There are well-developed faults in an open-pit molybdenum mine, which may lead to slope instability during open-pit mining. Three typical faults running through the ultimate boundary were selected to investigate progressive instability mechanism of the slopes in the open-pit mine under coupling effect of multiple faults by integrating Rhino+Griddle 3D geological modeling and FLAC^3D multi-field coupling simulation. Furthermore, the safety factors of slopes under different working conditions were calculated, and relatively unstable slopes were also identified for addressing the subsequent safety of the ultimate boundary. The simulation results show that, plastic failure and tensile stress concentration will occur in the slope at the junction of Zone C and Zone D, while local plastic failure will occur in the faults, which, however, won't induce instability failure in the surrounding slopes; there won't be much displacement in the slope at the ultimate boundary, with the maximum displacement (2.44 cm) at the toe of slope; the safety factors of the ultimate boundary of open pit under natural working conditions and rainfall are 2.48 and 2.36 respectively, indicating that the slope is stable. However, slope at Zone E with F1 fault is considered to be with a relatively instability, for which comprehensive measures including real-time monitoring, prevention and reinforcement can be taken to ensure the safety.

Based on the causes of historical dam failure accidents in tailings ponds, a safety grade evaluation index system was built for tailings ponds. The correlation between safety evaluation indexes was demonstrated with Pearson correlation analysis method. An in-depth analysis was conducted for the distribution of sample data to confirm that the sample data met the requirements of the Kriging method for data distribution. Then, the weight of each safety evaluation index was calculated with entropy weight method. The two safety evaluation indexes with higher sensitivity were selected as the X-axis and Y-axis of the Kriging geostatistical model. A geostatistical model was established by using the Kriging method and was verified with empirical case data. The results show that the predicted safety grades of 12 groups of tailings pond data with known safety grades in the geostatistical model are completely consistent with the actual safety grades, which verifies the feasibility and high accuracy of this method.

To analyze the damage effect of drilling and blasting in construction on the surrounding rock of roadway, a Hopkinson pressure bar (SHPB) test and numerical simulation were carried out to investigate the dynamic response characteristics of saturated mudstone under one-dimensional impact load. Under impact with different strain rate, rock samples undergo four deformation stages, including compaction, quasi-elasticity, yielding and unloading. The saturated mudstone has its dynamic elastic modulus in a linear relationship with the strain rate. As the strain rate increases, dynamic fragmentation of the rock samples is significantly enhanced, and the energy dissipation of saturated mudstone also correspondingly increases. The mean error between the peak stress obtained from the numerical simulation and the experiment is 7.9%, which accurately reflects the dynamic characteristics, such as morphology of mudstone fragmentation and dynamic stress-strain relationship, under impact with different strain rates. The strain rate effect of mudstone indicates that blasting disturbance will continuously change its energy state, so dynamic monitoring should be strengthened for the deformation of surrounding rock during tunneling.

Based on the mining environment and stress characteristics of backfill in the upward horizontal slicing and filling method, the influence of wetting-drying (WD) cycles on the mechanical properties of cemented tailings backfill (CTB) was explored. An ultrasonic test and a mechanical property test were conducted for CTB samples with different cement-tailings (c/t) ratios after different WD cycles to investigate variation in the velocity of longitudinal waves, compressive strength, and failure mode. The results show that after WD cycles, the CTB samples with c/t ratios of 1∶4, 1∶6 and 1∶8 all had a decreased ultrasonic wave velocity, and as the number of WD cycle increases, CTB samples with different c/t ratios had enhanced plastic deformation capacity but decreased strength. It is shown that WD cycle has a small impact on backfill with a higher c/t ratio. And the CTB without WD cycle mainly experiences tension-shear failure, while CTB after WD cycles has multiple axially parallel cracks penetrating vertically, which significantly increase after more WD cycles. It is found that the c/t ratio is a key factor for CTB to resist deterioration caused by WD cycles. In mining operation, the c/t ratio for different filling areas can be reasonably optimized to reduce filling costs.

