hazards in advance in the event that the design is flawed and people need to be evacuated. We all know that safety is important to everyone; however, our particular set of technical skills play a significant role in ensuring the safety of a large portion of the global mining workforce."

Jan 1, 2018

By S. Al-Thyabat

Conditioning of flotation feed is an important prerequisite for efficient flotation. This process depends on parameters such as conditioning time, slurry concentration and slurry pH. The objective of this paper is to optimize the conditioning process for the flotation feed. This was achieved by first conducting preliminary experiments at constant slurry pH to determine the significant ranges of slurry concentration and conditioning time. Then, factorial experimental design (32 experiments) was used to investigate the significance of the above parameters and their interactions on flotation recovery and grade. Statistical analyses using ANOVA approach showed that flotation recovery (R) and concentrate grade (g) were affected by feed conditioning. It was found that concentrate grade was linearly proportional to the slurry concentration and conditioning time, while flotation recovery was exponentially proportional to the conditioning parameters. A contour plot of the flotation response (recovery and grade) against conditioning parameters showed that the optimum operating parameters for flotation feed conditioning was a conditioning time of 1.5 to 2 minutes and a slurry conentration of 55% to 60%.

Jan 1, 2008

By Ronald E. Triplett, George K. Burke, Bernard H. Voor

To accommodate larger double-stack rail cars through a tunnel, the existing active tracks were to be lowered. The tunnel is an elliptical, horseshoe-shape with concrete base support, removal of which would jeopardize the tunnel integrity, particularly since the soils beneath part of the tunnel floor included soft, saturated silt overlying rock. Grouted bar anchors combined with a SuperJet soilcrete strut replaced the existing support in situ prior to floor lowering. Close cooperation and detailed timely engineering support provided a quality project. Project design and construction, ground improvement methods, and the performance of the work are described.

Jan 1, 2001

By Xuehui Li, Xinshuai Shi, Jun Wang, Jianguo Ning, Shang Yang, Zhaohui Zhang

Pressure relief is an important step in preventing and controlling rock burst disasters. Hydraulic slotting has emerged as a strategy for pressure relief and energy release in roadway surrounding rock. There is no qualitative description for the response relationship between hydraulic slotting parameters and the degree of pressure and energy release in a roadway, which is a scientific problem. In this paper, theoretical analysis, numerical simulation, and on-site industrial practice were comprehensively used. First, a structural model for the pressure relief and energy release of coal by hydraulic slotting in a deep roadway was established, and the pressure relief and energy release principle by hydraulic slotting was revealed. The principles and methods for determining hydraulic slotting parameters in the roadway were proposed. Three evaluation indices for the pressure relief and energy release effect of coal after hydraulic slotting were proposed, namely, the stress ratio of coal, elastic energy density ratio of coal, and displacement rate of the roadway rib. According to the evaluation indices, the degree of pressure relief and energy release was divided into three levels: reverse pressurization, full pressure relief, and transitional pressure relief. Then, a FLAC3D numerical model was constructed, the levels of pressure relief and energy release under different hydraulic slotting parameters were analyzed, and the optimal slotting parameters were determined. Finally, on-site industrial practices were carried out in Panel 7312 of the Xinhe Coal Mine. An effect test showed that the effect of pressure relief and energy release was remarkable after hydraulic slotting was adopted on the roadway surrounding rock, which can effectively reduce the risk of rock bursts. These research results can further improve hydraulic slotting technology systems and provide a theoretical basis and technical reference for the stability control of deep roadway surrounding rock.