In order to explore the stability and durable service of a blend of construction waste and red clay under load, a California Bearing Ratio (CBR) test, rebound modulus test and failure strength test were performed to determine the appropriate content of red clay in the blend. The influence of compaction degree, confining pressure, stress ratio and loading times on the macroscopic permanent deformation of the blend was comprehensively analyzed by triaxial tests, and the evolution of internal structure during the deformation process of the blend was simulated by discrete element method. The results show that with red clay at an appropriate content of 65%, the blend of construction waste and red clay can have its permanent deformation increase gradually as the stress level becomes higher, and decrease gradually with the increase of confining pressure and compaction degree. During the deformation process, the internal shear stress, coordination number and slip rate of the blend all increase gradually as the stress level becomes higher. Based on the test results and in consideration of stress state, including confining pressure, failure strength and loading stress, physical state, such as compaction degree, and loading times, a permanent deformation prediction model was proposed and validated. With a correlation coefficient of 0.89, the model is considered to have a “relatively good fit”.

As for fracture development and potential instability in the open-pit bench slopes in Dabaoshan in Guangdong Province, the blockiness was proposed to be taken as a quantitative measure of rock mass integrity. Based on the existing data, 72 potentially unstable zones on the western slope were identified. The structural characteristics of the rock mass were analyzed based on investigation of fracture occurrence characteristics and 3D fracture network modeling, and rock mass integrity was also evaluated by blockiness analysis. It is found that there are fully-developed fractures with good ductility in all the zones, creating the structural conditions for block formation; in 84.9% of the zones, the inclination angle of superior fractures and the slope inclination angle are supplementary, which can significantly reduce sliding risks; in 97.3% of the zones, the blockiness of rock mass is less than 27, indicating mild blockiness, good integrity of rock mass, and overall high stability. There are only rock mass with blockiness over 27 in Zone 6^# and Zone 46^#, presenting moderate blockiness. In these two zones, the rock mass structure is significantly degraded, leading to reduction in damage tolerance. It is concluded that there are potential risk of slope instability in those two zones, which should be taken as the focus in the subsequent disaster prevention and control for the slopes in Dabaoshan.

Homogenization of backfill slurry cannot be automatically evaluated during stirring process. To solve this problem, the Yolo-v7 algorithm was improved by using the CBAM and the SPD-Conv, and the improved Yolo-v7 algorithm was then employed to automatically evaluate the homogenization of backfill slurry. The research results show that the improved Yolo-v7 algorithm demonstrates a significantly improved performance, with the accuracy, recall rate and mean average precision (mAP) all increased by 17.5, 28.8, and 32.4 percentage points respectively. Analysis of slurry parameter sensitivity indicates that slurry concentration is the principal factor affecting evaluation of slurry inhomogeneity, followed by the cement-tailings ratio. Moreover, the homogeneity of backfill slurry with high concentration can be significantly improved by prolonging stirring time.

To reveal the relationship among uniaxial compressive strength, tensile strength and shear strength, 159 sets of rock mechanics data collected from 38 mines were taken as samples. Firstly, the relationship between uniaxial compressive strength and tensile strength was characterized by data analysis, and the range of uniaxial compressive and tensile strength in the formulae for calculating shear strength was calibrated. Then, a BP neural network prediction model was constructed, with which the shear strength of rocks can be predicted with uniaxial compressive strength and tensile strength as the input layer. The results show that: the uniaxial compressive strength of rock is 7.6-27.8 times tensile strength; the cohesion value obtained from multiple comparison tests is 1.0-2.2 times that from the original empirical formula, and the tangent value of internal friction angle is 0.35-1.00 times that from the original empirical formula, presenting high confident and reliable results; based on comparison between the predicted value with BP neural network and the actual value, the cohesion deviation ranges from -10.53 MPa to 20.12 MPa, and the internal friction angle deviation ranges from -18.43° to 28.97°, exhibiting an overall normal distribution.