Sep 20, 2023

By W. A. Haley

The use of diesel engines in underground hard-rock mines dates back to the late 1940s. For the first several years, they were used only occasionally, being limited to a few metal mines that experimented with crawler¬mounted front-end loaders, tractor-trailer hauling units, a few tractors for drill-compressor mounts, and utility cleanup machines. By the mid-1950s, track loaders had become commonplace in limestone mines and uranium mines on the Colorado Plateau in the United States, as well as in Canada. Use of crawler-mounted tractors as drill and compressor mounts also increased. By the end of the 1950s, rubber-tired loaders and some haulers began to replace the track-type machines and rail-mounted cars that had been in use. About 1960, the rubber-tired machines brought about a new era of underground mining mobility and flexibility, centered on a method commonly known as "trackless mining." Ultimately, many of the underground rail-type systems for loading and hauling were replaced by the trackless mining technique. ECONOMIC CONSIDERATIONS The size and nature of mineral deposits, plus ground control techniques, historically had dictated small open¬ings to the surface from many underground mines. The small mine openings led to the development of special rubber-tired loaders and haulers designed specifically for access through the small openings. However, some mines, particularly those in massive mineral deposits, are able to excavate and maintain very large openings, and some use modified room-and-pillar systems. With the large mine openings, the use of larger, more produc¬tive equipment such as that commonly found in surface mining becomes economical. In fact, productivity gen¬erally increases at a more pronounced rate than machine size increases because many of the larger machines were designed for heavy-duty shot-rock applications in surface mines and construction sites where the handling of blasted rock is common. Table 1 can be used as a very Table 1. General Productivity Comparison for Conventional Machines In Underground Use (Shot-Rock Conditions) 2.3 m3 (3 cu yd) 4.6 m3 (6 cu yd) Loader Loader Expected Surface 230 t/h 540 t/h Production (250 stph) (600 stph) Expected Underground 90 t/h 270 t/h Production (100 stph) (300 stph) Expected Total Efficiency: Surface 40%-60% 50%75% Underground 25%.-40% 30%50% Expected Useful Machine 8000 hr 12,000 hr Life Before Replacement general comparison of the production and efficiency between small and large machines. Combining greater productivity often inherent in larger machines, with reduced downtime resulting from using fully developed machines with fast parts and service backup, some mine operators have been able to reduce material handling costs appreciably while reduc¬ing manpower requirements for operators and main¬tenance men. Large mine openings increase the amount of rock that must be handled in the development work, and they sometimes increase the dilution in stopes or rooms, de¬pending upon the dimensions of the ore zone. Providing adequate space for the unrestricted operation of large surface mining machines could, therefore, lead to more waste segregation and handling costs. It could also cause greater ore dilution that would result in a lower grade of ore being delivered to the processing plant. The tradeoffs between opposing cost factors must be reconciled and balanced to achieve the best overall cost of the crude ore, concentrates, or product. EQUIPMENT MODIFICATIONS Loaders and haulers designed for surface mining are seldom used underground in their standard con¬figurations without some modifications. If done, the modifications generally are made by the equipment dealer and/or the user, and the modifications usually include one or more of the following items: 1) The exhaust stack is lowered, and its direction is changed. Usually, it is repositioned horizontally to the rear, or it is fed into the engine fan to diffuse the exhaust gases. 2) The operator's position is lowered by either lowering the seat or changing the seat to a side mount. 3) The operator controls are adjusted to fit the new operator position. 4) Other components, such as the radiator and loader tower, are lowered. 5) Special bumper guards are mounted at the base of the radiator area. 6) An exhaust conditioner is mounted and con¬nected, using either a catalytic or a water-type condi¬tioner, or both. This usually is controlled by the safety and health regulatory authority having jurisdiction. 7) The positions of other components are rearranged

Jan 1, 1982

By Henning E. Soderberg, Robert Brien, O&apos

Coal may be the answer to US energy problems, but it also has its drawbacks. If it is crushed, screened, dumped, or moved, the environmental engineer is faced with a dust problem. To control coal dust, at both minimum initial cost and reasonable operating cost, demands a working knowledge of dust control equipment design, as well as the coal handling operation. Here are some guidelines to reduce air volume requirements and cut overall system costs.