To enhance the positioning accuracy of seismic source in underground mining and solve the problem of spatiotemporal changes in the wave velocity field caused by excavation disturbance and complex structures, seismic source relocation with real-time inversion of wave velocity field as a prior was proposed, which combined ray tracing and quasi-Newton iteration algorithm to realize dynamic update of wave velocity field and high-precision positioning of seismic source. The validity of this method was verified by synthetic monitoring and field testing. The synthetic monitoring results showed that compared to the least squares procedure, the real-time inversion resulted in the average positioning accuracy enhanced by 49.8% and the accuracy of wave velocity inversion exceeding 95%. In the field testing, a 180 m×180 m to-be-filled area in mining operation was taken for imaging target. Compared to the least squares procedure, this method brought the average error in positioning reduced by 7.074 m, and the accuracy of wave velocity inversion in each region exceeding 95%. The research indicates that this method not only is suitable for seismic source positioning in microseismic monitoring, but also can be used as a geophysical method for detecting passive imaging of goaf areas.

The polymetallic ore from Shizhuyuan of Hunan Province was taken for research. Based on the characteristics of uneven coarse-fine dissemination of minerals and insufficient deliberation of Mo-Bi sulfide ore for flotation, an enhanced recovery technology by stepwise separation was developed for the low-grade Mo-Bi sulfide ore. The results show that an experiment by adopting such stepwise separation technique produced a bulk flotation concentrate grading 1.31% Mo and 2.83% Bi at corresponding recoveries of 88.43% and 85.99%, presenting 5.34 and 4.80 percentage points up, respectively, compared to on-site conventional bulk flotation process. This creates favorable conditions for the following tungsten flotation from the bulk flotation tailings. It is shown that a closed-circuit tungsten flotation can produce a tungsten concentrate grading 44.64% WO₃ at 84.12% recovery.

The tailings from gravity separation of cassiterite in a polymetallic ore from Guangxi Zhuang Autonomous Region were taken for experimental research. Due to its low grade, fine particle size of target minerals therein and similar floatability of tourmaline and cassiterite, a multiple ligand metal-group collector, CSC-BHA-Pb, and a high-efficiency depressant, XY1, were adopted to selectively inhibit gangue minerals dominated by tourmaline while efficiently enrich fine-grained cassiterite. On the basis of iron removal and desulfurization, an open-circuit flotation test of cassiterite, consisting of one roughing, one scavenging and seven cleaning, finally produced a tin concentrate with a Sn grade of 14.05% and a recovery of 28.79%, achieving efficient tin enrichment.

With a copper sulfide ore from Jiangxi Province taken in the research, an experimental study was carried out for enhancing the recovery of associated Au by flash flotation. In a laboratory test, the underflow of secondary-stage classification was subjected to an open-circuit flash flotation consisting of two-stage roughing and two-stage scavenging, leading to the Au grade up from 16.60 g/t to 140.60 g/t, presenting good floatability. Additionally, the Cu grade in the concentrate from flash flotation cell is higher than that in the main process, which proves that flash flotation of classification underflow will not affect the Cu recovery from the main process. Industrial practice demonstrates that with the ores of different Au grade, the Au recovery rate has increased after technical transformation. A size analysis by sieving reveals that those fully liberated minerals within conventional size fraction in classification underflow can be mainly recovered by flash flotation, thus being prevented from overgrinding due to returning to the grinding mill.

The flotation reagent system was optimized for a lead-zinc ore in Qinghai based on experimental studies. It is shown that a new chelating agent (PT-8) combined with the conventional 25^# xanthate can effectively enhance the flotation recovery of copper-lead sulfide minerals. A new collector (SXF-10) at a low dosage can efficiently recover marmatite from the ore, significantly improving the recoveries of Cu, Pb and Zn from flotation process. With raw ore at a grinding fineness of -0.074 mm 70%, 25^# xanthate and PT-8 are used for Pb flotation, and SXF-10 is used for Zn flotation. A closed-circuit test with such process flow can produce a Cu-Pb concentrate grading 3.52% Cu and 62.53% Pb at corresponding recoveries of 70.08% and 93.75%, respectively, and a Zn concentrate grading 48.19% Zn at 87.97% recovery.