Jan 10, 1981

By R. L. Grayson, J. R. Warneke

The mining industry has searched for a message that will resonate with the public while focusing on the industry’s positive contributions to society. Thus far environmental groups remain unconvinced, and the public is still largely uninformed. In the authors’ opinion, a number of recent events and experiences have converged to help shape a more effective public communication on mining’s important role. This paper presents a fresh and cogent perspective on coal’s role in sustaining society by anchoring the message in a transparent material flow analysis that links coal’s impacts across a broad range of social and environmental aspects.

Jan 1, 2006

By Isabelle Lamb, Manfred K. Lechner, Daniel N. Adams

Sydney Australia’s M5 East Motorway connects southwestern suburbs to the airport. Central to the project are twin 3.8 km roadway tunnels which are mostly driven through rock. In order to come up with the winning tender, the project team developed several unique approaches for tunnel excavation and support. The project’s flat roof with thin shotcrete lining, 20 m wide underground intersections, thixotropic-grout sheathed rockbolts and cablebolts, and new shotcrete testing methods all differ significantly from traditional concepts. This paper presents the design basis and construction methods used to achieve the 100-year design life required by the project’s owner.

Jan 1, 2001

By H. EI-shall

Bench top centrifuge is used to study the consolidation behavior of phosphatic clays. A mathematical model (based on first order kinetics) is developed to describe the relationship between the void ratio (at any time) and the effective stress. Permeability of the sediment as a function of the void ratio is also described. The obtained data are compared with that obtained by other consolidation testing methods. Centrifuge results indicate that consolidation data may be developed in a short period of time (hours) as compared to days or weeks when other techniques are used.

Jan 1, 1995

By H. B. Charmbury

In Pennsylvania there are two distinct coal regions, bituminous and anthracite, in which the mining conditions and production trends are different. Bituminous mining is more mechanized, and production is increasing. This report is concerned specifically with the bituminous area. The bituminous coal industry apparently is facing a shortage of miners and mine officials which may become acute by 1970. Since 1949, when Pennsylvania had more than 100,000 bituminous coal miners, the number has steadily decreased. There are two principal reasons for the decrease of certified miners and mine officials. These reasons are the fact that apprentices are almost a thing of the past and the fact that the image of the coal miner is a poor one to people in general and to parents in particular. This has contributed greatly to the high average age of our miners today. The average age of all those employed in the mines is 50. Yet it is known that the progress in mine mechanization requires younger and more alert men to operate the newer type equipment.

Jan 1, 1966

By S. Komar Kawatra, J. J. Carlson

"It is widely known that steel made from iron ore pellets high in phosphorus is brittle. Removal of phosphorus from iron ores such as hematite has proven to be difficult. The majority of research on phosphorus removal from hematite ores has been focused on flotation. The purpose of this study was to assess phosphorus distribution under current operations at an existing hematite concentrator and determine how efficient the plant is at removing phosphorus. Objectives included: (1) Perform a phosphorus balance at an operating plant to assess phosphorus distribution under current operations; (2) Determine phosphorus removal efficiency of each separating unit and (3) Perform a quantitative liberation analysis to determine max phosphorus removal by physical means. Liberation was quantitatively determined using wet high intensity magnetic separation (WHIMS) to separate iron ore from liberated apatite (and other nonmagnetic compounds).It was observed that 61.2% of the total phosphorus entering Plant F was rejected during current operations. The desliming process was the most efficient step at removing phosphorus (45.1% P removed), whereas the flotation step was much less effective (16.1% P removed). Liberation analyses indicated that the deslime thickener may have been more efficient at removing phosphorus than the flotation circuit due to the greater abundance of liberated phosphorus in the deslime thickener feed. The scavenger returns stream was identified as having a high concentration of phosphorus and a possible point of phosphorus removal. However, liberation analyses indicated that only 32.4% of the total phosphorus in the stream was liberated. Grinding the scavenger returns to 80% passing 5.2 µm increased the amount of liberated phosphorus in the stream to 70.4%."

Jan 1, 2014