A polymetallic sulfide ore in Qinghai assays 1.45% lead and 2.64% zinc, and is associated with copper, sulfur, iron, plus rare and precious metals (gold and silver). Process mineralogy of the ore was studied by multi-elemental analysis, phase analysis, and MLA (Mineral Liberation Analyzer) analysis. A closed-circuit test was performed after condition tests on this basis. In the general mineral processing of lead-zinc sulfide ores, lime is added through the whole flow, which is not conducive to comprehensive utilization and recovery of copper, gold and silver. Aiming at this problem, a processing concept was proposed for comprehensive utilization of the valuable elements therein, which consists of preferential flotation of lead under natural pH value, zinc flotation by inhibiting sulfur from the obtained lead flotation tailings, and finally iron recovery after sulfur removal of the tailings from the previous step. Results show that with the polymetallic sulfide ore ground to a fineness of -0.074 mm 70%, the valuable elements therein can be recovered for comprehensive utilization by adopting a combined flotation and magnetic separation process, consisting of copper-lead flotation, zinc recovery from the obtained copper-lead flotation tailings, sulfur recovery from the obtained zinc flotation tailings, and finally iron separation from the obtained tailings. It is shown that the lead concentrate grading 60.12% at recovery of 94.72%, zinc concentrate grading 46.99% Zn at recovery of 88.26% and iron concentrate grading 67.22% Fe at recovery of 9.33% are produced.

Based on analysis of chemical composition, mineral composition, main mineral properties and disseminated grain size of slag from smelting of nickel laterite ore, a series of exploratory experiments were carried out. The iron in the slag was recovered by magnetizing roasting and magnetic separation. The magnetizing roasting is performed for 40 min at 750 ℃ with a coal powder at a ratio of 2%. The roasted ore is subjected to a two-stage low-intensity magnetic separation (LIMS) after being ground to a fineness of -0.045 mm 80%, and an iron concentrate grading 60.38% Fe can be obtained at recovery of 71.55%.

An experimental study on flotation separation was conducted for a gold-antimony concentrate with grade of Sb and Au at 20.66% and 98.95 g/t, respectively, from Russia. In an experiment on antimony floatation while suppressing sulfur by adding efficient sulfur depressant of BK526, activator of lead nitrate, and collector of ammonium dibutyl dithiophosphate, the gold-antimony concentrate was ground to a fineness of -0.045 mm 86%, sodium sulfide and activated carbon was added to remove reagent and sulfuric acid was used to adjust the pulp pH to be acidic. Finally, it produced an antimony concentrate grading 51.56% Sb at 86.06% recovery, and a gold concentrate grading 108.35 g/t Au and 4.41% Sb, with gold recovery of 71.20%. It is shown that stibnite and gold-bearing pyrite can be efficiently separated.

An experimental study was carried out to address problems of higher grade of iron in the tailings from 1^st-stage high intensity magnetic separation (HIMS) and higher metal losses in the production of Lilou Iron Mine in Anhui Province. In the study, the low-intensity magnetic separator (LIMS) used before 1^st-stage HIMS was replaced with a CCT-A high-gradient permanent magnetic drum separator, and the HIMS process consisting of one roughing and one scavenging was optimized to a HIMS process consisting of one roughing and two scavenging. With these measures, both strongly and weakly magnetic minerals can be effectively recovered. The 1^st-stage scavenging tailings have the iron grade decreased by around 1 percentage point, and nearly 40 000 tons of iron concentrate can be recovered annually, which can bring remarkable economic benefits to the mine.

In Meishan Iron Mine, the iron concentrate produced on the site has SiO₂ grade more than 6.0%. Aiming at this problem, an experimental study was carried out based on the microscopic identification of different concentrate products by using a new process flow, in which the magnetic field intensity of low-intensity magnetic separation (LIMS) was reduced for roughing, and a process of regrinding and re-separation was added for the scavenging concentrate by high-intensity magnetic separation (HIMS). The effects of those technical transformation on the TFe grade of iron concentrate and the content of impurity SiO₂ therein were explored. The results show that after such technical transformation, the LIMS can produce a concentrate grading 63.57% TFe and 3.37% SiO₂ with corresponding recoveries of 72.79% and 12.94%, with a yield of 52.21%; while the HIMS can produce a concentrate grading 44.79% TFe and 9.07% SiO₂ with corresponding recoveries of 17.20% and 11.67%, with a yield of 17.51%. This new flowchart can produce the concentrate grading 58.85% TFe and 4.80% SiO₂ at corresponding recoveries of 89.98% and 24.61%, with a total yield of 69.72%. It is shown the SiO₂ content in the total concentrate of Meishan Mine can be greatly reduced.

Based on flotation tests of a low-grade copper sulfide ore, a processing technique was developed, composed of flash flotation of Cu, Cu-S bulk flotation, and Cu-S separation after regrinding of roughing concentrate from bulk flotation. Effects of factors on flotation were also explored, including grinding fineness, flotation reagent, regrinding fineness of roughing concentrate from bulk flotation. Ammonium dibutyl dithiophosphate was used as the collector for Cu in the flash flotation, a combination of ammonium dibutyl dithiophosphate and butyl xanthate as the collector for Cu-S bulk flotation, lime as a sulfur depressant, and terpineol oil as a frother. With grinding fineness of -0.074 mm 60% and regrinding fineness of -0.045 mm 70%, a closed-circuit test for the raw ore assaying 0.40% Cu and 2.45% S produced a bulk copper concentrate grading 19.49% Cu at 87.68% recovery, and a sulfur concentrate grading 48.65% S at 35.11% recovery. It is concluded that efficient separation and enrichment of copper and comprehensive recovery of sulfur from this low-grade copper ore can be actualized by this technique.

A roasting and acid leaching process was adopted to extract lithium from a low-grade clay-type lithium ore with a Li₂O grade of 0.13%. After exploration of the effects of various factors on lithium leaching rate, including roasting temperature, roasting time, and mass fraction of sulfuric acid, liquid-solid ratio, leaching temperature and leaching time in acid leaching process, the optimal experimental conditions were determined as follows: roasting at 500 ℃ for 1 h, leaching at 95 ℃ for 60 min with sulfuric acid at a mass fraction of 15%, and liquid-solid ratio of 6 mL/g. Under these conditions, the lithium leaching rate can reach 85.26%, indicating that efficient leaching of lithium from such low-grade clay-type lithium ore can be actualized. XRD and SEM analyses of the samples before and after leaching show that lithium leaching is attributed to the exchange between hydrogen ions in the sulfuric acid solution and lithium ions in the mineral.

Vanadium slag and sodium carbonate, as raw materials, were mixed and grounded for pretreatment. The obtained mixture was then taken in a roasting experiment for investigating effects of factors, such as holding time, temperature and feeding materials for roasting process with different alkali ratios (a mass ratio of sodium carbonate to the converted content of V₂O₅ in vanadium slag) on transformation and leaching rate of vanadium from clinker, and also exploring phase transformation of vanadium slag during the roasting with a low alkali ratio. The results show that if roasting process at higher temperatures is held for 90 minutes, the alkali ratio in the roasting process can be reduced by properly increasing the temperature. After roasting at a temperature of 880 ℃ with an alkali ratio of 1.0, water leaching can result in the total vanadium (TV) content in the residue falling down to 0.49% and the leaching rate of vanadium up to 94.70%. This method is suitable for sodium roasting in a rotary kiln. Under the existing process conditions for sodium roasting of vanadium slag in a multiple hearth furnace, an addition of some clinker into the material for roasting is beneficial to reducing the alkali ratio. It is shown that after roasting is held for 45 minutes at 780 ℃ with an alkali ratio of 1.2, the TV content in the final leaching residue can fall to 0.96% and the leaching rate can reach 89.19%. The microscopic analysis of the clinker after roasting with a low alkali ratio show that there is a high aggregation of sodium vanadate phase, which is in good consistency with NaVO₃ in the elemental composition.

After an introduction of the current status of gallium resources and analysis of the occurrence state of gallium in sodium aluminate solution, zinc smelting slag, tailings from vanadium extraction and coal-based solid waste, progress in the research of gallium extraction processes is elaborated and the adaptability, advantages and disadvantages of each technique are also summarized. Finally, based on the discussion of problems faced by the gallium industry and the development direction in the future, it is suggested that iteration and upgrading of gallium extraction processes should be vigorously promoted based on China's scale advantage in alumina industry.

An extraction system composed of P507 as the extractant and sulfonated kerosene as the diluent was used to separate zinc from a sulfuric acid leachate containing cobalt, manganese and zinc. The effects of key parameters on zinc separation efficiency were investigated, including the volume fraction of P507 in organic phase, saponification degree, pH value of feed, O/W ratio, reaction time, and reaction temperature. The results show that under the following optimal process conditions, including P507 at 20% by volume, saponification degree of 80%, feed pH of 5, O/W ratio of 1∶1, reaction time of 10 min and reaction temperature of 25 ℃, the extraction rate of Zn reaches 99.92%, while the extraction rates of Mn and Co are just 17.41% and 3.69%, respectively, presenting an excellent separation effect.

Heavy metal pollution in Dalong River was assessed by field sampling and detection analysis. Hg(II) and Pb(II) in the water and sediment were adsorbed with sludge-derived biochar for ecosystem restoration in Dalong River, and the restoration effect was then compared with that by using activated carbon. The results indicate that heavy metal pollution in Dalong River is relatively slight, and the contents of heavy metals in the surface water at all sampling points are within a safe range, only the sediment at some sampling points has the heavy metal pollution at a light pollution level. It is shown that sludge-derived biochar has a good adsorption for Hg(II) and Pb(II), with the adsorption process conforming to the Langmuir isotherm (with R² of 0.932 2 and 0.998 9). Compared with activated carbon, the sludge-derived biochar can bring better restoration effect, and its desorption of Hg(II) and Pb(II) also shows that it can have a significantly better fixation effect of heavy metals.

To extract rare earth elements (REE) of yttrium (Y) and cerium (Ce) from yttrium aluminium garnet (YAG) fluorescent powder, experiments were performed for YAG by adopting direct acid leaching and roasting-acid leaching process, respectively. The effects of various factors on the leaching rates of Y and Ce were investigated, including roasting agent types and dosage, roasting temperature, roasting time, as well as leaching temperature and leaching time. It is shown that YAG fluorescent powder is stable in structure, while Y and Ce elements therein, not in the form of simple oxides, cannot be effectively extracted by direct acid leaching; a roasting pretreatment can effectively destroy the structure of YAG fluorescent powder, and convert the rare earth elements into oxides, which is beneficial to the following acid leaching process. After 2 h roasting at 900 ℃ with YAG and Na₂CO₃ in a mass ratio of 1∶0.5, and then 1 h leaching at 60 ℃ with HCl at a concentration of 3 mol/L, a liquid-solid ratio of 20 mL/g, and 1.2 mL/g of hydrogen peroxide, the leaching rates of Y and Ce can reach 97.23% and 84.91%, respectively.

Extraction of rhenium from low-grade molybdenum concentrate by adopting a process of oxidative volatilization followed by ion exchange was explored in experiments. Firstly, a two-stage high-temperature oxidative roasting was adopted to oxidize rhenium into Re₂O₇. Rhenium and molybdenum were effectively separated due to rhenium volatilizing into the flue gas. The volatilization rate of rhenium reached 85.42% after roasting process at 675 ℃. Then, the rhenium-containing flue gas was eluted to get rhenium-containing solution, in which rhenium was selectively extracted with D201×7 ion exchange resin. With the solution pH of 9, liquid/solid ratio of 50 mL/g, a 20 min adsorption at temperature of 35 ℃ brought the adsorption rate of Re up to 97.52% and the adsorption rate of Mo less than 20%. Finally, a product of ammonium rhenate was prepared by stepwise desorption followed by concentration and crystallization. By this process, the total recovery of Re can reach 70.68%.

With nickel laterite ore from Morowali of Indonesia as the raw material, effects of various factors, including reaction time and temperature, stirring speed, acid-to-ore ratio, and slurry concentration, on the leaching rate were systematically investigated. Additionally, the leaching mechanism was explored by using the PHREEQC, a thermodynamic calculation software. The results show that after one hour reaction at 250 ℃, with acid-to-ore ratio of 300 kg/t and slurry concentration of 25%, the leaching rates of Ni and Co can reach 98.40% and 99.90%, respectively, while the leaching rates of Fe and Al are just 3.84% and 40.25%, respectively. The high selectivity of high pressure acid leaching (HPAL) is attributed by the factor that at high temperatures, not only the hydrolysis reaction of Fe and Al can be promoted, but also the pOH can be decreased from 14 to 10, which thereby inhibits leaching of Fe and Al. As the temperature rises to above 200 ℃, the number of H⁺ increases, which can ensure a sufficient amount of H⁺ for leaching of nickel and cobalt metals.

Based on an overview of progress in the research on iron recovery from high-iron red mud by alkaline electrolysis, the research findings at home and abroad are introduced in terms of principle and advantages of electrolysis, influencing factors for the process, and the obtained iron products. The advantages of red mud in the process for preparing metallic iron by electrolytic reduction of red mud in alkaline media at low temperatures are elaborated, and some problems to be solved in the future are proposed, which can provide references for innovation and development of this process.

In combination with experiments and multi-step finite element simulation, the residual stress distribution in components manufactured by laser powder bed fusion (LPBF) and the influence of the initial residual stress field on the laser shock peening (LSP)-induced compressive residual stress field. The validity of the simulation results was verified by the residual stress values measured with X-ray diffraction. The results show that the thermal stress of the lower and middle layer materials in the component fabricated by LPBF undergoes a transition from zero stress, compressive stress, tensile stress to compressive stress, while the top layer material shows a transition from zero stress, compressive stress to tensile stress, thus leading to tensile residual stress occurred on the surface layer of LPBF fabricated component, while compressive residual stress in the lower part. The action of initial residual stress field results in the reduced peak value but increased depth of LSP-induced compressive residual stress. The initial residual tensile stress field of LPBF components can exert a suppressive and dragging effect on the reverse plastic deformation caused by surface waves, leading to reduced intensity and changed position of the “residual stress hole”, which can improve the uniformity of LSP-induced residual stress distribution.

Manganese-containing master alloy was prepared by vacuum induction melting. Effects of factors, such as manganese source, Mn/Al mass ratio, preheating time, and refining time on the composition of the manganese-containing master alloy were studied, and the influence of manganese-containing master alloys replacing the traditional “Mn + Al” feed on the prepared manganese-containing AB₅-type rare-earth-based hydrogen storage alloy was also discussed. The experimental results show that compared to manganese powder, manganese flakes as manganese source can bring higher melting yield. Manganese-containing master alloys can be formed with different Mn/Al mass ratios. As the preheating time is prolonged, the melting yield firstly increases and then decreases, and the optimal condition for preheating is 10 minutes at a power of 4 kW. As the refining time increases, the melting yield decreases slightly, and the optimal condition for refining is 1 minute at a power of 15 kW. The average melting yields of rare-earth-based hydrogen storage alloy prepared with MnAlNi alloy and MnAl alloy are increased by 0.37% and 0.78% respectively than that with “Mn + Al” as raw material. The preparation of AB₅-type rare-earth-based hydrogen storage alloy with manganese-containing master alloy can optimize the yield, with the hydrogen storage performance not influenced. The application of this method in practical industrial production brings good effect.

The effect of ultrasonic melt processing time on the microstructures and hardness of as-cast and homogenized Al-Cu-Mg-Ag-Fe-Ni alloys was studied. The results show that ultrasonic melt processing leads to more uniform distribution of Fe/Ni-rich phases. Two-minute ultrasonic processing can result in grain refinement of the alloy. However, ultrasonic processing, if extended to 6 minutes, will result in grain coarsening. It is found that ultrasonic processing promotes the dissolution of Al₂Cu phase and enhances the solid-solution strengthening effect, resulting in improvement in hardness of the as-cast alloy. After homogenization, Fe/Ni-rich phase is detected as the residual secondary phase in the structure, which exhibits discontinuous distribution along the grain boundaries after ultrasonic processing.

Influence of Al content on the microstructure, phases, hardness, tensile and impact properties of quenched and tempered 4Cr5MoVSi die steel was studied by using optical microscopy, transmission electron microscopy (TEM), tensile testing machine among other instruments. It is found martensite is formed in the quenched and tempered 4Cr5MoVSi die steel with Al content of 0-0.6% (mass fraction, the same below), while martensite plus δ ferrite are formed in the quenched and tempered 4Cr5MoVSi die steel with Al content of 1.2%-1.8%. Without adding Al or with an addition of 0.3%-1.8% Al, 4Cr5MoVSi die steel after quench and tempering has α-Fe and M₇C₃ formed as its main phases. The 4Cr5MoVSi die steel with an Al content of 0.3%-1.2% has its relatively small variation in its hardness after quenching and tempering; however, with Al content up to 1.8%, the quenched and tempered die steel has its hardness significantly reduced. It is concluded that an appropriate addition of Al is beneficial to improvement in the room-temperature tensile strength, elongation at break, and impact toughness of quenched and tempered 4Cr5MoVSi die steel; with an addition of 1.2% Al, the quenched and tempered 4Cr5MoVSi die steel can have higher hardness, strength, and impact toughness.

To investigate the rheological properties of heat-treatment-free AlSi9MnMg alloy specifically for high-pressure die casting (HPDC) process, influence of temperature and shear rate on the viscosity of molten alloy was studied by using a rotational rheometer based on the Searle principle, and the microstructure of the samples in rheological testing was analyzed. The results indicate that the viscosity of molten alloy decreases as the temperature rises in the test. At a given temperature, the viscosity decreases as the shear rate increases. With the shear rate exceeding 800 s^-1, the viscosity remains unchanged with shear time at the same shear rate. Under the action of shear force, the dendritic grains in the melt undergo fragmentation, agglomeration and spheroidization. Moreover, as the temperature rises in the test, the particle agglomerates become smaller in size.

Based on the Derwent Innovations Index (DII), the technologies for recycling cathode materials in spent lithium-ion batteries were analyzed in terms of patent indicators, including the number of patent applications and patent holders, the number of priority rights in countries/regions, patent citations, patent H-index among others. It is found that the recycling technologies are in rapid development, and the Chinese research organizations have created a patent portfolio plan, including a large number of high-quality patents that can improve competition. Over the past five years, the technologies, including pretreatment processes of discharge and separation, recovery of lithium, nickel, cobalt and manganese, extraction of precious metals by hydrometallurgical and pyrometallurgical processes, as well as recycling equipment, have become hotpots in this field. China has a relatively large number of core patents, mainly involving hydrometallurgical process, direct recycling process, and a combination of pyrometallurgical and hydrometallurgical processes. Finally, some countermeasures and suggestions are proposed for enhancing the competitiveness of China's industry in recycling cathode materials in spent lithium-ion batteries, which provides references for promoting the healthy development of China's lithium battery recycling industry.

For phosphorus-iron waste left after lithium extraction from cathode powder of spent lithium iron phosphate (LFP) batteries, a selective leaching with sulfuric acid was adopted to remove impurities therein. The results indicate that after a 4-stage leaching at 80 ℃ for 2 h, with a sulfuric acid at a concentration of 0.32 mol/L and a liquid-to-solid ratio of 5 mL/g, the total leaching rates of elemental impurities, including Cu, Mn, Al, Ca, Na and Mg reach 70.81%, 68.33%, 65.57%, 68.03%, 67.85% and 64.28%, respectively, and the total leaching rates of P and Fe are only 2.52% and 2.24%, respectively. It is concluded that the main elemental impurities can be effectively leached out from such phosphorus-iron waste, with elements of P and Fe left for subsequent resource utilization.